6-K

UNITED STATES
SECURITIES AND EXCHANGE COMMISSION
Washington, D.C. 20549

FORM 6-K

REPORT OF FOREIGN PRIVATE ISSUER
PURSUANT TO RULE 13A-16 OR 15D-16 OF
THE SECURITIES EXCHANGE ACT OF 1934

For the month of April 2017

Commission File Number: 001-34244

HUDBAY MINERALS INC.
(Translation of registrant’s name into English)

25 York Street, Suite 800
Toronto, Ontario
M5J 2V5, Canada
(Address of principal executive offices)

Indicate by check mark whether the registrant files or will file annual reports under cover of Form 20-F or Form 40-F.

Form 20-F [  ]           Form 40-F [X]

Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(1): [  ] 

Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(7): [  ] 

Indicate by check mark whether the registrant by furnishing the information contained in this Form is also thereby furnishing the information to the Commission pursuant to Rule 12g3-2(b) under the Securities Exchange Act of 1934.

Yes [  ]              No [X]

If “Yes” is marked, indicate below the file number assigned to the registrant in connection with Rule 12g3-2(b):

EXPLANATORY NOTE

On March 30, 2017, Hudbay Minerals Inc. (“Hudbay”) filed on the Canadian Securities Administrators’ System for Electronic Document Analysis and Retrieval (SEDAR) website at www.sedar.com the following documents: (1) News Release announcing Update on Operations and Growth Projects; (2) Technical Report on the Lalor Mine, Snow Lake, Manitoba, Canada dated effective March 30, 2017; (3) Certificate of Qualified Person - Robert Carter; (4) Consent of Qualified Person - Robert Carter; (5) Technical Report on the Rosemont Project, Pima County, Arizona, USA issued and effective March 30, 2017; (6) Certificate of Qualified Person - Cashel Meagher; (7) Consent of Qualified Person - Cashel Meagher.

Copies of the filings are attached to this Form 6-K and incorporated herein by reference, as follows:

• News Release announcing Update on Operations and Growth Projects
• Technical Report on the Lalor Mine, Snow Lake, Manitoba, Canada dated effective March 30, 2017
• Certificate of Qualified Person - Robert Carter
• Consent of Qualified Person - Robert Carter
• Technical Report on the Rosemont Project, Pima County, Arizona, USA issued and effective March 30, 2017
• Certificate of Qualified Person - Cashel Meagher
• Consent of Qualified Person - Cashel Meagher

SIGNATURE

Pursuant to the requirements of the Securities Exchange Act of 1934, the registrant has duly caused this report to be signed on its behalf by the undersigned, thereunto duly authorized.

HUDBAY MINERALS INC.
(registrant)
By:	/s/ Patrick Donnelly
Name: Patrick Donnelly
Title: Vice President and General Counsel

Date: April 3, 2017

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EXHIBIT INDEX

The following exhibit is furnished as part of this Form 6-K:

Exhibit	Description
99.1	News Release announcing Update on Operations and Growth Projects
99.2	Technical Report on the Lalor Mine, Snow Lake, Manitoba, Canada dated effective March 30, 2017
99.3	Certificate of Qualified Person - Robert Carter
99.4	Consent of Qualified Person - Robert Carter
99.5	Technical Report on the Rosemont Project, Pima County, Arizona, USA issued and effective March 30, 2017
99.6	Certificate of Qualified Person - Cashel Meagher
99.7	Consent of Qualified Person - Cashel Meagher

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EX-99.1

 

	TMX, NYSE – HBM
	2017 No. 4

Hudbay Provides Update on Operations and Growth Projects

Toronto, Ontario, March 30, 2017 – Hudbay Minerals Inc. (“Hudbay” or the “company”) (TSX, NYSE: HBM) is pleased to announce an optimized mine plan for its 100%-owned Lalor mine in Manitoba, Canada as well as a feasibility study for its Rosemont project in Arizona, United States. The company has also provided its annual mineral reserve and resource update for all of its properties. All amounts are in U.S. dollars, unless otherwise noted.

Lalor highlights:

• Updated Lalor mine plan incorporates a throughput rate of 4,500 tonnes per day at the Stall concentrator, an increase from the current 3,000 tonnes per day

• Planned Lalor zinc production increases to 90 thousand tonnes contained in concentrate in 2017, from 71 thousand tonnes in 2016

• Feasibility work is ongoing for the Lalor gold zone and copper-gold zone targeting an additional 1,500 tonnes per day through the New Britannia mill to fully utilize Lalor’s 6,000 tonnes per day shaft capacity

Rosemont highlights:

• Rosemont is expected to have a 19-year mine life and demonstrates robust economics with a projected 15.5% after-tax project IRR on the estimated $1.9 billion project capital cost (100% basis) at a copper price of $3.00 per pound

• Rosemont is expected to have average annual production over the first 10 years of 127 thousand metric tonnes of copper at an average annual cash cost of $1.14 per pound of copper and sustaining cash cost of $1.59 per pound of copper¹

• Development of Rosemont is conditional upon receipt of final permits and the approval of Hudbay's Board of Directors

“Our enhancements to the Lalor mine plan offer low-cost, near term zinc production growth with potential future gold production upside, and positions our Manitoba business unit to be a strong contributor to Hudbay’s results for many years to come,” said Alan Hair, president and chief executive officer. “Our Peru and Manitoba operations are expected to generate strong free cash flow that we can reinvest in the Rosemont project in order to grow the long-term copper production profile of the company. The Rosemont project is expected to be one of the first new copper mines to be built when copper prices improve and, once approved d, has the capacity to generate strong returns for Hudbay shareholders. ”

^{____________________________________________________
1} Cash cost and sustaining cash cost, net of by-product credits per pound of copper are not recognized under IFRS. Rosemont’s byproduct credits are calculated using $11.00 per pound molybdenum and precious metal stream price of $3.90 per ounce silver, subject to 1% annual inflation adjustment after three years. Cash cost include the impact of capitalized stripping. For a detailed description of each of these non-IFRS financial performance measures, please see the discussion under "Non-IFRS Financial Performance Measures" on page 8 of this news release.

 

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National Instrument 43-101 (“NI 43-101”) technical reports in respect of the Lalor mine (“Lalor Technical Report”) and the Rosemont project (“Rosemont Technical Report”) have been filed on SEDAR at www.sedar.com and will be filed on EDGAR at www.sec.gov.

Lalor Mine Plan – Stall Base Metal Mill

Hudbay’s updated mine plan for Lalor enables cost-effective production growth by optimizing the use of existing infrastructure. When full development of Lalor was approved in 2010, the original plan was to build a new concentrator at the mine site to process all of the material from Lalor, while utilizing Hudbay’s nearby Stall concentrator to process base metal zone ore until the new concentrator was completed. Since then, however, the performance of the Stall concentrator has exceeded expectations and, with a modest capital investment, the optimized throughput rate for the Stall mill is expected to be 4,500 tonnes per day on a sustainable basis starting in the third quarter of 2018. In addition, the 2015 acquisition of the New Britannia gold mill in Snow Lake is expected to provide a low-cost solution to process the gold zone and copper-gold zone ore at optimal gold recovery rates, with the potential to augment the base metal production from the Stall mill and utilize the full 6,000 tonnes per day capacity of the Lalor mine shaft.

The updated Lalor mine plan incorporates the increased base metal throughput and includes the processing of the base metal zone, the copper-gold zone and portions of the gold zone when in contact with base metal ore at the Stall base metal mill. Pending the completion of engineering work on the New Britannia mill, the updated Lalor mine plan assumes that copper-gold zone reserves will be mined and processed at the Stall base metal mill at an average rate of 130,000 tonnes per annum between 2020 and 2023. An updated mine plan incorporating the New Britannia mill is expected to enable the copper-gold zone and gold zone material to be processed at New Britannia, given the significantly higher potential gold recoveries at New Britannia. The redirection of copper-gold zone material to New Britannia would also permit accelerated processing of zinc-rich ore and higher zinc production through Stall during those years.

A summary of the updated Lalor mine plan for the Stall mill is shown below.

Lalor Mine Plan Summary – Stall Base Metal Mill		LOM1 Total / Average
Production
Total ore mined	million tonnes	14.2
Peak daily throughput	tonnes per day	4,500
Mine life	years	10.5
Zinc grade	% Zn	5.12%
Copper grade	% Cu	0.69%
Gold grade	g/t Au	2.61
Silver grade	g/t Ag	26.50
Zinc recovery	%	91.8%
Copper recovery	%	86.8%
Gold recovery	%	58.2%
Silver recovery	%	54.0%
Average annual zinc production2	thousand tonnes	63.8
Average annual copper production2	thousand tonnes	8.1
Average annual gold production2	thousand ounces	66.1
Average annual silver production2	thousand ounces	623.3
Mining unit cost3	C$/tonne mined	C$78.32
Milling unit cost3	C$/tonne milled	C$21.51
Cash Cost4
Cash cost	$/lb Zn	$0.37
Sustaining cash cost	$/lb Zn	$0.50
Capital Expenditures
Total project capital (2017 and 2018)5	C$ million	C$117
Total sustaining capital	C$ million	C$220

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¹ Life-of-mine (“LOM”).
² Production refers to contained metal in concentrate.
³ G&A costs related to shared services incurred in Flin Flon and allocated between 777, Reed and Lalor mines are not included in unit costs.
⁴ Cash cost and sustaining cash cost are reported on per pound of zinc contained in concentrate and are net of by-product credits, which are calculated using the following assumptions: copper price per pound - $2.60 in 2017, $2.75 in 2018, $3.00 in 2019 to 2020 and long-term; gold price per ounce - $1,300 in 2017 to 2020 and $1,260 long-term; silver price per ounce - $18.00 in 2017 to 2020 and long-term; CAD/USD exchange rate - 1.35 in 2017, 1.25 in 2018, 1.20 in 2019, 1.15 in 2020 and 1.10 long-term. Cash cost includes on-site and off-site costs, and sustaining cash cost includes the addition of royalties and sustaining capital.
⁵ Includes capital spending for the paste backfill plant, $40 million of which was included in Hudbay’s previously disclosed 2017 annual growth capex guidance.

Construction of a paste backfill plant is expected to be completed in the first quarter of 2018 for a total estimated cost of $50 million (C$68 million), of which $40 million was included in Hudbay’s 2017 growth capital guidance as announced on January 17, 2017. The paste backfill plant is intended to increase mining rates and maximize ore recovery, while reducing capitalized development costs and better maintaining the integrity of ground conditions. Capital spending of $36 million (C$49 million) is planned for refurbishments to the Stall concentrator and underground ore handling to enable sustainable throughput rates of 4,500 tonnes per day, of which $15 million will be spent in 2017 over and above initial 2017 guidance estimates.

Current mineral reserves for Lalor as of January 1, 2017 are summarized below. Reserves include the base metal zone, the copper-gold zone and portions of the gold zone in contact with the base metal ore, which represent approximately 80%, 4% and 16%, respectively, of the total reserve tonnage.

Lalor Mineral Reserve Estimates 1	Tonnes	Zn Grade (%)	Au Grade (g/t)	Cu Grade (%)	Ag Grade (g/t)
Proven	4,383,000	6.76	2.37	0.76	27.33
Probable	9,849,000	4.39	2.72	0.65	26.12
Total proven and probable	14,232,000	5.12	2.61	0.69	26.50

Note: totals may not add up correctly due to rounding.
¹ Mineral reserves calculated using metal prices of $1.07 per pound zinc (includes premium), $1,260 per ounce gold, $3.00 per pound copper and $18.00 per ounce of silver, and using a CAD/USD exchange rate of 1.10.

Current mineral resources, exclusive of reserves, for the Lalor base metal zone as of September 30, 2016 are summarized below.

Lalor Base Metal Zone  Mineral Resource Estimates 1	Tonnes	Zn Grade (%)	Au Grade (g/t)	Cu Grade (%)	Ag Grade (g/t)
Indicated	2,100,000	5.34	1.69	0.49	28.10
Inferred	545,300	8.15	1.45	0.32	22.28

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Note: totals may not add up correctly due to rounding.
¹ Mineral resources calculated using metal prices of $1.19 per pound zinc (includes premium), $1,300 per ounce gold, $18.00 per ounce of silver and $2.67 per pound copper.

Lalor Mine Plan – New Britannia Gold Mill

Hudbay’s current expectation is that a portion of the material mined from the gold zone and copper-gold zone at Lalor will be processed through the New Britannia mill at a rate of up to 1,500 tonnes per day starting in 2019. When combined with processing capacity at the Stall base metal mill, this is expected to achieve an aggregate throughput rate of up to 6,000 tonnes per day. The New Britannia mill is expected to achieve significantly higher gold recoveries than the Stall mill, and Hudbay is examining the potential to install a copper pre-float facility in the mill to maximize copper recoveries. Work on gold zone production and the New Britannia refurbishment is ongoing, and the focus of the remaining engineering work is on finalizing the New Britannia flowsheet and optimizing the utilization of the existing tailings management facilities.

Current mineral resources, exclusive of reserves, for the Lalor gold zone and copper-gold zone as of September 30, 2016 are summarized below.

Lalor Gold and Copper-Gold Mineral Resource Estimates 1	Tonnes	Zn Grade (%)	Au Grade (g/t)	Cu Grade (%)	Ag Grade (g/t)
Indicated	1,750,000	0.40	5.18	0.34	30.61
Inferred	4,124,000	0.31	5.02	0.90	27.61

Note: totals may not add up correctly due to rounding.
¹ Mineral resources calculated using metal prices of $1.19 per pound zinc (includes premium), $1,300 per ounce gold, $18.00 per ounce of silver and $2.67 per pound copper.

Rosemont Feasibility Study

Since the acquisition of the Rosemont project, Hudbay has completed an extensive work program, including in-fill drilling, detailed metallurgical test work, and a bottom-up approach to cost estimation, along with other feasibility-level work, as summarized in the Rosemont Technical Report.

The Rosemont project will be a traditional open pit, shovel and truck operation with an expected 19-year mine life. Rosemont is expected to generate an after-tax, unlevered internal rate of return of 15.5%, using a long-term copper price of $3.00 per pound of copper.

A summary of the Rosemont mine plan is shown below. References to “tons” refer to short tons, not metric tonnes, except where noted.

		Years 1-10	LOM	LOM
Rosemont Feasibility Study Summary		Average	Average1	Total1
Production
Ore mined	million tons	37	31	592
Waste mined2	million tons	95	61	1,155
Strip ratio2	waste:ore	2.5	2.0	2.0
Ore milled	million tons	32	31	592
Copper grade milled3	% TCu	0.53%	0.45%	0.45%
Copper recovery	%	82%	80%	80%
Copper production4	thousand tons	140	112	2,129
Copper production 4	thousand metric tonnes	127	102	1,932
Total on-site unit costs5	$/ton milled	$8.01	$7.92	$7.92
Cash Cost6
Cash cost	$/lb Cu	$1.14	$1.29	$1.29
Sustaining cash cost	$/lb Cu	$1.59	$1.65	$1.65
Capital Expenditures
Development capital	$ million	-	-	$1,921
Sustaining capital	$ million	$29	$20	$387
Capitalized stripping	$ million	$71	$41	$781
Total sustaining capital
(including capitalized stripping)	$ million	$100	$61	$1,168
Economics – Project Basis (100%)7
Net present value at 8%	$ million	-	-	$769
Net present value at 10%	$ million	-	-	$496
After-tax internal rate of return	%	-	-	15.5%
Payback period	years	-	-	5.2
Economics – Hudbay Basis (80%)7
Net present value at 8%	$ million	-	-	$719
Net present value at 10%	$ million	-	-	$499
After-tax internal rate of return	%	-	-	17.7%
Payback period	years	-	-	4.9

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¹ Life-of-mine (“LOM”) average and total calculated over years 1 to 19.
² Waste and strip ratio excludes pre-stripping tons.
³ Production refers to contained metal in concentrate.
⁴ Total copper grade includes both the sulfide and acid-soluble copper in the ore.
⁵ On-site unit costs include mining, milling, G&A, reclamation and severance tax costs, and are after deducting capitalized stripping.
⁶ Cash cost and sustaining cash cost are reported net of by-product credits, which are calculated using $11.00 per pound molybdenum and precious metal streaming prices of $3.90 per ounce silver and $450 per ounce gold, and include the impact of capitalized stripping. Cash cost includes on-site and off-site costs, and sustaining cash cost includes the addition of royalties and sustaining capital.
⁷ Economic analysis assumes $3.00 per pound copper, $11.00 per pound molybdenum, and precious metal streaming price of $3.90 per ounce silver, subject to 1% annual inflation adjustment after three years. Hudbay basis adjusts for joint venture partner expected payments to earn into their minority interest and outstanding joint venture loan owed to Hudbay.

The Rosemont project capital cost estimate of $1,921 million (100% basis) was developed based on realistic assumptions and informed by Hudbay’s recent experience in successfully building and ramping up the Constancia copper mine in Peru. The cost estimate is based on mid-cycle cost expectations, and does not factor in potential savings that could be available if construction begins during the current period of muted new mine construction activity.

Total project development capital is expected to be spent over an approximate three year construction period and will be offset by existing funding sources. The precious metals stream agreement with Silver Wheaton (Caymans) Ltd. provides for a payment of a $230 million deposit upon achievement of certain milestones. Up to $200 million in mobile equipment included in the Rosemont project cost estimate is expected to be financed using conventional equipment financing. In addition, the joint venture agreement with a Korean consortium contemplates $106 million in cash payments from the joint venture partners in order to complete their earn-in for 20% of the project and contributions of 20% of the remaining funding required net of the precious metals stream, equipment financing and joint venture earn-in proceeds. Combined, these funding sources are expected to provide approximately $800 million in project-level funding. Hudbay expects that a significant portion of the remaining required funding can be met through free cash flow generation from the company’s operations in Peru and Manitoba.

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	2017 No. 4

The permitting process for Rosemont is well advanced and continues to progress. The key federal permits outstanding are the Final Record of Decision from the U.S. Forest Service and the Section 404 Water Permit from the U.S. Army Corps of Engineers. These federal permits are currently in the final stages of the review process. All State of Arizona permits and approvals have been issued for Rosemont and remain in force and are current. The project design included in the Rosemont Technical Report is specifically intended to meet the impacts analyzed and commitments outlined in the federal and state permits.

Current mineral reserves and resources for Rosemont as of March 30, 2017 are summarized below.

Rosemont Project Mineral Reserve and Resource Estimates	Short Tons	Cu Grade1 (%)	Mo Grade (%)	Ag Grade (oz/T)
Mineral Reserves2
Proven	469,708,117	0.48%	0.012%	0.14
Probable	122,324,813	0.31%	0.010%	0.09
Total proven and probable	592,032,930	0.45%	0.012%	0.13
Mineral Resources3
Measured	177,700,000	0.38%	0.01%	0.079
Indicated	413,200,000	0.25%	0.01%	0.076
Total measured and indicated	591,000,000	0.29%	0.01%	0.077
Inferred	68,700,000	0.30%	0.01%	0.046

Note: totals may not add up correctly due to rounding.
¹ Total copper grade includes both the sulfide and acid-soluble copper in the ore.
² Mineral reserves calculated using metal prices of $3.15 per pound copper, $11.00 per pound molybdenum and $18.00 per ounce silver.
³ Mineral resources are exclusive of mineral reserves. Mineral resources include oxide, mix and hypogene resources. Mineral resources calculated using metal prices of $3.15 per pound copper, $11.00 per pound molybdenum and $18.00 per ounce of silver.

Constancia Mine

Current mineral reserves and resources for Constancia and Pampacancha as of January 1, 2017 are summarized below.

Constancia Mine Mineral Reserve and Resource Estimates	Tonnes	Cu Grade (%)	Mo Grade (g/t)	Au Grade (g/t)	Ag Grade (g/t)
Constancia Reserves1
Proven	431,300,000	0.30	95	0.037	2.88
Probable	109,900,000	0.23	62	0.034	2.55
Total proven and probable - Constancia	541,200,000	0.28	88	0.037	2.81
Pampacancha Reserves1
Proven	22,800,000	0.53	149	0.299	4.44
Probable	20,200,000	0.44	164	0.250	3.85
Total proven and probable - Pampacancha	43,000,000	0.49	156	0.276	4.17
Total proven and probable	584,200,000	0.30	93	0.054	2.91
Constancia Resources2
Measured	161,800,000	0.19	55	0.031	2.26
Indicated	287,800,000	0.17	50	0.026	1.89
Total measured and indicated - Constancia	449,600,000	0.18	52	0.028	2.02
Inferred	138,100,000	0.17	40	0.018	1.70
Pampacancha Resources2
Measured	7,500,000	0.35	57	0.235	4.13
Indicated	15,200,000	0.18	90	0.180	2.85
Total measured and indicated - Pampacancha	22,700,000	0.23	79	0.198	3.27
Total measured and indicated	472,300,000	0.18	53	0.036	2.08

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Note: totals may not add up correctly due to rounding.
¹ Mineral reserves calculated using metal prices of $3.00 per pound copper, $11.00 per pound molybdenum, $18.00 per ounce silver, and $1,260 per ounce gold.
² Mineral resources are exclusive of mineral reserves. Mineral resources calculated using metal prices of $3.00 per pound copper, $11.00 per pound molybdenum, $18.00 per ounce silver, and $1,260 per ounce gold.

777 and Reed Mines

Current mineral reserves and resources for 777 and Reed as of January 1, 2017 are summarized below.

777 Mine  Mineral Reserve and Resource Estimates 1	Tonnes	Cu Grade (%)	Zn Grade (%)	Au Grade (g/t)	Ag Grade (g/t)
Mineral Reserves
Proven	3,080,000	1.98	4.93	2.01	31.53
Probable	1,386,000	1.16	5.09	2.04	30.96
Total proven and probable	4,466,000	1.73	4.98	2.02	31.35
Mineral Resources2
Indicated	736,000	0.99	3.53	1.82	26.24
Inferred	673,000	1.01	4.26	1.72	30.95

Note: totals may not add up correctly due to rounding.
¹ Mineral reserves and resources calculated using metal prices of $2.67 per pound copper, $1.24 per pound zinc (includes premium), $1,300 per ounce gold, and $18.00 per ounce silver, and using a CAD/USD exchange rate of 1.25.
² Mineral resources are exclusive of mineral reserves.

Reed Mine Mineral Reserve and Resource Estimates 1	Tonnes	Cu Grade (%)	Zn Grade (%)	Au Grade (g/t)	Ag Grade (g/t)
Mineral Reserves
Proven	362,000	3.35	0.68	0.39	5.35
Probable	337,000	3.95	0.31	0.52	5.26
Total proven and probable	699,000	3.64	0.50	0.45	5.30
Mineral Resources2
Inferred	88,000	3.13	0.42	0.79	6.00

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Note: totals may not add up correctly due to rounding.
¹ Mineral reserves calculated using metal prices of $2.50 per pound copper, $1.22 per pound zinc (includes premium), $1,300 per ounce gold, and $18.00 per ounce silver, and using a CAD/USD exchange rate of 1.28.
² Mineral resources are exclusive of mineral reserves and are calculated using metal prices of $2.67 per pound copper, $1.24 per pound zinc (includes premium), $1,300 per ounce gold, and $18.00 per ounce silver, and using a CAD/USD exchange rate of 1.25.

The focus for the 777 and Reed mines is maximizing value as the mines approach the end of their lives. Hudbay has re-sequenced the 777 mine plan to prioritize stopes containing higher zinc grades in order to take advantage of favourable expected zinc prices.

Additional detail on the Constancia, Lalor, Rosemont, 777 and Reed properties, including a year-over-year reconciliation of reserves and resources, is included in Hudbay’s Annual Information Form for the year ended December 31, 2016, which is available on SEDAR at www.sedar.com and will be filed on EDGAR at www.sec.gov.

Non-IFRS Financial Performance Measures

Cash cost and sustaining cash cost per pound of copper produced are shown because the company believes they help investors and management assess the performance of its operations, including the margin generated by the operations and the company. These measures do not have a meaning prescribed by IFRS and are therefore unlikely to be comparable to similar measures presented by other issuers. These measures should not be considered in isolation or as a substitute for measures prepared in accordance with IFRS and are not necessarily indicative of operating profit or cash flow from operations as determined under IFRS. Other companies may calculate these measures differently. For further details on how Hudbay calculates these measures in respect of its operating assets, please refer to page 40 of Hudbay’s management’s discussion and analysis for the three months and year ended December 31, 2016 available on SEDAR at www.sedar.com and EDGAR at www.sec.gov.

Qualified Person

The technical and scientific information in this news release has been approved by Cashel Meagher, P. Geo, Hudbay’s Senior Vice President and Chief Operating Officer, and Robert Carter, P. Eng., Hudbay’s Lalor Mine Manager. Messrs. Meagher and Carter are qualified persons pursuant to NI 43-101. For a description of the key assumptions, parameters and methods used to estimate mineral reserves and resources, as well as data verification procedures and a general discussion of the extent to which the estimates of scientific and technical information may be affected by any known environmental, permitting, legal title, taxation, sociopolitical, marketing or other relevant factors, please refer to the NI 43-101 technical reports as filed by Hudbay on SEDAR at www.sedar.com.

Forward-Looking Information

This news release contains forward-looking information within the meaning of applicable Canadian and United States securities legislation. All information contained in this news release, other than statements of current and historical fact, is forward-looking information. Often, but not always, forward-looking information can be identified by the use of words such as “plans”, “expects”, “budget”, “guidance”, “scheduled”, “estimates”, “forecasts”, “strategy”, “target”, “intends”, “objective”, “goal”, “understands”, “anticipates” and “believes” (and variations of these or similar words) and statements that certain actions, events or results “may”, “could”, “would”, “should”, “might” “occur” or “be achieved” or “will be taken” (and variations of these or similar expressions). All of the forward-looking information in this news release is qualified by this cautionary note.

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Forward-looking information includes, but is not limited to, production, cost and capital and exploration expenditure guidance, including anticipated capital and operating cost savings and anticipated production at the company’s mines and processing facilities, the anticipated timing, cost and benefits of developing the Pampacancha deposit and Lalor paste backfill plant, anticipated mine plans, anticipated metals prices and the anticipated sensitivity of the company’s financial performance to metals prices, events that may affect its operations and development projects, the permitting, development and financing of the Rosemont project, the potential to increase throughput at the Stall mill and to refurbish the New Britannia mill and utilize it to process ore from the Lalor mine, anticipated cash flows from operations and related liquidity requirements, the potential outcome of labour negotiations in Peru, the anticipated effect of external factors on revenue, such as commodity prices, economic outlook, government regulation of mining operations, and business and acquisition strategies. Forward-looking information is not, and cannot be, a guarantee of future results or events. Forward-looking information is based on, among other things, opinions, assumptions, estimates and analyses that, while considered reasonable by Hudbay at the date the forward-looking information is provided, inherently are subject to significant risks, uncertainties, contingencies and other factors that may cause actual results and events to be materially different from those expressed or implied by the forward-looking information.

The material factors or assumptions that Hudbay identified and were applied by the company in drawing conclusions or making forecasts or projections set out in the forward-looking information include, but are not limited to:

•   the success of mining, processing, exploration and development activities;

•   the scheduled maintenance and availability of Hudbay’s processing facilities;

•   the sustainability and success of Hudbay’s cost reduction initiatives;

•   the accuracy of geological, mining and metallurgical estimates;

•   anticipated metals prices and the costs of production;

•   the supply and demand for metals that Hudbay produces;

•   the supply and availability of all forms of energy and fuels at reasonable prices;

•   no significant unanticipated operational or technical difficulties;

• the execution of Hudbay’s business and growth strategies, including the success of its strategic investments and initiatives;

•   the availability of additional financing, if needed;

• the ability to complete project targets on time and on budget and other events that may affect Hudbay’s ability to develop its projects;

•   the timing and receipt of various regulatory and governmental approvals;

• the availability of personnel for Hudbay’s exploration, development and operational projects and ongoing employee and union relations;

•   the ability to secure required land rights to develop the Pampacancha deposit;

• maintaining good relations with the communities in which Hudbay operates, including the communities surrounding its Constancia mine and Rosemont project and First Nations communities surrounding its Lalor and Reed mines;

•   no significant unanticipated challenges with stakeholders at Hudbay’s various projects;

• no significant unanticipated events or changes relating to regulatory, environmental, health and safety matters;

• no contests over title to Hudbay’s properties, including as a result of rights or claimed rights of aboriginal peoples;

•   the timing and possible outcome of pending litigation and no significant unanticipated litigation;

• certain tax matters, including, but not limited to current tax laws and regulations and the refund of certain value added taxes from the Canadian and Peruvian governments; and

• no significant and continuing adverse changes in general economic conditions or conditions in the financial markets (including commodity prices and foreign exchange rates).

9

	TMX, NYSE – HBM
	2017 No. 4

The risks, uncertainties, contingencies and other factors that may cause actual results to differ materially from those expressed or implied by the forward-looking information may include, but are not limited to, risks generally associated with the mining industry, such as economic factors (including future commodity prices, currency fluctuations, energy prices and general cost escalation), uncertainties related to the development and operation of Hudbay’s projects (including risks associated with the permitting, development and economics of the Rosemont project and related legal challenges), risks related to the maturing nature of the 777 and Reed mines and their impact on the related Flin Flon metallurgical complex, dependence on key personnel and employee and union relations, risks related to the schedule for mining the Pampacancha deposit (including the timing and cost of acquiring the required surface rights), risks related to the cost, schedule, permitting and economics of the capital projects intended to increase processing capacity for Lalor ore, risks related to political or social unrest or change, risks in respect of aboriginal and community relations, rights and title claims, operational risks and hazards, including unanticipated environmental, industrial and geological events and developments and the inability to insure against all risks, failure of plant, equipment, processes, transportation and other infrastructure to operate as anticipated, compliance with government and environmental regulations, including permitting requirements and anti-bribery legislation, depletion of Hudbay’s reserves, volatile financial markets that may affect Hudbay’s ability to obtain additional financing on acceptable terms, the failure to obtain required approvals or clearances from government authorities on a timely basis, uncertainties related to the geology, continuity, grade and estimates of mineral reserves and resources, and the potential for variations in grade and recovery rates, uncertain costs of reclamation activities, the company’s ability to comply with its pension and other post-retirement obligations, Hudbay’s ability to abide by the covenants in its debt instruments and other material contracts, tax refunds, hedging transactions, as well as the risks discussed under the heading “Risk Factors” in the company’s most recent Annual Information Form.

Should one or more risk, uncertainty, contingency or other factor materialize or should any factor or assumption prove incorrect, actual results could vary materially from those expressed or implied in the forward-looking information. Accordingly, you should not place undue reliance on forward-looking information. Hudbay does not assume any obligation to update or revise any forward-looking information after the date of this news release or to explain any material difference between subsequent actual events and any forward-looking information, except as required by applicable law.

Note to United States Investors

This news release has been prepared in accordance with the requirements of the securities laws in effect in Canada, which may differ materially from the requirements of United States securities laws applicable to U.S. issuers.

Information concerning Hudbay’s mineral properties has been prepared in accordance with the requirements of Canadian securities laws, which differ in material respects from the requirements of the Securities and Exchange Commission (the “SEC”) set forth in Industry Guide 7. Under the SEC’s Industry Guide 7, mineralization may not be classified as a “reserve” unless the determination has been made that the mineralization could be economically and legally produced or extracted at the time of the reserve determination, and the SEC does not recognize the reporting of mineral deposits which do not meet the SEC Industry Guide 7 definition of “Reserve”.

10

	TMX, NYSE – HBM
	2017 No. 4

In accordance with National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”) of the Canadian Securities Administrators, the terms “mineral reserve”, “proven mineral reserve”, “probable mineral reserve”, “mineral resource”, “measured mineral resource”, “indicated mineral resource” and “inferred mineral resource” are defined in the Canadian Institute of Mining, Metallurgy and Petroleum (the “CIM”) Definition Standards for Mineral Resources and Mineral Reserves adopted by the CIM Council on May 10, 2014. While the terms “mineral resource”, “measured mineral resource”, “indicated mineral resource” and “inferred mineral resource” are recognized and required by NI 43-101, the SEC does not recognize them. You are cautioned that, except for that portion of mineral resources classified as mineral reserves, mineral resources do not have demonstrated economic value. Inferred mineral resources have a high degree of uncertainty as to their existence and as to whether they can be economically or legally mined. It cannot be assumed that all or any part of an inferred mineral resource will ever be upgraded to a higher category. Therefore, you are cautioned not to assume that all or any part of an inferred mineral resource exists, that it can be economically or legally mined, or that it will ever be upgraded to a higher category. Likewise, you are cautioned not to assume that all or any part of measured or indicated mineral resources will ever be upgraded into mineral reserves.

11

	TMX, NYSE – HBM
	2017 No. 4

About Hudbay

Hudbay (TSX, NYSE: HBM) is an integrated mining company producing copper concentrate (containing copper, gold and silver) and zinc metal. With assets in North and South America, the company is focused on the discovery, production and marketing of base and precious metals. Directly and through its subsidiaries, Hudbay owns four polymetallic mines, four ore concentrators and a zinc production facility in northern Manitoba and Saskatchewan (Canada) and Cusco (Peru), and a copper project in Arizona (United States). The company’s growth strategy is focused on the exploration and development of properties it already controls, as well as other mineral assets it may acquire that fit its strategic criteria. Hudbay’s vision is to become a top-tier operator of long-life, low-cost mines in the Americas. Hudbay’s mission is to create sustainable value through the acquisition, development and operation of high-quality and growing long-life deposits in mining-friendly jurisdictions. The company is governed by the Canada Business Corporations Act and its shares are listed under the symbol "HBM" on the Toronto Stock Exchange, New York Stock Exchange and Bolsa de Valores de Lima. Hudbay also has warrants listed under the symbol “HBM.WT” on the Toronto Stock Exchange and “HBM/WS” on the New York Stock Exchange. Further information about Hudbay can be found on www.hudbay.com.

For further information, please contact:

Candace Brûlé

Director, Investor Relations

(416) 814-4387

candace.brule@hudbay.com

12

EX-99.2

NI 43-101 Technical Report

Lalor Mine

Snow Lake, Manitoba, Canada

Dated Effective March 30, 2017

 

25 York Street, Suite 800

Toronto, Ontario

Canada M5J 2V5

Prepared by:

Robert Carter, P.Eng.

Lalor Mine Manager, Hudbay Manitoba Business Unit

	Lalor Mine
	Form 43-101F1 Technical Report

CAUTIONARY NOTE REGARDING FORWARD-LOOKING INFORMATION

This Technical Report contains "forward-looking statements" and "forward-looking information" (collectively, "forward-looking information") within the meaning of applicable Canadian and United States securities legislation. All information contained in this Technical Report, other than statements of current and historical fact, is forward-looking information. Often, but not always, forward-looking information can be identified by the use of words such as “plans”, “expects”, “budget”, “guidance”, “scheduled”, “estimates”, “forecasts”, “strategy”, “target”, “intends”, “objective”, “goal”, “understands”, “anticipates” and “believes” (and variations of these or similar words) and statements that certain actions, events or results “may”, “could”, “would”, “should”, “might” “occur” or “be achieved” or “will be taken” (and variations of these or similar expressions). All of the forward-looking information in this Technical Report is qualified by this cautionary note.

Forward-looking information includes, but is not limited to, our objectives, strategies, intentions, expectations, production, cost, capital and exploration expenditure guidance, including the anticipated capital and operating cost savings and anticipated production at our mines and processing facilities, events that may affect Hudbay’s operations and development projects, the anticipated timing, cost and benefits of developing the Lalor growth projects, anticipated mine plans, anticipated metals prices and the anticipated sensitivity of our financial performance to metals prices, the potential to increase throughput at the Stall mill and to refurbish the New Britannia mill and utilize it to process ore from the Lalor mine, anticipated cash flows from operations and related liquidity requirements, the anticipated effect of external factors on revenue, such as commodity prices, estimation of mineral reserves and resources, mine life projections, reclamation costs, economic outlook, government regulation of mining operations, and expectations regarding business and acquisition strategies. Forward-looking information is not, and cannot be, a guarantee of future results or events. Forward-looking information is based on, among other things, opinions, assumptions, estimates and analyses that, while considered reasonable by us at the date the forward-looking information is provided, inherently are subject to significant risks, uncertainties, contingencies and other factors that may cause actual results and events to be materially different from those expressed or implied by the forward-looking information.

The material factors or assumptions that we identified and were applied by us in drawing conclusions or making forecasts or projections set out in the forward-looking information include, but are not limited to:

• the success of mining, processing, exploration and development activities;

• the accuracy of geological, mining and metallurgical estimates;

• anticipated metals prices and the costs of production;

• the supply and demand for metals we produce;

• the supply and availability of concentrate for our processing facilities;

• the supply and availability of third party processing facilities for our concentrate;

• the supply and availability of all forms of energy and fuels at reasonable prices;

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	Lalor Mine
	Form 43-101F1 Technical Report

• the availability of transportation services at reasonable prices;

• no significant unanticipated operational or technical difficulties;

• the execution of our business and growth strategies, including the success of our strategic investments and initiatives;

• the availability of additional financing, if needed;

• the ability to complete project targets on time and on budget and other events that may affect our ability to develop our projects;

• the timing and receipt of various regulatory and governmental approvals;

• the availability of personnel for our exploration, development and operational projects and ongoing employee and union relations;

• maintaining good relations with the communities in which we operate, including First Nations communities surrounding our Lalor mine;

• no significant unanticipated challenges with stakeholders at our various projects;

• no significant unanticipated events or changes relating to regulatory, environmental, health and safety matters;

• no contests over title to our properties, including as a result of rights or claimed rights of aboriginal peoples;

• the timing and possible outcome of pending litigation and no significant unanticipated litigation;

• certain tax matters, including, but not limited to current tax laws and regulations; and

• no significant and continuing adverse changes in general economic conditions or conditions in the financial markets (including commodity prices and foreign exchange rates).

The risks, uncertainties, contingencies and other factors that may cause actual results to differ materially from those expressed or implied by the forward-looking information may include, but are not limited to, risks generally associated with the mining industry, such as economic factors (including future commodity prices, currency fluctuations, energy prices and general cost escalation), uncertainties related to the development and operation of our projects, dependence on key personnel and employee and union relations, risks related to the cost, schedule and economics of the capital projects intended to increase processing capacity for Lalor ore, risks related to political or social unrest or change, risks in respect of aboriginal and community relations, rights and title claims, operational risks and hazards, including unanticipated environmental, industrial and geological events and developments and the inability to insure against all risks, failure of plant, equipment, processes, transportation and other infrastructure to operate as anticipated, compliance with government and environmental regulations, including permitting requirements and anti-bribery legislation, depletion of Hudbay’s reserves, volatile financial markets that may affect our ability to obtain additional financing on acceptable terms, the failure to obtain required approvals or clearances from government authorities on a timely basis, uncertainties related to the geology, continuity, grade and estimates of mineral reserves and resources, and the potential for variations in grade and recovery rates, uncertain costs of reclamation activities, Hudbay’s ability to comply with its pension and other post-retirement obligations, our ability to abide by the covenants in our debt instruments and other material contracts, tax refunds, hedging transactions, as well as the risks discussed under the heading “Risk Factors” in our most recent Annual Information Form and our management’s discussion and analysis of Hudbay for the year ended December 31, 2016.

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	Lalor Mine
	Form 43-101F1 Technical Report

Should one or more risk, uncertainty, contingency or other factor materialize or should any factor or assumption prove incorrect, actual results could vary materially from those expressed or implied in the forward-looking information. Accordingly, you should not place undue reliance on forward-looking information. We do not assume any obligation to update or revise any forward-looking information after the date of this Technical Report or to explain any material difference between subsequent actual events and any forward-looking information, except as required by applicable law.

Page iii

	Lalor Mine
	Form 43-101F1 Technical Report

TABLE OF CONTENTS

SECTION			PAGE
TABLE OF CONTENTS			iv
LIST OF TABLES			viii
LIST OF FIGURES			xiii
1	SUMMARY		1-1
	1.1	Summary	1-1
	1.2	Property Description and Location	1-2
	1.3	Geological Setting and Mineralization	1-2
	1.4	Exploration	1-4
	1.5	Drilling	1-5
	1.6	Sample Preparation, Analyses and Security	1-6
	1.7	Data Validation	1-9
	1.8	Mineral Processing and Metallurgical Testing	1-9
	1.9	Mineral Resource Estimates	1-10
	1.10	Mineral Reserve Estimates	1-14
	1.11	Mining Methods	1-17
	1.12	Recovery Methods	1-24
	1.13	Project Infrastructure	1-25
	1.14	Market Studies and Contracts	1-28
	1.15	Environmental Studies, Permitting and Social or Community Impact	1-30
	1.16	Capital and Operating Cost	1-32
	1.17	Economic Analysis	1-34
	1.18	Other Relevant Data and Information	1-34
	1.19	Conclusions and Recommendations	1-35
2	INTRODUCTION AND TERMS OF REFERENCE		2-1
	2.1	Qualified Person (QP) and Site Visit	2-2
	2.2	Sources of Information	2-2
	2.3	Unit Abbreviations	2-4
	2.4	Acronyms and Abbreviations	2-1
3	RELIANCE ON OTHER EXPERTS		3-1
4	PROPERTY DESCRIPTION AND LOCATION		4-1
	4.1	Land Tenure	4-1
	4.2	Land Use Permitting	4-5
5	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		5-1
	5.1	Accessibility	5-1
	5.2	Climate	5-1
	5.3	Local Resources	5-3
	5.4	Infrastructure	5-3
	5.5	Physiography	5-4
6	HISTORY		6-1
	6.1	Exploration in the Chisel Basin Area	6-1

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	Lalor Mine
	Form 43-101F1 Technical Report

	6.2	Historical Mining in the Snow Lake Area	6-1
	6.3	Lalor Mine Production	6-1
7	GEOLOGICAL SETTING AND MINERALIZATION		7-3
	7.1	Regional Geology	7-3
	7.2	Property Geology	7-5
	7.3	Base Metal Mineralization	7-8
	7.4	Gold Mineralization	7-9
8	DEPOSIT TYPE		8-1
9	EXPLORATION		9-1
	9.1	Underground Exploration	9-2
	9.2	Borehole Electromagnetic (EM) Surveys	9-2
	9.3	Surface Electromagnetic (EM) Surveys	9-2
	9.4	Airborne Electromagnetic (EM) Survey	9-3
10	DRILLING		10-1
	10.1	Surveying of Property Grid and Drill Hole Collars	10-3
	10.2	Downhole Surveying	10-4
11	SAMPLING PREPARATION, ANALYSES, AND SECURITY		11-1
	11.1	Laboratory/Laboratories Used	11-1
	11.2	Sample Collection	11-1
	11.3	Sample Preparation	11-2
	11.4	Assay Methodology	11-3
	11.5	Assay Certificates	11-8
	11.6	Security	11-8
12	DATA VERIFICATION		12-1
	12.1	Bureau Veritas Assay Methods and QAQC	12-1
	12.2	Hudbay Laboratory Methods and QAQC	12-9
	12.3	Check Assaying	12-17
	12.4	Site Visit	12-20
	12.5	Core Review	12-20
	12.6	Drilling Database	12-21
	12.7	Mineral Resource Database Management	12-22
13	MINERAL PROCESSING AND METALLURGICAL TESTING		13-1
	13.1	Summary	13-1
	13.2	Plant Metallurgical Performance	13-1
	13.3	Metallurgical Testing	13-3
14	MINERAL RESOURCE ESTIMATES		14-1
	14.1	Key Assumptions of Model	14-1
	14.2	Wireframe Models and Mineralization	14-1
	14.3	Density Assignation	14-4
	14.4	Exploratory Data Analysis	14-5
	14.5	Estimation and Interpolation Methods	14-25
	14.6	Block Model Validation	14-25
	14.7	Classification of Mineral Resource in Base Metal Lenses	14-49
	14.8	Classification of Mineral Resource in Gold Zones	14-51
	14.9	Third Party Review	14-54

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	Lalor Mine
	Form 43-101F1 Technical Report

	14.10	Reasonable Prospects of Economic Extraction	14-55
	14.11	Mineral Resource Statement	14-57
	14.12	Mine Reconciliation of Block Model	14-60
	14.13	Factors That May Affect the Mineral Resource Estimate	14-61
	14.14	Conclusions	14-61
	14.15	Recommendations	14-62
15	MINERAL RESERVE ESTIMATES		15-1
	15.1	Summary	15-1
	15.2	Dilution and Recovery	15-1
	15.3	Conversion of Mineral Resources to Mineral Reserves	15-4
16	MINING METHODS		16-1
	16.1	Introduction	16-1
	16.2	Lateral Development	16-1
	16.3	Vertical Development	16-2
	16.4	Stope Mining	16-2
	16.5	Backfill	16-9
	16.6	Ore Handling	16-10
	16.7	Surface Infrastructure	16-10
	16.8	Geotechnical Design	16-12
	16.9	Support Systems	16-14
	16.10	Underground Development	16-16
	16.11	Diamond Drilling	16-18
	16.12	Drainage System	16-18
	16.13	Mining Operations	16-18
	16.14	Workforce	16-26
	16.15	Mine Safety and Health	16-27
	16.16	MINING METHOD OPPORTUNITIES	16-28
17	RECOVERY METHODS		17-1
	17.1	Introduction	17-1
	17.2	Stall Concentrator	17-1
18	PROJECT INFRASTRUCTURE		18-1
	18.1	Lalor Infrastructure	18-1
	18.2	Lalor Ore Handling Improvements	18-3
	18.3	Paste Plant	18-3
	18.4	Stall Concentrator	18-4
	18.5	Stall Concentrator Expansion	18-5
	18.6	Anderson TIA	18-6
19	MARKET STUDIES AND CONTRACTS		19-1
20	ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT		20-1
	20.1	Environmental Studies and Planning	20-1
	20.2	Waste, Tailings Disposal and Water Management	20-1
	20.3	Permitting Requirements	20-2
	20.4	Mineral Lease and Surface Lease	20-4
	20.5	Community Support	20-4
	20.6	Aboriginal People and First Nations	20-5
	20.7	Heritage Resources	20-5
	20.8	Mine Closure Requirements and Plans	20-5

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	Lalor Mine
	Form 43-101F1 Technical Report

21	CAPITAL AND OPERATING COSTS		21-1
	21.1	Introduction	21-1
	21.2	Capital Costs	21-1
	21.3	Operating Costs	21-2
22	ECONOMIC ANALYSIS		22-1
23	ADJACENT PROPERTIES		23-1
24	OTHER RELEVANT DATA AND INFORMATION		24-1
	24.1	Gold Bulk Sample Program	24-1
	24.2	Taxes and Royalties	24-2
25	INTERPRETATION AND CONCLUSIONS		25-1
26	RECOMMENDATIONS		26-1
27	REFERENCES		27-1
28	SIGNATURE PAGE		28-1
29	CERTIFICATES OF QUALIFIED PERSONS		29-1

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	Lalor Mine
	Form 43-101F1 Technical Report

LIST OF TABLES

TITLE		PAGE
Table 1-1:	Summary Surface Diamond Drill Holes with Assay Results as of January 1, 2017	1-6
Table 1-2:	Expected LOM Recoveries at Stall Concentrator	1-10
Table 1-3:	Base Metal Mineral Resource, Inclusive of Mineral Reserves by Category and Mineralized Zone with a Cut-Off of 4.1% Zn Eq, as of September 30, 2016 (1)(2)(3)(4)(5)(6)(7)(8)(9)	1-12
Table 1-4:	Gold Mineral Resource, Inclusive of Mineral Reserves by Category and Mineralized Zone with a Cut-Off of 2.4 g/t Au Eq, as of September 30, 2016 (1)(2)(3)(4)(5)(6)(7)(8)(9)	1-12
Table 1-5:	Mined Out Areas from the Block Model	1-13
Table 1-6:	Ore Received by Year at the Stall Concentrator	1-13
Table 1-7:	Stoping Parameters by Mining Method	1-14
Table 1-8:	Summary of Mineral Reserves as of January 1, 2017	1-15
Table 1-9:	Base Metal Indicated Resource, Exclusive of Reserves, with a Cut-Off of 4.1% Zn Eq, as of September 30, 2016 (1)	1-16
Table 1-10:	Gold Indicated Resource, Exclusive of Reserves, with a Cut-Off of 2.4 g/t Au Eq, as of September 30, 2016 (1)	1-17
Table 1-11:	Mine Equipment	1-20
Table 1-12:	LOM Production Schedule	1-21
Table 1-13:	LOM Concentrate Production by Year	1-21
Table 1-14:	LOM Contained Metal In concentrate	1-22
Table 1-15:	Mine Operations Workforce	1-23
Table 1-16:	Key long-term zinc metal and zinc concentrate assumptions	1-29
Table 1-17:	Key long-term copper concentrate assumptions	1-29
Table 1-18:	Development Capital Cost Summary	1-32
Table 1-19:	Sustaining Capital Cost Summary	1-32
Table 1-20:	Unit Operating Cost Summary	1-33
Table 1-21:	Cash costs (net of by-product credits)	1-33
Table 2-1:	Contributors and Responsible Parties for this Report	2-3
Table 2-2:	Unit Abbreviations	2-4
Table 2-3:	Acronyms and Abbreviations	2-1
Table 4-1:	Mineral and oic lease Properties	4-2
Table 4-2:	General Permits and Quarry Lease	4-6
Table 6-1:	Lalor mine actual production	6-1
Table 7-1:	Summary of Zinc Rich Interpreted Wireframes	7-9
Table 7-2:	Summary of Gold Interpreted Wireframes	7-10
Table 9-1:	Underground Exploration Development	9-2

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	Lalor Mine
	Form 43-101F1 Technical Report

Table 10-1:	Summary Surface diamond drill holes with assay results as of january 1, 2017	10-1
Table 10-2:	Summary of LD Underground diamond Drill Holes as of January 1, 2017	10-2
Table 10-3:	Summary of LE underground diamond drill holes as of January 1, 2017	10-2
Table 10-4:	Summary of LP Underground Diamond Drill Holes as of January 1, 2017	10-2
Table 10-5:	Summary of lq underground diamond drill holes as of january 1, 2017	10-3
Table 10-6:	Summary of LX Underground diamond drill holes as of January 1, 2017	10-3
Table 11-1:	Hudbay Laboratory Detection Limits	11-3
Table 11-2:	Bureau Veritas Elemental Detection Limits AQ370	11-5
Table 11-3:	Bureau Veritas Elemental Detection Limits AQ270	11-6
Table 11-4:	Bureau Veritas Elemental Detection Limits and Range for Over Range Codes	11-7
Table 11-5:	Bureau Veritas Elemental Detection Limits for Legacy Codes (Group 7AR) and (Group 601)	11-7
Table 12-1:	Bureau Veritas Assays Specifications	12-1
Table 12-2:	OREAS Certified Blanks	12-2
Table 12-3:	Summary of Blank Performance	12-3
Table 12-4:	Certified Reference Materials	12-4
Table 12-5:	Summary of CRM Performance at Bureau Veritas	12-5
Table 12-6:	Summary of Coarse Duplicate Performance	12-7
Table 12-7:	Hudbay Laboratory Assays Specifications	12-9
Table 12-8:	Summary of Blank Performance	12-10
Table 12-9:	Summary of CRM Performance at Hudbay Laboratory	12-13
Table 12-10:	Summary of Regression Parameters	12-14
Table 12-11:	Summary of Coarse Duplicate Performance	12-15
Table 12-12:	Summary of RMA Regression Analysis	12-18
Table 13-1:	Historical Plant Head Assay and Metal Recovery	13-1
Table 13-2:	Historical Plant Concentrate Produced	13-2
Table 13-3:	EXPECTED LOM Recoveries at Stall Concentrator	13-3
Table 13-4:	Make-up of Variability Composites	13-4
Table 13-5:	Locked Cycle Test Conditions – Variability Composites	13-5
Table 13-6:	Locked Cycle Test Results – Variability Composites	13-6
Table 14-1:	Drilling Data by Year	14-1
Table 14-2:	Legend of Interpreted Wireframes	14-4
Table 14-3:	Samples and Length Analyzed	14-5
Table 14-4:	Zinc Assay Statistics by Lense	14-8
Table 14-5:	Gold Assay Statistics by Lense	14-8
Table 14-6:	Copper Assay Statistics by Lense	14-8

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	Lalor Mine
	Form 43-101F1 Technical Report

Table 14-7:	Silver Assay Statistics by Lense	14-9
Table 14-8:	Lead Assay Statistics by Lense	14-9
Table 14-9:	Iron Assay Statistics by Lense	14-9
Table 14-10:	Arsenic Assay Statistics by Lense	14-10
Table 14-11:	Density Assay Statistics by Lense	14-10
Table 14-12:	Capping Thresholds by Lense of Gold and Silver (g/t)	14-13
Table 14-13:	High-Yield Restriction Thresholds by Lense	14-13
Table 14-14:	Precious Metal Removed by Capping	14-14
Table 14-15:	Length Weighted 1.25 m Composite Statistics, Zinc (%)	14-16
Table 14-16:	Length Weighted Uncapped and Capped 1.25 m Composite Statistics, Gold (g/t) .	14-17
Table 14-17:	Length Weighted 1.25 m Composite Statistics, Copper (%)	14-18
Table 14-18:	Length Weighted Uncapped and Capped 1.25 m Composite Statistics, Silver (g/t)	14-18
Table 14-19:	Length Weighted 5 m Composite Statistics, Zinc (%)	14-19
Table 14-20:	Length Weighted Uncapped and Capped 5 M Composite Statistics, Gold (g/t)	14-19
Table 14-21:	Length Weighted 5 m Composite Statistics, Copper (%)	14-20
Table 14-22:	Length Weighted Uncapped and Capped 5 m Composite Statistics, Silver (g/t)	14-20
Table 14-23:	Variogram Models and Rotation Angles in Base Metal Lenses	14-23
Table 14-24:	Variogram Models and Rotation Angles in Gold Zones	14-24
Table 14-25:	NN, IDW and OK Model, Zinc Removed by Restriction in Base Metal Lenses	14-29
Table 14-26:	NN, IDW and OK Model, Gold Removed by Capping in Base Metal Lenses and Gold Zones	14-29
Table 14-27:	NN, IDW and OK Model, Copper Removed by Restriction in Base Metal Lenses and Gold Zones	14-30
Table 14-28:	NN, IDW and OK Model, Silver Removed by Capping in Base Metal Lenses and Gold Zones	14-31
Table 14-29:	NN, IDW and OK Model Statistics Mean Block Grade Comparisons for Zinc	14-32
Table 14-30:	NN, IDW and OK Model Statistics Mean Block Grade Comparisons for Gold	14-32
Table 14-31:	NN, IDW and OK Model Statistics Mean Block Grade Comparisons for Copper	14-33
Table 14-32:	NN, IDW and OK Model Statistics Mean Block Grade Comparisons Silver	14-33
Table 14-33:	Quality Control Statistics of the Zinc Interpolation Within the Base Metal Lenses	14-41
Table 14-34:	Quality Control Statistics of the Gold Interpolation Within the Base Metal Lenses	14-41
Table 14-35:	Quality Control Statistics of the Gold Interpolation Within the Gold Zones	14-41
Table 14-36:	Quality Control Statistics of the Copper Interpolation Within the Gold Zones	14-42
Table 14-37:	Zinc Grade-Tonnage Statistics in Base Metal Lenses	14-43
Table 14-38:	Gold Grade-Tonnage Statistics in Base Metal Lenses	14-43
Table 14-39:	Gold Grade-Tonnage Statistics in Gold Zones	14-44

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	Lalor Mine
	Form 43-101F1 Technical Report

Table 14-40:	Copper Grade-Tonnage Statistics in Gold Zones	14-44
Table 14-41:	RCI Values with Average ADIST, CDIST and MDIST	14-51
Table 14-42:	Lalor Metal Equivalencies	14-55
Table 14-43:	Metal Prices – 2017 Mineral Resource ($US)	14-55
Table 14-44:	Metal Recoveries	14-55
Table 14-45:	Manitoba Business Unit – General Management, Administration and Unallocated Services Costs	14-56
Table 14-46:	Copper and Zinc Concentrate Terms	14-56
Table 14-47:	Gold Doré Terms	14-57
Table 14-48:	Mining	14-57
Table 14-49:	Milling	14-57
Table 14-50:	Base Metal Mineral Resource, Inclusive of Mineral Reserves by Category and Mineralized Zone with a Cut-off of 4.1% Zn EQ, as of September 30, 2016 (1)(2)(3)(4)(5)(6)(7)(8)(9)	14-58
Table 14-51:	Gold Mineral Resource, Inclusive of Mineral Reserves, by Category and Mineralized Zone with a Cut-off of 2.4 g/t Au EQ, as of September 30, 2016 (1)(2)(3)(4)(5)(6)(7)(8)(9)	14-59
Table 14-52:	Mined-Out Areas from the Block Model	14-60
Table 14-53:	Ore Received by Year at the Stall Concentrator	14-60
Table 15-1:	Summary of Mineral Reserves as of January 1, 2017	15-1
Table 15-2:	Stoping Parameters by Mining Method	15-2
Table 15-3:	Average Dilution Factors by Lense Grouping	15-3
Table 15-4:	Recovery Factors by Mining Method and Lense	15-3
Table 15-5:	Average Recovery Factors by Lense Grouping	15-4
Table 15-6:	Metallurgical Assumptions	15-5
Table 15-7:	Copper and Zinc Concentrate Terms	15-5
Table 16-1:	Lateral Jumbo Development	16-13
Table 16-2:	Lateral Jumbo Development	16-17
Table 16-3:	Vertical Development	16-17
Table 16-4:	Mine Equipment	16-19
Table 16-5:	LOM Production Schedule	16-20
Table 16-6:	LOM Concentrate Production by Year	16-21
Table 16-7:	Mine Operations Workforce	16-27
Table 17-1:	Stall Mill Water Balance Data for 2016	17-10
Table 17-2:	Expansion Estimate Cost per Tonne	17-11
Table 19-1:	Key Long-Term Zinc Metal and Zinc Concentrate Assumptions	19-1
Table 19-2:	Key Long-Term Copper Concentrate Assumptions	19-2
Table 21-1:	Development Capital Cost Summary	21-1

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Table 21-2:	Sustaining Capital Cost Summary	21-1
Table 21-3:	Unit Operating Cost Summary	21-2
Table 21-4:	Cash costs (net of by-product credits)	21-2

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LIST OF FIGURES

TITLE		PAGE
Figure 1-1:	Lalor Mine Location	1-2
Figure 1-2:	Lalor Annual Zinc Production and C1 Cash Costs	1-34
Figure 4-1:	Location map of Hudbay MInes	4-3
Figure 4-2:	Mineral Claims and Lease Map	4-4
Figure 5-1:	Snow Lake Regional Map	5-2
Figure 7-1:	Geology of Manitoba	7-4
Figure 7-2:	Geology of the Flin Flon Greenstone Belt, Manitoba	7-5
Figure 7-3:	Volcanic Stratigraphy of the Snow Lake Area	7-6
Figure 7-4:	Geology of the Snow Lake Area	7-7
Figure 12-1:	Copper Coarse Duplicate Minimum and Maximum Plot	12-7
Figure 12-2:	Zinc Coarse Duplicate Minimum and Maximum Plot	12-8
Figure 12-3:	Silver Coarse Duplicate Minimum and Maximum Plot	12-8
Figure 12-4:	Gold Coarse Duplicate Minimum and Maximum Plot	12-9
Figure 12-5:	XY Comparison of the Average Gold Values Determined by Hudbay and Bureau Veritas Laboratories on CRMs Versus the CRM Recommended Best Value*	12-14
Figure 12-6:	Copper Coarse Duplicate Minimum and Maximum Plot	12-16
Figure 12-7:	Zinc Coarse Duplicate Minimum and Maximum Plot	12-16
Figure 12-8:	Silver Coarse Duplicate Minimum and Maximum Plot	12-17
Figure 12-9:	Gold Coarse Duplicate Minimum and Maximum Plot	12-17
Figure 12-10:	XY Plots of Check Assay Data Comparing Primary Laboratory Hudbay to Secondary Laboratory SGS*	12-19
Figure 12-11:	XY Plots of Check Assay Data Comparing Primary Laboratory Bureau Veritas to Secondary Laboratory SGS	12-20
Figure 13-1:	Copper and Gold Recovery to Concentrate, LCT and Plant Data	13-7
Figure 13-2:	Copper Concentrate Grades, LCT and Plant Data	13-8
Figure 14-1:	3D View of Wireframes, Looking west	14-3
Figure 14-2:	OK, IDW and NN Specific Gravity Distribution	14-5
Figure 14-3:	Box Plot of Zinc (%) by Lense	14-6
Figure 14-4:	Box plot of gold (G/T) by Lense	14-6
Figure 14-5:	Box Plot of Copper (%) by Lense	14-7
Figure 14-6:	Box Plot of Silver (G/T) by Lense	14-7
Figure 14-7:	Log Probability Plot of Au (g/t) in Zone 25	14-11
Figure 14-8:	Histograms and Basic Statistics of the Population Above and Below Capping in Zone 25 for Au (g/t)	14-12

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Figure 14-9:	Sample Length in Mineralized Zones	14-15
Figure 14-10:	Downhole Variogram Zinc, Lense 32	14-21
Figure 14-11:	Downhole Variogram Gold, Lense 25	14-22
Figure 14-12:	EW Cross Section N630 Showing OK Model and Composites Zinc Grade of Base Metal Lense 10	14-27
Figure 14-13:	EW Cross Section 1250N Showing OK Model and Composites Gold Grade in Gold Zone 25	14-28
Figure 14-14:	Zinc Swath Plot in Base Metal Lenses (by Easting)	14-34
Figure 14-15:	Gold Swath Plot in Base Metal Lenses (by Easting)	14-35
Figure 14-16:	Silver Swath Plot in Base Metal Lenses (by Easting)	14-36
Figure 14-17:	Zinc Swath Plot in Gold Zones (by Easting)	14-37
Figure 14-18:	Gold Swath Plot in Gold Zones (by Easting)	14-38
Figure 14-19:	Copper Swath Plot in Gold Zones (by Easting)	14-39
Figure 14-20:	Silver Swath Plot in Gold Zones (by Easting)	14-40
Figure 14-21:	Grade Tonnage of Restricted Zinc in the Base Metal Lenses	14-45
Figure 14-22:	Grade Tonnage of Capped Gold in the Base Metal Lenses	14-46
Figure 14-23:	Grade Tonnage of Capped Gold in the Gold Zones	14-47
Figure 14-24:	Grade Tonnage of Restricted Copper in the Gold Zones	14-48
Figure 14-25:	3D View (Looking SW) Displaying the Resource Classification in Lense 10	14-50
Figure 14-26:	3D View (Looking SW) Displaying the Resource Classification in Zone 25	14-53
Figure 16-1:	Typical Cut and Fill Mining Cross Section	16-3
Figure 16-2:	Typical Mechanized Cut and Fill Mining Cross Section	16-4
Figure 16-3:	Typical Post Pillar Cut and Fill Mining Cross Section	16-5
Figure 16-4:	Typical Post Pillar Cut and Fill Mining Plan View	16-5
Figure 16-5:	Typical Drift and Fill Mining Cross Section	16-6
Figure 16-6:	Typical Isometric View – Transverse Longhole Open Stoping	16-7
Figure 16-7:	Typical Long Section – Longitudinal Retreat Longhole Open Stoping	16-8
Figure 16-8:	Typical Long Section – Uppers Retreat Longhole Open Stoping	16-9
Figure 16-9:	Site General Arrangement	16-12
Figure 16-10:	Lalor Mine Ventilation Plan	16-23
Figure 17-1:	Lalor Concentrator Simplified Block Flow Diagram	17-2
Figure 17-2:	Stall Concentrator Crushing Process Flow Diagram	17-4
Figure 17-3:	Stall Concentrator Flotation Process Flow Diagram	17-7
Figure 17-4:	Stall Concentrator Dewatering Process Flow Diagram	17-9
Figure 18-1:	Site Access Road and Services	18-1
Figure 20-1:	Anderson TIA Stage 1A Construction Plan (AECOM, 2016)	20-3

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Figure 21-1:  Lalor Annual Zinc Production and C1 Cash Costs 21-3

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1 SUMMARY

1.1 Summary

This Technical Report (the Technical Report) has been prepared for Hudbay Minerals Inc. (Hudbay) to support the public disclosure of Mineral Resources and Mineral Reserves at the Lalor Mine and to provide an updated mine plan that contemplates 4,500 tpd of base metal, gold and copper-gold zone ore to Stall concentrator.

Hudbay is a Canadian integrated mining company with assets in North and South America principally focused on the discovery, production and marketing of base and precious metals. Hudbay’s objective is to maximize shareholder value through efficient operations, organic growth and accretive acquisitions, while maintaining its financial strength.

Hudbay operates multiple properties in the Province of Manitoba. Operations near Flin Flon include the 777 Mine, which also consists of an ore concentrator and zinc plant, and the Reed mine, which is located approximately 120 kilometres (km) by road southeast of Flin Flon. Operations near Snow Lake include the Lalor underground mine.

The Lalor mine consists of an ore concentrator, a tailings impoundment area and other ancillary facilities that support the operation. The property is located approximately 16 km by road west of the town of Snow Lake, Manitoba.

As of the date of this Technical Report, the Lalor mine is operating at approximately 3,000 to 3,500 metric tonnes per day (tpd) and is ramping-up production to 4,500 tpd by 2018. The production ramp-up is supported by the underground ore handling circuit capable of 4,500 tpd, transitioning to more bulk mining methods (65% of reserves) with additional mining fronts and design changes to improve mining efficiencies, developing ore passes and transfer raises to reduce truck haulage cycle times from the upper potions of the mine and commissioning of a paste backfill plant in the first quarter of 2018.

Autonomous operation of a Load Haul Dump loader underground is currently being trialed from the surface by tele-remote monitoring with changes to standard designs to allow isolation of autonomous areas and buffer storage for in-between shift mucking.

The increase in production to 4,500 tpd at Lalor is complemented by the Stall concentrator expansion to 4,500 tpd, which is currently underway and is expected to be commissioned in the third quarter of 2018.

The Qualifed Person (the “QP”) who supervised the preparation of this Technical Report is Robert Carter, P. Eng., Lalor Mine Manager, Hudbay Manitoba Business Unit.

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1.2 Property Description and Location

The Lalor mine is located approximately 208 km by road east of Flin Flon and 16 km by road west of Snow Lake in the Province of Manitoba at 54°52’N latitude, 100°08’W longitude and 303 metres above sea level (m ASL). Access to Lalor mine is from Provincial Road (PR) #395, a gravel road off PR #392, which joins the town of Snow Lake and PR #39 (Figure 1-1). From PR #395, there is an all-weather permanent road into the mine site.

FIGURE 1-1: LALOR MINE LOCATION

Hudbay owns 100% interest in the Lalor property through one Mineral Lease and eight Order In Council Leases to the south of the property.

Hudbay holds the exclusive right to the minerals, other than quarry minerals, and the mineral access rights required for the purpose of working the lands and mining and producing minerals from the Lalor mine. Surface tenure, currently necessary to accommodate buildings and/or structures, required for the efficient and economical performance of the mining operations has been applied for and approved.

1.3 Geological Setting and Mineralization

The Lalor deposit is interpreted as a volcanic massive sulphide (VMS) deposit that precipitated at or near the seafloor in association with contemporaneous volcanism, forming a stratabound accumulation of sulphide minerals. The depositional environment for the mineralization at Lalor is similar to that of present and past producing base metal deposits in felsic to mafic volcanic and volcaniclastic rocks in the Snow Lake mining camp. The deposit appears to have an extensive associated hydrothermal alteration pipe.

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The Snow Lake arc assemblage that hosts the producing and past-producing mines in the Snow Lake area is a 20 km wide by 6 km thick section that records a temporal evolution in geodynamic setting from ‘primitive arc’ to ‘mature arc’ to ‘arc-rift’. The ‘mature arc’ Chisel sequence that hosts the zinc rich Chisel, Ghost, Chisel North, and Lalor deposits typically contains thin and discontinuous volcaniclastic deposits and intermediate to felsic flow-dome complexes.

The Chisel sequence is lithologically diverse and displays rapid lateral facies variations and abundant volcaniclastic rocks. Mafic and felsic flows both exhibit evolved geochemical characteristics consistent with one of, or a combination of, the following: within-plate enrichment, derivation from a more fertile mantle source, lower average extents of melting at greater depths, and contamination from older crustal fragments. These rocks have undergone metamorphism at the lower to middle almandine-amphibolite facies.

Rock units in the hanging wall of the Lalor deposit typically reflect this diversity and variation in rock types that include mafic and felsic volcanic and volocaniclastic units, mafic wacke, fragmental units of various grain sizes, and crystal tuff units.

1.3.1 Base Metal Mineralization

Lalor base metal mineralization begins at approximately 600 metres (m) from surface and extends to a depth of approximately 1,100 m. The mineralization trends about 320° to 340° azimuth and dips between 30° and 45° to the north. It has a lateral extent of about 900 m in the north-south direction and 700 m in the east-west direction.

Sulphide mineralization is pyrite and sphalerite. In the near solid (semi-massive) to solid (massive) sulphide sections, pyrite occurs as fine to coarse grained crystals ranging one to six millimetres and averages two to three millimetres in size. Sphalerite occurs interstitial to the pyrite. A crude bedding or lamination is locally discernable between these two sulphide minerals. Near solid coarse grained sphalerite zones occur locally as bands or boudins that strongly suggest that remobilization took place during metamorphism.

Disseminated blebs and stringers of pyrrhotite and chalcopyrite occur locally within the massive sulphides, adjacent to and generally in the footwall of the massive sulphides. The hydrothermally altered rocks in the footwall commonly contain some very low concentrations of sulphide minerals.

Seven distinct stacked zinc rich mineralized zones have been interpreted within the Lalor deposit based on the zinc equivalency of 4.1% over a minimum three metre interval. The top two lenses of the stacked base metal zones (coded as Zone 10 and 11) have higher grade zinc and iron content. The footwall lenses coded as Zones 20, 30, 31, 32 and 40 have moderate to high zinc grades hosted in near solid sulphides containing higher grade gold and locally appreciable amounts of copper.

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Overall, Zones 10 and 20 have the largest extent and volume of mineralization. Zone 10 extends approximately 400 m in the east-west and 550 m in the north-south direction and Zone 20, 250 m in the east-west and 700m in the north-south direction.

1.3.2 Gold Mineralization

Gold and silver enriched zones occur near the margins of the zinc rich sulphide lenses and as lenses in local silicified alteration. Remobilization is illustrated in some of the gold-rich zones by late veining that is more or less restricted to the massive lenses. Some of the footwall zones tend to be associated with silicification and the presence of gahnite. These zones are often characterized by copper-gold association, and are currently interpreted as being associated with higher temperature fluids below a zone of lower temperature base-metal accumulations.

Footwall gold mineralization is typical of any VMS footwall feeder zone with copper-rich, disseminated and vein style mineralization overlain by a massive, zinc-rich lens. The fact that the footwall zone is strongly enriched in gold suggests a copper-gold association which is comparable to other gold enriched VMS camps and deposits.

Seven lense groups have been interpreted within the deposit area and are present between 750 m to 1,480 m below surface. Their general shape is similar to the base metals. However, the current interpretation suggests the deeper copper-gold lense tends to have a much more linear trend to the north than the rest of the zones. The gold mineralization associated with each zone was interpreted into three-dimensional wireframes based on a 2.4 grams per metric tonne (g/t) gold equivalent over a minimum 3 metre (m) interval.

1.4 Exploration

Exploration drilling since Hudbay’s NI 43-101 technical report on the Lalor mine dated March 29, 2012 has focused on delineation of the inferred resource, confirming the continuity of the mineralization down plunge and testing for new mineralization peripheral to the known deposit. Surface drilling since the previous disclosure is limited to four drill holes while the focus was on underground exploration, definition, and delineation drilling, which has continued to expand the resource.

1.4.1 Underground Exploration Development

Since 2014 one exploration drift and one exploration ramp were developed at Lalor for a total of 1,891 m. The development was undertaken to establish underground platforms to conduct exploration drilling on targets that could not be drilled from existing mine infrastructure. Prudent care was taken in the placement and size of both the exploration ramp and drift to assure the selected locations can accommodate future mining equipment and related infrastructure.

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1.4.2 Underground Exploration Drilling

In 2015, thirty-one drill holes were completed for a total of 10,395 m focusing on the copper-gold, Zone 27. Exploration drilling continued on gold Zone 25 from March to July of 2016 for a total of sixty nine drill holes and 16,098 m. The purpose of the exploration programs was to upgrade inferred resources, specifically focused on identifying areas of enriched gold and copper-gold mineralization. Due to the low angle and stacking nature of the mineralization at Lalor, holes were extended beyond the gold target depths to explore the on-strike and plunge potential of known base metal lenses, which led to increases in mineral resource inventory.

1.4.3 Borehole Electromagnetic (EM) Surveys

Time-domain borehole EM surveys with three-dimensional probes are routinely conducted on surface and underground drill holes. The survey results identify any off-hole conductors that were missed, indicate direction to the target, as well as the dimensions and the attitude of the conductor. The surveys can also detect possible conductors which may lie past the end of the hole allowing decisions to extend holes to be made.

1.4.4 Surface EM Surveys

Two time-domain surface EM surveys, for a total of approximately 35 line km, were completed north-northeast and south-southwest of the Lalor mine. Neither survey has identified any new significant targets of interest in the general Lalor mine area.

1.4.5 Airborne EM Survey

During the summer of 2014 an airborne EM survey was conducted to test the capabilities of the HeliSAM system for a total of 97.5 line km. This was performed by GAP Geophysics, based in Perth Australia, using a ground based transmitting loop and airborne total field magnetic sensor. The testing was aimed to identify the Lalor mine at a depth beyond the capabilities of conventional airborne EM systems. The test was successful and has led to further surveys of this type elsewhere in the mining camp.

1.5 Drilling

The Lalor mine was discovered by Hudbay drilling a surface exploration hole testing a electromagnetic geophysical anomaly in March 2007, which intersected appreciable widths of zinc-rich massive sulphides in hole DUB168. Surface drilling continued through July 2012.

A limited surface exploration drill program was conducted from August to October 2015 to explore for potential down plunge extensions of copper-gold Zone 27 and to test near mine geophysical conductors that could not be drilled from underground workings. As of January 1, 2017 a total of 203,037 m of surface drilling was completed at Lalor and Table 1-1 provides a summary by year.

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TABLE 1-1: SUMMARY SURFACE DIAMOND DRILL HOLES WITH ASSAY RESULTS AS OF JANUARY 1, 2017

			Number of	Core
Year	Hole Type	Operator	Holes	Size	Length (m)	Drilling Company
2007	Parent	Hudbay	2	BQ	2,342	Major Drilling Ltd.
	Parent	Hudbay	26	NQ	29,600	Major Drilling Ltd.
2008	Parent	Hudbay	41	NQ	45,454	Major Drilling Ltd.
	Wedge	Hudbay	32	NQ/AQ	12,112	Major Drilling Ltd.
2009	Parent	Hudbay	29	NQ	35,390	Major Drilling Ltd.
	Wedge	Hudbay	47	NQ/AQ	22,884	Major Drilling Ltd.
2010	Parent	Hudbay	13	NQ	17,438	Major Drilling Ltd.
	Wedge	Hudbay	17	NQ/AQ	11,576	Major Drilling Ltd.
2011	Parent	Hudbay	10	NQ	15,458	Major Drilling Ltd.
	Wedge	Hudbay	5	NQ/AQ	3,139	Major Drilling Ltd.
2012	Parent	Hudbay	3	NQ	4,688	Major Drilling Ltd.
2015	Parent	Hudbay	2	NQ	2,956	Rodren Drilling Ltd.
Total			222		203,037

1.6 Sample Preparation, Analyses and Security

Since the start of exploration at Lalor the following different laboratories and sample shipment/preparation procedures have been in use:

• Discovery to November 1, 2009, Lalor samples were prepared and analyzed at Hudbay laboratory in Flin Flon, Manitoba. As part of Hudbay Quality Assurance and Quality Control (QAQC) procedures, pulp duplicates were sent to ACME Analytical Laboratories Ltd. (ACME) in Vancouver, BC.

• November 1, 2009 to March 12, 2012, Lalor samples were received, crushed and pulverized at Hudbay laboratory and pulps shipped to ACME for analysis. Pulp duplicates were analyzed at Hudbay laboratory as part of QAQC procedures.

• March 13, 2012 to May 21, 2014, all samples were prepared and analyzed at Hudbay laboratory. As part of Hudbay QAQC procedures, pulp duplicates were sent to ACME.

• May 22, 2014 to present, parts of the sample stream were and are shipped to ACME, (re- named to Bureau Veritas after January 1, 2015). The remainder of sample stream is shipped to the Hudbay laboratory.

• A set of 303 drill core pulps from samples originating from within known resource envelopes were submitted for check assaying at a SGS Laboratory in Burnaby, BC as part of the QAQC program for the 2017 resource estimate.

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1.6.1 Sample Preparation

All samples arriving at the Hudbay analytical laboratory are checked against the geologist’s sample submission sheets. Laboratory analytical work sheets are generated for the analysis areas. Any wet samples are dried at 105°C as per industry standard. The core samples are crushed to (-)10 mesh then split to approximately 250 g and pulverized with 90% passing (-)150 mesh before being deposited into labelled bags. Crusher and pulverizer checks are conducted daily to ensure there is no excessive wear on the crusher plates and pulverizer pots.

All samples arriving at ACME (Bureau Veritas) are checked against chain of custody information on sample submittal form and prepped according to codes WGHT and PR80-250. The sample preparation includes weighing of sample, crushing 1 kg to minimum 80% passing 2 mm. A 250 g split crushed to minimum 85% passing 75 µm.

1.6.2 Analyses

Samples sent to the Hudbay laboratory were analyzed for the following elements: gold, silver, copper, zinc, lead, iron, arsenic and nickel. Base metal and silver assaying was completed by aqua regia digestion and read by a simultaneous Inductively Coupled Plasma (ICP) unit. The gold analysis was completed on each sample by atomic absorption spectrometry (AAS) after fire assay lead collection. All samples with gold values (AAS) > 10 g/t were re-assayed using a gravimetric finish.

Two different assay methods are used for samples shipped to Bureau Veritas: AQ270 and AQ370. AQ370 was the only method used on samples submitted from May 2014 to the second quarter of 2016 after which the AQ270 method was applied to selected holes. After the fourth quarter of 2016 all samples submitted to Bureau Veritas were assayed using the AQ270 method. All samples using method AQ270 and AQ370 were run for gold using method FA430. The following elements were run for over range as necessary: gold, copper, zinc and lead using methods FA530, GC820, GC816, MA404 (assays returning lead values above 20% using MA404 were also run using method GC817) respectively.

Samples shipped to Bureau Veritas from November 1, 2009 to March 12, 2012, were run using the legacy codes (Group 7AR) and (Group 601) with over range samples being run with gravimetric finish (Group 612). The sample preparation for these legacy codes is essentially similar to those listed for the current AQ270 and AQ370 codes used after May 22, 2014.

For the multi-element methods AQ270 and AQ370, aliquots of 1.000 ± 0.002 g are weighed into 100 mL volumetric flasks. Bureau Veritas QAQC protocol requires one pulp duplicate to monitor analytical precision, a blank, and an aliquot of in-house reference material to monitor accuracy in each batch of 36 samples. 30 mL of Aqua Regia, a 1:1:1 mixture of ACS grade concentrated HCl, concentrated HNO₃ and de-mineralised H₂O, is added to each sample. Samples are digested for one hour in a hot water bath (> 95°C). After cooling for 3 hours, solutions are made up to volume (100 mL) with dilute (5%) HCl. Very high-grade samples may require a 1 g to 250 mL or 0.25 g to 250 mL sample/solution ratio for accurate determination. Bureau Veritas QAQC protocol requires simultaneous digestion of a reagent blank inserted in each batch.

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For both AQ270 and AQ370 sample solutions are aspirated into an ICP emission spectrograph (ES) to determine 24 elements. For method AQ270 the solution is also run through an ICP mass spectrometer (MS) to provide values for an additional 10 elements bring the total number of elements to 34. Raw and final data from the ICP-ES/ICP-MS undergoes a final verification by a British Columbia Certified Assayer who then signs the Analytical Report before it is released to the client.

For the gold analysis FA430, 30 g charges are weighed into fire assay crucibles. The sample aliquot is custom blended with fire assay fluxes, PbO litharge and a silver inquart. Firing the charge at 1050°C liberates Au, Ag ± PGEs that report to the molten Pb-metal phase. After cooling the lead button is recovered, placed in a cupel, and fired at 950°C to render an Ag ± Au ± PGEs dore bead. The bead is weighed and parted (i.e. leached in 1 mL of hot HNO₃) to dissolve silver leaving a gold sponge. Adding 10 mL of HCl dissolves the Au ± PGE sponge. Solutions are analysed for gold on an ICP emission spectrometer. Gold in excess of 10 g/t forms a large sponge that can be weighed (gravimetric finish, method FA530).

1.6.3 Security

Security measures taken to ensure the validity and integrity of the samples collected include:

• Chain of custody of drill core from the drill site to the core logging area

• All facilities used for core logging and sampling located on a secure mine site

• Core sampling is undertaken by Hudbay geologists

• Sample splitting and shipping conducted by technicians under the supervision of Hudbay geologists

• Chain of custody for core cutting through to delivery of samples to laboratories

• Well documented and implemented receiving and processing procedures at the Hudbay and Bureau Veritas laboratories

• The Hudbay laboratory samples results are stored on a secure mainframe based Laboratory Information Management System (LIMS)

• The diamond drill hole database is stored on the secure Hudbay network, using the acQuire database management system with strict access rights

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1.7 Data Validation

Drill core is logged, sample intervals selected and either whole core or split core with remaining half core kept for reference is submitted to the laboratory for assaying. As part of Hudbay’s QAQC program, the following QAQC samples are inserted into the sample stream for every 100 samples:

1.	Two blanks
2.	Five duplicates
3.	Five base metal standards, each of differing grade thresholds
4.	Two gold standards of differing grade

It is concluded that the analytical accuracy and reproducibility of copper, zinc, and silver as indicated by the QAQC samples submitted at the Hudbay laboratory is appropriate for resource estimation. High grade gold standards are being under assayed at the Hudbay laboratory. This under assaying in turn affects the high-grade Lalor samples and will likely led to an underestimation of gold in the resource estimate, since the proportion of samples assayed at the Hudbay laboratory is approximately 80% of the total samples assayed between 2012 and 2016.

The accuracy and reproducibility of copper, zinc, silver, and gold assays, as indicated by the QAQC samples submitted at Bureau Veritas laboratories, is of good quality for resource estimation.

1.8 Mineral Processing and Metallurgical Testing

The Stall concentrator began processing ore from Lalor in August 2012, initially producing only a zinc concentrate and by October 2012 a copper concentrate. As Lalor increased ore production the mill underwent an initial expansion and by the summer of 2014 was capable of processing at 2,800 tpd throughput rate. Modifications and improvements at the Stall concentrator over the last few years has increased the milling performance to an average rate of 3,000 tpd in 2016.

The properties of the Lalor base metal ore are not expected to vary significantly from the previous four years of milling and it is appropriate to assume that the metal recoveries will remain in the 80 to 85% range for copper, with the exception of a few higher grade copper years and 90 to 95% range for zinc for the remaining Life of Mine (LOM). The yearly LOM metal recoveries, shown in Table 1-2, were calculated using Hudbay’s in-house metallurgical model that considers the relationship of metal grade versus recovery from historical data at optimal operating days. Optimal operating days are considered to be steady run and an appropriate control of parameters.

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TABLE 1-2: EXPECTED LOM RECOVERIES AT STALL CONCENTRATOR

	Metal Recoveries				Concentrate Grade
Year	Au (%)	Ag (%)	Cu (%)	Zn (%)	Zn (%)	Cu (%)
2017	59.6	51.9	83.9	93.6	51.0	21.0
2018	53.5	46.7	83.9	91.8	51.0	21.0
2019	55.8	48.2	83.1	91.7	51.0	21.0
2020	57.9	57.5	88.2	90.0	51.0	21.0
2021	61.9	66.1	89.1	90.5	51.0	21.0
2022	62.3	66.9	88.6	91.8	51.0	21.0
2023	60.1	60.8	88.2	91.8	51.0	21.0
2024	55.3	47.6	86.1	93.5	51.0	21.0
2025	59.7	52.2	87.7	91.7	51.0	21.0
2026	52.6	38.8	84.7	92.0	51.0	21.0
2027	55.8	39.5	81.7	90.6	51.0	21.0

Although it is the author’s opinion that actual plant performance overrides previous metallurgical testing it is appropriate to summarize the relevant testing completed on Lalor mineralization prior to processing ore at Stall concentrator.

The primary objectives of the test programs in 2009 and 2011 were to develop an appropriate flowsheet for either the design of a new concentrator or modifications to the existing Stall concentrator, and to determine expected concentrate grades and metal recoveries.

Mineralogical analysis showed that chalcopyrite in Lalor ore is mostly coarse grained and liberated at grind sizes of approximately 100 microns. However, 15 to 20% of the chalcopyrite remains locked, primarily with sphalerite, at sizes below 20 microns. Copper is present almost exclusively as chalcopyrite with minor bornite. Zinc is present mainly as sphalerite, with minor amounts of gahnite. The sphalerite is coarse-grained and liberated at a grind size of 250 microns. Lead is present as fine-grained galena and would require a grind size of 70 microns for liberation. However, there is insufficient galena in the ore to warrant a primary grind this fine.

Zinc recoveries achieved in the plant are generally higher than the laboratory recoveries while zinc concentrate grades in the plant have been lower than the laboratory results. This is a preferred operating option due to Hudbay’s short concentrate haul to their Flin Flon metallurgical site.

1.9 Mineral Resource Estimates

The mineral resources for Lalor are estimated either as base metal lenses or gold zones and classified as Measured, Indicated and Inferred resources, inclusive of mineral reserves, as of September 30, 2016. The Qualified Person for the mineral resource estimate is Robert Carter, P. Eng., Lalor Mine Manager, Hudbay Manitoba Business Unit.

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The resource is based on integrated geological and assay interpretation of information recorded from diamond drill core logging and assaying and underground mapping and is comprised of the following steps: exploratory data analysis, high-grade capping, high yield grade restrictions, and estimation and interpolation parameters consistent with industry standards. A total of 420,310 m in 1,707 holes have been drilled at Lalor deposit.

Mineral resources were classified in accordance to CIM Definition Standards on Mineral Resources and Mineral Reserves, into measured, indicated and inferred, depending upon the confidence level of the resource based on experience at Lalor mine, with similar deposits and spatial continuity of the mineralization.

The base metal resources were categorized as measured when a distance to an underground development drift is generally less than 10 m, indicated when the closest distance is less than or equal to 50 m to a composite, and the remainder of the interpolated blocks within the interpreted lenses are classified as inferred.

The gold zone resources were classified using the relative difference between the kriged grade and the composite grade. The Resource Classification Index (RCI) uses the ordinary kriging combined variance, block model grade and a calibration factor based on the distance of the composite, number of composites, number of quadrants, and the number of drill holes are used in the formula. An RCI value that corresponds to the 50^th percentile was used as a threshold for indicated resource, except in Zone 27 where a RCI of 70^th percentile was used. Measured blocks were classified based on the approximate distance of 10 m to an underground development drift. All remaining blocks with minimum criteria of one drill hole to interpolate the grades were classified as inferred.

The Lalor block model was validated to ensure appropriate honouring of the input data by the following methods:

• Visual inspection of the ordinary kriging (OK) block model grades in plan and section views in comparison to composites grade

• Metal removed via grade capping and high yield restriction methodology

• Comparison between the interpolation methods of nearest neighbour and inverse distance squared weighted to confirm the absence of global bias in the OK grade model

• Swath plot comparisons of the estimation methods to investigate local bias

• Review of block model ordinary kriging quality control parameters

• Comparison of grade tonnage curves and statistics by estimation method

• Third party review of the block model and estimation process

The base metal mineral resources, inclusive of reserves by category with a zinc equivalency cut-off of 4.1% zinc, is shown in Table 1-3 and the gold mineral resources, inclusive of reserves by category with a gold equivalency cut-off of 2.4 g/t gold, is shown in Table 1-4, as of September 30, 2016.

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TABLE 1-3: BASE METAL MINERAL RESOURCE, INCLUSIVE OF MINERAL RESERVES BY
CATEGORY AND MINERALIZED ZONE WITH A CUT-OFF OF 4.1% ZN EQ, AS OF SEPTEMBER 30, 2016 ^{(1)(2)(3)(4)(5)(6)(7)(8)(9)}

Category	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
Measured	5,126,000	8.34	2.46	0.86	31.34
Indicated	8,842,000	6.67	2.00	0.59	30.44
Measured + Indicated	13,967,000	7.28	2.17	0.69	30.77
Inferred	545,300	8.15	1.45	0.32	22.28

Notes:

1.	Domains were modelled in 3D to separate mineralized zones from surrounding waste rock. The domains were based on core logging, grade, structural and geochemical data.
2.	Raw drill hole assays were composited to 1.25 m lengths, honouring lithology boundaries.
3.	Capping of high gold and silver grades was considered necessary and was completed on assays prior to compositing.
4.	High yield restriction of base metal high grade and density was completed for each domain after compositing.
5.	Block grades for zinc, gold, silver, copper, lead, iron, arsenic and density were estimated from the composites using ordinary kriging interpolation into 5 m x 5 m x 5 m blocks coded by domain.
6.	Density values are from a multi elements regression formula based on 65,792 measurements
7.	Blocks were classified as Measured, Indicated or Inferred in accordance with CIM Definition Standards 2014.
8.	Metal prices of $US 1.19/lb zinc, $US 1,300/oz gold, $US 2.67/lb copper, and $US 18.00/oz silver with a CAD / US foreign exchange of 1.25 were used to calculate a zinc equivalence (Zn Eq) cut-off of 4.1%, where Zn Eq = Zn% + (1.98 x Cu%) + (1.11 x Au g/t) + (0.01 x Ag g/t) – (0.01 x Pb%). The Zn Eq considers the ratio of milling recovery, payability and value of metals after application of downstream processing costs. The Zn Eq cut-off of 4.1% covers administration overhead, mining removal, milling and general and administration costs.
9.	Totals may not add up correctly due to rounding.

TABLE 1-4: GOLD MINERAL RESOURCE, INCLUSIVE OF MINERAL RESERVES BY
CATEGORY AND MINERALIZED ZONE WITH A CUT-OFF OF 2.4 G/T AU EQ, AS OF SEPTEMBER 30, 2016 ^{(1)(2)(3)(4)(5)(6)(7)(8)(9)}

Category	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
Measured	332,000	0.44	6.34	0.40	33.03
Indicated	4,108,000	0.50	6.23	0.89	34.80
Measured + Indicated	4,440,000	0.50	6.24	0.86	34.67
Inferred	4,124,000	0.31	5.01	0.90	27.61

Notes:

1.	Domains were modelled in 3D to separate mineralized zones from surrounding waste rock. The domains were based on core logging, grade, structural and geochemical data.
2.	Raw drill hole assays were composited to 1.25 m lengths, honouring lithology boundaries.
3.	Capping of high gold and silver grades was considered necessary and was completed on assays prior to compositing.
4.	High yield restriction of base metal high grade and density was completed for each domain after compositing.
5.	Block grades for zinc, gold, silver, copper, lead, iron, arsenic and density were estimated from the composites using ordinary kriging interpolation into 5 m x 5 m x 5 m blocks coded by domain.
6.	Density values are from a multi elements regression formula based on 65,792 measurements collected by Hudbay.
7.	Blocks were classified as Measured, Indicated or Inferred in accordance with CIM Definition Standards 2014.
8.	Metal prices of $US 1,300/oz gold, $US 2.67/lb copper and $US 18.00/oz silver with a CAD / US foreign exchange of 1.25 were used to calculate a gold equivalence (Au Eq) cut-off of 2.4 g/t Au Eq, where Au Eq = Au g/t + (1.34 x Cu %) + (0.01 x Ag g/t). The Au Eq considers the ratio of milling recovery, payability and value of metals after application of downstream processing costs. Au Eq cut-off of 2.4 g/t covers administration overhead, mining removal, milling and general and administration costs.
9.	Totals may not add up correctly due to rounding.

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	Form 43-101F1 Technical Report

1.9.1 Mine Reconciliation of Block Model

A mine reconciliation of the block model was carried out on the mined out areas. The process involved selecting all mined out blocks and comparing to the actual metal balance reported as ore received at the Stall concentrator. Mined out areas from the block model do not include dilution or pillars left behind after mining extraction. Table 1-5 list the mined out areas by lense from the block model and Table 1-6 is the ore reported by year at the Stall concentrator since Lalor commenced production in August 2012 until September 2016.

TABLE 1-5: MINED OUT AREAS FROM THE BLOCK MODEL

Lense	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
10	1,560,246	9.63	1.94	0.71	19.39
11	55,355	14.46	0.14	0.30	17.22
20	336,600	8.80	2.19	0.82	30.71
21	99,594	0.65	7.70	0.65	30.85
23	24,210	1.50	5.37	0.61	39.89
24	25,849	0.89	6.44	0.42	23.24
25	56,481	0.64	8.47	0.40	31.93
30	9,098	4.25	1.82	0.26	35.67
31	64,845	4.73	0.95	0.20	15.14
32	183,404	7.98	5.51	1.60	54.22
Total	2,415,684	8.60	2.65	0.75	24.52
		Zn (tonnes)	Au (ounces)	Cu (tonnes)	Ag (ounces)
In-Situ Metal		207,641	205,689	18,196	1,904,304

TABLE 1-6: ORE RECEIVED BY YEAR AT THE STALL CONCENTRATOR

Year	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
2012	72,294	11.83	1.68	0.63	19.30
2013	400,589	9.44	1.20	0.84	19.41
2014	551,883	8.52	2.29	0.88	23.83
2015	934,278	8.18	2.53	0.71	21.39
2016 Sep YTD	814,207	6.88	2.30	0.64	21.62
Stockpile as of Sep 2016	11,114	5.38	2.14	0.41	21.42
Total	2,784,365	8.13	2.20	0.74	21.60
		Zn (tonnes)	Au (ounces)	Cu (tonnes)	Ag (ounces)
Metal		226,438	196,908	20,525	1,933,617
	Tonnes	Zn	Au	Cu	Ag
Variance to Block Model	115%	109%	96%	113%	102%

The block model compared very well to the ore reported at the Stall concentrator. The mine reconciliation concludes a 15% mining dilution and a metal variance reported at the Stall concentrator of 109% for zinc, 96% for gold, 113% for copper and 102% for silver. A mine reconciliation of 5 to 10% variance is well within industry standard. The precious metals reconciled very well, while there might be some conservatism of the zinc and copper grade estimates in the block model. The conservatism of the zinc and copper grade is believed to be linked to the high yield radius parameter. A previous model selected a 40 m high yield radius compared to the current 20 m radius used for this block model.

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	Form 43-101F1 Technical Report

1.10 Mineral Reserve Estimates

The mineral reserves were estimated based on a LOM plan prepared; using Deswik mine design software that generated mining inventory based on stope geometry parameters and mine development sequences. Appropriate dilution and recovery factors were applied based on cut and fill and longhole open stoping mining methods with a combination of paste and unconsolidated waste backfill material. The Qualified Person for the mineral reserve estimate is Robert Carter, P. Eng., Lalor Mine Manager, Hudbay Manitoba Business Unit.

The shallow dipping nature of the deposit and stacking of lenses results in multiple lenses being grouped together for mining purposes in the stope optimizer routines of Deswik so that they can be extracted as a single mining unit, based on stope mining parameters by mining method as shown in Table 1-7. Parameters most sensitive to Lalor mine are the minimum and maximum dip angles, which affects the dilution and recovery amounts of the optimized mining shape. The stope optimizer in Deswik generated an economic shape that honoured the geometric parameters.

TABLE 1-7: STOPING PARAMETERS BY MINING METHOD

Stope Shape Parameters	Unit	Longhole	Cut and Fill
Length
Minimum	Metres	23	150
Maximum		10	20
Width
Minimum	Metres	3.5	5
Maximum		50	50
Waste Pillar Width	Metres	5	-
Stope Height
Minimum	Metres	10	5
Maximum		20	5
Stope Dip	Degrees
Minimum		35	75
Hanging Wall Dip
Minimum	Degrees	20	70
Maximum		90	90
Footwall Dip
Minimum	Degrees	50	70
Maximum		90	90
Dilution
Hanging Wall Fixed	Metres	0.5	0.5
Footwall Fixed		0.5	0.5

The space between the lenses is treated as internal dilution and external dilution is set at a fixed distance of 0.5 m into the footwall and hanging wall after the stope geometry shape is finalized. Internal dilution and external dilution are included as part of the optimized mining shape. Dilution, set at zero grade and a bulk density of 2.8 t/m³, is based on the full mining shape with internal and external dilution. Average dilution of the mineral resources that are in the LOM production plan is 18.9% .

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	Form 43-101F1 Technical Report

Mining recovery is defined as the ratio of mineral resource tonnes delivered to the concentrator to the in-situ mineral resource tonnes. Some of the mineral resources are not recovered due to mining design, inefficiencies in mining and inefficiencies in mucking. Average recovery of the mineral resources that are in the LOM production plan is 81.1%

Diluted and recovered mineral resources exceeding a Net Smelter Return (NSR) cut-off of $88/t for longhole open stoping and $111/t for cut and fill mining method are included in the reserves. NSR’s are based on metal grades from the stope optimizer and block model, long-term metal prices, concentrator recoveries, smelter treatment, refining and payabilities and a Hudbay Manitoba Business Unit administration cost.

Metal prices of $US 1.07/lb zinc (includes premium), $US 1,260/oz gold, $US 3.00/lb copper, and $US 18.00/oz silver with an CAD / US foreign exchange of 1.10 was used to estimate mineral reserves.

The orebody is polymetallic with economically significant metals being zinc, gold, copper and silver. There are two different ore type, both of which are assumed to be treated using conventional flotation at Hudbay Stall concentrator:

• Base metals ores. Near solid to solid sulphide ores, with dominant pyrite and sphalerite with minor blebs and stringers of chalcopyrite and pyrrhotite.

• Gold rich ores. Silicified gold and silver enriched ores with stringers to disseminated chalcopyrite and sphalerite mineralization.

Metallurgical performance at Stall concentrator indicates that the base metal and gold rich ores can be blended and two concentrates will be produced, a zinc concentrate that will be shipped to the Hudbay Flin Flon metallurgical complex for production of refined zinc, and a gold enriched copper concentrate that will be shipped to third party smelters.

The Lalor mine mineral reserves as of January 1, 2017 are summarized in Table 1-8. The mine plan was prepared using measured and indicated mineral resources from the block model. Inferred resources were assumed as waste.

TABLE 1-8: SUMMARY OF MINERAL RESERVES AS OF JANUARY 1, 2017

Category	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
Proven	4,383,000	6.76	2.37	0.76	27.33
Probable	9,849,000	4.39	2.72	0.65	26.12
Proven + Probable	14,232,000	5.12	2.61	0.69	26.50

Notes:

1. CIM definitions were followed for mineral reserve

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	Form 43-101F1 Technical Report

2.	Mineral reserves are estimated at an NSR cut-off of $88/t for longhole open stoping mining method and $111/t for cut and fill mining method
3.	Metal prices of $US 1.07/lb zinc (includes premium), $US 1,260/oz gold, $US 3.00/lb copper, and $US 18.00/oz silver with a CAD / US foreign exchange of 1.10 were used to estimate mineral reserves.
4.	Bulk density of the resource is reported in the block model is from a multi elements regression formula based on 65,792 measurements. Stope geometry shapes include waste dilution based on a bulk density of 2.8 t/m3.
5.	Totals may not add up correctly due to rounding.

The conversion of resources to reserves is based on the LOM plan and NSR cut-offs that primarily focussed on capturing base metal resources for processing at the Stall concentrator. The secondary focus was to capture gold zone resources when in contact with or close proximity to base metal resources. In areas where a large separation existed between base metal and gold lenses, mining blocks were evaluated for economic stope mining shapes. When a non-economic shape was generated in a first pass, a second pass was evaluated for only base metal lenses and if an economic shape was generated the gold zone portion was removed. However, due to this larger separation, majority of these isolated gold lenses could have been evaluated independently of the base metal lenses and could potentially provide feed to a gold processing facility. Below approximately the 950 m level no attempt was made to generate an economic stope mining shape for gold zones 25 and 26 as the separation distance became too large. The author’s opinion is that these resources are potentially better suited for a gold processing facility and should be re-evaluated when Hudbay has a better understanding of their New Britannia gold mill and Birch Tailings Impoundment Area in Snow Lake.

Of the current 14,232,000 tonnes of mineral reserves, approximately 80% is converted from base metal resources and approximately 20% is converted from gold zone resources. Of the total reserves approximately 3.8% is represented by the indicated resources of copper-gold zone 27, which is inclusive of the 20% from the gold zone, noted above. Although the indicated resource of copper-gold zone 27 as shown in Table 14-51, is converted to reserves and is planned to be processed at the Stall base metal concentrator, it has the potential to be milled at the New Britannia gold mill if the refurbishment plan and the installation of a copper pre-float facility proves to be economically viable.

The author recommends that base metal indicated resources, exclusive of reserves, as shown in Table 1-9 remain as indicated resources until such a time that detailed mine planning is completed. Furthermore, the author recommends that gold zone indicated resources, exclusive of reserves, as shown in Table 1-10 still have reasonable prospects of economic extraction at either Stall base metal concentrator or a gold processing facility.

TABLE 1-9: BASE METAL INDICATED RESOURCE, EXCLUSIVE OF RESERVES,
WITH A CUTOFF OF 4.1% ZN EQ, AS OF SEPTEMBER 30, 2016 (1)

Category	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
Indicated	2,100,000	5.34	1.69	0.49	28.10

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	Form 43-101F1 Technical Report

Notes:

1. Refer to the Notes for Table 1-3 of this Technical Report for more information

TABLE 1-10: GOLD INDICATED RESOURCE, EXCLUSIVE OF RESERVES,
WITH A CUT-OFF OF 2.4 G/T AU EQ, AS OF SEPTEMBER 30, 2016 ⁽¹⁾

Category	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
Indicated	1,750,000	0.40	5.18	0.34	30.61

Notes:

1. Refer to the Notes for Table 1-4 of this Technical Report for more information

1.11 Mining Methods

The Lalor mine is a multi-lens flat lying orebody with ramp access from surface and shaft access to the 955 m level. Internal ramps located in the footwall of the orebody provide access between mining levels. Stopes are accessed by cross cuts from the major mining levels.

The mining method process includes underground lateral advance (development rounds), production mining, backfilling and transporting ore to surface. Geotechnical information, orebody geometry interpreted from diamond drill core and recent experience mining within the deposit were the major considerations for selection of mining methods.

Mining methods that are currently in use or planned in the immediate future include: mechanized cut and fill, post pillar cut and fill, drift and fill and longhole open stoping (transverse and longitudinal retreat).

Ore is mucked using Load Haul and Dump (LHD) loaders which are operated remotely in inaccessible areas. The ore is then loaded into underground haul trucks or ore passes and transported to the ore handling system at the production shaft for hoisting to surface.

A paste backfill plant will be constructed on site, planned for the first quarter of 2018. Paste backfill will be used in higher grade areas to increase recovery and accelerate the mining cycle. Lower grade areas will be filled with waste rock from waste development. No waste is planned to be hoisted.

Ore delivered to the production shaft is sized to less than 0.55 m at one of the two rockbreaker/grizzly arrangements and hoisted from the mine by two 16 tonne capacity bottom dump skips in balance. Ore is truck hauled to a primary crusher at the Chisel North mine site, crushed to less than 0.15 m, and then trucked to the Stall concentrator for further processing.

1.11.1 Stope Mining

Two main mining methods are used at Lalor mine, cut and fill and longhole open stoping. Cut and fill methods include: mechanized cut and fill, post pillar cut and fill and drift and fill. Longhole open stoping methods include: transverse, longitudinal retreat and uppers retreat. Each mining area is evaluated to determine the most economic stoping method. In general, where the dip exceeds 35° and the orebody is of sufficient thickness, longhole open stoping is used and lateral based cut and fill mining methods are used in flatter areas.

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	Form 43-101F1 Technical Report

Approximately 65% of the mineral reserves are mined using the longhole open stoping methods and 35% are mined with cut and fill methods.

1.11.2 Support Systems

Except when using cut and fill mining methods, all other drifts have arched backs for optimized shape and safety.

Ground support is broken down into primary support and secondary support.

Primary support refers to reinforcement of the rockmass immediately following excavation (first pass) to ensure safe working conditions before taking the next round. Primary support is typically undertaken with resin grouted rebar and #6 gauge galvanized welded wire mesh. Standard drift, <7 m wide: 2.2 m long #6 (for jackleg/stopper installations) or #7 (for bolter installations) resin rebar on a 1.2 m x 1.2 m square pattern. Rebar and screen are extended down the walls to within 1.8 m of the sill. Generally, this can be applied to a drift span up to 7.0 m if no major geological structures are encountered. Special design is needed for drift spans larger than 7.0 m or when major geological structure is present. Intersections, 7 m to 10.8 m wide: ground support uses the same support as for standard drifts except in the back where 3.6 m long #7 resin rebar on a 1.2 m x 1.2 m square pattern is installed.

Secondary support is additional support applied after the installation of primary support to provide further support in large spans, long term infrastructure excavations and structurally controlled areas where wedge failures may be a concern. Secondary support is installed at a later stage (second pass) and typically is a batch process. Examples of secondary support are rebar, cable bolts, strandlok bolts, inflatable rock anchors, split sets and shotcrete.

Secondary ground support is installed when excavation spans are larger than 10.8 m, major unfavorable ground conditions or rock structures are present, and/or after a site ground condition evaluation indicates it is required. Secondary ground support uses heavy duty longer bolts, such as cement grouted cable bolts. Typically, single cable bolts on a 1.8 m x 1.8 m pattern for long term excavations, or high strength inflatable rock anchors for temporary or short term excavations. The minimum bolt length should be equal to one-third of the final drift span.

1.11.3 Backfill

All stopes at Lalor mine are backfilled to maintain long term stability and to provide a floor to work from for subsequent mining. Unconsolidated waste rock fill is used in stopes where pillar or wall confinement is not required and the value of the adjacent pillars does not warrant the added expenditure of consolidated backfill. Consolidated backfill currently consists of cemented waste rock backfill and is planned to be primarily as paste backfill after commissioning of the paste plant in the first quarter of 2018. Where economically feasible consolidated backfill is used by adding cement to waste rock using a spray bar and placing it in stopes with LHDs or when paste is available using the underground distribution system to transport paste (via gravity) directly to the stope. Consolidated backfill is required to maintain long term stability and allow future recovery of sill pillars.

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	Form 43-101F1 Technical Report

The majority of consolidated backfill will be paste. Paste backfill is an engineered product comprised of mill tailings and a binder (3-5% cement by weight) mixed with water to provide a thickened paste that is delivered by borehole and pipes to stopes.

1.11.4 Ore Handling

Ore is mucked by LHD, loaded into underground haul trucks and hauled to one of the two ore passes that feed the shaft. Ore is dumped onto a grizzly at 910 m level for sizing to less than 0.55 m by a rockbreaker and grizzly. A 40 m raise and bin below the grizzly provides approximately 1,200 tonnes of coarse ore storage. A chute at the bottom of the raise at 955 m level feeds ore to a conveyor that loads a measuring flask with approximately 14 tonnes of ore. Ore is then skipped to surface by two 16-tonne capacity Bottom Dump skips in balance. The ore enters the head frame chute from the skips and is deposited into the surface ore bin or to the exterior concrete bunker via gravity. From the surface bin or bunker, ore is truck hauled to a primary crusher at the Chisel North mine site, crushed, and then trucked to the Stall concentrator to process. Opportunities to increase ore handling capacity and installation of additional shaft ore passes are currently being reviewed. However, based on an internal review the current ore handling system has the capacity to move 4,500 tpd.

1.11.5 Mining Operations

Typical development crew equipment consists of a two-boom electric hydraulic jumbo, and a mechanical bolter sized to excavate all lateral development (typical sizes include: 5.0 m x 5.0 m, 6.0 m x 5.0 m and 7.0 m x 5.0 m). The crew also uses LHDs, scissor lifts and backhoes for face preparation and extending services.

Current production is approximately 3,000 to 3,500 tonnes per day. At steady state by 2018, Lalor mine will produce 4,500 tonnes per day and be approximately 35% development based and 65% longhole. Development based production methods will produce approximately 4.2 ore rounds (1,600 tonnes) per day. Cut and fill mining areas are assumed to be in the ore producing portion of the mining cycle 75% of the time and in the backfill portion of the mining cycle or otherwise unavailable for mining 25% of the time. Longhole mining based production will produce approximately 2,900 tonnes per day.

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1.11.6 Mine Equipment

Lalor mine is a ramp and shaft accessible mine with production and development done by rubber tired underground mining equipment. The mine equipment fleet required to achieve 4,500 tonnes per day is shown in Table 1-11.

TABLE 1-11: MINE EQUIPMENT

Description	Fleet
Underground Trucks 65 tonne	4
Underground Trucks 42 tonne	4
LHD 8yd	5
LHD 10yd	5
Two Boom Jumbo	4
Bolters (includes require for cable bolting)	8
Longhole Drills	3
Powder Trucks	3
Scissor Lift Trucks	8
Grader	1
Boom truck	2
Shotcrete Sprayer	1
Trans-mixers	2
Personnel Carriers Toyota	26
Miscellaneous Underground (Minecats, forklifts, etc.)	19
Miscellaneous Surface (Loader, forklift, pickups, etc.)	22
Total Mobile Equipment.	117
Ventilation Fans – Surface fans (250 HP – 2500 HP)	6
Ventilation Fans – U/G fans (50 HP – 400 HP)	54
U/G Submersible Pumps 100 HP	7
U/G Submersible Pumps 60 HP	1
U/G Submersible Pumps 50 HP	1
U/G Submersible Pumps 40 HP	13
U/G Submersible Pumps 34 HP	1
U/G Submersible Pumps 20 HP	7
U/G Submersible Pumps <20 HP	5
Portable Refuge Stations	3
Shotcrete Machine - Wet Mix	1
Grout Pump c/w Mixer	3
Portable Welder	1
Total Stationary Equipment	103

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	Form 43-101F1 Technical Report

An allowance for replacement equipment has been included in the mine plan. As part of the mobile equipment fleet management plan, major mobile equipment will be replaced at approximately 15,000 operating hours.

1.11.7 Production Schedule

The LOM production schedule, shown in Table 1-12, is currently set to ramp-up to 4,500 tonnes tpd by 2018 and continues at that rate to the end of 2021 when a ramp-down begins to 3,000 tpd in 2026. Deswik software was used to assist with the LOM planning and generate the basis for the production schedule. The geologic block model was imported to the software, where a stope optimizer algorithm was applied to create economical mining shapes. These mining shapes were then linked to the development drifts and sequenced individually by their respective locations and geometrical limits. Mine resources and rates, realized through historical data, were applied and levelled through activity priority labels. From this output, adjustments were made to further balance resources and scheduling to create an improved plan.

TABLE 1-12: LOM PRODUCTION SCHEDULE

Year	Tonnes	Au (g/t)	Ag (g/t)	Cu (%)	Zn (%)
2017	1,278,282	1.67	22.68	0.59	7.52
2018	1,616,285	2.13	24.37	0.52	5.71
2019	1,620,000	1.86	21.43	0.48	5.62
2020	1,603,652	2.79	28.43	0.79	4.61
2021	1,620,000	2.86	26.39	0.92	4.83
2022	1,473,657	3.16	26.72	0.95	5.72
2023	1,267,267	3.21	29.87	0.89	5.72
2024	1,212,738	3.14	28.35	0.60	4.49
2025	1,212,739	2.89	27.35	0.66	3.41
2026	1,022,918	2.78	32.83	0.49	3.55
2027	304,098	1.83	23.93	0.37	2.68
Total	14,231,636	2.61	26.50	0.69	5.12

Lalor mine will produce a total of 1,312,563 tonnes of zinc concentrate and 404,864 tonnes of copper concentrate in milling the ore from the LOM production plan, as shown in Table 1-13. The LOM contained metal in concentrate is shown in Table 1-14.

TABLE 1-13: LOM CONCENTRATE PRODUCTION BY YEAR

	Zinc Concentrate		Copper Concentrate
Year	Tonnes	Zn (%)	Tonnes	Au (g/t)	Ag (g/t)	Cu (%)
2017	176,396	51.0	30,158	42.2	499.1	21.0
2018	166,124	51.0	33,298	55.3	552.6	21.0
2019	163,716	51.0	30,863	54.5	541.8	21.0
2020	130,580	51.0	53,179	48.7	492.7	21.0
2021	138,843	51.0	63,025	45.4	448.7	21.0
2022	151,843	51.0	58,903	49.2	446.9	21.0
2023	130,488	51.0	47,567	51.5	483.5	21.0
2024	99,888	51.0	29,786	70.7	549.8	21.0
2025	74,437	51.0	33,629	62.3	514.7	21.0
2026	65,591	51.0	20,064	74.6	649.6	21.0
2027	14,658	51.0	4,394	70.8	653.4	21.0
Total	1,312,563	51.0	404,864	53.4	502.8	21.0

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TABLE 1-14: LOM CONTAINED METAL IN CONCENTRATE

Year	Zn (tonnes)	Cu (tonnes)	Au (oz)	Ag (oz)
2017	89,962	6,333	40,917	483,928
2018	84,723	6,993	59,202	591,589
2019	83,495	6,481	54,079	537,611
2020	66,596	11,168	83,265	842,391
2021	70,810	13,235	91,994	909,201
2022	77,440	12,370	93,174	846,328
2023	66,549	9,989	78,760	739,424
2024	50,943	6,255	67,705	526,512
2025	37,963	7,062	67,359	556,492
2026	33,451	4,213	48,122	419,039
2027	7,476	923	10,002	92,306
Total	669,408	85,022	694,578	6,544,821

1.11.8 Mine Ventilation

The Chisel North mine ventilation system in sequence with the Lalor mine Downcast Raise, provide 400,000 cubic feet per minute (cfm) down the Lalor mine Access Ramp, with 150,000 cfm exhausting to surface via the Chisel North mine Ramp. An additional 555,000 cfm is downcast via the Lalor mine Production Shaft for a total of 955,000 cfm exhausting up the Main Exhaust Shaft. In the summer total volume of air increases slightly. Three heaters heat mine air in the winter: the 36M BTU Chisel North Mine Heater, the 30M BTU Lalor Mine Ramp Heater and an 80M BTU heater at the production shaft.

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With the increase in mining rate from 3,000 to 4,500 tpd several new areas are being brought into production. As the footprint of the mine expands, the ventilation system will also require expansion to allow fresh air to be delivered to active mining areas.

1.11.9 Mine Power

Grid electricity is supplied by Manitoba Hydro, the provincial power utility. Manitoba Hydro’s 115 kV power line terminates at the Chisel North mine site, approximately 3.5 road km from the Lalor mine site. This feeds power to the Hudbay owned Main distribution substation consisting of two (2) 115-25 kV 24 MVA transformers. Substation is completely equipped with an E-house complete with 4 GE Powervac circuit breakers and Tie breaker. Lalor mine underground mine electrical distribution to the mine workings consist 13.8 kV that is further stepped down to 600 V.

1.11.10 Workforce

Lalor mine is operated on a continuous cycle. The majority of operations and maintenance personnel work 11.5 hour shifts on a 5-5-4 day cycle or a 7-7 day cycle. Operations support, technical and administrative personnel work 8 hour day shifts, 40 hours per week. The mine is operated under Collective Bargaining Agreements between Hudbay management and local unions.

Mine operations workforce is comprised of Hudbay hourly operations and maintenance personnel as well as salaried supervision, mine administration and technical staff, plus contractor personnel for specialized work and manpower shortages. Personnel will vary year to year. Steady state personnel requirements are shown in Table 1-15.

TABLE 1-15: MINE OPERATIONS WORKFORCE

Discipline	Personnel
Direct Operations	180
Supervision and Administration	47
Health and Safety	4
Mine Maintenance	74
Mine Technical	34
Total Lalor Mine	339

1.11.11 Mine Safety and Health

All personnel are required to work under the applicable laws of the Province of Manitoba, Canada. All contractors working on site are required to have an approved health and safety program in place and have on site representation. Hudbay Plant Safety Rules and Regulations are used at Lalor mine operations including, but not limited to:

a)	Positive Attitude Safety System (PASS) safety program
b)	Health monitoring programs (hearing and lung)

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c)	Dust monitoring
d)	Ongoing water and environmental monitoring
e)	Personal Protective Equipment (ie. reflective outerwear, eye protection, hearing protection, respirators)
f)	Task analysis and job procedures

1.11.12 Mining Method Opportunities

Lalor mine is considering different opportunities to improve mining efficiencies:

• Autonomous operation of LHDs are currently being trialed from surface by tele-remote with changes to standard designs to allow isolation of autonomous areas and buffer storage (transfer raises) for in between shift mucking

• A main ore pass from 755 m level to 910 m level is planned for 2017 to reduce trucking time from the upper levels of the mine

• Alternative truck loading systems are being investigated as an alternative to LHD loading

• Stoping block design changes are being considered to allow box hole primary mucking and circle route loading of trucks

1.12 Recovery Methods

The Stall concentrator complex is located approximately 16 km east of the Lalor mine. Conventional crushing, grinding and flotation operations are used to process the ore. The nominal throughput rate will be expanded from the current 3,000 tpd rate to 4,500 tpd and the mill will operate 24 hours per day, 365 days per year, with scheduled downtime for maintenance as required.

The concentrator produces a copper concentrate with gold and silver credits and a zinc concentrate, both are shipped by truck to Flin Flon, from there the copper concentrate is loaded onto rail cars and shipped to third party smelters. Tailings from the flotation circuit will be utilized to produce a cemented paste backfill for use underground at Lalor. Tailings not required for paste backfill will continue to be pumped to the existing Anderson TIA.

Run of mine ore as large as 0.55 m in one dimension is withdrawn from the head frame ore bin at Lalor and is transported to a crushing plant located at the Chisel North mine. The crushing plant reduces the ore to a range of 10 to 15 cm and is transported to the Stall concentrator coarse ore bins or to a stockpile at the mill. A stockpile of 36,000 tonnes, equivalent to 8 days production at the future expansion rate, is required to blend high-grade zinc to ensure a more consistent zinc feed.

The Stall concentrator consists of the following areas: stockpiles, crushing and screening, grinding circuits, copper and zinc flotation process, and concentrate dewatering. Flotation tailings currently pumped to the Anderson TIA will be pumped to the future paste plant at Lalor or to the Anderson TIA depending on the demand for paste. Water for the ore-grade mineral extraction process utilizes two sources: fresh and reclaimed water. Approximately 25% of the water usage is withdrawn from the fresh water of Snow Lake while the remaining approximately 75% of the water is reclaimed from Anderson TIA.

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Engineering for the Stall concentrator expansion is currently underway and construction is planned to commence in the third quarter of 2017 with commissioning in the third quarter of 2018. This schedule is offset to the Lalor production ramp-up by about 6 months. Lalor is planned to be at a nominal mining rate of 4,500 tpd in the first quarter of 2018. The additional ore tonnage from Lalor is planned for either stockpiling or will be transported to the Flin Flon concentrator for processing.

1.13 Project Infrastructure

This section addresses the infrastructure facilities that support the current operation at the Lalor mine and process facilities with discussion on projects, expansions and improvements. The current infrastructure facilities include Lalor (underground, onsite and offsite), power, Stall concentrator, Anderson TIA. Projects, expansions and improvements are related to the Stall concentrator, paste plant, Anderson TIA and underground ore handling system.

1.13.1 Lalor Infrastructure

Lalor mine is designed to hoist 6,000 tpd combined ore and waste. Primary access to the mine is a concrete lined 6.9 m diameter production shaft with a secondary ramp access from surface through the Chisel North mine. Ore is hoisted to surface and trucked to the Chisel North site where it is crushed then hauled to the Stall concentrator for processing into two concentrates zinc and copper.

General area infrastructure includes provincial roads and 115 kV Manitoba Hydro grid power to within four km of Lalor, and Manitoba Telecom land line and cellular phone service. The town of Snow Lake is a full service community with available housing, hospital, police, fire department, potable water system, restaurants and stores. The community is serviced by a 914 m gravel airstrip to provide emergency medical evacuation.

Lalor is located 3.5 km from the Hudbay Chisel North mine. Chisel North infrastructure includes a mined out open pit used for waste rock disposal, fresh (process) water sources, pumps and waterlines, 4160 V and 550 V power, mine discharge water lines, a 2,500 gallons per minute (gpm) water treatment plant with retention areas, plus mine buildings including offices and a change house.

The permitted Hudbay Anderson TIA, located approximately 12km from Lalor is used for tailings disposal.

Underground Infrastructure

• Main Production shaft 6.9 m diameter concrete lined with five compartments

• Three main shaft stations at 835 m, 910 m and 955 m levels

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• Lateral development consists of 6 m x 5 m ramps and level development

• Secondary egress consists of a ramp access from surface to the 810 m level where it joins the rest of the Lalor ramp system. Total distance is approximately 6.0 km.

• Power Distribution consists of 25 kV power down the shaft, 7.5 MVA 25 kV to 13.8 kV transformer to the 910 mL shaft station and primary distribution throughout the mine is 13.8 kV with transformers to 600 volt for local distribution

• Compressed air and process water are piped throughout the mine from surface

• Underground wireless radio communication throughout the mine is provided by a Leakey Feeder system

• Fiber-optic backbone for data and video

• Ore handling system consists of two rock breakers and bins (1,400 tonnes each) on 910 m level feeding chutes and conveyor system on 955 m level supplying the ore to the skips

• Mobile Maintenance shop is located in the Chisel North underground workings

• Discharge system consists of a series of drain holes and sumps with submersible pumps that feed the top of the two settling cones on 910 m level. Water from the 955 m level is pumped to surface by one 1,250 hp pump.

Onsite Infrastructure

• Office/change house complex with dry space for 341 personnel

• Hoisthouse consisting of the electrical distribution for the site, hoist and communication control room, production hoist, service hoist and compressors

• Headframe consisting of the utility hoist, bin house, external bunker, twin 250 hp downcast fans and mine heater

• Two 30,000 US gallon propane tanks

• Main pump station includes holding tanks and PAL water system and pumps for discharge, potable, process and fire water

• Bio Disk Sewage treatment plant for up to 38 m³/day

• Fuel tanks and pumps for diesel and gas.

• Two backup generators one for the utility hoist and one for the main pump station

• Temporary offices for health safety, training and mine rescue, temporary change house for contractors on site and temporary trailer for onsite contractor.

• Warehouse/shop

• Vent shaft and two exhaust fans

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Offsite Infrastructure

• 198 person camp in the Town of Snow Lake to house out of town personnel

• 3.5 km gravel access road connecting Provincial Road 395 to the mine site

• Two 24MVA - 115 kV to 25 kV power substations located on the Chisel North site

• The Chisel North complex is used for the diamond drilling core processing facility, shop to maintain the surface equipment fleet and offices for the project group

• Crushing of the Lalor ore is done at the Chisel North site with a maximum total stock pile capacity of 15,000 tonnes

• Booster pump station at Chisel North with holding tanks and pumps for process and discharge water. Equipped with a backup generator (350 kW).

• Two downcast fans, mine heaters with 30,000 US gallon propane tank each

• Discharge water from the Lalor site is pumped 3.5 km to the booster pump station at the Chisel North site where it is pumped the remaining 3.5 km to the settling ponds by the Chisel Pit

• A High Density Sludge Process Acidic Water Treatment Plant that can treat up to 2,500 gpm prior to being discharged to the environment

Ore Handling Improvements

Based on a review of Lalor’s underground ore handling circuit by Stantec in early 2017 Hudbay is planning capital improvements in 2017. These improvements will ensure Lalor is able to maintain a steady 4,500 tpd of production through the ore circuit. These improvements pertain to maintenance repair and replacement of liners in the ore circuit and to reduce potential hang-ups in the system.

1.13.2 Stall Concentrator

Hudbay operates the Stall concentrator approximately 16 km from Lalor. The mill is currently operating seven days per week at 3,000 tpd, processing ore from the Lalor mine. The mill has two circuits, with design capacities of 909 tpd and 2,182 tpd. The concentrator has two flotation circuits producing a zinc concentrate and a copper concentrate. The tailings associated with the Lalor mine are deposited in the Anderson TIA.

Produced copper concentrate is currently hauled by 40 ton trucks to Flin Flon, where the concentrate is loaded onto gondola rail cars for market. Produced zinc concentrate is hauled by 40 ton trucks to Flin Flon and is processed at Hudbay’s zinc processing facilities.

1.13.3 Stall Concentrator Expansion

The Stall concentrator expansion project to 4,500 tpd, slated for commissioning in third quarter of 2018, addresses the following areas: crushing, copper and zinc flotation, thickening, dewatering, electrical power distribution, compressed air systems and water distribution systems that will enable the concentrator to operate at 4,500 tpd.

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1.13.4 Anderson TIA

Anderson TIA is located in the Snow Lake area between the Stall concentrator and Lalor mine. The purpose of the Anderson TIA is environmental management (storage) of mine tailings produced in the Stall concentrator. Hudbay has submitted a Notice of Alteration to Manitoba Sustainable Development to expand the TIA within the existing limits to accommodate the future tailings produced through the entire Lalor mine operations. The construction of this expansion is anticipated to be complete prior to 2019.

1.13.5 Paste Plant

The Lalor paste plant project was approved in February 2017 and is critical for the sustainability of the mine production plan. The paste plant will be located northeast of the existing headframe complex and delivery capacity of the paste is 165 tph solids (tails) or 93 m³/hr paste. The paste plant is designed to fill voids left by mining of approximately 4,500 tpd. Taking into account waste generated from development in the LOM and the plan not to hoist waste from underground the combined paste/waste backfilling capacity is approximately 6,000 tpd. The paste plant will be capable of varying the binder content in the paste to provide flexibility in the strength gain of the paste where higher and early strength may be required depending on mining method.

Tails that are currently pumped from the Stall concentrator to the Anderson TIA will be diverted to the Anderson booster pump station. The tailings will be directed into the Anderson TIA when not required for the paste plant. Two pipelines will be installed between the Anderson booster pump station and the paste plant located at Lalor mine site, approximately a 13 km distance.

Paste will be delivered underground via one of two (nominal 8 inch diameter) cased boreholes from the surface to the 780 m level of the Lalor mine. The boreholes were drilled and cased in 2016. A network of underground lateral piping and level to level boreholes will transfer the paste from the base of the discharge hopper to the required underground locations. The majority of the underground distribution system will utilize existing drifts or planned future development.

1.14 Market Studies and Contracts

Hudbay has a marketing division that is responsible for establishing and maintaining all marketing and sales administrations of concentrates and metals. As well, Hudbay conducts ongoing research of metal prices and sales terms as part of normal business and long range planning process. Contract terms used in the Lalor financial evaluation are based on this research and the author has reviewed these results and they support the assumptions made in this technical report.

Lalor will produce a zinc concentrate and a copper concentrate with gold and silver credits. Zinc concentrates are trucked to Hudbay’s operations in Flin Flon where they are processed into refined zinc and sold to customers in North America. The key long-term assumptions for the sale of Lalor’s zinc metal and zinc concentrate are summarized in Table 1-16. This Technical Report assumes zinc concentrate will be processed at the Flin Flon zinc plant from 2017-2021 and after that time Lalor’s zinc concentrate will be sold to third party refineries.

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TABLE 1-16: KEY LONG-TERM ZINC METAL AND ZINC CONCENTRATE ASSUMPTIONS

	Units	LT Total / Average
Zinc Concentrate Grade	%	51%
Moisture Content of Zinc Concentrate	% H2O	9%
Zinc Concentrate Base Treatment Charge	US$/ tonne concentrate	$200
Zinc Concentrate Metal Price Basis	US$ / tonne Zinc metal	$2,204.6
Zinc Concentrate Escalator	%	6%
Zinc Concentrate De-escalator	%	3%
Zinc Concentrate Payability	%	85%
Zinc Concentrate Minimum Deduction	%	8%
Zinc Concentrate Freight Cost	C$/wmt	$118
Freight Allowance/Capture	US$/ wmt concentrate	$40
Zinc Metal Premium	US$/lb	$0.07
Zinc Metal Distribution Cost	US$/lb	$0.055

The copper concentrate produced at Lalor is sent to copper smelters in North America by rail. The key assumptions for the sale of Lalor’s copper concentrate are summarized in Table 1-17 below:

TABLE 1-17: KEY LONG-TERM COPPER CONCENTRATE ASSUMPTIONS

	Units	LT Total / Average
Copper Grade in Copper Concentrate	% Cu	21%
Moisture Content of Copper Concentrate	% H2O	9%
Copper Concentrate Base Treatment Charge	US$ / dry tonne con	$80
Copper Refining Charge	US $ / lb Cu	$0.08
Silver Refining Charge	US $ / oz Ag	$0.50
Gold Refining Charge	US$ / oz Au	$5.00
Copper Concentrate Freight Cost	C$ / wet tonne con	$213
Copper Payability	%	96.5%
Copper Minimum Deduction	%	1%
Gold Payability	%	96%
Silver Payability	%	90%

Engineering, supply and construction contracts are initiated, managed and administrated by Hudbay’s Manitoba Business Unit. Hudbay follows a standard contracting out process that specifies contractors’ requirements to be eligible to be considered for work. Contractor selection criteria include ability to complete the work within the required time, safety record and programs, price, and proposed alternatives. The Lalor contracts that are in place have rates and charges that are within industry norms.

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1.15 Environmental Studies, Permitting and Social or Community Impact

Commencing in 2007, AECOM carried out the environmental baseline investigations needed to conclude an environmental impact assessment for the Lalor project, including all necessary terrestrial and aquatic field studies. This baseline work was utilized in the Lalor Paste Plant Notice of Alteration (NoA) which was submitted to Manitoba Sustainable Development in the fourth quarter of 2016 and was approved in January 2017. AECOM also conducted baseline work and studies which are summarized in the Anderson TIA expansion NoA submitted in the third quarter of 2016 to Manitoba Sustainable Development for approval.

Due to the extensive work completed by AECOM and other existing studies completed as part of Environmental Effects Monitoring (EEM) programs at the various operations in the Snow Lake area, it is contemplated that no additional baseline studies are necessary for potential future improvement projects. There is no present indication that future approvals will not be obtained to meet potential future construction schedules.

There are no known environmental concerns which could adversely affect Hudbay’s ability to mine ore from Lalor mine. Because of its location in close proximity to the existing facilities in the Snow Lake area, Lalor was able to utilize existing infrastructure, services, and previously disturbed land that is associated with permitted, pre-existing and current mining operations in the Snow Lake area. The Lalor mine and associated projects are designed to minimize the potential impact on the surrounding environment by keeping the footprint of the operations as small as possible and by using existing licensed facilities for the withdrawal of water and disposal of wastes.

The NoA for the expansion of Anderson TIA was prepared by AECOM utilizing geotechnical tailings dam designs from Hudbay’s Engineer of Record; BGC Engineering Inc. As detailed in the Environmental Assessment of the Proposed TIA Expansion submitted as a NoA, the entire volume of tailings from Lalor LOM was to be stored in Anderson TIA (AECOM, 2016). This conceptual design did not discount the volume of tails that could be used for paste backfill at Lalor.

The existing Lalor mine Environment Act Licence (EAL) was obtained in the first quarter of 2014 and covers all facilities on the Lalor site, including sewage and mine wastewater treatment facilities and the pipelines which carry freshwater into the site and remove treated wastewater from it. The main permits required for the Lalor operation are presently valid licenses and permits for the Lalor mine, Stall concentrator, Anderson TIA, and New Britannia site. Applications for the Anderson TIA expansion NoA have been submitted for approval. Other upgrades and augmentation plans may require the submission of a NoA to an existing licensed operation but no new tailings impoundment area will be required. No federal permits are anticipated.

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Presently the New Britannia site inclusive of the Birch Lake Tailings Management Facility (BLTMF), in Snow Lake, although currently in care and maintenance, has a current EAL and the seasonal discharge from BLTMF is regulated under the Metal Mining Effluent Regulations (MMER). Hudbay is currently in the process of applying for a new water withdrawal licence for this site which is anticipated to be obtained before potential operational needs. Potential future use of the New Britannia site will require the submission of a NoA in order to process material from the Lalor mine.

Prior to commercial production of ore at Lalor mine a mineral lease was applied for and obtained from Manitoba Mines Branch. The mineral lease grants the holder the exclusive right to mine minerals within the lease area.

In specific areas associated with proposed pipeline routes and future improvements to the existing Anderson TIA, surface leases will be required. Activities are currently underway to apply for and obtain the required surface leases. There is no indication that theses leases cannot be obtained in the time lines of the expansion project.

The main settlement in the region of the Lalor mine is the town of Snow Lake, which is an important mining and service centre for the Ecodistrict and surrounding area. Snow Lake has a population of approximately 840 according to the 2006 data from Statistics Canada, with the majority of these residents employed at or supported by the mines located throughout the area. Many other Snow Lake residents are employed in the industries and services that support the region’s mining operations.

Hudbay and AECOM have carried out public consultation, including meetings to inform local communities about the progress of development of the Lalor mine and expansion of Anderson TIA and environmental effects of these projects. The projects will continue to provide jobs for both Flin Flon and the Town of Snow Lake during construction of upgrades and continued operation of the mine. The additional feed from the mine will also help ensure the continued employment of Hudbay employees in the Flin Flon and Snow Lake areas. Since the economies of both communities are based on mining, opposition to the projects is seen as unlikely.

Based on Hudbay’s long-term (more than 50 years) mining experience in the Snow Lake region, there is no known current First Nation or Aboriginal hunting, fishing, trapping or other traditional use in the zone of potential influence for the Lalor mine, other current operations, and potential future projects.

Operation of the Lalor mine and construction of potential future upgrades will not affect any known site of potential historical, archaeological or cultural significance.

The Manitoba Mines and Minerals Act requires a closure plan and financial assurance for any advanced exploration or mining project. Manitoba accepted Closure Plans prepared by SRK in 2005 and financial assurance to cover the cost of closure for all existing infrastructure that will continue to be used during operation of the Lalor mine. Existing facilities which support the Lalor mine include the Chisel North mine, which is connected by an underground ramp to Lalor, Stall concentrator and Anderson TIA, piping systems associated with milling and tailings deposition, the Chisel Open Pit and the Chisel North water treatment plant. Prior to commercial production at Lalor mine, Manitoba approved the Closure Plan for the Lalor AEP and accepted financial assurance in the amount of $1.5 million. NoA applications for the paste plant, expansion of the Anderson TIA, and potential upgrades to the New Britannia site also will require the submission of updated closure plans and financial assurance.

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1.16 Capital and Operating Cost

Capital and operating costs are estimated in constant 2017 Canadian dollars.

1.16.1 Capital Costs

The total development capital required to increase throughput at Lalor to the targeted 4,500 tpd is estimated to be C$117 million, as shown in Table 1-18, which includes approximately an 18% contingency. This capital is expected to be spent during 2017 and 2018.

TABLE 1-18: DEVELOPMENT CAPITAL COST SUMMARY

Development Capital	000 C$
Paste Backfill	67,786
Ore Handling Underground	3,250
Stall Mill Upgrades	45,870
Total Development Capital	116,906

The development capital costs were estimated internally by Hudbay with input from Golder Associates, Stantec Inc. and Boge & Boge Ltd.

The LOM sustaining capital costs are estimated to be C$220 million. The breakdown of the sustaining capital over the next 5 years and for the LOM is shown in Table 1-19.

TABLE 1-19: SUSTAINING CAPITAL COST SUMMARY

Sustaining Capital (000 C$)	2017	2018	2019	2020	2021	2017-LOM
Mine Capital and Development	17,927	18,381	13,442	13,849	10,635	86,771
Normal Capital	3,000	10,000	3,000	3,000	3,000	28,000
Replacement Equipment	11,629	14,739	13,390	11,004	9,776	91,518
Major Installations	3,640	5,924	1,461	982	835	13,803
Total Sustaining Capital	36,196	49,044	31,293	28,835	24,246	220,092

Normal capital includes C$7 million related to the expansion of the Anderson tailings facility and C$21 million related to the Stall concentrator. This sustaining capital estimate has made no consideration for the availability of used equipment from the 777 and Reed mines. When these mines eventually close, some equipment may be available for use at Lalor and may reduce the sustaining capital estimate above. No contingency has been included in the sustaining capital estimate.

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Reclamation costs, salvage value and severance costs have not been considered in this report.

1.16.2 Operating Costs

Operating costs were developed by Hudbay based on a bottom-up approach and utilizing budget quotes from local suppliers, Manitoba operations experience, labor costs within the region and actual costs at Lalor.

The mine plus mill unit operating costs are estimated to be C$99.83/tonne mined over the LOM. The addition of cemented rock fill and paste backfill has increased the mine unit costs, but maximizes recovery of the mineral resource and results in lower capitalized costs than prior years due to less underground development. Table 1-20 summarizes the mine plus mill unit operating costs over the next 5 years and for the LOM.

TABLE 1-20: UNIT OPERATING COST SUMMARY

(C$/tonne mined)	2017	2018	2019	2020	2021	2017-LOM
Mine Development	26.53	15.07	14.87	13.64	13.98	17.12
Ore Extraction	20.28	27.19	33.60	35.59	34.75	30.50
Ore Removal	29.87	30.24	28.22	28.20	28.47	30.70
Total Mine Operating Costs	76.67	72.49	76.68	77.43	77.20	78.32
Mill Operating Costs1	22.02	20.14	20.12	20.19	20.12	21.51
Total Mine + Mill Operating Costs	98.70	92.63	96.80	97.63	97.32	99.83

¹ Milling costs include concentrate haulage to Flin Flon

The total C1 cash costs and sustaining cash costs (net of by-product credits) per pound of zinc over the LOM and over the next 5 years are shown in Table 1-21. C1 cash costs include on-site and off-site costs. Sustaining cash costs include C1 costs plus sustaining capital.

TABLE 1-21: CASH COSTS (NET OF BY-PRODUCT CREDITS)

Cash Costs (Net of By-Product Credits1)	Units	Next 5 Years	LOM
C1 Cash Costs	US$ / lb Zn in con	$0.41	$0.37
C1 Cash Costs + Sustaining Capex	US$ / lb Zn in con	$0.57	$0.50

¹ By-product credits are calculated using the following assumptions: copper price per pound - US$2.60 in 2017, US$2.75 in 2018, US$3.00 in 2019 to 2020 and long-term; gold price per ounce - US$1,300 in 2017 to 2020 and US$1,260 long-term; silver price per ounce - US$18.00 in 2017 to 2020 and long-term; CAD/USD exchange rate - 1.35 in 2017, 1.25 in 2018, 1.20 in 2019, 1.15 in 2020 and 1.10 long-term.

Lalor’s annual zinc production (contained zinc in concentrate) and C1 cash costs (net of byproducts) are shown below in Figure 1-2. Over the first 5 years, annual production is expected to average 79 thousand tonnes of zinc at an average C1 cash cost of US$0.41/lb. Over the 10.5 year LOM, annual production is expected to average 64 thousand tonnes of zinc at an average C1 cash cost of US$0.37/lb. Lower C1 cash costs from years 2020 to 2023 are a result of mining the copper-gold (Zone 27).

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FIGURE 1-2: LALOR ANNUAL ZINC PRODUCTION AND C1 CASH COSTS

1.17 Economic Analysis

Hudbay is a producing issuer and has excluded information required by Item 22 of Form 43-101F1 as the updated mine plan does not represent a material increase of Hudbay’s current production.

1.18 Other Relevant Data and Information

1.18.1 Gold Bulk Sample Program

In the fourth quarter of 2016, Hudbay developed 37 drift rounds at Lalor to assess the continuity and variability of non-contact gold mineralization within discrete areas of zones 21 and 25. Approximately 10,000 tonnes of bulk sample material was mined and hauled to surface. The material was primary crushed and processed through a sample tower to collect a representative subsample of each development round. The integrity of the material was maintained at all times through a rigorous chain of custody process. The mined material, stored on surface, is available for milling pending the potential economic viability of refurbishing the New Britannia gold mill in Snow Lake by Hudbay.

The preliminary results indicate that the gold grades from the bulk sample program are as expected with minor variations when compared to those modeled based on diamond drill data. The bulk sample program has increased the confidence and the understating of the gold zones and gold mineralization at Lalor.

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Following full assessment of the 2016 bulk sample data it is intended to collect additional subsamples in 2017 from other areas of gold zones 21 and 25 as well as from other previously untested gold zones for confirmation purposes.

1.18.2 Taxes and Royalties

Applicable Tax Rates

The Lalor mine is not directly taxable as Hudbay pays provincial and federal taxes on a legal entity basis. The combined federal and provincial tax rates are assumed to be approximately 27% for the LOM and Hudbay has approximately C$750 million in tax pools that can be used to offset future income taxes for federal and provincial purposes. Hudbay’s mining operations in Manitoba are also subject to the Manitoba Mining Tax. The Manitoba Mining Tax is not applied to a new mining project until the original capital expenditures are recovered.

Royalties

There are no royalties applicable to Lalor.

1.19 Conclusions and Recommendations

1.19.1 Conclusions

The Lalor mine operation has been mining ore since August 2012. Since then the mine has operated uninterrupted and been in a continuous production ramp-up cycle, achieving the highest annual tonnage of approximately 1.1 million tonnes in 2016, with complementary throughput at the Stall concentrator. The production ramp-up is planned to continue in 2017 to reach a steady state of 4,500 tpd by the first quarter of 2018.

The production increase of 50%, compared to current production, is supported by an underground ore handling circuit capable of 4,500 tpd, transitioning to more bulk mining methods (65% of reserves) with additional mining fronts and design changes to improve mining efficiencies, developing ore passes and transfer raises to reduce truck haulage cycle times from the upper potions of the mine and commissioning of a paste plant backfill plant in the first quarter of 2018. Autonomous operation of a Load Haul Dump loader underground is currently being trialed from surface by tele-remote monitoring with changes to standard designs to allow isolation of autonomous areas and buffer storage for in between shift mucking.

The increase in production to 4,500 tpd at Lalor is complemented by the Stall concentrator expansion to 4,500 tpd, which is currently underway and is expected to be commissioned in the third quarter of 2018.

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The mineral resources, as of September 30, 2016, are estimated as base metal lenses or gold zones based on geological and mineralization properties. The Hudbay validation process and third party review confirmed the resource block model is interpolated using industry accepted modelling techniques and classified in accordance with the 2014 CIM Definition Standards – For Mineral Resources and Mineral Reserves.

A mine reconciliation of the mined out areas compared to the ore reported at the concentrator was very close on the precious metals and a slight conservatism of the zinc and copper grades might be evident. This conservatism of the base metals is likely due to over constraining the high grade samples to 20 m as part of the high yield restriction step.

The mineral resources stated with a metal equivalency cut-offs provide for economic extraction of reserves from stated resources.

The mineral reserves, as of January 1, 2017, are based on a LOM plan that generated a mining inventory based on stope geometry parameters with appropriate dilution and recovery factors The conversion of resources to reserves is based on the LOM plan and NSR cut-offs that primarily focussed on capturing base metal resources for processing at the Stall concentrator. The secondary focus was to capture gold zone resources when in contact with or close proximity to base metal resources. In areas where a large separation existed between base metal and gold lenses, mining blocks were evaluated for economic stope mining shapes. When a non-economic shape was generated in a first pass, a second pass was evaluated for only base metal lenses and if an economic shape was generated the gold zone portion was removed. However, due to this larger separation, majority of these isolated gold lenses could have been evaluated independently of the base metal lenses and could potentially provide feed to a gold processing facility. Below approximately the 950 m level no attempt was made to generate an economic stope mining shape for gold zones 25 and 26 as the separation distance became too large. The author’s opinion is that these resources are potentially better suited for a gold processing facility and should be re-evaluated when Hudbay has a better understanding of their New Britannia gold mill and Birch Tailings Impoundment Area in Snow Lake.

The author considers that the mineral reserves as classified and reported comply with all disclosure in accordance with requirements and CIM Definitions. The author is not aware of any mining, metallurgical, infrastructure, permitting or other relevant factors that could materially affect the mineral reserve estimate.

The production and compilation of this technical report was supported by the capable and professional management and staff at Hudbay. The supervision, revision and approval of the assembly of this Technical Report is by the QP Robert Carter, P. Eng., Lalor Mine Manager, Hudbay Manitoba Business Unit.

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1.19.2 Recommendations

Hudbay recommends the following:

• Investigate the impact of under assaying of high grade standards at Hudbay’s Flin Flon laboratory and whether this in turn affects the high grade Lalor samples submitted and has potentially led to an underestimations of gold in the resource estimate, since the proportion of samples assayed at the laboratory was approximately 80% of the total samples assayed between 2012 and 2016.

• Investigate the high yield restriction parameters of the high grade base metal samples, and consider whether the restriction distance is suitable or were they over constrained, based on the conservatism noted in the mine reconciliation for zinc and copper.

• Pursue the option of a temporary paste backfill plant to utilize the boreholes from surface prior to commissioning of the permanent plant in the first quarter of 2018. This option provides assurance to achieve the ramp-up in production and is another source of backfill rather than relying on waste development.

• Due to the approximate 6 month timing offset of the production ramp-up at Lalor to 4,500 tpd and the Stall concentrator expansion to 4,500 tpd, Hudbay should pursue transporting of ore from Lalor to their Flin Flon concentrator for earlier processing.

• Finalize the evaluation of the gold bulk sample program conducted in the fourth quarter of 2016 and since Hudbay owns a sample tower consider collecting additional subsamples from other areas of gold zones 21 and 25 as well as from other previously untested gold zones for confirmation purposes.

• Hudbay owns the New Britannia mill, a gold leach plant on care and maintenance, in Snow Lake. Hudbay should continue to assess the feasibility of processing a portion of the material mined from the gold zone and copper-gold zone at Lalor at the New Britannia mill at a rate of 1,500 tpd starting in 2019. When combined with the processing capacity of the Stall concentrator, this would enable an aggregate throughput rate of up to 6,000 tpd and utilize the full capacity of the Lalor mine shaft.

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2 INTRODUCTION AND TERMS OF REFERENCE

This Technical Report has been prepared for Hudbay to support the public disclosure of Mineral Resources and Mineral Reserves at the Lalor Mine and to provide an updated mine plan that contemplates 4,500 tpd of base metal, gold and copper-gold zone ore to Stall concentrator.

Hudbay is a Canadian integrated mining company with assets in North and South America, principally focused on the discovery, production and marketing of base and precious metals. Hudbay’s objective is to maximize shareholder value through efficient operations, organic growth and accretive acquisitions, while maintaining its financial strength.

Hudbay operates multiple properties in the Province of Manitoba. Operations near Flin Flon include the 777 Mine, which also consists of an ore concentrator and zinc plant, and the Reed mine, which is located approximately 120 km by road southeast of Flin Flon. Operations near Snow Lake include the Lalor underground mine.

The Lalor mine consists of an ore concentrator, a tailings impoundment area and other ancillary facilities that support the operation. The property is located approximately 16 km by road west of the town of Snow Lake, Manitoba. Hudbay owns 100% interest in the Lalor property through one Mineral Lease and eight Order In Council Leases to the south of the mine.

As of the issue date of this Technical Report, the Lalor mine is operating at a processing rate of approximately 3,000 to 3,500 tpd and is ramping up production to 4,500 tpd by 2018. The production ramp-up is supported by the underground ore handling circuit capable of 4,500 tpd, further transitioning to more bulk longhole open stoping mining methods with additional mining fronts and design changes to improve mining efficiencies, developing ore passes and transfer raises to reduce truck haulage cycle times from the upper potions of the mine and commissioning of a paste backfill plant in the first quarter of 2018. Autonomous operation of Load Haul Dump loaders underground are currently being trialed from surface by tele-remote monitoring with changes to standard designs to allow isolation of autonomous areas and buffer storage for in between shift mucking.

The increase in production to 4,500 tpd at Lalor is complemented by the Stall concentrator expansion to 4,500 tpd, which is currently underway and is expected to be complete in the second quarter of 2018.

This Technical Report represents an update of information pertaining to the Lalor mine previously disclosed in a Pre-Feasibility Study technical report dated March 2012. At the time of the previous technical report, Lalor was a development project, but the property has since commenced production starting in August 2012.

Major works completed at Lalor include the construction of a concrete-lined 6.9 m diameter production shaft capable of 6,000 tpd (which is the primary access), the development of a secondary ramp access from the surface through the Chisel North mine shaft, and the establishment of two shaft stations at 835 mL and 910 mL. The facilities at the Lalor mine have mined and processed more than 3 million tonnes of ore as of December 31, 2016.

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2.1 Qualified Person (QP) and Site Visit

This Technical Report has been prepared in accordance with National Instrument Form NI 43-101 F1. The QP who supervised the preparation of this Technical Report is Robert Carter, P. Eng., Lalor Mine Manager at Hudbay’s Manitoba Business Unit.

Robert Carter is not independent of Hudbay, and this is not an independent technical report. Nevertheless, Hudbay is a “producing issuer” as defined in NI 43-101. As such, this technical report is not required to be prepared by or under the supervision of an independent QP.

Mr. Carter is directly involved with the Lalor mine on a permanent basis because of his role as mine manager and was involved in the early exploration, project evaluation and pre-feasibility study. Mr. Carter visits and inspects the Lalor operation on a routine basis and has overseen the mineral resource and reserve estimation process. Mr. Carter has acted as the Qualified Person for the overall Technical Report. Prior to the publication of this Technical Report, Mr. Carter’s last site visit was on March 29, 2017.

2.2 Sources of Information

Geology and mineral resources sources of information include: core drilling and sampling data, underground development and mapping, assay and geochemistry analysis.

Mineral reserve sources of information are the mineral resource block model, actual production and cost data, budget projections, life of mine inventory based on stope geometry parameters and mine development sequence.

Metallurgy, processing and economic sources of information are the actual operating data since production and concentrating commenced in 2012 and operating budget estimates.

Multiple participants have worked on this Technical Report. Discussions were held with personnel from Hudbay Manitoba Business Unit (MBU) and Hudbay Toronto:

• Tony Scheres, Manager of Technical Services and Business Development, Lalor mine

• Sarah Bernauer, Chief Geologist, Lalor mine

• Jennifer Pakula, Chief Engineer, Lalor mine

• Jason Lanteigne, Mine Engineer, Lalor mine

• Doug Salahub, Mines Analyst, MBU

• Johan Krebs, Geologist, Lalor mine

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• Jay Cooper, Superintendent Environment, MBU

• Jarid Medina, Stall Mill Superintendent, Stall concentrator

• Karl Hoover, Process Manager, Snow Lake Projects

• Marc-Andre Brulotte, Manager Project Evaluation, Toronto

• Mark Gupta, Manager Corporate Development, Toronto

• Juan Carlos, Ordóñez Calderón, Exploration Geochemist, Toronto

• Matthew Holden, Senior Geophysicists, Toronto

Table 2-1 lists the participants by section.

TABLE 2-1: CONTRIBUTORS AND RESPONSIBLE PARTIES FOR THIS REPORT

Section	Description	Participants	Responsible QP
1	Summary	Robert Carter	Robert Carter
2	Introduction	Robert Carter	Robert Carter
3	Reliance on Other Experts	Robert Carter	Robert Carter
4	Property Description and Location	Johan Krebs, Sarah Bernauer	Robert Carter
5	Accessibility, Climate, Local Resources, Infrastructure and Physiography	Johan Krebs	Robert Carter
6	History	Johan Krebs	Robert Carter
7	Geological Setting and Mineralization	Johan Krebs	Robert Carter
8	Deposit Type	Johan Krebs	Robert Carter
9	Exploration	Matthew Holden	Robert Carter
10	Drilling	Johan Krebs	Robert Carter
11	Sample Preparation Analyses and Security	Johan Krebs, Sarah Krebs	Robert Carter
12	Data Verification	Juan Carlos, Ordóñez Calderón	Robert Carter
13	Mineral Processing and Metallurgical Testing	Jarid Medina, Karl Hoover	Robert Carter
14	Mineral Resource Estimates	Marc-Andre Brulotte, Robert Carter	Robert Carter
15	Mineral Reserve Estimates	Jason Lanteigne, Robert Carter	Robert Carter
16	Mining Methods	Jennifer Pakula, Jason Lanteigne	Robert Carter
17	Recovery Methods	Jarid Medina	Robert Carter
18	Project Infrastructure	Tony Scheres, Jarid Medina, Jay Cooper, Robert Carter	Robert Carter
19	Market Studies and Contracts	Mark Gupta	Robert Carter

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Section	Description	Participants	Responsible QP
20	Environmental Studies, Permitting, and Social or Community Impact	Jay Cooper	Robert Carter
21	Capital and Operating Costs	Mark Gupta, Doug Salahub	Robert Carter
22	Economic Analysis	-	Robert Carter
23	Adjacent Properties	-	Robert Carter
24	Other Relevant Data and Information	Robert Carter	Robert Carter
25	Interpretation and Conclusions	Robert Carter	Robert Carter
26	Recommendations	Robert Carter	Robert Carter
27	References	Robert Carter	Robert Carter

2.3 Unit Abbreviations

Units of measurement in this report conform to the SI (metric) system unless otherwise noted. Table 2-2 lists the notable unit abbreviations utilized in this report.

TABLE 2-2: UNIT ABBREVIATIONS

Abbreviation	Term	Abbreviation	Term
$C or C$	Canadian dollars	M	million
%	Percent	m ASL	metres above sea level
°C	degree Celsius	m2	squared metre
µm	micrometre or micron	m3	cubic metre
BTU	British thermal unit	m3/hr	cubic metre per hour
CFM or cfm	Cubic feet per minute	mL	metre level
dmt	Dry metric tonnes	mm	millimetre
g	gram	ml	millilitres
g/t	gram per metric tonne	MVA	Mega volt amp
Ga	billion years	MW	Megawatt
gpm	gallon per minute	nT	nanotesla
Ha	hectare	oz	Troy ounces
HP or hp	Horsepower	ppm	parts per million
hr	hour	psi	Pounds per square inch
kg	kilogram	QP	Qualified Person
km	kilometre	t	metric tonne
km/hr	kilometre per hour	tpd	metric tonnes per day
kV	kilovolt	US GPM	United States gallon per minute
kW	kilowatt	US$ or $US	United States dollar
L/min	litres per minute	V	Volt
lb	pound (unit of weight)	wmt	Wet metric tonne
m	Metre	Zn Eq	Zinc equivalent

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2.4 Acronyms and Abbreviations

Abbreviations of company names and other notable terms used in the report are as shown in Table 2-3.

TABLE 2-3: ACRONYMS AND ABBREVIATIONS

Abbreviation	Term	Abbreviation	Term
3D	Three-Dimensional		System
AAS	Atomic Absorption Spectrometry	EAL	Environmental Act Licence
ACME	ACME Analytical Laboratories	EDA	Exploratory data analysis
	Ltd.	EDM	Electronic distance
acQuire	Drillhole Database Management		measurement
	Program	EEM	Environmental Effects
ADIST	Average distance of composites		Monitoring
AEP	Advanced Exploration Project	EM	Electromagnetic
Ag	Silver	ES	Emission spectrograph
ANFO	Ammonium nitrate	Fe	Iron
As	Arsenic	FOB	Fine ore bin
ASL	Above Sea Level	GPSS	Global Positioning Satellite
Au	Gold		System
Au Eq	Gold Equivalency	HCl	Hydrogen chloride
AV	average	HNO3	Nitric acid
		Hudbay	Collectively all Hudbay Minerals
BLTMF	Birch Lake Tailings		Inc. subsidiaries and business
	Management Facility		groups
BQ	BQ drill core size 36.4mm	ICP	Inductively Coupled Plasma
BV	Bureau Veritas	LCT	Lock cycle test
CBV	Certified Best Value	IDW	Inverse Distance Squared
CDIST	Closest distance of a composite		Weighted
CIM	Canadian Institute of Mining,	KSTD	Standard deviation of kriging
	Metallurgy and Petroleum	LHD	Load Haul and Dump
CIP/CIL	Carbon-in-pulp / Carbon-in-	LIMS	Information management
	leach		system
CI	Confidence interval	LOD	Limit of detection
CRF	Cemented waste rock backfill	LOM	Life of Mine
CRM	Certified Reference Materials	MBU	Manitoba Business Unit
Cu	Copper	MCCN	Mathias Colomb First Nation
CV	Coefficient of Variation	MDIST	Maximum distance of a
DH	Drill hole		composite
DE	Data entry	MIBC	Methyl isobutyl carbinol
DGPS	Differential Global Positioning	ML	Mineral Lease

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Abbreviation	Term	Abbreviation	Term
MS	Mass Spectrometry	RMA	Reduced-to-Major-Axis
NCOMP	Number of composites	Reflex	Reflex E-Z Shot
NoA	Notice of Alteration	ROM	Run of mine
NI	National Instrument	RSLOP	Regression slope
NN	Nearest Neighbour	R squared or	Coefficient of determination
NSR	Net smelter return	R2
NQ	NQ drill core size 47.6mm	RTK	Real Time Kinematic
NSS	Near solid sulphide	SG	Specific Gravity
OIC	Order In Council	SS	Solid sulphide
OES	Optical Emission Spectrometry	TIA	Tailings Impoundment Area
OK	Ordinary Kriging	RE	Relative Error
OLS	Ordinary least square	TMI	Total Magnetic Intensity
OREAS	Ore Research and Exploration	UCS	Unconfined compressive strength
P. Eng.	Professional Engineer	URF	Unconsolidated waste rock
Pb	Lead		backfill
PR	Provincial Road	UTM	Universal Transverse Mercator
QAQC	Quality Assurance and Quality	VMS	Volcanogenic Massive Sulphide
	Control	Zn	Zinc
QP	Qualified Person	Zn Eq	Zinc Equivalency
RCI	Resource classification index

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3 RELIANCE ON OTHER EXPERTS

Standard professional procedures were followed in preparing the contents of this Technical Report. Data used in this report has been verified where possible and the author has no reason to believe that the data was not collected in a professional manner and no information has been withheld that would affect the conclusions made herein.

The information, conclusions, opinions, and estimates contained herein are based on:

• Information available to Hudbay at the time of preparation of this Technical Report,

• Assumptions, conditions, and qualifications as set forth in this Technical Report

For the purpose of the report, the author has relied on title and property ownership information provided by Hudbay Manitoba Business Unit, Land Manager, Janelle Toffan on March 13, 2017.

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4 PROPERTY DESCRIPTION AND LOCATION

The Lalor mine is located approximately 208 km by road east of Flin Flon and 16 km by road west of Snow Lake in the Province of Manitoba at 54°52’N latitude, 100°08’W longitude and 303 m ASL (Figure 4-1).

Hudbay owns 100% interest of the Lalor property through one (1) Mineral Lease (ML) and eight (8) Order in Council (OIC) Leases to the south (Figure 4-2).

4.1 Land Tenure

Mineral and Order in Council Leases are issued and administered by the Province of Manitoba Mines Branch. Annual payments of $10.50/ha, with a $193 minimum, for producing leases is due for each lease over the 21-year term.

ML334:

• ML334 was issued in 2012 and replaces mineral claims CB5361, CB10605, CB10606, CB10607 and CB10608 (Figure 4-2).

• Covers an area of approximately 796 hectares and encompasses the majority of the Lalor deposit.

• The required amount of expenditures, on approved work, within the ML334 area shall be no less than $1,250/ha during the 21-year term if a term renewal will be required. No later than 60 days after each of the 5^th, 10^th, 15^th and 21^st anniversaries of the issuance of the Mineral Lease, the Mineral Lessee shall submit a report to the Director of Mines setting out all work carried out on the mineral lease area for the applicable period.

OIC LEASES (M5778, M5779, M5780, M5781, M7278, M7279, M7280 and M7281):

• Covers an area of approximately 152 hectares. The southerly up-plunge extension of the mineralization lies within these OIC Leases.

• In addition to the annual payments mentioned above, an annual tax of $10.00/OIC is due by December 31^st of each year.

• There are no work commitments due on OIC Leases.

Mineral and Order in Council Lease status for the Lalor property is shown in Table 4-1.

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TABLE 4-1: MINERAL AND OIC LEASE PROPERTIES

Lease No.	Lease Name	Holder	Hectares	Annual Fees (excludes annual $10/OIC tax)	Anniversary Date	Termination Date
M5778	OX NO. 153	Hudbay	15.90	193.00	Apr 08, 2017	Apr 08, 2023
M5779	OX NO. 154	Hudbay	17.99	193.00	Apr 08, 2017	Apr 08, 2023
M5780	OX NO. 155	Hudbay	18.25	199.50	Apr 08, 2017	Apr 08, 2023
M5781	OX NO. 156	Hudbay	20.20	220.50	Apr 08, 2017	Apr 08, 2023
M7278	OX NO. 143	Hudbay	21.70	231.00	Sep 06, 2017	Sep 06, 2023
M7279	OX NO. 144	Hudbay	20.55	220.50	Sep 06, 2017	Sep 06, 2023
M7280	OX NO. 145	Hudbay	21.60	231.00	Sep 06, 2017	Sep 06, 2023
M7281	OX NO. 146	Hudbay	14.84	193.00	Sep 06, 2017	Sep 06, 2023
ML334		Hudbay	795.55	8,358.00	Mar 29, 2018	Mar 29, 2033
Totals			946.58	$ 10,039.50

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FIGURE 4-1: LOCATION MAP OF HUDBAY MINES

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FIGURE 4-2: MINERAL CLAIMS AND LEASE MAP

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4.2 Land Use Permitting

4.2.1 General Permits

General Permits are issued and administered by the Province of Manitoba Crown Lands and Property Agency. Provided all terms and conditions of the General Permit are met, including payment of annual fees, the permit is automatically renewed for a 1-year term on an annual basis.

Two General Permits (GP59093 and GP63483) and one Quarry Lease are held by Hudbay and are required to carry on mining activities at Lalor:

GP59093:

• Specific Use: 4.0 km x 5 m wide all weather access road (to accommodate a 4 km 25 kV transmission line, a 4 km discharge line and a 4 km fresh water line), a 200 m x 200 m parking lot and an additional access road in PT. NW 9-68-18W (0.25 km x 30 m).

• The annual fee for the All Weather Road is set at $100 plus one additional dollar for every kilometre of road. The annual fee for the Commercial Lot (parking) for one acre or less is set at $210 plus an additional $10 for each additional acre or portion of an acre.

• Annual Permit Renewal Fee is set at $10/permit.

GP63483:

• Specific Use: Mine Site

• The annual fee for Commercial Lot (mine site) is set at $2/acre or portion of an acre.

• Annual Permit Renewal Fee is set at $10/permit.

4.2.2 Quarry Lease

Quarry Leases are issued and administered by the Province of Manitoba Mines Branch.

One Quarry Lease is held by Hudbay and is required to carry on mining activities at Lalor:

• QL-1928 was issued November 29, 2007 for a 10-year term and provides the holder with the exclusive right to explore for, develop, and produce clay, gravel, rock or stone. QL-1928 will expire November 2017 and provided this lease is still required it can be renewed for another 10-year term.

• The annual fee for a Quarry Lease is set at $27/ha, or fraction thereof, plus royalty and rehabilitation levies on extractions as prescribed by regulations.

General Permit and Quarry Lease status for the Lalor property is shown in Table 4-2.

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TABLE 4-2: GENERAL PERMITS AND QUARRY LEASE

Permit / Quarry Number	Holder	Hectares	Issue Dates	Annual Fees (Excludes GST)	Anniversary Date
GP59093	Hudbay	4.05	Dec 31, 2007	415.00	Dec 31, 2017
GP63483	Hudbay	159.37	Jun 10, 2010	798.00	Dec 31, 2017
QL-1928	Hudbay	11.00	Nov 26, 2007	297.00	Nov 26, 2017
Total		174.42		$ 1,510.00

Hudbay holds the exclusive right to the minerals, other than quarry minerals, and the mineral access rights required for the purpose of working the lands, mining, and producing minerals from the Lalor mine. Surface tenure, currently necessary to accommodate buildings and/or structures, required for the efficient and economical performance of the mining operations has been applied for and approved.

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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Accessibility

The Lalor deposit is located approximately 208 km by road east of Flin Flon and 16 km by road west of the community of Snow Lake, Manitoba. Access to the deposit is from Provincial Road (PR) #395, a gravel road off PR #392, which joins the town of Snow Lake and PR #39 (Figure 5-1). From PR #395 there is an all weather permanent road into the mine site.

5.2 Climate

The Snow Lake area has a typical mid-continental climate, with short summers and long, cold winters. Climate generally has only a minor effect on local exploration and mining activities.

The nearest Environment Canada weather station is located near Baker’s Narrows at the Flin Flon airport, approximately 16 km southeast of Flin Flon, and approximately 100 km west of the Lalor deposit. The average annual temperature at the Baker’s Narrows weather station is 0.1°C. The average summer temperature is approximately 17°C, and the average winter temperature is -14°C. The lowest monthly average temperature occurs in January at -21.1°C, and the highest monthly average temperature is in July at 18.3°C. Freeze-up of small bays and lakes occurs in mid-November, with breakup occurring in mid-May. There is an average of 115 frost-free days.

On average 45.7 cm of precipitation falls annually, 35% as snow. Since 1960, extreme monthly precipitations have been zero to a high of 18.11 cm, with a maximum daily precipitation of 7.82 cm. Average monthly winds for the area range from 10km/hr to 13km/hr, with 40% of the winds originating from the northwest, northeast or north.

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FIGURE 5-1: SNOW LAKE REGIONAL MAP

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5.3 Local Resources

The nearest community is the town of Snow Lake, Manitoba, located approximately 16 km from Lalor. The community of 899 (Statistics Canada, 2016 census) has 498 private dwellings. There are two cottage subdivisions located on Wekusko Lake along PR #392, as well as residences at Herb Lake Landing, approximately 40 km south of the town. There are also a small number of seasonal remote cottages located near lakes throughout the area.

Snow Lake community services include a health facility staffed with two doctors, ambulance, fire truck, a grocery store, two hotels/motels, three service stations, a kindergarten to grade 12 school, a hockey arena, a five-sheet curling rink and a nine-hole golf course.

The nearest larger centres (5,000+) are Flin Flon (208 km), The Pas (219 km) and Thompson (260 km), all accessible by paved highway. There is a 1,100 m by 20 m unserviced gravel municipal airstrip located approximately 30 km from Lalor along PR #393. A small craft charter service operates out of the community of Snow Lake, where small planes and helicopters can be chartered. Rental vehicles are available at the Flin Flon airport. The nearest full service commercial airport is located at Baker’s Narrows, near Flin Flon, approximately 191 km from Lalor. The nearest international airport is located in Winnipeg, approximately 700 km from Snow Lake.

There is no rail in the immediate area of Lalor or Snow Lake. The nearest rail access is at Wekusko siding, approximately 65 km southeast of Lalor. Wekusko is accessible by an all-weather road. A gravel rail bed (ties and rail removed) connects the Stall concentrator to Chisel Lake mine, and continues to a rail line at Optic Lake siding, approximately 65 km west of Chisel Lake. Optic Lake is not road accessible.

5.4 Infrastructure

Hudbay operates a zinc metallurgical plant in Flin Flon, Manitoba, approximately 215 km from Lalor. Present capacity is 115,000 tpa refined zinc.

Hudbay operates an ore concentrator approximately 16 km from Lalor. The mill is currently operating seven days per week at 3,000 to 3,500 tonnes per day, processing ore from the Lalor mine. The mill has two circuits, with design capacities of 909 tpd and 2,182 tpd.

The concentrator has two flotation circuits producing a zinc concentrate and a copper concentrate. The tailings are deposited in the Anderson Tailings Impoundment Area. Concentrates are hauled by truck to Hudbay metallurgical facilities in Flin Flon.

General area infrastructure includes provincial roads and 115 kV Manitoba Hydro grid power to within four kilometres of Lalor, and Manitoba Telecom land line and cellular phone service.

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The Town of Snow Lake is a full service community with available housing, hospital, police, fire department, potable water system, restaurants and stores. The community is serviced by a 914 m gravel airstrip to provide emergency medical evacuation.

Lalor is located 3.5 km from the Hudbay Chisel North mine. Chisel North infrastructure includes a mined out open pit used for waste rock disposal, fresh (process) water sources, pumps and waterlines, 4160V and 550V power, mine discharge water lines, a 2,500 gpm water treatment plant with retention areas, plus mine buildings including offices and a change house. These facilities are used for geology core processing, surface mobile equipment shop, project offices, and the crushing of Lalor ore. The Chisel site is also the location of two electrical transformers 115 kV to 25 kV that feed Lalor mine. The infrastructure on site at Lalor includes, the main office change house, headframe, hoistroom, down cast fans, exhaust fans, main pump station, potable water treatment plant, sewage treatment plant and several other smaller buildings for purposes such as health and safety, additional change house space, etc. The permitted Hudbay Anderson TIA, located approximately 12 km from Lalor, is used for tailings disposal.

5.5 Physiography

The deposit is located in the Boreal Shield Ecozone, the largest ecozone in Canada, extending as a broad inverted arch from northern Saskatchewan east to Newfoundland. The area of Lalor and surrounding water bodies (Snow, File, Woosey, Anderson and Wekusko lakes) are located in the Churchill River Upland Ecoregion in the Wekusko Ecodistrict. The dominant soils are well to excessively drained dystic brunisols that have developed on shallow, sandy and stony veneers of water-worked glacial till overlying bedrock. Significant areas consist of peat-filled depressions with very poorly drained Typic and Terric Fibrisolic and Mesisolic Organic soils overlying loamy to clayey glaciolacustrine sediments.

The property area is approximately 300 m ASL, with depressional lowlands, and has gentle relief that rarely exceeds 10 m, consisting of ridged to hummocky sloping rocks.

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6 HISTORY

The Snow Lake area has a long exploration and mining history. The Lalor deposit was discovered in 2007.

6.1 Exploration in the Chisel Basin Area

Exploration in the Chisel Basin has been active since 1955. The Chisel Basin area has hosted three producing mines; namely, Chisel Lake, Chisel Open Pit and Chisel North. All three mines have very similar lithological and mineralogical features. This basin is also the host of the Lalor deposit.

In early 2007, drill hole DUB168 was drilled almost vertically to test a 2003 surveyed Crone Geophysics deep penetrating pulse electromagnetic anomaly and intersected a band of conductive mineralization between 781.74 m and 826.87 m (45.13 m). Assay results include 0.30% Cu and 7.62% Zn over the 45.13 m including 0.19% Cu and 17.26% Zn over 16.45 m. Drilling at Lalor has been continuous since the discovery of mineralization on the property.

6.2 Historical Mining in the Snow Lake Area

The Snow Lake area has had an active mining history for more than 50 years. Hudbay has played an integral part in this history since the late 1950s by operating nine mines in the area including Photo Lake, Rod, Chisel Lake Chisel North and Chisel Open Pit, Stall Lake, Osborne Lake, Spruce Point, Ghost Lake, and Anderson Lake.

The Stall concentrator was commissioned in 1979 and operated continuously until shutdown in early 1993 as a result of the depletion of the Chisel Open Pit and Stall Lake mines. The concentrator was reopened in 1994 to process ore from the Photo Lake Mine and continued to process ore from the Chisel North mine until February 2009. With the reopening of Chisel North in 2010, the concentrator reopened and has remained open and is today the processing facility for all ore produced at Lalor.

6.3 Lalor Mine Production

Lalor commenced initial ore production from the ventilation shaft in August 2012 and achieved commercial production from the main shaft in the third quarter of 2014. Table 6-1 summarizes the actual ore production by year at Lalor from 2012 to 2016.

TABLE 6-1: LALOR MINE ACTUAL PRODUCTION

Lalor Mine Actual Production
	Units	2016	2015	2014	2013	2012
Ore Mined	tonnes	1,086,362	934,277	551,883	400,590	72,293
Gold	g/t	2.24	2.53	2.29	1.21	1.67
Silver	g/t	21.63	21.38	23.83	19.39	19.29
Copper	%	0.62	0.71	0.88	0.84	0.63
Zinc	%	7.01	8.18	8.52	9.44	11.83

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Notes: Lalor ore production in 2014 includes partial production from the ventilation shaft, which began production in August 2012. Lalor ore production in each of 2012 and 2013 was from the ventilation shaft.

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7 GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology

The Lalor property lies in the eastern (Snow Lake) portion of the Paleoproterozoic Flin Flon Greenstone Belt (Figure 7-1) and is overlain by a thin veneer of Pleistocene glacial/fluvial sediments. Located within the Trans-Hudson Orogen, the Flin Flon Greenstone Belt consists of a variety of distinct 1.92 to 1.87 Ga (billion years ago, or giga-annums) tectonostratigraphic assemblages including juvenile arc, back-arc, ocean-floor and ocean-island and evolved volcanic arc assemblages that were amalgamated to form an accretionary collage (named the Amisk Collage) prior to the emplacement of voluminous intermediate to granitoid plutons and generally subsequent deformation (Syme et al., 1998). The volcanic assemblages consist of mafic to felsic volcanic rocks with intercalated volcanogenic sedimentary rocks. The younger plutons and coeval successor arc volcanics, volcaniclastic, and sedimentary successor basin rocks include the older, largely marine turbidites of the Burntwood Group and the terrestrial metasedimentary sequences of the Missi Group.

The Flin Flon belt is in fault and / or gradational contact with the Kisseynew Domain metasedimentary gneisses to the north and is unconformably overlain by the Phanerozoic cover of sandstone and dolostones to the south (Figure 7-2). Regional metamorphism at 1.82 to 1.81 Ga formed mineral assemblages in the Flin Flon belt that range from prehnite-pumpellyite to middle amphibolite facies in the east and upper amphibolite facies in the north and west (David and Machado, 1996; Froese and Moore, 1980; Syme et al., 1998).

The Snow Lake portion of the Flin Flon belt is dominated by fold-thrust style tectonics that is atypical of western and central portions of the belt. It is a south-verging, northeast dipping imbricate that was thrust over the previously amalgamated collage of oceanic and arc rocks to the west (Bailes and Galley, 1999). The thrust package of the Snow Lake area has been modified by 1.82 to 1.81 Ga regional metamorphism of lower to middle almandine-amphibolite facies mineral assemblages (David and Machado, 1996; Froese and Moore, 1980).

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FIGURE 7-1: GEOLOGY OFF MANITOBA

Source: http://www.gov.mb.ca/iem/geo/exp-sup/files/fig1.pdf, Department of Growth, Enterprise and Trade (GET)

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FIGURE 7-2: GEOLOGY OF THE FLIN FLON GREENSTONE BELT, MANITOBA

Source: http://www.gov.mb.ca/iem/geo/exp-sup/files/fig6.pdf. Manitoba Department of Growth, Enterprise and Trade (GET)

Intrusions in the belt are divided into pre-, syn- and post tectonic varieties where the pre-tectonic group includes intrusions that are coeval with the volcanic rocks, as well as those that crosscut volcanic and Missi supracrustal rocks. Numerous mafic to ultramafic dykes intrude the volcanic rocks.

7.2 Property Geology

The Snow Lake arc assemblage (Figures 7-3 and 7-4) that hosts the producing and past-producing mines in the Snow Lake area is a 20km wide by 6km thick section that records a temporal evolution in geodynamic setting from ‘primitive arc’ (Anderson sequence to the south) to ‘mature arc’ (Chisel sequence) to ‘arc-rift’ (Snow Creek sequence to the northeast, Bailes and Galley, 2007). The ‘mature arc’ Chisel sequence that hosts the zinc rich Chisel, Ghost, Chisel North, and Lalor deposits typically contains thin and discontinuous volcaniclastic deposits and intermediate to felsic flow-dome complexes.

The Chisel sequence is lithologically diverse and displays rapid lateral facies variations and abundant volcaniclastic rocks. Mafic and felsic flows both exhibit evolved geochemical characteristics (relative to the unevolved underlying Anderson sequence) consistent with one of, or a combination of, the following: within-plate enrichment, derivation from a more fertile mantle source, lower average extents of melting at greater depths, and contamination from older crustal fragments. These rocks have undergone metamorphism at the lower to middle almandine-amphibolite facies.

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Rock units in the hanging wall of the Lalor deposit typically reflect this diversity and variation in rock types that include mafic and felsic volcanic and volocaniclastic units, mafic wacke, fragmental units of various grain sizes, and crystal tuff units.

FIGURE 7-3: VOLCANIC STRATIGRAPHY OF THE SNOW LAKE AREA

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FIGURE 7-4: GEOLOGY OF THE SNOW LAKE AREA

Source: http://web33.gov.mb.ca/mapgallery/mgg-gmm.html [January 31, 2013]. Manitoba Department of Growth, Enterpriise and Trade (GET)

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The Lalor deposit is similar to other massive sulphide bodies in the Chisel sequence (Chisel Lake, Ghost Lake, Chisel North, and Photo Lake), and lies along the same stratigraphic horizon as the Chisel Lake and Chisel North deposits. It is interpreted that the top of the zone is near a decollement contact with the overturned hanging wall rocks.

The most common dyke intrusion throughout these rocks is a fine grained feldspar-phyric gabbro to diorite. The Chisel Lake pluton, a late 1.8 km by 9.8 km layered ultramafic intrusion (Bailes and Galley, 2007), truncates the main lens of the Chisel Lake massive sulphide deposit but is not seen in any of the Lalor drill core.

The extensive hydrothermal alteration and metamorphic recrystallization of the footwall rocks has produced some exotic aluminous mineral assemblages. These assemblages include chlorite and sericite dominant schists and cordierite+anthophyllite gneisses. Other minerals indicative of hydrothermal alteration that occur extensively throughout these rock assemblages include quartz, feldspar, kyanite, biotite, garnet, staurolite, hornblende, and carbonate. Clinopyroxene, gahnite and anhydrite also occur locally. These assemblages are typical of metamorphosed footwall hydrothermal alteration commonly associated with volcanogenic massive sulphide (VMS) deposits and are similar to that at the other massive sulphide deposits in the Chisel Lake area.

7.3 Base Metal Mineralization

The Lalor VMS deposit is flat lying, with zinc mineralization beginning at approximately 600 m from surface and extending to a depth of approximately 1,100 m. The mineralization trends about 320° to 340° azimuth and dips between 30° and 45° to the north. It has a lateral extent of about 900 m in the north-south direction and 700 m in the east-west direction.

Sulphide mineralization is pyrite and sphalerite. In the near solid (semi-massive) to solid (massive) sulphide sections, pyrite occurs as fine to coarse grained crystals ranging one to six millimetres and averages two to three millimetres in size. Sphalerite occurs interstitial to the pyrite. A crude bedding or lamination is locally discernable between these two sulphide minerals. Near solid coarse grained sphalerite zones occur locally as bands or boudins that strongly suggest that remobilization took place during metamorphism.

Disseminated blebs and stringers of pyrrhotite and chalcopyrite occur locally within the massive sulphides, adjacent to and generally in the footwall of the massive sulphides. The hydrothermally altered rocks in the footwall commonly contain some very low concentrations of sulphide minerals.

Some sections of massive pyrrhotite occur, but these tend to give way to pyrite-sphalerite-dominant zones.

Seven distinct stacked zinc rich mineralized zones have been interpreted within the Lalor deposit based on the zinc equivalency of 4.1% over a minimum three metre interval (Table 7-1).

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The top two lenses of the stacked base metal zones (coded as Zone 10 and 11) have higher grade zinc and iron content. The footwall lenses coded as Zones 20, 30, 31, 32 and 40 have moderate to high zinc grades hosted in near solid sulphides containing higher grade gold and locally appreciable amounts of copper.

Overall, Zones 10 and 20 have the largest extent and volume of mineralization. Zone 10 extends approximately 400 m in the east-west and 550 m in the north-south direction and Zone 20, 250 m in the east-west and 700 m in the north-south direction.

TABLE 7-1: SUMMARY OF ZINC RICH INTERPRETED WIREFRAMES

Zone	Area (m2)	Volume (m³)	Average Thickness of Mineralization (m)	Number of Drillholes	Assayed Length of Drill Core (m)	Volume (m³)/ Number of Drillholes
10	198,253	1,898,251	9.6	557	7,923.07	3,408
11	45,079	127,763	2.8	117	472.61	1,092
20	140,481	1,435,583	10.2	491	6,141.25	2,924
30	36,694	317,443	8.7	87	801.26	3,649
31	42,656	305,428	7.2	119	979.49	2,567
32	79,474	585,311	7.4	327	2,698.28	1,790
40	74,218	693,170	9.3	102	1,223.09	6,796
		4,669,779			20,239.05

7.4 Gold Mineralization

Gold and silver enriched zones occur near the margins of the zinc rich sulphide lenses and as lenses in local silicified alteration. Remobilization is illustrated in some of the gold-rich zones by late veining that is more or less restricted to the massive lenses. Some of the footwall zones tend to be associated with silicification and the presence of gahnite. These zones are often characterized by copper-gold association, and are currently interpreted as being associated with higher temperature fluids below a zone of lower temperature base-metal accumulations.

Footwall gold mineralization is typical of any VMS footwall feeder zone with copper-rich, disseminated and vein style mineralization overlain by a massive, zinc-rich lens. The fact that the footwall zone is strongly enriched in gold suggests a copper-gold association which is comparable to other gold enriched VMS camps and deposits (Mercier-Langevin, 2009).

General observations of the known gold zones indicate areas which are coarse grained and porphyroblastic in nature are gold poor, while fine grain siliceous (± veins ± sulphide traces) and strained looking stratigraphy tend to be gold rich. To date no definitive structural controls of the gold mineralization has been interpreted. However, the intensity and style of alteration can vary strongly over short distances and may suggest that the alteration was forming discordant stockwork like zones that are now strongly transposed in the main foliation (Mercier-Langevin, 2009).

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Seven lense groups have been interpreted within the deposit area and are present between 750 m to 1,480 m below the surface (Table 7-2). Their general shape is similar to the base metals. However, the current interpretation suggests the deeper copper-gold lense tends to have a much more linear trend to the north than the rest of the zones. The gold mineralization associated with each zone was interpreted into three-dimensional wireframes based on a 2.4 g/t gold equivalent or 1.9% copper equivalent over minimum 3 metre interval.

A structural study with external consultant Jean-Francois Ravenelle, from SRK Consulting (Canada) Inc. is currently ongoing. As of March 2017, four, one week site visits have been completed. Purposes of the study are to:

• Define potential structural controls on higher grade sections within the gold lenses of 21 and 25

• Investigate the suspected presence of a mine scale folded geometry

TABLE 7-2: SUMMARY OF GOLD INTERPRETED WIREFRAMES

Zone	Area (m2)	Volume (m³)	Average Thickness of Mineralization (m)	Number of Drillholes	Assayed Length of Drill Core (m)	Volume (m³)/ Number of Drillholes
21	122,358	668,776	5.5	370	2,882.30	1,808
23	46,137	225,844	4.9	177	1,152.51	1,276
24	83,483	423,505	5.1	269	1,589.02	1,574
25	247,164	1,470,650	6.0	454	4,393.73	3,239
26	60,374	504,483	8.4	55	784.56	9,172
27	72,125	462,584	6.4	36	511.31	12,850
28	26,528	247,365	9.3	13	105.81	19,028
		3,755,842			11,419.24

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8 DEPOSIT TYPE

Lalor is interpreted as a VMS deposit that precipitated at or near the seafloor in association with contemporaneous volcanism, forming a stratabound accumulation of sulphide minerals. VMS deposits typically form during periods of rifting along volcanic arcs, fore arcs, and in extensional back arc basins. Rifting causes extension and thinning of the crust, providing the high heat source required to generate and sustain a high-temperature hydrothermal system (Franklin et al., 2005).

The location of VMS deposits is often controlled by synvolcanic faults and fissures, which permit a focused discharge of hydrothermal fluids. A typical deposit will include the massive mineralization located proximal to the active hydrothermal vent, footwall stockwork mineralization, and distal products, which are typically thin but extensive. Footwall, and less commonly, hanging wall semiconformable alteration zones are produced by high temperature water-rock interactions (Franklin et al., 2005).

The depositional environment for the mineralization at Lalor is similar to that of present and past producing base metal deposits in felsic to mafic volcanic and volcaniclastic rocks in the Snow Lake mining camp. The deposit appears to have an extensive associated hydrothermal alteration pipe.

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9 EXPLORATION

Exploration in the Lalor deposit area was previously conducted by Hudbay through the main exploration office located in Flin Flon, Manitoba. More recently exploration is managed and conducted by the Lalor mine geology department with core logging and storage facilities, located at the Chisel North mine site, adjacent to the Lalor mine.

In 2003, a Crone Geophysics high power time-domain electromagnetic (EM) system experimental survey was conducted over the deepest portion (approximately 600 m vertical depth) of the Chisel North Mine. The survey was designed and interpreted by Hudbay and was conducted by Koop Geotechnical Services Inc. The survey provided conclusive evidence that the system could detect conductive bodies at depths greater than 500 m and it was decided to extend the survey coverage further down-dip and down-plunge of the known mineralized lenses. A double-wired transmitter loop measuring two km by two km was used to maximize the EM field strength. The survey results were interpreted using three-dimensional computer modeling software. The model indicated a highly conductive, shallow-dipping zone at a vertical depth of 800 m. The Lalor drilling began in March 2007 to test the geophysical anomaly, and diamond drill hole DUB168 intersected conductive sulphides at a depth of approximately 780 m. Drilling is ongoing and has been continuous since the discovery hole.

Exploration drilling, since the disclosed March 29, 2012 NI 43-101 Technical Report, has focused on delineation of the inferred resource, confirming the continuity of the mineralization down plunge and testing for new mineralization peripheral to the known deposit. Surface drilling since the previous disclosure is limited to four drill holes while the focus was on underground exploration, definition, and delineation drilling, which has continued to expand the resource.

The first of the four surface drill holes, DUB288, was completed 500 m east-northeast of the Lalor mine in April of 2012 as a follow-up to copper and gold mineralization intersected in drill holes DUB283 and DUB283W2 that were completed in the fall of 2011. The results suggest further investigation in that area is warranted.

In mid-2015, two surface drill holes (DUB289 and DUB290) were collared approximately 750 m east of the deposit testing a conductive horizon identified by earlier BHEM surveys. A barren sulphide horizon was intersected at a depth of 930 m in DUB290. No further investigation was warranted.

Hole DUB291, 1,709 m in length, was completed in the summer of 2015 and tested the down plunge potential of the deposit. Several zones of mineralization were intersected suggesting future exploration in this area is warranted.

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9.1 Underground Exploration

9.1.1 Exploration Development

Since 2014, one exploration drift and one exploration ramp were developed at Lalor for a total of 1,891 m. The development was undertaken to establish underground platforms to conduct exploration drilling on targets that could not be drilled from existing mine infrastructure. Prudent care was taken in the placement and size of both the exploration ramp and drift to assure the selected locations can accommodate future mining equipment and related infrastructure. Table 9-1 summarizes the underground exploration related development and intended future uses of the ramp and drift.

TABLE 9-1: UNDERGROUND EXPLORATION DEVELOPMENT

Location	Year	Meters	Exploration Target (Zone)	Definition Target (Zone)	Possible Future Use
865 Exploration Drift	2014	98	21, 25, 26, 28	20, 30, 40	Ventilation Drift
865 Exploration Drift	2015	396
865 Exploration Drift	2016	282
1025 Exploration Ramp	2014	637	27, 21, 25	20	Haulage and Ventilation
1025 Exploration Ramp	2015	478

9.1.2 Exploration Drilling

In 2015, thirty-one drill holes were completed for a total of 10,395 m focusing on the copper-gold, Zone 27. Exploration drilling continued on Zone 25 from March to July of 2016 for a total of sixty-nine drill holes and 16,098 m. The purpose of the exploration programs was to upgrade inferred resources, specifically focused on identifying areas of enriched gold and copper-gold mineralization. Due to the low angle and stacking nature of the mineralization at Lalor, holes were extended beyond the gold target depths to explore the on-strike and plunge potential of known base metal lenses, which led to increases in mineral resource inventory.

9.2 Borehole Electromagnetic (EM) Surveys

Time-domain borehole EM surveys with three dimensional probes are routinely conducted on surface and underground drill holes. The survey results identify any off-hole conductors that were missed, indicate direction to the target, as well as the dimensions and the attitude of the conductor. The surveys can also detect possible conductors which may lie past the end of the hole allowing decisions to extend holes to be made.

9.3 Surface Electromagnetic (EM) Surveys

Two time-domain surface EM surveys, for a total of approximately 35 line km, were completed north-northeast and south-southwest of the Lalor mine. Neither survey has identified any new significant targets of interest in the general Lalor mine area.

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9.4 Airborne Electromagnetic (EM) Survey

During the summer of 2014, an airborne EM survey was conducted to test the capabilities of the HeliSAM system for a total of 97.5 line km. Performed by GAP Geophysics, based in Perth Australia, using a ground based transmitting loop and airborne total field magnetic sensor. The testing was aimed to identify the Lalor mine at a depth beyond the capabilities of conventional airborne EM systems. The test was successful and has lead to further surveys of this type elsewhere in the mining camp.

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10 DRILLING

The Lalor mine was discovered by drilling a surface exploration hole testing an electromagnetic geophysical anomaly in March 2007, which intersected appreciable widths of zinc-rich massive sulphides in hole DUB168. Surface drilling continued to July 2012. A limited surface exploration drill program was conducted from August to October 2015 to explore for potential down plunge extensions of 27 lens and to test near mine geophysical conductors that could not be drilled from underground workings. As of January 1, 2017, a total of 203,037 m of surface drilling was completed at Lalor and Table 10-1 provides a summary by year.

TABLE 10-1: SUMMARY SURFACE DIAMOND DRILL HOLES WITH ASSAY
RESULTS AS OF JANUARY 1, 2017

Year	Hole Type	Operator	Number of Holes	Core Size	Length (m)	Drilling Company
2007	Parent	Hudbay	2	BQ	2,342	Major Drilling Ltd.
	Parent	Hudbay	26	NQ	29,600	Major Drilling Ltd.
2008	Parent	Hudbay	41	NQ	45,454	Major Drilling Ltd.
	Wedge	Hudbay	32	NQ/AQ	12,112	Major Drilling Ltd.
2009	Parent	Hudbay	29	NQ	35,390	Major Drilling Ltd.
	Wedge	Hudbay	47	NQ/AQ	22,884	Major Drilling Ltd.
2010	Parent	Hudbay	13	NQ	17,438	Major Drilling Ltd.
	Wedge	Hudbay	17	NQ/AQ	11,576	Major Drilling Ltd.
2011	Parent	Hudbay	10	NQ	15,458	Major Drilling Ltd.
	Wedge	Hudbay	5	NQ/AQ	3,139	Major Drilling Ltd.
2012	Parent	Hudbay	3	NQ	4,688	Major Drilling Ltd.
2015	Parent	Hudbay	2	NQ	2,956	Rodren Drilling Ltd.
Total			222		203,037

Underground drilling began at Lalor with hole LP0001 in January 2012 and drilling has been continuous to date. Holes are drilled at all dips and azimuths needed to provide adequate coverage of the orebody for interpretation and mining purposes. Holes with dips steeper than +70° are preferably avoided due to poor ergonomics and the increased risk for the drill crews.

Underground drilling at Lalor is divided into five different categories based on the primary planned purpose of the hole. The hole categories are as follows:

LD Prefix

Lalor definition holes with LD prefix are drilled into known lenses to upgrade inferred resources to a higher category and to identify mineralization contacts for preliminary mine design purposes. Typical hole spacing is about 15 to 20 m. Drill costs are allocated to the yearly operational budget.

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TABLE 10-2: SUMMARY OF LD UNDERGROUND DIAMOND DRILL HOLES AS OF JANUARY 1, 2017

Year	Hole Type	Operator	Number of Holes	Core Size	Length (m)	Drilling Company
2012	LD	Hudbay	48	BQ/NQ	5,264/2,194	Major Drilling Ltd.
2013	LD	Hudbay	142	BQ/NQ	4,660/13,873	Major Drilling Ltd.
2014	LD	Hudbay	315	BQ/NQ	9,995/22,377	Major Drilling Ltd.
2015	LD	Hudbay	258	BQ/NQ	14,144/13,038	Major Drilling Ltd.
2016	LD	Hudbay	174	BQ/NQ	10,643/3,411	Major Drilling Ltd.
Total			937		99,599

LE Prefix

Lalor engineering holes with LE prefix are drilled for mine infrastructure purposes such as drain holes, holes for electrical cables and service holes for break through rounds. Drill costs are allocated to the yearly operational engineering budget.

TABLE 10-3: SUMMARY OF LE UNDERGROUND DIAMOND DRILL HOLES AS OF JANUARY 1, 2017

Year	Hole Type	Operator	Number of Holes	Core Size	Length (m)	Drilling Company
2012	LE	Hudbay	2	NQ	81	Major Drilling Ltd.
2015	LE	Hudbay	1	NQ	62	Major Drilling Ltd.
2016	LE	Hudbay	6	NQ	239	Major Drilling Ltd.
Total			9		382

LP Prefix

Lalor project holes with LP prefix were holes drilled before the projects team handover to the operations team in 2012 and prior to mine production at Lalor. Project holes are similar to LD prefix holes noted above and costs of these holes were allocated to the project group budget in 2012.

TABLE 10-4: SUMMARY OF LP UNDERGROUND DIAMOND DRILL HOLES AS OF JANUARY 1, 2017

Year	Hole Type	Operator	Number of Holes	Core Size	Length (m)	Drilling Company
2012	LP	Hudbay	32	NQ	5,551	Major Drilling Ltd.
Total			32		5,551

LQ Prefix

Lalor delineation holes with LQ prefix are drilled from within the mineralization. Holes are typically drilled to establish exact ore contacts for detailed mine planning and stope design purposes. Typical hole spacing is 10 to 15 m and costs are allocated to the yearly operational budget.

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TABLE 10-5: SUMMARY OF LQ UNDERGROUND DIAMOND DRILL HOLES AS OF JANUARY 1, 2017

Year	Hole Type	Operator	Number of Holes	Core Size	Length (m)	Drilling Company
2012	LQ	Hudbay	2	BQ	61	Major Drilling Ltd.
2013	LQ	Hudbay	159	BQ	4,195	Major Drilling Ltd.
2014	LQ	Hudbay	80	BQ	2,692	Major Drilling Ltd.
2015	LQ	Hudbay	231	BQ	9,445	Major Drilling Ltd.
2016	LQ	Hudbay	74	BQ	2,113	Major Drilling Ltd.
Total			546		18,506

LX Prefix

Lalor exploration holes with LX prefix are drilled for targets outside known lenses or in areas of inferred resources and low drill density. Costs are allocated to the yearly capital budget.

TABLE 10-6: SUMMARY OF LX UNDERGROUND DIAMOND DRILL HOLES AS OF JANUARY 1, 2017

Year	Hole Type	Operator	Number of Holes	Core Size	Length (m)	Drilling Company
2015	LX	Hudbay	31	NQ	10,395	Major Drilling Ltd.
2016	LX	Hudbay	69	BQ/NQ	16,098	Major Drilling Ltd.
Total			100		26,493

10.1 Surveying of Property Grid and Drill Hole Collars

All surveying at Lalor mine is conducted in Universal Transverse Mercator projection using NAD83 datum Zone 14. The mine survey was tied into the Canadian Spatial Reference System grid through the use of the Global Navigation Satellite System (GNSS) and an initial reference point at surface. Using the initial reference point, a reference azimuth was established by the use of a gyro compass. Based on the initial GNSS point and gyro azimuth a traverse was conducted to carry the survey down the ramp to the 810 m level. Once the main underground development workings were completed, another traverse was done to bring the survey down to 955 m level. All surveys into the individual levels are based on the control points created by the initial survey located at 810 m level and 955 m level. All surveying is done using electronic theodolites using the resection/free station technique.

Diamond drill lines are marked up according to layouts issued by Lalor Geology. An electronic version of the layout is created for viewing on the electronic theodolite. In the field, the electronic theodolite is set up using resection or free station technique. Reflectorless Electronic Distance Measurement (EDM) is used to locate the planned drill azimuth line location according to the electronic drill hole layout. Collar location and a drill azimuth back sight is spray painted on the walls of the drift to function as front and back sights for diamond drillers to line up the drill. An anchor is installed on the painted lines to provide a permanent reference line. Once drilling is completed the collar survey is recorded using the resection setup method. Collar location is surveyed using reflectorless EDM surveying and collar location stored as single point data.

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10.2 Downhole Surveying

Downhole surveys were completed using a Reflex EZ-Shot®, EZ-A® or EZ-Trac® (Reflex) instrument. Surveys were completed at regular intervals of 30 m down the hole.

The Reflex instruments measure the azimuth relative to the earth’s magnetic field and records magnetic azimuth, dip of the hole and Total Magnetic Intensity (TMI) in nanotesla (nT). Magnetic interference is likely to occur in areas where the TMI deviates significantly from the regional value. Regional TMI for the mine is 58,000 nT to 60,000 nT. The Reflex instrument is calibrated to notify user when the TMI deviates more than 1,000 nT from the programmed regional TMI value. All down hole survey results are validated by a geologist before being entered into the acQuire drill hole database.

A down hole gyro survey was conducted at the Lalor Mine in June and July of 2016 to check the accuracy of the magnetic Reflex readings. The survey was conducted by company technicians using a Reflex TN14 gyro compass and a MEMS down hole gyro probe. Thirty-eight underground holes were surveyed by gyro method for a total of 11,358 m. No significant discrepancies were identified between azimuth values measured by magnetic Reflex survey methods and azimuth values measured by the gyro survey methods.

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11 SAMPLING PREPARATION, ANALYSES, AND SECURITY

11.1 Laboratory/Laboratories Used

Since the start of exploration at Lalor the following different laboratories and sample shipment/prep procedures have been in use:

• Discovery to November 1, 2009, Lalor samples were prepared and analyzed at Hudbay laboratory in Flin Flon, Manitoba. As part of Hudbay QAQC procedures, pulp duplicates were sent to ACME Analytical Laboratories Ltd. (ACME) in Vancouver, BC.

• November 1, 2009 to March 12, 2012, Lalor samples were received, crushed and pulverized at Hudbay laboratory and pulps shipped to ACME for analysis. Pulp duplicates were analyzed at Hudbay laboratory as part of the QAQC procedure.

• March 13, 2012 to May 21, 2014, all samples were prepared and analyzed at Hudbay laboratory. As part of Hudbay QAQC procedures, pulp duplicates were sent to ACME.

• May 22, 2014 to present, parts of the sample stream were and are shipped to ACME, (re- named to Bureau Veritas after January 1, 2015). The remainder of the sample stream is shipped to Hudbay laboratory.

• A set of 303 drill core pulps from samples originating from within known resource envelopes were submitted for check assaying at a SGS Laboratory in Burnaby, BC as part of the QAQC program for the 2017 resource estimate.

Bureau Veritas and SGS are certified independent laboratories, while the certified laboratory in Flin Flon is owned by Hudbay.

11.2 Sample Collection

Drill core is logged, sample intervals selected and marked clearly on the core with the following information written directly on the core using a red grease pen:

1)	Sample from: start depth of sample interval measured from collar of drill hole
2)	Alphanumeric unique sample number consisting of:

a) A letter prefix of either M or N.

i)	M prefix: Whole core is sampled and submitted to lab for assaying, rejects are discarded and pulps retained for reference
ii)	N prefix: pulps and rejects are to be stored for reference (bulk of holes are split with half core samples kept for reference)

b) A unique six digit number

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3) Sample to: end depth of sample interval measured from collar of drill hole

Visual metal mineral estimates are recorded for each individual sample and noted under sample number in sample book. The current practice for every 100 samples the following QAQC samples are inserted into the sample stream:

1)	Two blanks
2)	Five duplicates
3)	Five base metal standards, each of differing grade thresholds
4)	Two gold standards of differing grade

Once logging is complete all data from sample book including QAQC samples are entered into Hudbay’s acQuire drill hole database by hole number. Before samples are split/bagged for shipment the core is photographed. The photographing of the core is the last step of the logging process as to assure that a full photographic record of exact locations of all contacts, sample locations and numbers is captured. A standard setup as well as a single camera type with standardized settings is use to ensure photographs of consistent quality. Each photo covers a maximum of five boxes and includes in the frame a placard with the following information in clear legible writing:

1)	Hole ID
2)	Project name
3)	Name and number of any mineralized zones as well as start and end of zone given as distances in meter down hole
4)	Current date
5)	Box number for first and last box in photograph
6)	Depth range of hole displayed in photograph measured in meters from collar with 2 significant digits

Once photographs have been captured they are saved on the company server and linked to acQuire drill hole database for network access. This procedure applies to all drill core at Lalor mine including drill core to be sampled by destruction. If intervals of a drill hole scheduled to be sampled by destruction are deemed not to be sampled by logging geologist the core is kept until assays of submitted samples from drill hole have been received and reviewed by the geologist. If the geologist deems that no further sampling is needed core is discarded.

11.3 Sample Preparation

11.3.1 Hudbay

All samples arriving at the Hudbay analytical laboratory are checked against the geologist’s sample submission sheets. Laboratory analytical work sheets are generated for the analysis areas. Any wet samples are dried at 105°C as per industry standard. The core samples are crushed to (-)10 mesh then split to approximately 250g and pulverized with 90% passing (-)150 mesh before being deposited into labelled bags. Crusher and pulverizer checks are conducted daily to ensure there is no excessive wear on the crusher plates and pulverizer pots.

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11.3.2 ACME/Bureau Veritas

All samples arriving at ACME (Bureau Veritas) are checked against chain information on sample submittal form and prepped according to codes WGHT and PR80-250. The sample preparation includes weighing of sample, crushing 1kg to minimum 80% passing 2 mm. A 250 g split crushed to minimum 85% passing 75 µm.

11.3.3 SGS

All samples arriving at SGS Canada Inc. laboratory are checked against chain information on sample submittal form upon arrival. No sample preparation was conducted at SGS as the laboratory was only used for check assays done on pulps previously assayed at either Hudbay or Bureau Veritas laboratories.

11.4 Assay Methodology

11.4.1 Hudbay

Samples sent to the Hudbay laboratory were analyzed for the following elements: gold, silver, copper, zinc, lead, iron, arsenic and nickel. Base metal and silver assaying was completed by aqua regia digestion and read by a simultaneous ICP unit. The gold analysis was completed on each sample by atomic absorption spectrometry (AAS) after fire assay lead collection. All samples with gold values (AAS) > 10 g/t were re-assayed using a gravimetric finish. Detection limits of the ICP and AAS are shown in Table 11-1.

TABLE 11-1: HUDBAY LABORATORY DETECTION LIMITS

Element	Detection Limit
Ag	0.439 g/t
As	0.002%
Au	0.034 g/t
Cu	0.003%
Fe	0.007%
Ni	0.001%
Pb	0.002%
Zn	0.010%

All analytical balances are certified annually by a third party. Check weights are used daily to verify calibration of balances. All metal standards used to make the calibration standards for the AAS and ICP are certified and traceable. Each is received with a certificate of analysis. Both the AAS and ICP are serviced twice per year by the instrument manufacturer’s qualified service representative to ensure that the instruments meet original design specifications.

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The Flin Flon assay laboratory has been participating in CANMET PTP/MAL round robin testing since 2000. PTP/MAL is a requirement for laboratories that are ISO 17025 certified. The laboratory has also been participating since 2002 in round robin testing conducted by GEOSTATS of Australia.

Fine sample pulps are kept in secure storage at the laboratory after analysis. Pulps are only released after all data is validated.

11.4.2 Acme/Bureau Veritas

Two different assay methods are used for samples shipped to Bureau Veritas: AQ270 and AQ370. AQ370 was the only method used on samples submitted from May 2014 to the second quarter of 2016 after which the AQ270 method was applied to selected holes. After the fourth quarter of 2016 all samples submitted to Bureau Veritas were assayed using the AQ270 method. All samples using method AQ270 and AQ370 were run for gold using method FA430. The following elements were run for over range as necessary: gold, copper, zinc and lead using methods FA530, GC820, GC816, MA404 (assays returning lead values above 20% using MA404 were also run using method GC817) respectively. Detection limits for over range methods, AQ270 and AQ370 are listed in Table 11-2 to Table 11-4.

Samples from selected holes were also submitted for determination of specific gravity using SPG02 method (volume determination by submersion followed by drying).

Samples shipped to Bureau Veritas from November 1, 2009 to March 12, 2012, were run using the legacy codes (Group 7AR) and (Group 601) with over range samples being run with gravimetric finish (Group 612). The sample preparation for these legacy codes is essentially similar to those listed for the current AQ270 and AQ370 codes used after May 22, 2014.

For the multi-element methods AQ270 and AQ370, aliquots of 1.000 ± 0.002 g are weighed into 100 mL volumetric flasks. Bureau Veritas QAQC protocol requires one pulp duplicate to monitor analytical precision, a blank, and an aliquot of in-house reference material to monitor accuracy in each batch of 36 samples. 30 mL of Aqua Regia, a 1:1:1 mixture of ACS grade concentrated HCl, concentrated HNO₃ and de-mineralised H₂O> is added to each sample. Samples are digested for   one hour in a hot water bath (> 95°C). After cooling for 3 hours, solutions are made up to volume (100 mL) with dilute (5%) HCl. Very high-grade samples may require a 1 g to 250 mL or 0.25 g to 250 mL sample/solution ratio for accurate determination. Bureau Veritas QAQC protocol requires simultaneous digestion of a reagent blank inserted in each batch.

For both AQ270 and AQ370 sample solutions are aspirated into an ICP emission spectrograph (ES) to determine 24 elements. For method AQ270 the solution is also run through an ICP mass spectrometer (MS) to provide values for an additional 10 elements bring the total number of elements to 34. Raw and final data from the ICP-ES/ICP-MS undergoes a final verification by a British Columbia Certified Assayer who then signs the Analytical Report before it is released to the client. The 24 element assay method AQ370 has detection limits displayed in Table 11-2 and the 34 element assay method AQ270 has detection limits displayed in Table 11-3.

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For the gold analysis FA430, 30 g charges are weighed into fire assay crucibles. The sample aliquot is custom blended with fire assay fluxes, PbO litharge and a silver inquart. Firing the charge at 1050°C liberates Au, Ag ± PGEs that report to the molten Pb-metal phase. After cooling the lead button is recovered, placed in a cupel, and fired at 950°C to render an Ag ± Au ± PGEs dore bead. The bead is weighed and parted (i.e. leached in 1 mL of hot HNO₃) to dissolve silver leaving a gold sponge. Adding 10 mL of HCl dissolves the Au ± PGE sponge. Solutions are analysed for gold on an ICP emission spectrometer. Gold in excess of 10 g/t forms a large sponge that can be weighed (gravimetric finish, method FA530).

As part of Bureau Veritas QAQC protocol, a sample-prep blank is inserted as the first sample and carried through all stages of preparation to analysis as well as a pulp duplicate to monitor analytical precision. Two reagent blanks are inserted in each batch to measure background, and aliquots of Certified Reference Materials are used to monitor accuracy of the obtained gold assays. Raw and final data undergo a final verification by a British Columbia Certified Assayer who signs the Analytical Report before it is released to the client. Bureau Veritas is currently registered with ISO 9001 accreditation.

Fine sample pulps are kept in secure storage at the laboratory after analysis. Pulps are only released after all data is validated.

TABLE 11-2: BUREAU VERITAS ELEMENTAL DETECTION LIMITS AQ370

Element	Detection Limit	Upper Limit
g	2 g/t	300 g/t
Al	0.01%
As	0.01%	10%
Bi	0.01%
Ca	0.01%
Cd	0.001%
Co	0.001%
Cr	0.001%
Cu	0.001%	10%
Fe	0.01%
Hg	0.001%
K	0.01%
Mg	0.01%
Mn	0.01%
Mo	0.001%	20%
Na	0.01%

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Element	Detection Limit	Upper Limit
Ni	0.001%
P	0.001%
Pb	0.01%	4%
S	0.05%
Sb	0.001%
Sr	0.001%
W	0.001%
Zn	0.01%	20%

TABLE 11-3: BUREAU VERITAS ELEMENTAL DETECTION LIMITS AQ270

Element	Detection Limit	Upper Limit
Ag	0.5ppm	300ppm
Al	0.01%
As	5ppm	100000ppm
Ba	5ppm
Bi	0.5ppm
Ca	0.01%
Cd	0.5ppm
Co	0.5ppm
Cr	0.5ppm
Cu	0.5ppm	100000ppm
Fe	0.01%
Ga	5ppm
Hg	0.05%
K	0.01%
La	0.5ppm
Mg	0.01%
Mn	5ppm
Mo	0.5ppm	200000ppm
Na	0.01%
Ni	0.5ppm
P	0.00%
Pb	0.5ppm	40000ppm
S	0.05%
Sb	0.5ppm
Sc	0.5ppm
Se	2ppm	500ppm
Sr	5ppm
Th	0.5ppm

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Element	Detection Limit	Upper Limit
Ti	0.00%
Tl	0.5ppm
U	0.5ppm
V	10ppm
W	0.5ppm
Zn	5ppm	200000ppm

TABLE 11-4: BUREAU VERITAS ELEMENTAL DETECTION LIMITS AND RANGE FOR OVER RANGE CODES

Element	Over Range Codes	Technique	Detection Limit	Upper Limit
Au	FA530	Fire Assay/Gravemetric Finish	0.9ppm	1,000,000 ppm
Ag	FA530	Fire Assay/Gravemetric Finish	50ppm	1,000,000 ppm
Cu	GC820	Titration	0.01%	100%
Zn	GC816	Titration	0.01%	100%
Pb	GC817	Titration	0.01%	100%
Fe	GC818	Titration	0.01%	100%

TABLE 11-5: BUREAU VERITAS ELEMENTAL DETECTION LIMITS FOR LEGACY CODES
(GROUP 7AR) AND (GROUP 601)

Element	Detection Limit
Ag	2.000 g/t
Al	0.010 %
As	0.010%
Au	0.010 g/t
Bi	0.010 %
Ca	0.010 %
Cd	0.001 %
Co	0.001 %
Cr	0.001 %
Cu	0.001 %
Fe	0.010 %
Hg	0.001 %
K	0.010 %
Mg	0.010 %
Mn	0.010 %
Mo	0.001 %
Na	0.010 %
Ni	0.001 %
P	0.001 %

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Element	Detection Limit
Pb	0.010 %
Sb	0.001 %
Sr	0.001 %
W	0.001 %
Zn	0.010 %

11.5 Assay Certificates

Assay certificates since the discovery of Lalor have had two sources:

1)	Hudbay laboratory in Flin Flon, Manitoba
2)	ACME laboratory (renamed to Bureau Veritas as of January 1, 2015) in Vancouver, British Columbia

Assay certificates are received from both laboratories in digital form via e-mail. Data for import arrives as CSV files. Each CSV file is accompanied by a PDF certificate covering the sample number listed in the CSV file. The files are sent directly from the laboratory and independently to Lalor Senior Geologist and Hudbay database manager. Digital copies are archived with write privileges given only to the Hudbay database manager.

11.6 Security

Security measures taken to ensure the validity and integrity of the samples collected include:

• Chain of custody of drill core from the drill site to the core logging area

• All facilities used for core logging and sampling located on a secure mine site

• Core sampling is undertaken by Hudbay geologists

• Sample splitting and shipping conducted by technicians under the supervision of Hudbay geologists

• Chain of custody for core cutting through to delivery of samples to laboratories

• Well documented and implemented receiving and processing procedures at the Hudbay and Bureau Veritas laboratories

• The Hudbay Laboratory samples results are stored on a secure mainframe based Laboratory Information Management System (LIMS)

• The diamond drill hole database is stored on the secure Hudbay network, using the acQuire database management system with strict access rights

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The author believes that there are no factors that could have materially impacted the accuracy and reliability of the sample preparation, security, and analytical procedures and that those used are appropriate and adequate for VMS type mineralization.

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12 DATA VERIFICATION

Data verification procedures and results since the previous technical report are included in this section based on information collected and reviewed from 2012 to 2016. For data verification procedures and results prior to 2012, refer to previous technical report dated March 29th, 2012.

12.1 Bureau Veritas Assay Methods and QAQC

Over the past five years, 2012 to 2016, a total of 23,822 drill core samples were analyzed at Bureau Veritas laboratories. Copper, zinc, and silver were digested in aqua regia and analyzed by inductively coupled plasma optical emission spectrometry (ICP-OES) and more recently in 2016 by inductively coupled plasma mass spectrometry (ICP-MS) (Table 12-1). Samples with copper and zinc over the upper limit of detection (ULD) were analyzed by titration, whereas those samples with silver values over the ULD were analyzed by fire assay and gravimetric finish (Table 12-1).

TABLE 12-1: BUREAU VERITAS ASSAYS SPECIFICATIONS

Element	Unit	Lower Detection Limit	Upper Detection Limit	Digestion	Instrumental Finish	Method Code	Legacy Method Code
Cu1	%	0.001	10	Aqua Regia	ICP-OES	AQ370	7AR
Cu2	%	0.00005	10	Aqua Regia	ICP-OES and ICP-MS	AQ270	7AX
Cu overlimits	%	0.01	100		Cu Titration	GC820	G820
Zn1	%	0.01	20	Aqua Regia	ICP-OES	AQ370	7AR2
Zn2	%	0.0005	20	Aqua Regia	ICP-OES and ICP-MS	AQ270	7AX
Zn overlimits	%	0.01	100		Zn Titration	GC816
Ag1	ppm	2	300	Aqua Regia	ICP-OES	AQ370	7AR1
Ag2	ppm	0.5	300	Aqua Regia	ICP-OES and ICP-MS	AQ270
Ag overlimits	ppm	50		Fire Assay	Gravimetric	FA530	G603-G612
Au	ppm	0.01 - 0.005	10	Fire Assay	AAS	FA430	G601
Au overlimits	ppm	0.17 - 0.9	16600	Fire Assay	Gravimetric	FA530	G6 Grav

Gold was determined by lead-collection fire assay fusion, for total sample decomposition, followed by atomic absorption spectroscopy (AAS) instrumental analysis (Table 11-1). Fire assays were performed on 15 to 30g subsample pulps to circumvent problems due to potential nugget effect.

As part of Hudbay QAQC program, QAQC samples were systematically introduced in the sample stream to assess sub-sampling procedures, potential cross-contamination, precision, and accuracy. Hudbay commonly includes 5% certified reference materials (CRM), 2% certified blanks, and 5% coarse duplicates. Blanks and CRMs were prepared mostly by Ore Research and Exploration (OREAS). However, a few high-grade gold standards are from Rocklabs. All QAQC samples were analyzed following the same analytical procedures as those used for the drill core samples.

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12.1.1 Blanks

Certified OREAS blanks were inserted into the sample stream commonly one every fifty samples to monitor potential cross-contamination (Table 12-2). Between 2012 and 2016, 841 blanks were analyzed representing 3.5% of the samples submitted to Bureau Veritas.

TABLE 12-2: OREAS CERTIFIED BLANKS

Element	Cu	Zn	Ag	Au
Unit	%	%	ppm	ppm
F61	0.0035	0.0054	<0.1	0.009
F72	0.005	0.0107	0.068*	<0.001
Certified Method	Aqua regia1 - 4 acids2	Aqua regia1 - 4 acids2	Aqua regia1 - 4 acids2	Fire assay

*Indicative value

Blanks F6 and F7 are coarse blanks, <7mm chips, prepared exclusively for Hudbay by Ore Research and Exploration (Table 11-2). Both blanks have the same matrix consisting of fresh alkali olivine basalt. They are packaged in 100 g laminated foil pouches.

Gold assays in both blanks were determined by fire assay (Table 12-2). Copper, zinc, and silver were determined following an aqua regia digestion for blank F6 and a 4 acid digestion for blank F7. The 4 acid digestion is a stronger acid attack than the aqua regia digestion used for assaying material from the Lalor mine. However, to assess contamination both blanks are appropriate given their low concentration of base and precious metals. Silver is not certified in these blanks but their indicative values are a good guide to track potential contamination.

Blank failure thresholds due to potential contamination issues are set to values that exceed the certified best value (CBV) plus three standard deviations. A summary of the blank performance is shown on Table 12-3. The potential impact of contamination on resource estimation is assessed by (1) quantifying the amount of contamination relative to the certified blanks, and (2) the percent rate of failed blanks. An assay program is considered successful when the blank failure rate is <10%, and the average grades of contaminated blanks are not economic.

A total of 320 F6 blanks and 521 F7 blanks were systematically inserted along with the drill core samples analyzed at Bureau Veritas (Table 11-3). Contamination with copper and zinc was insignificant with an average contamination of 25 to 30 ppm for copper and 230 to 330 ppm for zinc. The blank failure rate for these metals is <6% and there are no cases of contamination at economic grade levels.

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TABLE 12-3: SUMMARY OF BLANK PERFORMANCE

Coarse Blank F6
Analyte	No. Blanks	Failed Blanks	Contamination Rate	Maximum Contamination	Average Contamination
Cu	320	17	5.3%	100 ppm	25 ppm
Zn	320	9	2.8%	440 ppm	230 ppm
Ag	320	0	0.0%	0	0
Au	320	12	4%	7 ppb	3 ppb

Coarse Blank F7
Analyte	No. Blanks	Failed Blanks	Contamination Rate	Maximum Contamination	Average Contamination
Cu	521	16	3.1%	60 ppm	30 ppm
Zn	521	24	4.6%	0.25%	330 ppm
Ag	521	3	0.6%	1.9	1 ppm
Au	521	31	6.0%	85 ppb	13 ppb

Blank F6 shows no contamination with silver, whereas blank F7 indicates an average contamination with silver of 1ppm. However, the failure rate for silver on blank F7 is <1% which renders the contamination an isolated case and not significant (Table 12-3).

The blank failure rate for gold is 4 to 6%. The contamination recorded by the blanks is on average 3 to 13ppb which is below grades of economic interest. Blank F7 carries contamination of up to 85ppb gold. The contamination is not economically significant, but it is recommended to visit the laboratory and review the sample preparation facilities since cleaner laboratory conditions are achievable and desired for gold assays (Table 12-3).

The performance of blanks indicates that there are no significant problems with contamination at the Bureau Veritas laboratories, samples were handled with care, and the assay results are free of contamination and adequate for the resource estimation.

12.1.2 Standards

From 2012 to 2016 a total of 1,601 OREAS and Rocklabs CRMs were analyzed at Bureau Veritas representing 6.7% of the sample stream. Table 12-4 presents a list of all CRMs used by Hudbay for quality control over the past five years.

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TABLE 12-4: CERTIFIED REFERENCE MATERIALS

Element	Cu	Zn	Ag	Au
Unit	%	%	ppm	ppm
OREAS A5	0.0945	0.506	1.91	0.141
OREAS B5	1.33	0.125	2.86	0.476
OREAS C5	3.37	5.28	21.1	2.493
OREAS D5	9.42	2.46	90.0	7.344
OREAS E5	0.393	23.70	19.7	0.780
OREAS A6	0.055	0.039	0.535	0.121
OREAS B6	0.872	0.702	2.94	0.689
OREAS C6	2.05	2.41	5.18	0.987
OREAS D6	4.20	3.31	25.17	3.04
OREAS E6	0.245	18.14	15.23	0.316
Rocklabs SN60	-	-	-	8.595
Rocklabs SN75	-	-	-	8.671
Rocklabs SP59	-	-	-	18.12
Rocklabs SP73	-	-	-	18.17
Certified Method	Aqua Regia	Aqua Regia	Aqua Regia	Fire assay

OREAS standards are a series of matrix-matched polymetallic CRMs prepared exclusively for Hudbay (Table 12-4). The geological materials were sourced from Hudbay Manitoba mine ore bodies. The mines are located in northern Manitoba within the Flin Flon Greenstone Belt in the Canadian Shield. These ore bodies are typical of polymetallic copper, zinc, silver, and gold rich volcanogenic massive sulphide (VMS) deposits. The gangue materials comprise felsic and mafic volcanic rocks; commonly the host rocks of economic mineralization. These CRMs are packaged into 50g laminated foil pouches sealed under dry nitrogen. The OREAS series covers a wide range of low-, medium-, and high-grade metal values adequate for the QAQC program.

Rocklabs standards are a series of commercially available high-grade gold CRMs prepared from pulverized feldspar minerals, basaltic rocks, and barren pyrites (Table 12-4). These materials are blended with pulverized gold-bearing minerals screened to avoid nugget effect.

More than 50 samples were analyzed per CRM which provide sufficient information to set acceptance criteria relative to the average (AV) and standard deviation (SD) of the actual assay values of the CRMs. However, the acceptance criteria is re-set to the CBV and SD recommended by the CRM in those cases where (1) the absolute bias is >10%, relative to the certified best value (CBV) provided by the CRM, and (2) the laboratory determinations have unexpected larger variance than the recommended SD of the CRM. Table 12-5 summarizes the performance of the CRM assays at Bureau Veritas laboratories.

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TABLE 12-5: SUMMARY OF CRM PERFORMANCE AT BUREAU VERITAS

Cu (%)
CRM	No. of Samples	No. of Failures	Failure Rate (%)	CRM Value (%)	ACME Average	CRM CV	Lab CV	Relative Bias
OREAS A5	83	2	2.4%	0.0945	0.0915	3%	2%	-3.1%
OREAS B5	74	1	1.4%	1.33	1.30	3%	2%	-1.9%
OREAS C5	56	0	0%	3.37	3.29	6%	2%	-2.3%
OREAS D5	58	0	0%	9.42	9.22	4%	4%	-2.1%
OREAS E5	59	1	1.7%	0.393	0.389	8%	2%	-1.1%
OREAS A6	191	1	0.5%	0.055	0.053	4%	3%	-3.9%
OREAS B6	190	0	0.0%	0.872	0.870	2%	2%	-0.3%
OREAS C6	185	2	1.1%	2.05	2.03	3%	2%	-1.0%
OREAS D6	184	2	1.1%	4.20	4.12	4%	2%	-2.0%
OREAS E6	183	2	1.1%	0.245	0.247	5%	2%	0.7%
Zn (%)
CRM	No. of Samples	No. of Failures	Failure Rate (%)	CRM Value (ppm)	ACME Average	CRM CV	Lab CV	Relative Bias
OREAS A5	83	1	1.2%	0.506	0.542	5%	3%	7.0%
OREAS B5	74	0	0%	0.125	0.132	7%	3%	5.7%
OREAS C5	56	0	0%	5.28	5.35	11%	3%	1.2%
OREAS D5	58	1	2%	2.46	2.55	7%	3%	3.8%
OREAS E5	59	1	1.7%	23.70	25.10	4%	2%	5.9%
OREAS A6	191	0	0%	0.039	0.037	5%	14%	-6.0%
OREAS B6	190	1	0.5%	0.702	0.703	3%	3%	0.2%
OREAS C6	185	1	0.5%	2.41	2.43	3%	3%	1.0%
OREAS D6	184	0	0.0%	3.31	3.37	4%	2%	1.8%
OREAS E6	183	1	0.5%	18.14	18.07	4%	2%	-0.4%
Au (ppm)
CRM	No. of Samples	No. of Failures	Failure Rate (%)	CRM Value (ppm)	ACME Average	CRM CV	Lab CV	Relative Bias
OREAS A5	83	2	2.4%	0.141	0.132	7%	9%	-6.4%
OREAS B5	74	1	1.4%	0.476	0.460	3%	6%	-3.3%
OREAS C5	56	2	4%	2.493	2.488	2%	4%	-0.2%
OREAS D5	58	2	3%	7.344	7.367	4%	8%	0.3%
OREAS E5	59	2	3.4%	0.780	0.758	3%	6%	-2.8%
OREAS A6	190	3	1.6%	0.121	0.126	4%	4%	3.9%
OREAS B6	190	4	2.1%	0.706	0.720	2%	4%	1.9%
OREAS C6	185	5	2.7%	0.987	0.988	2%	3%	0.1%
OREAS D6	184	1	0.5%	3.04	3.040	4%	4%	0.0%
OREAS E6	183	3	1.6%	0.316	0.325	4%	4%	2.8%
Rocklabs SN75	175	2	1.1%	8.671	8.626	2%	2%	-0.5%
Rocklabs SP73	163	3	1.8%	18.17	18.00	2%	2%	-1.0%
Ag (ppm)
CRM	No. of Samples	No. of Failures	Failure Rate (%)	CRM Value (ppm)	ACME Average	CRM CV	Lab CV	Relative Bias
OREAS B5	73	4	5.5%	2.86	2.95	15%	16%	3.0%
OREAS C5	56	1	2%	21.1	22.45	11%	5%	6.4%
OREAS D5	58	0	0%	90.0	98.21	13%	3%	9.1%
OREAS E5	59	1	1.7%	19.7	20.80	10%	5%	5.6%
OREAS B6	182	2	1.1%	2.94	2.92	18%	16%	-0.6%
OREAS C6	185	4	2.2%	5.18	5.12	10%	11%	-1.2%
OREAS D6	184	4	2.2%	25.17	26.32	6%	5%	4.5%
OREAS E6	183	2	1.1%	15.23	15.66	6%	7%	2.8%

The performance gates were set such that CRM assayed values within AV±2SD and isolated values between AV±2SD and AV±3SD were accepted. In contrast, two consecutive assayed values between AV±2SD and AV±3SD and all values outside the AV±3SD were rejected.

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To evaluate the accuracy of the assays using the CRMs, the relative bias was calculated after excluding the outlier values located outside the AV±3SD:

Bias (%) = 100*[(AVeo/CBV)-1]

AVeo represents the average of the actual assay values after excluding outliers. The analytical bias was assessed according to the following ranges: good between 0 and ±5%, reasonable between ±5% and ±10%, and unacceptable for values ±10%.

A successful assaying program aims at having good to reasonable bias and good to reasonable reproducibility of the CRM assays. The reproducibility can be assessed by the failure rate of the CRM assays considered to be good at <5%, reasonable between 5 and 10%, and unacceptable at >10%. In addition, comparison of the coefficient of variation (CV) of the CRM assays with the reported CV in the CRM certificate is a good guide to assess the variability of the assays at a given laboratory.

The analytical bias is good for copper and good to reasonable for zinc. The CRM failure rate for both base metals is <3%. The laboratory coefficient of variation (CV) for copper and zinc is in most cases better than or comparable to the variation documented by the CRM certificates.

The bias for gold and silver is good to reasonable. In addition, the CRM failure rate is <4% for gold and <6% for silver. The laboratory coefficient of variation (CV) for gold and silver is comparable to the variation reported in the CRM certificates.

It is concluded that the accuracy and reproducibility of copper, zinc, silver, and gold assays, as indicated by the CRMs assayed at Bureau Veritas laboratories, is of good quality for resource estimation.

12.1.3 Duplicates

The following analysis is based on coarse duplicates analyzed during 2015 and 2016. There is no record of coarse duplicates between 2012 and 2014. Coarse duplicates, approximately one in every twenty samples, were submitted to Bureau Veritas laboratory in order to monitor sub-sampling precision. After crushing to 10 mesh (2mm), a coarse duplicate sub-sample was riffle split and pulverized to ≥85% passing through 200 mesh (75μm). The duplicate sample was analyzed immediately after its paired sample. Quarter-core twin sample duplicates and pulp duplicates were not analyzed in the sample stream.

Coarse duplicates were evaluated using the hyperbolic method developed by AMEC (Simón, 2004). Minimum and maximum element concentrations of the sample pairs are plotted in the y and x axis, respectively. In the Minimum-Maximum diagrams all samples plot along and above the y = x line and the failure boundary is given by the hyperbolic equation y²=m²x²+b².

The coarse duplicates were evaluated using a failure boundary that asymptotically approaches the line with slope m corresponding to a 20% absolute relative error (RE), and an intercept b representing the practical detection limit, set at 5 times the lower limit of detection (LOD) of the laboratory. The RE, expressed in percentage, is calculated as the absolute value of the pair difference divided by the pair average. An acceptable level of sub-sampling variance is achieved when the failure rate does not exceed 10% of the total pairs (Table 12-6).

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TABLE 12-6: SUMMARY OF COARSE DUPLICATE PERFORMANCE

Analyte	Duplicate Samples	Duplicate Failures	Total Rate of Failures	Practical Detection Limit	Accepted Absolute Relative Error (RE)
Cu	1035	17	1.6%	50 ppm	20%
Zn	677	7	1.0%	500 ppm	20%
Ag	436	30	6.9%	2.5 ppm	20%
Au	962	53	5.5%	0.05 ppm	20%

Table 12-6 and Figure 12-1 to Figure 12-4 show the results of the coarse duplicates submitted to Bureau Veritas. The duplicate pairs display failure rates ranging between 1 and 7% for copper, zinc, silver, and gold when evaluated by the hyperbolic method for an absolute relative error of 20%. The low failure rates indicate that a good level of sub-sampling variance was achieved by Bureau Veritas laboratories for copper, zinc, silver, and gold assays. Therefore, it is concluded that the sub-sampling procedures were adequate for all metals used in the resource model.

FIGURE 12-1: COPPER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

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FIGURE 12-2: ZINC COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

FIGURE 12-3: SILVER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

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FIGURE 12-4: GOLD COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

12.2 Hudbay Laboratory Methods and QAQC

From 2012 to 2016 a total of 104,024 drill core samples were analyzed at the Hudbay laboratory in Flin Flon, Manitoba. Copper, zinc, and silver were digested in aqua regia and analyzed by ICP-OES (Table 12-7). Gold was determined by lead-collection fire assay fusion, for total sample decomposition, followed by atomic absorption spectroscopy (AAS) analysis. Fire assays were performed on 15 to 30 g subsample pulps to avoid problems due to potential nuggetty gold.

TABLE 12-7: HUDBAY LABORATORY ASSAYS SPECIFICATIONS

Element	Unit	Lower Detection Limit	Upper Detection Limit	Digestion	Instrumental Finish	Method Code
Cu	%	0.01	16.5	Aqua Regia	ICPA-OES	LAI-006
Cu overlimits	%	0.05	80	Aqua Regia	ICPA-OES	LAI-006
Zn	%	0.01	33	Aqua Regia	ICPA-OES	LAI-006
Zn overlimits	%	0.05	100	Aqua Regia	ICPA-OES	LAI-006
Ag	ppm	0.446	500	Aqua Regia	ICPA-OES	LAI-006
Ag overlimits	ppm	2.23	2500	Aqua Regia	ICPA-OES	LAI-006
Au	ppm	0.103	6.857	Fire Assay	AAS	LAI-044
Au overlimits	ppm	6.857		Litharge FA	Gravimetric	LAI-115

Quality control samples were systematically introduced in the sample stream to assess sub-sampling procedures, potential cross-contamination, precision, and accuracy. Hudbay sampling program commonly consists of 5% certified reference materials (CRM), 2% certified blanks, and 5% coarse duplicates. Blanks and CRMs are mostly from OREAS and a few high gold grade standards from Rocklabs. All QAQC samples were analyzed following the same analytical procedures as those used for drill core samples.

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12.2.1 Blanks

Certified OREAS blanks F6 and F7 were inserted into the sample stream approximately one every fifty samples to monitor potential cross-contamination (Table 12-2). A total of 2,278 blanks were inserted, representing 2.2% of the total number of samples analyzed at the Hudbay laboratory. The performance of the certified blanks was assessed using the same protocols explained in detail on Section 12.1.1.

A total of 218 F6 blanks and 2,060 F7 blanks were systematically inserted along with the drill core samples analyzed at the Hudbay laboratory in Flin Flon. A summary of the blank performance is shown on Table 12-8.

TABLE 12-8: SUMMARY OF BLANK PERFORMANCE

Coarse Blank F6
Analyte	No. Blanks	Failed Blanks	Contamination Rate	Maximum Contamination	Average Contamination
Cu	218	5	2.3%	160 ppm	80 ppm
Zn	218	67	30.7%	1040 pm	170 ppm
Ag	218	7	3.2%	0.929 ppm	0.51 ppm
Au	218	18	8.3%	0.71 ppm	0.12 ppm

Coarse Blank F7
Analyte	No. Blanks	Failed Blanks	Contamination Rate	Maximum Contamination	Average Contamination
Cu	2060	36	1.7%	1340 ppm	130 ppm
Zn	2060	240	11.7%	2680 ppm	195 ppm
Ag	2060	12	0.6%	1.82 ppm	0.72 ppm
Au	2060	76	3.7%	2.53 ppm	0.23 ppm

Copper displays a low blank failure rate of <3%. The failure rate for zinc is high with 12 to 31% failed blanks. However, the grades of contaminated blanks are low with an average of 80 to 130 ppm for copper and 170 to 195 ppm for zinc. There are a few cases of contamination with moderate grades in the range of 0.1 to 0.13% copper, and 0.1 to 0.27% Zn.

The blank failure rate for silver is <5%. On average, silver contamination is in the range 0.5 to 0.7 ppm. A few blanks are contaminated with 0.9 to 1.8 ppm silver.

The blank failure rate for gold is 3 to 8% which is acceptable. The contamination with gold is high, relative to the blanks, with average grades ranging from 0.5 to 0.7 ppm. A few cases of contamination with up to 1.8 ppm gold are documented.

It is recommended to audit the laboratory and request a cleaner assaying protocol to (1) reduce the contamination with copper and zinc to levels below 0.1%, and (2) to reduce the average contamination with gold and silver to levels below 0.1 ppm. These cleaner laboratory conditions are achievable and desired for base and precious metals.

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Overall the performance of blanks indicates no significant issues with contamination at the Hudbay laboratory in Flin Flon. The contamination rates are generally low for copper, silver and gold, and high for zinc but at grade levels that are not economic. Therefore the results are acceptable for the resource estimation.

12.2.2 Standards

From 2012 to 2016 a total of 5,570 OREAS and Rocklabs CRMs were analyzed representing 5.4% of the total samples submitted to the Hudbay laboratory. Table 12-4 presents a list of all CRMs used by Hudbay for quality control over the past five years.

OREAS standards are a series of matrix-matched polymetallic CRMs prepared exclusively for Hudbay and sourced from the Flin Flon mine ore bodies. Rocklabs standards are a series of commercially available high-grade gold standards.

Section 12.1.2 explains the methodology for estimation of bias and assessment of the CRM performance. In summary, a successful assaying program aims at having a bias better than ±10%, and good reproducibility of the CRM assays given by a rate of failed CRMs <10%. Table 11-9 summarizes the performance of the CRMs at the Hudbay laboratory in Flin Flon.

The analytical bias is good for copper, and good to reasonable for zinc (Table 12-9). The CRM failure rate is <8% for copper and <5% for zinc. The laboratory coefficient of variation for copper and zinc is better than or comparable to the variation indicated by the CRM certificates (Table 12-9). These results suggest that the copper and zinc assays at the Hudbay laboratory have good accuracy and reproducibility for resource estimation.

The analytical bias for silver is good to reasonable for all standards except OREAS A6 which has a bias of 20%. However, the certified SD of OREAS A6 indicates a coefficient of variation of 20% for this standard, and the average silver value at Hudbay for OREAS A6 is within 2SD of the certified value. In addition, the silver grade of this standard (0.535 ppm) is close to the lower limit of detection for silver (0.4 ppm) at the Hudbay laboratory (Table 12-7 and Table 12-9). The failure rates for CRMs assayed for silver, including OREAS A6, are <4%. The coefficients of variation of the silver determinations at the laboratory are better than or comparable to the variation indicated in the CRM certificate. It is concluded that the accuracy and reproducibility of the silver assays at the Hudbay laboratory is of good quality for resource estimation.

Several problems were documented by the analysis of CRMs assayed for gold at the Hudbay laboratory in Flin Flon.

First, the relative bias ranges from good to reasonable, better than ±7%, for all OREAS CRMs (Table 12-9). However, negative biases of up to -20% are estimated for the high gold grade Rocklab standards. Rocklab standard SN75 was also analyzed at Bureau Veritas where the estimated bias was -0.5%, versus a bias of -10.8% at the Hudbay laboratory. Therefore, the large negative bias in the high-grade gold standards needs to be investigated with the laboratory in Flin Flon.

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Second, the variability of gold indicated by the CRMs assayed at the Hudbay laboratory is relatively large. For instance, the CRM gold values determined by Hudbay have coefficients of variation (CV) ranging from 8 to 24%, which is high relative to the indicated CVs in the CRM certificates ranging from 2 to 7%, and the CRMs assayed for gold at Bureau Veritas with CVs in the range of 2 to 9%.

The variability in the low gold grade (<1 ppm gold) OREAS CRMs can be explained by the relatively high lower limits of detection (0.1 ppm LOD) for gold at the Hudbay laboratory. Most of these low grade standards are within 10 times the LOD and therefore large variance is expected. However, the laboratory LOD does not satisfactorily explain the large variability of gold recorded by the assays of high gold grade Rocklab CRMs with CVs ranging from 10 to 17%. These variations are larger than the expected variation based on the CRMs certificates. The large variation in the determination of gold at the Hudbay laboratory results in a large number of CRM failures exceeding acceptable thresholds. This problem with the high-grade gold standards needs to be addressed with the laboratory.

To further investigate the impact of the poor performance of gold CRMs at the laboratory in Flin Flon, an ordinary least square regression of the average CRM values obtained at the laboratory was fit onto the recommended best values reported in the CRM certificates. For comparison, the same regression analysis was conducted for the average CRM assays from Bureau Veritas. The results are summarized on Figure 12-5 and Table 12-10.

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TABLE 12-9: SUMMARY OF CRM PERFORMANCE AT HUDBAY LABORATORY

Cu (%)
CRM	No. of Samples	No. of Failures	Failure Rate (%)	CRM Value (%)	HBMS Average	CRM CV	Lab CV	Relative Bias
OREAS A5	78	6	7.7%	0.0945	0.0908	3%	3%	-3.9%
OREAS B5	78	0	0%	1.33	1.31	3%	2%	-1.9%
OREAS C5	76	1	1.3%	3.37	3.28	6%	2%	-2.7%
OREAS D5	76	1	1.3%	9.42	9.47	4%	2%	0.5%
OREAS E5	76	0	0%	0.393	0.384	8%	2%	-2.2%
OREAS A6	761	2	0.3%	0.055	0.053	4%	9%	-4.3%
OREAS B6	747	6	0.8%	0.872	0.855	2%	2%	-2.0%
OREAS C6	748	7	0.9%	2.05	2.01	3%	2%	-2.1%
OREAS D6	747	4	0.5%	4.20	4.00	4%	2%	-4.8%
OREAS E6	736	5	0.7%	0.245	0.239	5%	3%	-2.3%
Zn (%)
CRM	No. of Samples	No. of Failures	Failure Rate (%)	CRM Value (ppm)	HBMS Average	CRM CV	Lab CV	Relative Bias
OREAS A5	78	0	0%	0.506	0.496	5%	3%	-2.1%
OREAS B5	78	2	3%	0.125	0.130	7%	8%	3.7%
OREAS C5	76	0	0%	5.28	5.09	11%	2%	-3.6%
OREAS D5	76	1	1%	2.46	2.28	7%	3%	-7.2%
OREAS E5	76	1	1.3%	23.70	23.77	4%	2%	0.3%
OREAS A6	761	17	2%	0.039	0.040	5%	10%	3.2%
OREAS B6	747	2	0.3%	0.702	0.665	3%	2%	-5.3%
OREAS C6	748	3	0.4%	2.41	2.30	3%	2%	-4.5%
OREAS D6	747	4	0.5%	3.31	3.10	4%	2%	-6.3%
OREAS E6	736	6	0.8%	18.14	18.06	4%	2%	-0.4%
Au (ppm)
CRM	No. of Samples	No. of Failures	Failure Rate (%)	CRM Value (ppm)	HBMS Average	CRM CV	Lab CV	Relative Bias
OREAS A5	72	39	54.2%	0.141	0.140	7%	24%	-0.7%
OREAS B5	79	58	73.4%	0.476	0.460	3%	20%	-3.4%
OREAS C5	76	1	1%	2.493	2.524	2%	5%	1.2%
OREAS D5	76	1	1%	7.344	7.107	4%	8%	-3.2%
OREAS E5	76	1	1.3%	0.780	0.813	3%	8%	4.3%
OREAS A6	569	569	100%	0.121	0.128	4%	23%	6.2%
OREAS B6	738	487	66.0%	0.706	0.676	2%	15%	-4.2%
OREAS C6	742	371	50.0%	0.987	1.007	2%	13%	2.0%
OREAS D6	741	24	3.2%	3.04	2.98	4%	8%	-2.0%
OREAS E6	731	428	58.5%	0.316	0.328	4%	13%	3.9%
Rocklabs SN60	181	170	93.9%	8.595	6.804	3%	13%	-20.8%
Rocklabs SN75	497	261	52.5%	8.671	7.734	2%	17%	-10.8%
Rocklabs SP59	187	127	67.9%	18.12	15.891	2%	15%	-12.3%
Rocklabs SP73	490	134	27.3%	18.17	17.28	2%	10%	-4.9%
Rocklabs SQ48	92	50	54.3%	30.25	27.84	2%	10%	-8.0%
Ag (ppm)
CRM	No. of Samples	No. of Failures	Failure Rate (%)	CRM Value (ppm)	HBMS Average	CRM CV	Lab CV	Relative Bias
OREAS A5	78	0	0%	1.91	2.03	7%	15%	6.4%
OREAS B5	72	2	2.8%	2.86	2.98	15%	8%	4.2%
OREAS C5	76	0	0%	21.1	21.33	11%	6.1%	1.1%
OREAS D5	76	2	3%	90.0	88.61	13%	6.0%	-1.5%
OREAS E5	76	1	1.3%	19.7	20.37	10%	6.6%	3.4%
OREAS A6	634	4	0.6%	0.535	0.642	20%	21.2%	20.1%
OREAS B6	747	4	0.5%	2.94	2.81	18%	6.6%	-4.3%
OREAS C6	748	4	0.5%	5.18	5.02	10%	4.6%	-3.1%
OREAS D6	747	3	0.4%	25.17	23.72	6%	3.2%	-5.7%
OREAS E6	736	7	1.0%	15.23	15.03	6%	3.4%	-1.3%

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FIGURE 12-5: XY COMPARISON OF THE AVERAGE GOLD VALUES DETERMINED BY
HUDBAY AND BUREAU VERITAS LABORATORIES ON CRMS VERSUS THE CRM
RECOMMENDED BEST VALUE*

The 95% confidence interval (CI) of the regression intercepts for both Hudbay and Bureau Veritas laboratories contain the zero value indicating that the intercepts are not significant and do not carry a major weight in the analysis of bias (Table 12-10).

TABLE 12-10: SUMMARY OF REGRESSION PARAMETERS

Analyte	Laboratory	Method	Intercept	Slope	95% CI-Intercept		95% CI-Slope		Bias
Au	HBMS	OLS	0.016	0.91	-0.285	0.317	0.89	0.94	-9%
Au	Bureau Veritas	OLS	0.011	0.99	-0.012	0.034	0.99	1.00	-1%

The average bias was calculated from the regression analysis as Bias (%) = OLSS-1; in which OLSS is the slope of the ordinary least square regression (OLS). Evidently, the Hudbay laboratory carries an average negative bias for gold of -9%, which may range between -11% and -6% (95% CI). In contrast, Bureau Veritas laboratory carries a very small negative bias of -1%, ranging from 0 to -1%. In summary, the laboratory in Flin Flon may underestimate the gold grades by -9% ±2.5 in the range of 0.1 to 30 ppm gold (Table 12-10, Figure 12-5).

It is concluded that the analytical accuracy and reproducibility of copper, zinc, and silver as indicated by the CRM analysis, at the Hudbay laboratory is appropriate for resource estimation. Gold grades are being under assayed and this issue needs to be discussed with the laboratory. This under assaying of gold standards will likely lead to an underestimation of gold in the resource estimate, since the proportion of samples assayed at the Hudbay laboratory is approximately 80% of the total samples assayed between 2012 and 2016.

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12.2.3 Duplicates

The following analysis is based on coarse duplicates analyzed during 2015 and 2016. There is no record of coarse duplicates between 2012 and 2014. Coarse duplicates, approximately one in every twenty samples, were submitted to the Hudbay laboratory in Flin Flon to monitor sub-sampling precision. After crushing to 10 mesh (2 mm), a coarse duplicate sub-sample were split and pulverized to ≥95% passing through 150 mesh (105 μm). The laboratory in Flin Flon uses riffle splitting and rotator splitting. However, rotator splitting is a relatively new feature in the laboratory and most of the samples discussed here were riffle split. The duplicate sample was analyzed immediately after its paired sample. Quarter-core twin sample duplicates and pulp duplicates were not analyzed in the sample stream.

Coarse duplicates were evaluated using the hyperbolic method developed by AMEC (Simón, 2004) and described in detail in Section 12.1.3. The coarse duplicates submitted at Hudbay laboratory were also evaluated using a failure boundary of 20% absolute relative error (RE). An acceptable level of sub-sampling variance is achieved when the failure rate does not exceed 10% of all sample pairs. Table 12-11 summarizes the results of duplicate analysis at the Hudbay laboratory and minimum and maximum plots are presented on Figures 12-6 to 12-9.

TABLE 12-11: SUMMARY OF COARSE DUPLICATE PERFORMANCE

Analyte	Duplicate Samples	Duplicate Failures	Total Rate of Failures	Practical Detection Limit	Accepted Absolute Relative Error (RE)
Cu	2208	16	0.7%	500 ppm	20%
Zn	2124	37	1.7%	500 ppm	20%
Ag	2173	76	3.5%	2 ppm	20%
Au	1456	162	11.1%	0.5 ppm	20%

The duplicate pairs display failure rates of <5% for copper, zinc, and silver indicating that the sub-sampling procedures employed by the Hudbay laboratory are of good quality for base metals and silver (Table 12-11, Figures 12-6 to 12-8).

Gold duplicate analysis display a failure rate of 11% when evaluated at 20% RE and a practical detection limit of 0.5ppm corresponding to five times the lower limit of detection (Table 12-11, Figure 12-9). When the gold duplicates are evaluated at a practical detection limit of 1 ppm, 10 times the LOD, and 20% RE, the failure rate is 7% which is acceptable. It is concluded that an adequate sub-sampling variance for gold at the laboratory in Flin Flon is reached only for gold grades above 1ppm.

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FIGURE 12-6: COPPER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

FIGURE 12-7: ZINC COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

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FIGURE 12-8: SILVER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

FIGURE 12-9: GOLD COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

12.3 Check Assaying

The check assays of duplicate pulp samples evaluated here only represent sample batches submitted to Bureau Veritas and Hudbay between 2015 and 2016. Check assays are not available for samples submitted before these dates.

A total of 304 representative pulp samples (1.5%) were selected and re-analyzed at SGS Canada Inc. (SGS) laboratory in Vancouver to assess the accuracy of assay results reported by Bureau Veritas (BV) and the Hudbay laboratory in Flin Flon relative to the umpire laboratory SGS. Only samples with ≥0.5ppm gold were submitted for re-analysis at the secondary laboratory. Copper, zinc, and silver were digested in aqua regia and analyzed by ICP-OES. Gold was fire assayed and analyzed by AAS. These methods are comparable to those used by Bureau Veritas and the Hudbay laboratory.

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CRMs, certified blanks, and pulp duplicates were not inserted along with the check samples, which leaves the secondary laboratory, SGS, untested for its quality.

A Reduced-to-Major-Axis regression (RMA) was used to evaluate the check samples. Assays from the primary laboratory were regressed onto the umpire laboratory. The RMA regression calculates an unbiased fit for values that are independent from each other. The results of the regression analysis are presented on Table 12-12 and Figures 12-10 and 12-11.

TABLE 12-12: SUMMARY OF RMA REGRESSION ANALYSIS

Analyte	Laboratory	Method	R.square	Intercept	Slope	95% CI-Intercept		95% CI-Slope		Bias
Cu	Bureau Veritas	RMA	0.999	0.000	1.003	-0.015	0.015	0.995	1.011	0.3%
Zn	Bureau Veritas	RMA	0.997	0.005	0.992	-0.003	0.013	0.979	1.006	-0.8%
Ag	Bureau Veritas	RMA	0.981	0.992	0.997	0.066	1.887	0.962	1.032	-0.3%
Au	Bureau Veritas	RMA	0.992	0.188	0.990	-0.009	0.380	0.970	1.010	-1.0%
Cu	HBMS	RMA	0.998	-0.018	1.044	-0.025	-0.011	1.038	1.050	4.4%
Zn	HBMS	RMA	0.998	0.013	0.975	0.008	0.017	0.970	0.981	-2.5%
Ag	HBMS	RMA	0.989	-1.305	1.031	-1.962	-0.659	1.015	1.046	3.1%
Au	HBMS	RMA	0.989	-0.462	1.036	-0.586	-0.340	1.022	1.051	3.6%

The coefficient of determination (R squared, R²) is used to assess the variance explained by the linear relationship between the pairs. Samples analysed at Hudbay and Bureau Veritas laboratories show a very good fit (R² > 0.98), relative to SGS, for copper, zinc, silver, and gold.

For Bureau Veritas-SGS pairs and Hudbay-SGS pairs, the 95% confidence interval of the regression intercepts include zero or are within five times the lower limit of detection indicating that the effect of the intercept on the analysis of bias is not significant.

The bias, expressed as a percent, is calculated as Bias (%) = RMAS-1; in which RMAS is the slope of the RMA regression. The slope of the RMA regression for the pairs analyzed at Bureau Veritas and SGS ranges between 0.990 and 1.003, and between 0.975 and 1.044 for those analyzed at the Hudbay laboratory and SGS. The estimated bias, relative to SGS, for all base and precious metals is < ±2% for Bureau Veritas and < ±5% for Hudbay laboratory.

The RMA regression results are illustrated on Figures 12-10 and 12-11, where the black line represents the y = x function and the red dashed line represents the RMA regression line.

The overall bias estimated by the RMA regression analysis, regression intercepts, and r-squared indicates that the accuracy achieved by Bureau Veritas and Hudbay for copper, zinc, silver, and gold, during 2015 and 2016, is of good quality for resource estimation.

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FIGURE 12-10: XY PLOTS OF CHECK ASSAY DATA COMPARING PRIMARY LABORATORY HUDBAY TO SECONDARY LABORATORY SGS*

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FIGURE 12-11: XY PLOTS OF CHECK ASSAY DATA COMPARING PRIMARY LABORATORY
BUREAU VERITAS TO SECONDARY LABORATORY SGS

12.4 Site Visit

Robert Carter, P.Eng., Lalor Mine Manager at Hudbay Manitoba Business Unit is a regular employee at Lalor and in his role continually conducts personal site inspections to become familiar with conditions in the mine, to observe the geology and mineralization and verify work completed. He last visited the mine on March 29, 2017.

12.5 Core Review

Robert Carter, P.Eng., reviewed the geological data and verifies drill core mineralization during his personal site inspections.

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12.6 Drilling Database

Drill hole data (header, down hole surveys, geological intervals, sample intervals, QAQC samples, and geotechnical details) is entered into the local acQuire database through various data entry (DE) objects. The core logger enters geological data into the acQuire database via a customized core logging DE object. The DE object is designed to follow the work flow of the data entry process and applies built in business rules and pick lists to ensure all data is entered consistently. This is essentially a preliminary validation check before data is committed to the database. Therefore all data such as: reported lengths, geology intervals, and sample intervals are automatically validated on entry to prevent overlapping intervals, duplicate sample numbers, spelling errors, etc. Specific gravity (SG) results, if measured in the core shack, are stored within Excel spreadsheets and emailed to the database manager. The SG values are uploaded in acQuire using a customized interface tool.

Once the core logging is complete the drill hole data is reviewed and approved by the senior mine geologist or a designate. This includes signing approvals that are built in various validation/verification objects to ensure all data (collar details, surveys, geology, sample details, etc) is checked thoroughly and is complete. The Hudbay database manager also does routine checks and reports any inconsistencies/discrepancies to the attention of the senior geologist or the core loggers. Once the senior geologist is satisfied with the drill hole input data, diamond drill reports are generated within acQuire and stored in the local directory in PDF format.

On receipt of the analytical results from the Hudbay Flin Flon laboratory and Bureau Veritas laboratory, the assay files (certificates, csv files, etc) are inspected prior to uploading into the acQuire database. QC results are examined and any discrepancies are flagged and reported either back to the laboratory or to the database manager. The Hudbay database manager also compares the analytical results to the logged visual estimates for copper, zinc and iron. Discrepancies are brought to the attention of the senior geologist and re-assaying may be requested if the significance of the interval is warranted. Once the preliminary checks are complete, the analytical results (including laboratory measured SG values) for assay intervals are uploaded directly into the acQuire database using the sample number as the unique identifier. The database manager informs the senior geologist after the assays have been uploaded in the database. The senior geologist then completes the assay verification and approval process.

QC and duplicate assay results from both laboratories are kept in the acQuire database as unique identified variables. This information is readily available and can be easily charted externally or directly within the database system.

Drill core from the initial surface drilling program is stored at the Hudbay Hangar site near Flin Flon and a designated core storage location near the Stall concentrator. Drill core from the underground drilling programs is stored at the Lalor mine core storage area near the Chisel North mine site. Fine pulp rejects are kept in perpetuity in lidded plastic pails at the Hudbay Hangar and at the Lalor core storage area in sea cans.

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12.7 Mineral Resource Database Management

All information used in the estimation of the mineral resources was extracted directly from the Hudbay acQuire database management system managed by the database manager. The MineSight software package used for the 3D modeling and grade interpolation downloads information directly from this database.

In the author’s opinion, the drill hole and assay database is acceptable for resource estimation.

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13 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 Summary

The Stall concentrator, located 16km from Lalor mine was commissioned in 1979. Since then, it has processed ore from many of the Snow Lake area mines. Prior to processing ore from Lalor mine, Stall concentrator processed ore from the Chisel North mine from 2000 to 2013 with similar mineralogy to that of Lalor. The processed ore from Chisel North mine produced only a zinc concentrate, while Lalor ore with increased copper and precious metal content warranted the refurbishment of the defunct copper circuit and an overall throughput increase as the Lalor mine ramped up production.

The Stall concentrator began processing ore from Lalor in August 2012, initially producing only a zinc concentrate and by October 2012 a copper concentrate. As Lalor increased ore production the mill underwent an initial expansion and by the summer of 2014 was capable of processing at 2,800 tpd throughput rate.

It is the author’s opinion that actual plant metallurgical performance overrides previous metallurgical testing results, since the metallurgical blend from Lalor is not expected to materially change over the life of mine. It is appropriate to assume that previous actual performance recoveries are expected in the future.

13.2 Plant Metallurgical Performance

The Stall concentrator metallurgical performance milling of Lalor ore from October 2012 to December 2016 is shown in Table 13-1 and actual concentrate produced is shown in Table 13-2.

TABLE 13-1: HISTORICAL PLANT HEAD ASSAY AND METAL RECOVERY

		Head Assays				Metal Recoveries
Year	Tonnes	Au (g/t)	Ag (g/t)	Cu (%)	Zn (%)	Au (%)	Ag (%)	Cu (%)	Zn (%)
2012	59,787	1.33	15.31	0.60	9.23	47.6	55.5	68.0	94.0
2013	422,287	1.51	29.38	0.60	8.91	59.6	56.5	77.7	94.9
2014	526,015	2.31	24.05	0.89	8.49	53.3	50.8	71.4	93.7
2015	928,501	2.53	21.28	0.71	8.21	55.8	54.8	84.5	90.8
2016	1,089,530	2.25	21.67	0.63	7.03	57.4	56.4	82.1	92.8

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TABLE 13-2: HISTORICAL PLANT CONCENTRATE PRODUCED

	Zinc Concentrate		Copper Concentrate
Year	Tonnes	Zn (%)	Tonnes	Au (g/t)	Ag (g/t)	Cu (%)
2012	10,260	56.4	1,450	28.0	383.3	18.5
2013	67,346	51.5	9,784	39.2	386.9	20.2
2014	81,946	51.1	16,518	39.2	388.4	20.1
2015	133,808	51.8	26,848	48.7	403.1	20.7
2016	138,056	51.5	27,299	51.6	488.2	20.7

Currently the Stall concentrator is producing a copper concentrate grade of 21% copper at 83 to 85% recovery and a zinc concentrate grade of 51% zinc at 90 to 95% recovery. Gold and silver are recovered to the copper concentrate as co-products. Lalor ore has about 0.2 to 0.3% lead head grade and since the mill process is not configured to separate the lead from copper, lead reports to the copper concentrate. Lead grade in copper concentrate ranges from 5 to 10% and is dependent on the lead head grade and copper/lead ratio. The smelter charges a penalty for lead in copper concentrate and no economic value is received from lead.

The copper concentrate produced at Stall concentrator contains 3 to 6% zinc primarily due to the liberation. According to previous mineralogy studies there is a certain amount of sphalerite inevitably locked in chalcopyrite in ultrafine size fractions.

The Lalor ore base metal properties are not expected to vary significantly from the previous four years of milling and it is appropriate to assume that the metal recoveries will remain in the 80 to 85% range for copper and 90 to 95% range for zinc for the remaining LOM. The yearly LOM metal recoveries, shown in Table 13-3 were calculated using Hudbay’s in-house metallurgical model that considers the relationship of metal grade versus recovery from historical data at optimal operating days. Optimal operating days are considered to be steady run and appropriate control of parameters.

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TABLE 13-3: EXPECTED LOM RECOVERIES AT STALL CONCENTRATOR

	Metal Recoveries				Concentrate Grade
Year	Au (%)	Ag (%)	Cu (%)	Zn (%)	Zn (%)	Cu (%)
2017	59.6	51.9	83.9	93.6	51.0	21.0
2018	53.5	46.7	83.9	91.8	51.0	21.0
2019	55.8	48.2	83.1	91.7	51.0	21.0
2020	57.9	57.5	88.2	90.0	51.0	21.0
2021	61.9	66.1	89.1	90.5	51.0	21.0
2022	62.3	66.9	88.6	91.8	51.0	21.0
2023	60.1	60.8	88.2	91.8	51.0	21.0
2024	55.3	47.6	86.1	93.5	51.0	21.0
2025	59.7	52.2	87.7	91.7	51.0	21.0
2026	52.6	38.8	84.7	92.0	51.0	21.0
2027	55.8	39.5	81.7	90.6	51.0	21.0

13.3 Metallurgical Testing

Although it is the author’s opinion that actual plant performance overrides previous metallurgical testing it is appropriate to summarize the relevant testing completed on Lalor mineralization prior to processing ore at Stall concentrator.

The following reports were used in the preparation of this summary:

1.	A Report on the Recovery of Copper, Zinc, Gold and Silver from Lalor Samples SGS Vancouver Metallurgy July 20, 2009
2.	An Investigation into the Recovery of Copper, Lead, Zinc and Gold from Lalor Samples – Phase II
	SGS Vancouver Metallurgy
	February 25, 2011
3.	Pre-Feasibility Study Technical Report on the Lalor Deposit, Snow Lake, Manitoba, Canada. Effective Date: March 29, 2012

The primary objectives of the test programs were to develop an appropriate flowsheet for either the design of a new concentrator or modifications to the existing Stall concentrator, and to determine expected concentrate grades and metal recoveries.

All aspects of the test program including work on gold zone material were covered in the previous technical report, dated March 29^th, 2012. The current LOM plan does not specifically target mining of gold zone material unless it is in contact with base metals, which is consistent with previous mining practices at Lalor since 2012. This summary only includes laboratory test results on base metal ores from Zone 10 and Zone 20, focussing on the locked cycle flotation tests done on the variability composites from these two zones.

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These ore types have been processed at the Stall Concentrator for several years now, and plant recoveries from 2014 through 2016 will be compared to the earlier locked cycle flotation results.

Mineralogical analysis showed that chalcopyrite in Lalor ore is mostly coarse grained and liberated at grind sizes of approximately 100 microns. However, 15 to 20% of the chalcopyrite remains locked, primarily with sphalerite, at sizes below 20 microns. Copper is present almost exclusively as chalcopyrite with minor bornite. Zinc is present mainly as sphalerite, with minor amounts of gahnite. The sphalerite is coarse grained and liberated at a grind size of 250 microns. Lead is present as fine grained galena and would require a grind size of 70 microns for liberation. However there is insufficient galena in the ore to warrant a primary grind this fine.

Ore hardness was measured with Bond Work Index tests in the first phase of work. The average rod mill work index was 6.8 kwh/t on the base metals and 10.6 kwh/t on the contact gold material. The ball mill work indexes averaged 10.5 kwh/t on the base metals and 13.2 kwh/t on the contact gold material.

In the second phase of work at SGS, four variability composites were prepared from Zone 10 and Zone 20 ores as shown in Table 13-4.

TABLE 13-4: MAKE-UP OF VARIABILITY COMPOSITES

	Composite 1	Composite 2	Composite 3	Composite 4
	High grade Zn	High grade Zn	Low grade Zn	Low grade Zn
	High grade Cu	Low grade Cu	High grade Cu	Low grade Cu
	High grade Au	Low grade Au	High grade Au	Low grade Au
Percent by Zone
10	35	75	40	34
20	65	25	60	66
Head Assays, g/t, %
Au	2.98	0.29	4.29	1.14
Ag	21.10	12.30	21.30	16.70
Cu	0.95	0.22	0.96	0.42
Zn	11.00	10.30	5.23	4.42
Pb	0.23	0.23	0.21	0.26
Fe	22.90	23.20	17.10	16.30
S	24.60	22.80	16.60	14.70

The conditions for the locked cycle tests on the variability composites are shown in Table 13-5, and the results are shown in Table 13-6.

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TABLE 13-5: LOCKED CYCLE TEST CONDITIONS – VARIABILITY COMPOSITES

Process	Conditions	Composite 1	Composite 2	Composite 3	Composite 4
Primary Grind	P80 Size, microns	79	80	82	80
	Water	Tap Water	Tap Water	Tap Water	Tap Water
	pH	8.6	8.6	8.6	8.6
	NaCN/ZnSO4, g/t	20/60	20/60	20/60	20/60
Cu-Pb Bulk Roughing	3418A, g/t	40	40	40	40
	pH	9.5 - 9.6	9.5 - 9.6	9.5 - 9.6	9.5 - 9.6
Cu-Pb Bullk Conc Regrind	P80 Size, microns	30	30	30	30
Cu-Pb Bulk Cleaning	3418A, g/t	15	15	15	15
	pH	10.5	10.5	10.5	10.5
	No. of stages	3	3	3	3
Cu-Pb Separation	NaCN, g/t	300	400	400	400
	3418A, g/t	4	4	4	4
Pb Cleaning	3418A, g/t	4.5	4.5	4.5	4.5
	pH	10.5	10.5	10.5	10.5
	No. of stages	4	4	4	4
Zn Roughing	CuSO4, g/tt	550	500	210	180
	Xanthate (SIPX), g/t	55	50	30	25
	pH	11.5	11.5	11.5	11.5
Zn Cleaning and 1st Cleaner	Xanthate (SIPX), g/t	6	2	2	3
Scavenging	pH	11.5	11.5	11.5	11.5
	No. of stages	3	3	3	3

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TABLE 13-6: LOCKED CYCLE TEST RESULTS – VARIABILITY COMPOSITES

		Assays, % or g/t				Distribution, %
Product		Composite 1	Composite 2	Composite 3	Composite 4	Composite 1	Composite 2	Composite 3	Composite 4
Feed	Wt					100	100	100	100
	Cu	0.95	0.25	0.90	0.44	100	100	100	100
	Pb	0.22	0.23	0.19	0.27	100	100	100	100
	Zn	10.76	10.30	5.27	4.17	100	100	100	100
	Au	2.87	0.20	3.40	0.93	100	100	100	100
	Ag	22.1	11.3	22.6	15.7	100	100	100	100
Lead Concentrate	Wt					1.27	1.00	1.17	1.51
	Cu	0.71	0.21	0.54	0.36	0.95	0.84	0.71	1.25
	Pb	14.68	20.36	12.86	15.22	84.72	87.07	76.19	85.89
	Zn	0.98	0.68	0.72	0.54	0.12	0.07	0.16	0.20
	Au	24.07	0.93	32.75	2.04	10.66	4.56	11.25	3.30
	Ag	145.0	196.0	154.0	151.0	8.4	17.3	8.0	14.6
Copper Concentrate	Wt					3.06	0.72	3.05	1.38
	Cu	27.40	24.30	27.60	27.20	88.52	70.27	93.32	86.49
	Pb	0.44	1.24	0.66	1.01	6.09	3.81	10.46	5.21
	Zn	6.52	7.71	3.32	3.64	1.85	0.54	1.92	1.21
	Au	58.74	12.43	62.60	36.79	62.66	43.66	56.04	54.60
	Ag	409.0	447.0	434.0	567.0	56.7	28.4	58.6	50.1
Bulk Concentrate	Wt					4.33	1.72	4.21	2.89
	Cu	19.57	10.28	20.10	13.20	89.47	71.11	94.03	87.74
	Pb	4.62	12.37	4.04	8.42	90.81	90.88	86.65	91.10
	Zn	4.89	3.62	2.60	2.02	1.97	0.61	2.08	1.41
	Au	48.57	5.74	54.32	18.66	73.32	48.22	67.29	57.90
	Ag	331.6	300.9	356.4	349.9	65.0	45.7	66.6	64.7
Zinc Concentrate	Wt					16.17	15.75	7.92	6.71
	Cu	0.22	0.19	0.26	0.27	3.70	12.25	2.29	4.16
	Pb	0.02	0.02	0.03	0.04	1.22	1.35	1.14	1.09
	Zn	60.55	60.97	57.42	57.95	90.99	92.36	86.38	93.24
	Au	0.23	0.06	0.78	0.32	1.28	4.36	1.82	2.30
	Ag	13.5	15.5	17.9	16.3	9.9	21.6	6.3	7.0
Zinc Cleaner	Wt					5.46	11.25	4.22	3.53
Scavenger Tails	Cu	0.15	0.36	0.34	0.45	2.13	6.60	1.67	2.78
	Pb	0.04	0.08	0.10	0.17	1.24	1.68	1.65	1.32
	Zn	3.23	4.16	3.45	3.26	3.14	3.53	3.33	2.92
	Au	0.15	3.70	1.41	5.31	2.85	8.02	4.59	5.34
	Ag	7.5	26.2	22.3	28.4	4.2	7.4	4.9	5.0
Zinc Rougher Tails	Wt					74.05	71.28	83.65	86.86
	Cu	0.08	0.03	0.02	0.03	4.69	10.03	2.02	5.32
	Pb	0.02	0.02	0.02	0.02	6.73	6.09	8.56	6.49
	Zn	0.57	0.38	0.52	0.12	3.90	2.61	8.21	2.43
	Au	0.88	0.11	1.07	0.37	22.75	39.40	26.31	34.46
	Ag	6.20	4.00	6.00	4.20	20.82	25.33	22.24	23.30

A copper/lead separation stage was in the laboratory flowsheet, but will not be in the plant flowsheet.

The copper concentrate plant product is equivalent to the bulk concentrate in Table 13-5.

Comparisons between laboratory locked cycle test (LCT) results and monthly average plant operating results are shown in Figure 13-1 to Error! Reference source not found. and discussed below.

Recoveries for copper and gold (Figure 13-1) in the plant are generally lower than the laboratory recoveries. This is probably due to the combined effects of the following factors:

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• The ore blends in plant feeds are more variable and include small amounts of ore from zones that were not represented in the variability composites tested in the laboratory

• The plant grind is coarser than the laboratory grind (approximately 100 micron vs. 80 micron). The grind design criterion for the proposed concentrator upgrade project is 100 micron.

• The plant has been able to improve upon the concentrate grades that were achieved in the laboratory (Figure 13-2). Concentrate grades and recoveries are inversely related.

FIGURE 13-1: COPPER AND GOLD RECOVERY TO CONCENTRATE, LCT AND PLANT DATA

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FIGURE 13-2: COPPER CONCENTRATE GRADES, LCT AND PLANT DATA

Average gold and copper head grades in the plant feed have been a little higher than the laboratory head grades at 0.71% Cu and 2.36 g/t Au, compared to 0.64% Cu and 1.85 g/t Au in the laboratory head samples.

Zinc recoveries achieved in the plant are generally higher than the laboratory recoveries (Figure 13 - 3) while zinc concentrate grades in the plant have been lower than the laboratory results (Figure 13 – 4). This is a preferred operating option due to Hudbay’s short concentrate haul to their Flin Flon metallurgical site.

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FIGURE 13-3: ZINC RECOVERY TO CONCENTRATE, LCT AND PLANT DATA

FIGURE 13-4: ZINC CONCENTRATE GRADES, LCT AND PLANT DATA

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Zinc head grades in the plant feed have been slightly higher than the laboratory head grades at 7.76% Zn compared to 7.64% Zn in the laboratory head samples.

Metallurgical forecasts for this report will be based on recent plant data.

Penalty charges of the copper concentrate contract are as follows:

Zinc and Lead (Combined)

• US $3.00 per dmt for every 1% above 3% up to 8%

• US $5.00 per dmt for every 1% above 8% up to 12%

• US $8.00 per dmt for every 1% above 12%

Mercury

• US $2.00 per dmt for every 10ppm above 15ppm

Moisture

• US $2.00/wmt for every 1% above 10%

In 2016, approximate penalty charges of the deleterious elements of the copper concentrate in $US per tonne were 23.20, 2.21, and 15.06 for zinc and lead, mercury, and moisture respectively. The penalty charges do not have a significant effect on the potential economic extraction.

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14 MINERAL RESOURCE ESTIMATES

Hudbay prepared an update of the Lalor mine 3D block model using MineSight® version 11.60, industry standard commercial software that specializes in geologic modelling and mine planning. The 3D block model and determination of the updated mineral resources at the Lalor mine were performed by Hudbay personnel following Hudbay procedures. The work was reviewed and approved by Robert Carter, P.Eng., Lalor Mine Manager, Hudbay Manitoba Business Unit and Qualified Person of this Technical Report.

14.1 Key Assumptions of Model

As shown in Table 14-1, there are 1,707 assayed drill holes totalling approximately 420,310 m within the Lalor database used to support the mineral resource estimate.

TABLE 14-1: DRILLING DATA BY YEAR

The drill hole database was exported from acQuire®, validated and provided in Microsoft Excel® format with a cut-off date for mineral resource estimate purposes of September 30, 2016. The files were imported as collar, downhole survey and assay data into MineSight.

14.2 Wireframe Models and Mineralization

The Lalor mineralized envelopes trend along an azimuth of approximately 320° with a general dip of approximately 37° to the north east. The volcanogenic massive sulphide deposit is hosted in a Paleoproterozoic bimodal volcanic sequence. Seven zinc-rich massive sulphide lenses (i.e. 10, 11, 20, 30, 31, 32 and 40) and seven gold-rich disseminated to semi-massive zones (i.e. 21, 23, 24, 25, 26, 27 and 28) have now been identified. The base metal lenses and gold rich zones are stratigraphically stacked and are mostly aligned to the main regional deformation S2 foliation, averaging N320°/37°. The Lalor mine is continuous along a strike length of approximately 1,600 m in north-south direction, approximately 700 m in an east-west direction and with a vertical extent of approximately 830 m.

Four sets of structures were recognized by Hudbay and SRK geologists (Ravenelle, 2016), the Manitoba Geological Survey and the Geological Survey of Canada: the bedding S0 with an average N317°/36°, the dominant foliation S2 averaging N318°/37°, the late foliation S3 averaging N012°/81° and the fold F3 with a fold axis of 26° → 012°.

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The base metal grade shells were built using MineSight, from 2D cross sections linked to create solids and verified in plan, using an approximate 4.1% zinc equivalency cut-off. The gold-rich grade shell were built with Leapfrog® version 4.0 and were constrained with the logging, lithogeological data and an approximate 2.4 g/t gold equivalency or an approximate 1.9% copper equivalency. The gold-rich mineralized envelopes interpretation was stretched to follow the geological continuity of the zones. Wireframes of the mineralized envelopes are shown in Figure 14-1 (refer to Table 14-2 for the grade shells legend).

The term “Zone”, “Lense” and “Mineralized Envelope” is used interchangeably throughout this section and the report in general. Although Zone 27 is referred to as a gold-rich zone it contains appreciable amounts of copper and is amenable to processing in either a base metal or gold concentrator.

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FIGURE 14-1: 3D VIEW OF WIREFRAMES, LOOKING WEST

Note: Lense 23 is located under lense 32 and is hidden in this view.

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TABLE 14-2: LEGEND OF INTERPRETED WIREFRAMES

14.3 Density Assignation

Density values generated from two multi elements regression formulas based on 65,792 measurements collected by Hudbay and measured at Flin Flon laboratory, ACME/Bureau Veritas laboratory and at Hudbay logging facility, using a non-wax-sealed immersion technique to measure the weight of each sample in air and in water.

Two regression formulas were used to predict the density value of intervals without a measured density:

Regression formula in the base metal lenses

Regression formula in the gold zones

The hybrid density (i.e. measured values when available, predicted when missing) was interpolated to each mineralized envelop via ordinary kriging. The interpolation results were validated against a nearest neighbour model (NN) and an inverse distance weighting (IDW) models which show similar distribution (Figure 14-2).

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FIGURE 14-2: OK, IDW AND NN SPECIFIC GRAVITY DISTRIBUTION

Note: OK SG in blue, IDW SG in green and NN SG in red.

14.4 Exploratory Data Analysis

Exploratory data analysis (EDA) includes basic statistical evaluation of the assays and composites for zinc (Zn), gold (Au), silver (Ag), copper (Cu), lead (Pb), iron (Fe), arsenic (As), density (SG), and samples length.

14.4.1 Assays

The Table 14-3 presents the number of samples collected and total length analyzed.

TABLE 14-3: SAMPLES AND LENGTH ANALYZED

Box Plots

Box plots of the basic statistics for zinc, gold, copper and silver for each mineralized envelope are displayed in Figure 14-3 to Figure 14-6. These box plots confirm the occurrence of two mineralization systems at Lalor: higher grade zinc mineralization (zones 10, 11, 20, 30, 31, 32 and 40) and gold rich mineralization (zones 21, 23, 24, 25, 26, 27 and 28).

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FIGURE 14-3: BOX PLOT OF ZINC (%) BY LENSE

FIGURE 14-4: BOX PLOT OF GOLD (G/T) BY LENSE

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FIGURE 14-5: BOX PLOT OF COPPER (%) BY LENSE

FIGURE 14-6: BOX PLOT OF SILVER (G/T) BY LENSE

Assay Statistics

The EDA of assay statistics of zinc, gold, copper, silver, lead, iron, arsenic and density are summarized in Table 14-4 to Table 14-11.

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TABLE 14-4: ZINC ASSAY STATISTICS BY LENSE

TABLE 14-5: GOLD ASSAY STATISTICS BY LENSE

TABLE 14-6: COPPER ASSAY STATISTICS BY LENSE

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TABLE 14-7: SILVER ASSAY STATISTICS BY LENSE

TABLE 14-8: LEAD ASSAY STATISTICS BY LENSE

TABLE 14-9: IRON ASSAY STATISTICS BY LENSE

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TABLE 14-10: ARSENIC ASSAY STATISTICS BY LENSE

TABLE 14-11: DENSITY ASSAY STATISTICS BY LENSE

The high coefficient of variation (CV) values of copper and gold in the gold zones suggest that the gold and copper grades are variable and that a method of modeling using additional constraints may be better suited to these zones.

Contact Plots

Given the fact that for the most part, the different mineralized envelopes are separated by barren volcanic rocks, contact plots were not analyzed and a strict code matching system between the composites and block model has been used for the grade interpolation.

Grade Capping and High Yield Restriction

Since most of the mineralized envelopes show a high skewness in the statistical distribution of gold and silver grade, length weighted, log-scaled probability plots, deciles analysis and disintegration analysis of the assays were used to define high-grade outliers for gold and silver within each of the separately evaluated domains. These high-grade outliers can lead to overestimation of average grades unless some means of moderating the effect of the high-grade samples is applied. A common method for accomplishing this is by capping high assays at some predetermined level prior to grade estimation. In reviewing the statistics by zone it was decided to cap high grade gold and silver for all zones. Capping was completed on the assays prior to compositing. As an example, the log probability plot and the histograms of the gold distribution in zone 25 are displayed below in Figure 14-7 and Figure 14-8.

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FIGURE 14-7: LOG PROBABILITY PLOT OF AU (G/T) IN ZONE 25

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FIGURE 14-8: HISTOGRAMS AND BASIC STATISTICS OF THE POPULATION ABOVE AND
BELOW CAPPING IN ZONE 25 FOR AU (G/T)

In reviewing statistics of zinc, copper, lead, iron and density the composite data showed some high grade values were discontinuous from the remainder of the data set and there was justification of restricting these higher grade values by limiting their search distance. A high yield restriction of 20 m (i.e. four blocks) was applied at the interpolation stage based on the 95^th percentile of the 1.25 m composites values instead of capping. The capping and high yield thresholds are shown below in Table 14-12 and Table 14-13 respectively.

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TABLE 14-12: CAPPING THRESHOLDS BY LENSE OF GOLD AND SILVER (G/T)

TABLE 14-13: HIGH-YIELD RESTRICTION THRESHOLDS BY LENSE

Table 14-14 presents the overall gold and silver metal removed by capping of the assays. It is not uncommon to see large differences between the original assay and capped value as in zone 21, 24, 25 and 28. These gold zones could benefit from a more constraining modeling method which could generate a statistical distribution closer to stationarity, and could present a smaller impact on the overall metal loss.

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TABLE 14-14: PRECIOUS METAL REMOVED BY CAPPING

14.4.2 Composites

In order to normalize the weight of influence of each sample, assay intervals were regularized by compositing drill hole data using the interpreted mineralized envelopes to break composites. A 1.25 m length was selected as the composites length, which represents approximately 99% of all the samples intervals. Figure 14-9 presents the distribution of the assay intervals.

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FIGURE 14-9: SAMPLE LENGTH IN MINERALIZED ZONES

The 1.25 m intervals (+/- 0.62 m of threshold) were composited using “honour geology” from the coded drill hole file. The compositing was weighted both on length and density, and the weighting factor stored in the composite file (WFACT).

For bias assessment purposes, assay intervals were also composited into 5 m lengths (+/- 2.5 m of threshold) using the same methodology. The 5 m composites were used to estimate a nearest neighbour model (NN). EDA of the 1.25 m and 5 m composite statistics for Zn, Au Cu and Ag are shown in Table 14-15 to Table 14-22.

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TABLE 14-15: LENGTH WEIGHTED 1.25 M COMPOSITE STATISTICS, ZINC (%)

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TABLE 14-16: LENGTH WEIGHTED UNCAPPED AND CAPPED 1.25 M COMPOSITE STATISTICS, GOLD (G/T)

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TABLE 14-17: LENGTH WEIGHTED 1.25 M COMPOSITE STATISTICS, COPPER (%)

TABLE 14-18: LENGTH WEIGHTED UNCAPPED AND CAPPED 1.25 M COMPOSITE STATISTICS, SILVER (G/T)

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TABLE 14-19: LENGTH WEIGHTED 5 M COMPOSITE STATISTICS, ZINC (%)

TABLE 14-20: LENGTH WEIGHTED UNCAPPED AND CAPPED 5 M COMPOSITE STATISTICS, GOLD (G/T)

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TABLE 14-21: LENGTH WEIGHTED 5 M COMPOSITE STATISTICS, COPPER (%)

TABLE 14-22: LENGTH WEIGHTED UNCAPPED AND CAPPED 5 M COMPOSITE STATISTICS, SILVER (G/T)

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The length weighted mean grades of both 1.25 m and 5 m length composites are similar to those of the assays; therefore providing confidence that the compositing process is working as intended. The low to moderate CV values of zinc within the based metal zones suggest that no further domaining is warranted and that a linear interpolation method can be used. As for the gold zones, the relatively high CV values suggest that the gold and copper grades are variable and that a linear interpolation method may produce skewed results. Applying non-linear interpolation methods and/or revisions of the wire framing criteria should be further investigated for these zones in the future update of the resource model.

14.4.3 Variography

Down-hole and directional correlograms for Zn, Au, Cu, Ag, Pb, Fe, As and SG based on each individual grade shells were created using SAGE® software. Due to limited number of composite pairs in the raft portions of lenses 24, 25, 26, 27 and 28, the analysis was conducted on the primary lenses only. Lense 28 did not compute valid variograms and IDW was the interpolation method used.

The Zn variograms show low nugget effects in the base metal lenses with values lower or equal to 5% of the total sill. The ranges of correlation generally vary between 60 and 200 m in the major axis, 37 to 80 m in the semi-major axis and 12 to 54 m in the minor axis. As for gold, the nugget effect varies from low to high, up to 66% of the total sill. The ranges of correlation generally vary between 40 and 90 m in the major axis, 19 to 40 m in the semi-major axis and 7 to 25 m in the minor axis.

As an example, the downhole variogram of Zn in lense 32 and gold in zone 25 are shown in Figure 14-10 and Figure 14-11 respectively. A nugget and a nested model were fitted to the experimental correlograms. Correlogram model parameters for Zn, Cu, Au and Ag are shown in Table 14-23 and Table 14-24, both for the base metal lenses and the gold zones.

FIGURE 14-10: DOWNHOLE VARIOGRAM ZINC, LENSE 32

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FIGURE 14-11: DOWNHOLE VARIOGRAM GOLD, LENSE 25

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TABLE 14-23: VARIOGRAM MODELS AND ROTATION ANGLES IN BASE METAL LENSES

Note: Ranges are in metres and search ellipse orientations are given using GSLIB-MS rotation convention.

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TABLE 14-24: VARIOGRAM MODELS AND ROTATION ANGLES IN GOLD ZONES

Note: Ranges are in metres and search ellipse orientations are given using GSLIB-MS rotation convention.

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14.5 Estimation and Interpolation Methods

The mineralized envelopes were used to code blocks in the model using a multi ore percent method, where ORE1 and ORE%1 represent the majority block, ORE2 and ORE%2 represent the partial block and ORE3 and ORE%3 represent the minority block. Grade interpolation used a strict composite and block matching system based on the mineralized envelopes codes. For example, in the case of Lense 10, only composites coded as Lense 10 were used to interpolate blocks grades.

The block model consists of regular blocks (5 m along strike by 5 m across strike by 5 m vertically). The block dimensions were selected based on the resource wireframe widths and to match the smallest mining unit at Lalor mine. The ordinary kriging (OK) grade interpolation was completed on the uncapped (or unrestricted) and capped (or restricted) with composites of 1.25 m in length, using three passes with increasing requirements. The composite selection parameters for grade estimation in each domain (minimum, maximum, maximum number of composites per hole) were selected to minimize bias.

The first interpolation pass uses 100% of the variogram ranges and is restricted to a minimum of four composites, a maximum of 16 composites (with a maximum of four composites per hole), without quadrant declustering. The second interpolation pass uses 75% of the variogram ranges and is restricted to a minimum of six composites, a maximum of 16 composites (with a maximum of four composites per hole) and uses quadrant declustering. Finally, the third interpolation pass uses 50% of the variogram ranges and is restricted to a minimum of nine composites, a maximum of 16 composites (with a maximum of four composites per hole) and uses quadrant declustering. At the end of the interpolation, more than 95% of the blocks within the mineralized envelopes had been interpolated. In order to estimate the remaining un-interpolated blocks, a “fill pass” was added at the process, using a minimum of two composites to interpolate the blocks. The “fill pass” used ellipses sizes equivalent to 150% to 225% of the variograms ranges.

The interpolations use both the length and density of the composites (WFACT) to weight the grade of the nearest neighbour (NN using 5 m composites), the inverse distance (IDW) using and the ordinary kriging (OK) interpolations. The NN and IDW interpolations were used to monitor the quality of the OK interpolation.

14.6 Block Model Validation

The Lalor block model was validated to ensure appropriate honouring of the input data by the following methods:

• Visual inspection of the OK block model grades in plan and section views in comparison to composites grade

• Metal removed via grade capping and high yield restriction methodology

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• Comparison between the interpolation methods of NN and IDW to confirm the absence of global bias in the OK grade model

• Swath plot comparisons of the estimation methods to investigate local bias

• Review of block model ordinary kriging quality control parameters

• Comparison of grade tonnage curves and statistics by estimation method

• Third party review of the block model and estimation process

14.6.1 Visual Inspection

Visual inspection of block grade versus composited data was conducted in section view. The visual inspection of block grade versus composited data showed a good reproduction of the data by the model. As an example, two long sections (looking west) are presented. Figure 14-12 presents the zinc grade in Lense 10 while Figure 14-13 presents the gold grade in Lense 25.

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FIGURE 14-12: EW CROSS SECTION N630 SHOWING OK MODEL AND COMPOSITES ZINC GRADE OF BASE METAL LENSE 10

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FIGURE 14-13: EW CROSS SECTION 1250N SHOWING OK MODEL AND COMPOSITES GOLD GRADE IN GOLD ZONE 25

14.6.2 Metal Removed by Capping

The impact of capping was evaluated by estimating uncapped and capped grade and the unrestricted and restricted grade in the block model. Generally the amounts of metal removed by capping from the different interpolation methods are consistent with the difference of the capped and uncapped assays.

The percentages of metal removed by capping from the NN, IDW and OK models are shown in Table 14-25 to Table 14-28. The amount of capping for gold in the raft part of zone 26 is high. The limited drilling (35 samples only) is the likely source of difference between the un-capped and capped values. This zone is considered inferred resource and should be drilled in the future.

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The amount of zinc metal loss in the base metal lenses is lower than 5% for all the lenses except for lense 31 (-9%) and lense 40 (-13%). It is believed that the larger drill spacing in parts of these two lenses is the main source of difference between the un-restricted and restricted values.

TABLE 14-25: NN, IDW AND OK MODEL, ZINC REMOVED BY RESTRICTION IN BASE METAL LENSES

TABLE 14-26: NN, IDW AND OK MODEL, GOLD REMOVED BY CAPPING IN BASE METAL LENSES AND GOLD ZONES

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TABLE 14-27: NN, IDW AND OK MODEL, COPPER REMOVED BY RESTRICTION IN BASE METAL LENSES AND GOLD ZONES

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TABLE 14-28: NN, IDW AND OK MODEL, SILVER REMOVED BY CAPPING IN BASE METAL LENSES AND GOLD ZONES

14.6.3 Global Bias Checks

A comparison between the interpolation methods estimates was completed on all the blocks within the lenses. Differences between the NN, IDW and OK grades are acceptable in most lenses aside from the raft were only limited information is available. The differences are summarized in Table 14-29 to Table 14-32.

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TABLE 14-29: NN, IDW AND OK MODEL STATISTICS MEAN BLOCK GRADE COMPARISONS FOR ZINC

TABLE 14-30: NN, IDW AND OK MODEL STATISTICS MEAN BLOCK GRADE COMPARISONS FOR GOLD

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TABLE 14-31: NN, IDW AND OK MODEL STATISTICS MEAN BLOCK GRADE COMPARISONS FOR COPPER

TABLE 14-32: NN, IDW AND OK MODEL STATISTICS MEAN BLOCK GRADE COMPARISONS SILVER

14.6.4 Local Bias Checks

A local bias check was performed by plotting the average zinc, gold, copper and silver of the NN, IDW and OK models in swaths plots oriented along the model easting.

In reviewing the swath plots, only minor discrepancies were found between the different grade models. In areas where there is extrapolation beyond the drill holes, the swath plots indicate less agreement for all variables. The swath plots are shown below in Figure 14-14 to Figure 14-20.

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FIGURE 14-14: ZINC SWATH PLOT IN BASE METAL LENSES (BY EASTING)

Note: Line charts show the grades and histogram shows the tonnes. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-15: GOLD SWATH PLOT IN BASE METAL LENSES (BY EASTING)

Note: Line charts show the grades and histogram shows the tonnes. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-16: SILVER SWATH PLOT IN BASE METAL LENSES (BY EASTING)

Note: Line charts show the grades and histogram shows the tonnes. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-17: ZINC SWATH PLOT IN GOLD ZONES (BY EASTING)

Note: Line charts show the grades and histogram shows the tonnes. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-18: GOLD SWATH PLOT IN GOLD ZONES (BY EASTING)

Note: Line charts show the grades and histogram shows the tonnes. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-19: COPPER SWATH PLOT IN GOLD ZONES (BY EASTING)

Note: Line charts show the grades and histogram shows the tonnes. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-20: SILVER SWATH PLOT IN GOLD ZONES (BY EASTING)

Note: Line charts show the grades and histogram shows the tonnes. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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14.6.5 Block Model Quality Control

The closest distance of a composite (CDIST), the maximum distance of a composite (MDIST), the average distance of composites (ADIST), the number of composites (NCOMP) and the number of holes (NHOLE) used for the OK interpolation of zinc, gold, silver and copper were recorded in the block model.

The standard deviation of the kriging (KSTD) and the regression slope (RSLOP) were also recorded in the block model. Table 14-33, to Table 14-36 present the quality control parameters recorded in the block model from the OK interpolation.

TABLE 14-33: QUALITY CONTROL STATISTICS OF THE ZINC INTERPOLATION WITHIN THE BASE METAL LENSES

TABLE 14-34: QUALITY CONTROL STATISTICS OF THE GOLD INTERPOLATION WITHIN THE BASE METAL LENSES

Aside from lenses 31 and 40, the average standard deviation and the regression slope of the kriging are indicating a low variability of the zinc OK interpolation which is to be expected in a VMS deposit. As mentioned previously, it is believed that the overall drill spacing in lenses 31 and 40 is the source of the higher standard deviation and kriging regression slope.

TABLE 14-35: QUALITY CONTROL STATISTICS OF THE GOLD INTERPOLATION WITHIN THE GOLD ZONES

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TABLE 14-36: QUALITY CONTROL STATISTICS OF THE COPPER INTERPOLATION WITHIN THE GOLD ZONES

The average standard deviation and the regression slope of the kriging are indicating a relative high variability of the gold and copper OK interpolation. As mentioned previously, the interpretation of the gold mineralized envelopes could be reviewed to improve the stationarity.

14.6.6 Grade-Tonnage Statistics

Table 14-37 to Table 14-40 present the grade-tonnage statistics of zinc, gold and copper for each interpolation method at different cut-offs. The grade-tonnage curve for zinc, gold and copper are shown in Figure 14-21 to Figure 14-24 as a way to present the overall assessment of the resources.

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TABLE 14-37: ZINC GRADE -TONNAGE STATISTICS IN BASE METAL LENSES

TABLE 14-38: GOLD GRADE -TONNAGE STATISTICS IN BASE METAL LENSES

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TABLE 14-39: GOLD GRADE -TONNAGE STATISTICS IN GOLD ZONES

TABLE 14-40: COPPER GRADE -TONNAGE STATISTICS IN GOLD ZONES

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FIGURE 14-21: GRADE TONNAGE OF RESTRICTED ZINC IN THE BASE METAL LENSES

Note: Solid lines represent tonnes, dashed lines represent grades, green represents IDW model, red represents NN model and blue represents OK model

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FIGURE 14-22: GRADE TONNAGE OF CAPPED GOLD IN THE BASE METAL LENSES

Note: Solid lines represent tonnes, dashed lines represent grades, green represents IDW model, red represents NN model and blue represents OK model

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FIGURE 14-23: GRADE TONNAGE OF CAPPED GOLD IN THE GOLD ZONES

Note: Solid lines represent tonnes, dashed lines represent grades, green represents IDW model, red represents NN model and blue represents OK model

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FIGURE 14-24: GRADE TONNAGE OF RESTRICTED COPPER IN THE GOLD ZONES

Note: Solid lines represent tonnes, dashed lines represent grades, green represents IDW model, red represents NN model and blue represents OK model

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14.7 Classification of Mineral Resource in Base Metal Lenses

Mineral resources have been classified according to the 2014 CIM Definition Standards on Mineral Resources and Mineral Reserves (CIM definitions), as incorporated in NI 43-101. Resource blocks are classified as Measured, Indicated or Inferred, depending upon the confidence level of the resource based on experience at Lalor mine, with similar deposits and the spatial continuity of the mineralization.

The resource category classification relies on the following method:

Measured

• Distance to an underground development is generally less than 10 m

Indicated

• Distance to closest composite is less than or equal to 50 m

• Blocks interpolated from the last two interpolation pass (i.e. the most restrictive)

• Blocks estimated from at least two drill holes

Inferred

• The remainder of the interpolated blocks within the interpreted lenses are classified as Inferred Resources

A smoothing algorithm was applied to remove isolated blocks of measured within areas of mostly indicated category or isolated indicated blocks within areas of mostly measured category blocks and solids were created for all the different base metal lenses resource categories. Proportions of measured and indicated category blocks were not changed significantly by this process. Figure 14 25 presents a 3D view of the resource categories of lense 10.

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FIGURE 14-25: 3D VIEW (LOOKING SW) DISPLAYING THE RESOURCE CLASSIFICATION IN LENSE 10

Note: One block equals 5m by 5m by 5m. Measured = green blocks, Indicated = yellow blocks and Inferred = orange blocks.

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14.8 Classification of Mineral Resource in Gold Zones

The resource category classification relies on the relative difference between the kriged grade and the composites grades. The Resource Classification Index (RCI) uses the following formula¹:

Where C is a calibration factor based the distance of the composites, the number of composites, number of quadrants and number of drill holes using the following formula:

The RCI value corresponding to the 50^th (0.084) percentiles of the distribution of blocks with gold grade above 1 g/t contained within the interpreted gold zones was determined and used as a threshold for the indicated resource category in all the gold zones except for zone 27 (i.e. a copper-gold zone), where the RCI values corresponding to the 70^th (0.131) percentiles was used as a threshold for the indicated resource category. Table 14-41 presents the RCI deciles values along with their corresponding average distance, closest distance and maximum distance of the composites to the blocks centre.

TABLE 14-41: RCI VALUES WITH AVERAGE ADIST, CDIST AND MDIST

Under this classification system, in order for a block to be considered as measured, the blocks must be within approximately 10 m of an underground development. To be considered as indicated, a block must have a RCI value lower than 0.084 (or lower 0.131 in zone 27) and a CDIST of less than or equal to 25 m. All remaining blocks were classified as inferred resources with minimum criteria of one drill hole to interpolate the grades within one of the three interpolation passes.

^{____________________________________________
1} Arik, A. 2002,” Resource Classification Index”, MineSight in the Foreground.

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A smoothing algorithm was applied to remove isolated blocks of measured within areas of mostly indicated category or isolated indicated blocks within areas of mostly measured category blocks and solids were created for all the different gold zones resource categories. Proportions of measured and indicated category blocks were not changed significantly by this process. Figure 14 26 presents a 3D view of the resource categories of zone 25.

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FIGURE 14-26: 3D VIEW (LOOKING SW) DISPLAYING THE RESOURCE CLASSIFICATION IN ZONE 25

Note: One block equals 5m by 5m by 5m. Measured = green blocks, Indicated = yellow blocks and Inferred = orange blocks.

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14.9 Third Party Review

Hudbay requested that T. Maunula & Associates Consulting Inc., an independent consultant, perform a validation of gold zones and selected base metals lenses. The following minor issues were highlighted by the third party validation:

• Entry error noted and corrected in OK interpolation parameter.

• It was noted that some variograms ranges (example in lense 10 and 40) could be increased. The nugget effect was quite low for some of the variograms which seems inconsistent with the skewed data population and high CV, the next round of modelling should inspect this further.

• Swath plots demonstrated good correlation between composites and block models except in areas of lower data density or smaller resource volumes

• Based on the visual inspection, the block model grades appeared to honour the data well. The interpolated block model grades exhibit satisfactory consistency with the drill hole composites. However, it was noted that high coefficients of variation (CV) are present. This may cause a bias using linear estimation techniques. Compositing and capping reduced the CV to an acceptable level for most lenses.

• 0.24% of blocks were estimated with one drill hole and classified as Measured or Indicated Resource

The third party recommends the following:

• Review interpolation methodology to reduce number of passes. This could be evaluated as part of the change of support analysis. The interpolation parameters for minimum and maximum number of composites, use of octants, etc. and their impact could be evaluated against the Herco adjusted NN model.

• Assess the calculation of the density from the interpolated grades rather than interpolating density.

• Use reconciliation information from mined areas to validate the capping levels and density calculations.

Based on the review, the third party has concluded that the Lalor block model has been interpolated using industry accepted modeling techniques using MineSight desktop software. This included geologic input, appropriate block model cell sizes, grade capping, assay compositing, and reasonable interpolation parameters. The results have been verified by visual review and statistical comparisons between the estimated block grades and the composites used to interpolate. The OK model has been selected as the best representation of the grade distribution based on the current geological understanding and zone interpretation. The OK model has been validated with alternate estimation methods: NN and IDW. No biases have been identified in the model. Mineral resources were classified in accordance with the 2014 Canadian Institute of Mining, Metallurgy and Petroleum’s “CIM Definition Standards - For Mineral Resources and Mineral Reserves” incorporated by reference into National Instrument 43-101 “Standards of Disclosure for Mineral Projects.”

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14.10 Reasonable Prospects of Economic Extraction

The component of the mineralization within the block model that meets the requirements for reasonable prospects of economic extraction is based on the 2017 Lalor mine and Stall concentrator budgets. The following table presents the metal equivalency formulas used in the 2017 block model:

TABLE 14-42: LALOR METAL EQUIVALENCIES

Mineralization Type	Metal Equivalence	Formula
Base Metal Lenses	Zn Eq	Zn Eq = Zn% + (1.98 x Cu%) + (1.11 x Au g/t) + (0.01 x Ag g/t) – (0.01 x Pb%)
Gold Zones	Au Eq	Au Eq = Au g/t + (1.34 x Cu%) + (0.01 x Ag g/t)

Table 14-43 to Table 14-49 present the economic parameters and recoveries used to determine the metal equivalency formulas specified above. Metal equivalence considers the ratio of recovery, payability and value of metals after application of downstream processing costs.

TABLE 14-43: METAL PRICES – 2017 MINERAL RESOURCE ($US)

	Unit	Value
Gold	$US/oz	1,300
Silver	$US/oz	18.00
Copper	$US/lb	2.67
Zinc1	$US/lb	1.19
FX Rate	CAD / USD	1.25

¹Net zinc price includes premium and distribution costs based on
processing and refining at Hudbay's Flin Flon Zinc Plant

The cost and price inputs are considered approximation and were used to test the economic of the resource. The cost and price inputs may differ from the mineral reserve.

TABLE 14-44: METAL RECOVERIES

	Base Metal Concentrator	Gold Leach Concentrator
Gold to Copper Concentrate Silver to Copper Concentrate Copper to Copper Concentrate Copper Concentrate Grade	58.0% 55.0% 85.0% 21.0%
Gold to Dore Silver to Dore		92.0% 53.1%
Zinc to Zinc Concentrate Zinc Concentrate Grade	92.5% 51.0%

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TABLE 14-45: MANITOBA BUSINESS UNIT – GENERAL MANAGEMENT, ADMINISTRATION AND UNALLOCATED SERVICES COSTS

Administration Overhead	Unit	Value
Base Metal	$C/tonne mined	16.00
Gold Zone	per dollar of revenue	0.08

TABLE 14-46: COPPER AND ZINC CONCENTRATE TERMS

	Unit
Payables
Gold in Copper Concentrate	96% of content
Silver in Copper Concentrate	90% of content
Copper in Copper Concentrate	Minimum deduction 1 unit
Zinc in Zinc Concentrate	97.50%
Treatment
Copper Concentrate	$US 90.00/dmt
Zinc Concentrate (includes refining)	$US 320.00/dmt
Refining
Gold	$US 5.00/oz
Silver	$US 0.50/oz
Copper	$US 0.09/lb
Freight
Copper Concentrate (third party smelter)	$US 160.00/dmt
Zinc Concentrate	Freigth to Flin Flon included with milling

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TABLE 14-47: GOLD DORÉ TERMS

	Unit
Payable
Gold in Dore	98.5% of content
Silver in Dore	99% of content
Treatment
Dore	$US 0.40/oz
Refining
Gold	$US 5.00/oz
Silver	$US 0.50/oz
Freight
Dore	$US 0.50/oz

TABLE 14-48: MINING

	Unit	Value
Ore Removal	$C/tonne mined	20.00
General & Administration	$C/tonne mined	33.00

TABLE 14-49: MILLING

	Unit	Value
Base Metal (includes concentrate freight to Flin Flon)	$C/tonne milled	23.00
Gold Concentrator	$C/tonne milled	40.00

14.11 Mineral Resource Statement

Mineral resources for the Lalor mine were classified under the 2014 CIM Definition Standards for Mineral Resources and Mineral Reserves².

The mineral resources, classified as Measured, Indicated and Inferred, inclusive of mineral reserves are summarized in Table 14-50 and Table 14-51 as of September 30, 2016. The Qualified Person for the mineral resource estimate is Robert Carter, P. Eng., Lalor Mine Manager, Hudbay Manitoba Business Unit.

^{________________________________________
2} Ontario Securities Commission web site (http://www.osc.gov.on.ca/en/15019.htm)

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Mineral resources that are not mineral reserves do not have demonstrated economic viability. Due to the uncertainty thatmay be associated with Inferred mineral resources it cannot be assumed that all or any part of Inferred resources will be upgraded to an Indicated or Measured resource.

TABLE 14-50: BASE METAL MINERAL RESOURCE, INCLUSIVE OF MINERAL RESERVES BY
CATEGORY AND MINERALIZED ZONE WITH A CUT-OFF OF 4.1% ZN EQ, AS OF SEPTEMBER 30, 2016 ^{(1)(2)(3)(4)(5)(6)(7)(8)(9)}

Notes:

1.	Domains were modelled in 3D to separate mineralized zones from surrounding waste rock. The domains were based on core logging, grade, structural and geochemical data.
2.	Raw drill hole assays were composited to 1.25 metre lengths, honouring lithology boundaries.
3.	Capping of high gold and silver grades was considered necessary and was completed on assays prior to compositing.
4.	High yield restriction of base metal high grade and density was completed for each domain after compositing.
5.	Block grades for zinc, gold, silver, copper, lead, iron, arsenic and density were estimated from the composites using ordinary kriging interpolation into 5 m x 5 m x 5 m blocks coded by domain.
6.	Density values are from a multi elements regression formula based on 65,792 measurements.
7.	Blocks were classified as Measured, Indicated or Inferred in accordance with CIM Definition Standards 2014.
8.	Metal prices of $US 1.19/lb zinc, $US 1,300/oz gold, $US 2.67/lb copper, and $US 18.00/oz silver with a CAD / US foreign exchange of 1.25 were used to calculate a zinc equivalence (Zn Eq) cut-off of 4.1%, where Zn Eq = Zn% + (1.98 x Cu%) + (1.11 x Au g/t) + (0.01 x Ag g/t) – (0.01 x Pb%). The Zn Eq considers the ratio of milling recovery, payability and value of metals after application of downstream processing costs. The Zn Eq cut-off of 4.1% covers administration overhead, mining removal, milling and general and administration costs.

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9. Totals may not add up correctly due to rounding.

TABLE 14-51: GOLD MINERAL RESOURCE, INCLUSIVE OF MINERAL RESERVES, BY
CATEGORY AND MINERALIZED ZONE WITH A CUT-OFF OF 2.4 G/T AU EQ, AS OF SEPTEMBER 30, 2016 ^{(1)(2)(3)(4)(5)(6)(7)(8)(9)}

Notes:

1.	Domains were modelled in 3D to separate mineralized zones from surrounding waste rock. The domains were based on core logging, grade, structural and geochemical data.
2.	Raw drill hole assays were composited to 1.25 metre lengths, honouring lithology boundaries.
3.	Capping of high gold and silver grades was considered necessary and was completed on assays prior to compositing.
4.	High yield restriction of base metal high grade and density was completed for each domain after compositing.
5.	Block grades for zinc, gold, silver, copper, lead, iron, arsenic and density were estimated from the composites using ordinary kriging interpolation into 5 m x 5 m x 5 m blocks coded by domain.
6.	Density values are from a multi elements regression formula based on 65,792 measurements collected by Hudbay.
7.	Blocks were classified as Measured, Indicated or Inferred in accordance with CIM Definition Standards 2014.
8.	Metal prices of $US 1,300/oz gold, $US 2.67/lb copper and $US 18.00/oz silver with a CAD / US foreign exchange of 1.25 were used to calculate a gold equivalence (Au Eq) cut-off of 2.4 g/t Au Eq, where Au Eq = Au g/t + (1.34 x Cu %) + (0.01 x Ag g/t). The Au Eq considers the ratio of milling recovery, payability and value of metals after application of downstream processing costs. Au Eq cut-off of 2.4 g/t covers administration overhead, mining removal, milling and general and administration costs.
9.	Totals may not add up correctly due to rounding.

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14.12 Mine Reconciliation of Block Model

A mine reconciliation of the block model was carried out on the mined out areas. The process involved selecting all mined out blocks and comparing to the actual metal balance reported as ore received at the Stall concentrator. Mined out areas from the block model do not include dilution or pillars left behind after mining extraction. Table 14-52 list the mined out areas by lense from the block model and Table 14-53 is the ore reported by year at the Stall concentrator since Lalor commenced production in August 2012 until September 2016.

TABLE 14-52: MINED-OUT AREAS FROM THE BLOCK MODEL

Lense	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
10	1,560,246	9.63	1.94	0.71	19.39
11	55,355	14.46	0.14	0.30	17.22
20	336,600	8.80	2.19	0.82	30.71
21	99,594	0.65	7.70	0.65	30.85
23	24,210	1.50	5.37	0.61	39.89
24	25,849	0.89	6.44	0.42	23.24
25	56,481	0.64	8.47	0.40	31.93
30	9,098	4.25	1.82	0.26	35.67
31	64,845	4.73	0.95	0.20	15.14
32	183,404	7.98	5.51	1.60	54.22
Total	2,415,684	8.60	2.65	0.75	24.52
		Zn (tonnes)	Au (ounces)	Cu (tonnes)	Ag (ounces)
In-Situ Metal		207,641	205,689	18,196	1,904,304

TABLE 14-53: ORE RECEIVED BY YEAR AT THE STALL CONCENTRATOR

Year	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
2012	72,294	11.83	1.68	0.63	19.30
2013	400,589	9.44	1.20	0.84	19.41
2014	551,883	8.52	2.29	0.88	23.83
2015	934,278	8.18	2.53	0.71	21.39
2016 Sep YTD	814,207	6.88	2.30	0.64	21.62
Stockpile as of Sep 2016	11,114	5.38	2.14	0.41	21.42
Total	2,784,365	8.13	2.20	0.74	21.60
		Zn (tonnes)	Au (ounces)	Cu (tonnes)	Ag (ounces)
Metal		226,438	196,908	20,525	1,933,617
	Tonnes	Zn	Au	Cu	Ag
Variance to Block Model	115%	109%	96%	113%	102%

The block model compared very well to the ore reported at the Stall concentrator. The mine reconciliation concludes a 15% mining dilution and a metal variance reported at the Stall concentrator of 109% for zinc, 96% for gold, 113% for copper and 102% for silver. A mine reconciliation of 5 to 10% variance is well within industry standard. The precious metals reconciled very well, while there might be some conservatism of the zinc and copper grade estimates in the block model. The conservatism of the zinc and copper grade is believed to be linked to the high yield radius parameter. A previous model selected a 40 m high yield radius compared to the current 20 m radius used for this block model.

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14.13 Factors That May Affect the Mineral Resource Estimate

Areas of uncertainty that may materially impact the mineral resource estimate includes:

• Medium and long-term commodity price assumptions

• Operating cost assumptions

• Metal recovery assumptions used and changes to the metallurgical recovery assumptions as a result of new metallurgical information

• Changes to the tonnage and grade estimates may vary as a result of more drilling and new assay information

• Assumptions as to the ability to maintain mining claims and surface rights, access to the site, obtain environmental and other regulatory permits and obtain social license to operate

14.14 Conclusions

The mineral resource estimation in the base metal lenses is well-constrained by three-dimensional wireframes representing realistic volumes of mineralization. Exploratory data analysis conducted on assays and composites shows that the wireframes are suitable domains for mineral resource estimation. As for the gold mineralized zones, which are displaying high coefficient of variation, they could benefit from a modelling method with additional constraints.

As a result of validation steps conducted on the mineral resource block model the following was concluded:

• Visual inspection of block grade versus composited data shows a good reproduction of the data by the model.

• Checks for global bias in the grade estimates of the block model show differences within acceptable levels between the NN, IDW and the OK models. The variation are less than 7% for zinc and 4% for gold in the base metal lenses, while the variation of gold is less than 6% (aside from Zone 28) in the gold zones.

• Checks for local bias (swath plots) indicate good agreement between the NN, IDW and OK for all variables.

• A mine reconciliation of the mined out areas compared to the ore reported at the concentrator was very close on the precious metals and a slight conservatism of the zinc and copper grades might be evident.

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• Applying linear interpolation method to the gold zones method may not produce highly accurate results given the high coefficient of variation of gold.

The impact of grade capping was evaluated by estimating uncapped and capped grade models. Generally, the amounts of metal removed by capping in the models are consistent with the amounts calculated during the grade capping study on the assays.

Mineral resources are constrained and reported using economic and technical criteria such that the mineral resource has reasonable prospects of economic extraction.

The estimated mineral resources for the Lalor mine conform to the requirements of 2014 CIM Definition Standards – for Mineral Resources and Mineral Reserves and requirements in Form 43-101F1 of National Instrument (NI) 43-101, Standards of Disclosure for Mineral Projects.

14.15 Recommendations

The author recommends that the following points be further investigated in order to increase the geological knowledge and confidence.

• The host rocks are heavily altered, metamorphosed and deformed, making the protoliths and the relevant alteration assemblage often hard to identify. Therefore, full geochemistry analytical package should be used to assay the core. This would enable the utilization of geochemical proxies and would allow the identification of the host rocks and alterations, hence increasing the geological knowledge and confidence.

• Lithological, alteration and structural interpretations should be added to the current grade shell model to increase the geological knowledge and confidence.

• As suggested by the statistical analysis and the validation process, the gold zones could benefit from a modeling method that would use additional constraints instead of relying on the interpreted geological continuity.

• The gold raft zones should be the focus of exploration to increase the geological knowledge and confidence. For instance, the rafts of zones 24 and 25 have developments that could be use for mapping purposes.

• In the event that a revision of the wire framing method does not improve the statistics and stationarity of the gold zones, applying non-linear interpolation methods should be investigated in order to ensure that the grade interpolation produce accurate results in the gold zones.

• Review interpolation methodology to reduce number of passes. The interpolation parameters for minimum and maximum number of composites, use of octants, etc. And their impact could be evaluated against the Herco adjusted NN model.

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• Assess the calculation of the density from the interpolated grades rather than interpolating SG.

It is also recommended that Hudbay perform a check on grade smoothing using a global change-of-support correction. This should be performed to ensure that the grade smoothing is acceptable around the cut-off grades of interest.

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15 MINERAL RESERVE ESTIMATES

15.1 Summary

The Lalor mine mineral reserves as of January 1, 2017 are summarized in Table 15-1. The mine plan was prepared using measured and indicated mineral resources from the block model. Inferred resources were assumed as waste. The mineral reserves were estimated based on a Life of Mine (LOM) plan prepared; using Deswik mine design software that generated mining inventory based on stope geometry parameters and mine development sequences. Appropriate dilution and recovery factors were applied based on cut and fill and longhole open stoping mining methods with a combination of paste and unconsolidated waste backfill material. The Qualified Person for the mineral reserve estimate is Robert Carter, P. Eng., Lalor Mine Manager, Hudbay Manitoba Business Unit.

TABLE 15-1: SUMMARY OF MINERAL RESERVES AS OF JANUARY 1, 2017

Category	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
Proven	4,383,000	6.76	2.37	0.76	27.33
Probable	9,849,000	4.39	2.72	0.65	26.12
Proven + Probable	14,232,000	5.12	2.61	0.69	26.50

Notes:

1.	CIM definitions were followed for mineral reserve
2.	Mineral reserves are estimated at an NSR cut-off of $88/t for longhole open stoping mining method and $111/t for cut and fill mining method
3.	Metal prices of $US 1.07/lb zinc (includes premium), $US 1,260/oz gold, $US 3.00/lb copper, and $US 18.00/oz silver with a CAD / US foreign exchange of 1.10 were used to estimate mineral reserves.
4.	Bulk density of the resource is reported in the block model is from a multi elements regression formula based on 65,792 measurements. Stope geometry shapes include waste dilution based on a bulk density of 2.8t/m3.
5.	Totals may not add up correctly due to rounding.

The author considers that the mineral reserves as classified and reported comply with all disclosure in accordance with requirements and CIM definitions.

The author is not aware of any mining, metallurgical, infrastructure, permitting or other relevant factors that could materially affect the mineral reserve estimate.

15.2 Dilution and Recovery

There are two sources of dilution: internal (planned) dilution and external (unplanned) dilution. Dilution amounts are included in the conversion of mineral resource to mineral reserves. The shallow dipping nature of the deposit and stacking of lenses results in multiple lenses being grouped together for mining purposes in the stope optimizer routines of Deswik so that they can be extracted as a single mining unit, based on stope mining parameters by mining method as shown in Table 15-2.

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TABLE 15-2: STOPING PARAMETERS BY MINING METHOD

Stope Shape Parameters	Unit	Longhole	Cut and Fill
Length
Minimum	Metres	23	150
Maximum		10	20
Width
Minimum	Metres	3.5	5
Maximum		50	50
Waste Pillar Width	Metres	5	-
Stope Height
Minimum	Metres	10	5
Maximum		20	5
Stope Dip	Degrees
Minimum		35	75
Hanging Wall Dip
Minimum	Degrees	20	70
Maximum		90	90
Footwall Dip
Minimum	Degrees	50	70
Maximum		90	90
Dilution
Hanging Wall Fixed	Metres	0.5	0.5
Footwall Fixed		0.5	0.5

Parameters most sensitive to Lalor mine are the minimum and maximum dip angles, which affects the dilution and recovery amounts of the optimized mining shape. The stope optimizer in Deswik generated an economic shape that honoured the geometric parameters.

The space between the lenses is treated as internal dilution and external dilution is set at a fixed distance of 0.5 m into the footwall and hanging wall after the stope geometry shape is finalized. Internal dilution and external dilution are included as part of the optimized mining shape. Dilution, set at zero grade and a bulk density of 2.8 t/m³, is based on the full mining shape with internal and external dilution. Average dilution factors by lense grouping are shown in Table 15-3.

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TABLE 15-3: AVERAGE DILUTION FACTORS BY LENSE GROUPING

Lense Grouping	Diluted Tonnes	Average Dilution (%)
10, 11	4,278,000	19.8
20, 21, 24 ,25	6,277,000	19.7
23, 32	2,222,000	15.6
26, 28, 30, 40	2,639,000	14.8
27	626,000	29.4
31	618,000	11.3
Development	896,000	27.1
Total	17,556,000	18.9

Mining recovery is defined as the ratio of mineral resource tonnes delivered to the concentrator to the in-situ mineral resource tonnes. Mining recovery factors used for each mining method and by lense are shown in Table 15-4. Average recovery factors by lense grouping are shown in Table 15-5. Some of the mineral resources are not recovered due to:

• Mining design. This includes rib, post and sill pillars that are not recovered to maintain rock stability.

• Inefficiencies in mining. This includes small blocks of ore along ore/waste contacts and underbreak.

• Inefficiencies in mucking. This includes losses of broken rock in longhole stopes mucked by remote control LHD and broken rock that is mixed with waste backfill and is not mucked.

TABLE 15-4: RECOVERY FACTORS BY MINING METHOD AND LENSE

Mining Method	Lense	Criteria	Backfill	Recovery (%)
PPCF and CF	All	Standard 7 x 7m pillars	Waste	75
		Reduced pillar sizes	Paste	80
		>$250/t NSR ore	Paste	85
Drift and Fill	27	>$300/t NSR ore	Paste	90
Longhole	10 and 30	5m rib pillars	Waste	70
	10 and 30	No rib pillars	Paste	85
	20, 27, 32 and 10 above 755m level	4m rib pillars	Waste	72
	20, 27, 32 and 10 above 755m level	No rib pillars	Paste	90

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TABLE 15-5: AVERAGE RECOVERY FACTORS BY LENSE GROUPING

Lense Grouping	Recovered Tonnes	Average Recovery (%)
10, 11	3,310,000	77.4
20, 21, 24 ,25	5,196,000	82.8
23, 32	1,941,000	87.3
26, 28, 30, 40	1,908,000	72.3
27	548,000	87.6
31	433,000	70.0
Development	896,000	100.0
Total	14,232,000	81.1

15.3 Conversion of Mineral Resources to Mineral Reserves

The mine plan was prepared using measured and indicated mineral resources from the block model. Inferred resources were assumed as waste. Appropriate dilution and recovery factors were applied based on cut and fill or longhole open stoping mining methods and backfill or unconsolidated waste backfill. Upon determination of diluted and recovered mineral resources, stope economic criteria were applied.

Diluted and recovered mineral resources exceeding a Net Smelter Return (NSR) cut-off of $88/t for longhole open stoping and $111/t for cut and fill mining method are included in the mineral reserves. NSR’s are based on metal grades from the stope optimizer and block model, long-term metal prices, concentrator recoveries, smelter treatment, refining and payabilities and a Hudbay Manitoba Business Unit administration cost.

15.3.1 Metal Prices

Metal prices of $US 1.07/lb zinc (includes premium), $US 1,260/oz gold, $US 3.00/lb copper, and $US 18.00/oz silver with an CAD / US foreign exchange of 1.10 was used to estimate mineral reserves.

15.3.1 Metallurgy

The orebody is polymetallic with economically significant metals being zinc, gold, copper and silver. There are two different ore types, both of which are assumed to be treated using conventional flotation at the Hudbay Stall concentrator:

• Base metals ores. Near solid to solid sulphide ores, with dominant pyrite and sphalerite with minor blebs and stringers of chalcopyrite and pyrrhotite.

• Gold rich ores. Silicified gold and silver enriched ores with stringers to disseminated chalcopyrite and sphalerite mineralization.

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Metallurgical performance at Stall concentrator indicates that the base metal and gold rich ores can be blended and metallurgical assumptions are shown in Table 15-6. Two concentrates will be produced, a zinc concentrate that will be shipped to the Hudbay Flin Flon metallurgical complex for production of refined zinc, and a gold enriched copper concentrate that will be shipped to third party smelters.

TABLE 15-6: METALLURGICAL ASSUMPTIONS

	Base Metal Concentrator
Gold to Copper Concentrate Silver to Copper Concentrate Copper to Copper Concentrate Copper Concentrate Grade	57.6% 50.9% 82.8% 21.0%
Zinc to Zinc Concentrate Zinc Concentrate Grade	94.2% 51.0%

15.3.1 Payability, Treatment Charges, Refining Charges

The basis of the NSR cut-off is further determined by applying the copper and zinc concentrate terms as shown in Table 15-7.

TABLE 15-7: COPPER AND ZINC CONCENTRATE TERMS

	Unit
Treatment
Copper Concentrate	$US 85.00/dmt
Zinc Concentrate (includes refining)	$C 404.97/dmt
Freight
Copper Concentrate (third party smelter)	$C 215.00/wmt
Zinc Concentrate (Hudbay smelter)	Included with milling
Payables
Gold in Copper Concentrate	96% of content
Silver in Copper Concentrate	90% of content
Copper in Copper Concentrate	Minimum deduction 1 unit
Zinc in Zinc Concentrate	97.5%
Refining
Gold	$US 5.00/oz
Silver	$US 0.50/oz
Copper	$US 0.085/lb
Zinc Marketing
Distribution	$US 0.06/lb
Premium	$US 0.07/lb
Administration
Copper Concentrate	$C 106.92/dmt
Zinc Concentrate	$C 139.95/dmt

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The NSR cut-offs of $88/t for longhole open stoping and $111/t for cut and fill mining method are sufficient to cover all expenses related to mining and milling the mineral reserves as shown in Table 15-1.

The mineral reserves are supported by a LOM plan prepared using Deswik mine design software that generated mining inventory based on stope geometry parameters, and mine development and production sequences. This is further detailed in the LOM production plan as described in Section 16 of this report.

The author considers the dilution and loss factors to be appropriate for the mining methods selected and the stope geometry shape and applicable stoping parameters of the Lalor ore body and the methodology of converting mineral resources to mineral reserves.

The conversion of resources to reserves is based on the LOM plan and NSR cut-offs that primarily focussed on capturing base metal resources for processing at the Stall concentrator. The secondary focus was to capture gold zone resources when in contact with or close proximity to base metal resources. In areas where a large separation existed between base metal and gold lenses, mining blocks were evaluated for economic stope mining shapes. When a non-economic shape was generated in a first pass, a second pass was evaluated for only base metal lenses and if an economic shape was generated the gold zone portion was removed. However, due to this larger separation, majority of these isolated gold lenses could have been evaluated independently of the base metal lenses and could potentially provide feed to a gold processing facility. Below approximately the 950 m level no attempt was made to generate an economic stope mining shape for gold zones 25 and 26 as the separation distance became too large. The author’s opinion is that these resources are potentially better suited for a gold processing facility and should be re-evaluated when Hudbay has a better understanding of their New Britannia gold mill and Birch Tailings Impoundment Area in Snow Lake.

Of the current 14,232,000 tonnes of mineral reserves, approximately 80% is converted from base metal resources and approximately 20% is converted from gold zone resources. Of the total reserves approximately 3.8% is represented by the indicated resources of copper-gold zone 27, which is inclusive of the 20% from the gold zone, noted above. Although the indicated resource of copper-gold zone 27 as shown in Table 14-51, is converted to reserves and is planned to be processed at the Stall base metal concentrator, it has the potential to be milled at the New Britannia gold mill if the refurbishment plan and the installation of a copper pre-float facility proves to be economically viable.

The author recommends that base metal indicated resources, exclusive of reserves, as shown in Table 15-8 remain as indicated resources until such a time that detailed mine planning is completed. Furthermore, the author recommends that gold zone indicated resources, exclusive of reserves, as shown in Table 15-9 still have reasonable prospects of economic extraction at either Stall base metal concentrator or a gold processing facility.

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TABLE 15-8 BASE METAL INDICATED RESOURCE, EXCLUSIVE OF RESERVES,
WITH A CUT-OFF OF 4.1% ZN EQ, AS OF SEPTEMBER 30, 2016 ⁽¹⁾

Category	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
Indicated	2,100,000	5.34	1.69	0.49	28.10

Notes:

1. Refer to the Notes for Table 14-50 of this Technical Report for more information

TABLE 15-9 GOLD INDICATED RESOURCE, EXCLUSIVE OF RESERVES, WITH A
CUT-OFF OF 2.4 G/T AU EQ, AS OF SEPTEMBER 30, 2016 ⁽¹⁾

Category	Tonnes	Zn (%)	Au (g/t)	Cu (%)	Ag (g/t)
Indicated	1,750,000	0.40	5.18	0.34	30.61

Notes:

1. Refer to the Notes for Table 14-51 of this Technical Report for more information

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16 MINING METHODS

16.1 Introduction

The mining method process includes underground lateral advance (development rounds), production mining, backfilling and transporting ore to surface. Geotechnical information, orebody geometry interpreted from diamond drill core and recent experience mining within the deposit were the major considerations for selection of mining methods.

A geotechnical report, along with stoping recommendations was prepared in 2010. The ore zones, hanging wall and footwall rock is of good quality allowing the use of mechanized drilling and blasting techniques. The orebody dips at an average of -30°, with localized dips as flat as -10° and as steep as -55°. Mining methods that are currently in use or planned in the immediate future include: mechanized cut and fill, post pillar cut and fill, drift and fill and longhole open stoping (transverse and longitudinal retreat).

Ore is mucked using Load Haul and Dump (LHD) loaders which are operated remotely in inaccessible areas. The ore is then loaded into underground haul trucks or ore passes and transported to the ore handling system at the production shaft for hoisting to surface.

A paste backfill plant will be constructed on site, planned for the first quarter of 2018. Paste backfill will be used in high grade areas to increase recovery and accelerate the mining cycle. Lower grade areas will be filled with waste rock from waste development. No waste is planned to be hoisted.

Ore delivered to the production shaft is sized to less than 0.55 m at one of the two rock breaker/grizzly arrangements and hoisted from the mine by two 16 tonne capacity bottom dump skips in balance. Ore is truck hauled to a primary crusher at the Chisel North mine site, crushed to less than 0.15 m, and then trucked to the Stall concentrator for further processing.

16.2 Lateral Development

Lateral advance is made in 4 m long segments (rounds), with typical dimensions of 6 m wide by 5 m high. Lateral drilling is completed with two boom electric hydraulic jumbo drills, each round requires approximately 80 holes. Rounds mined in low sulphide areas or waste is blasted using ANFO, while rounds mined in high sulphide areas (ore or waste) are blasted using an emulsion with a sulphide blast inhibitor. Ore and waste are mucked using LHDs. Following mucking, standard ground support, consisting of resin grouted rebar and welded wire mesh to within 1.8 m of the sill, is installed. Mine services, including compressed air, process water and discharge water pipes, paste backfill pipeline, power cables, leaky feeder communications antenna and ventilation duct are installed in main levels and stope entrances.

Generally, main levels are developed parallel to and in the footwall of the ore zones. To optimize development, in some areas of the mine, main levels are located to provide access to multiple ore zones. As levels are developed, stope entrance crosscuts are stubbed off and used as temporary remucks. Main levels are connected by a haul ramp to allow mechanized equipment to travel from level to level.

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Stope access crosscuts to cut and fill stopes are driven at -15% to allow multiple cuts from a single crosscut. Cut and fill stope entrances are located approximately every 150 m along strike. For longhole stopes, transverse mining areas have drawpoint crosscuts developed in waste from the footwall and longitudinal mining areas have primary access points and then sill along the orebody with stopes being retreated to the access point.

16.3 Vertical Development

Main ventilation raises and ore pass raises are developed by a mining contractor using a raisebore and/or Alimak climber and hand held drills. Ground support and ladder ways, if required, are installed to Hudbay standards.

Longhole slot raises, transfer ore passes and auxiliary ventilation raises are limited to approximately 35 m long and are developed using longhole conventional drop raise techniques.

Drain, backfill and electrical cable holes are drilled using longhole or raisebore drills, and reamed to designed diameter.

16.4 Stope Mining

Two main mining methods are used at Lalor mine, cut and fill and longhole open stoping. Cut and fill methods include: mechanized cut and fill, post pillar cut and fill and drift and fill. Longhole open stoping methods include: transverse, longitudinal retreat and uppers retreat. Each mining area is evaluated to determine the most economic stoping method. In general where the dip exceeds 35° and the orebody is of sufficient thickness, longhole open stoping is used and lateral based cut and fill mining methods are used in flatter areas.

Approximately 65% of the mineral reserves are mined using the longhole open stoping methods and 35% are mined with cut and fill methods.

16.4.1 Cut and Fill Mining

Where the ore is flatter than 35°, single pass overhand mechanized cut and fill mining is generally the mining method chosen. The ore is accessed from a footwall drift by a crosscut developed at approximately -15%, a typical cross section is shown in Figure 16-1. Ore is mined 5 m high up to the hanging wall and footwall contacts. When ore mining is complete, ore remaining on the sill is mucked, pipe and ventilation duct is stripped, backfill is placed and the entrance crosscut is back slashed to provide access to the next cut.

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FIGURE 16-1: TYPICAL CUT AND FILL MINING CROSS SECTION

16.4.1.1 Mechanized Cut and Fill Mining

Where mining is planned to be less than 14 m wide and economics do not warrant consolidated fill, single pass overhand mechanized cut and fill mining is generally the cut and fill mining method chosen. Unconsolidated backfill is placed tight to the back and the entrance crosscut is then back slashed to provide access to the next cut.

Ground control used in mechanized cut and fill mining is 3.6 m resin rebar in the back for sections <10.8 m wide, 2.2 m resin rebar in the walls to within 1.8 m of the sill and welded wire mesh. In areas of excessive width, cement grouted cablebolts are installed to provide additional support and the area is mined in two passes or slashed and mucked/filled remotely. See Figure 16-2 for a typical cross section.

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FIGURE 16-2: TYPICAL MECHANIZED CUT AND FILL MINING CROSS SECTION

16.4.1.2 Post Pillar Cut and Fill Mining

Where mining is planned to be greater than 14 m wide and economics do not warrant consolidated fill, overhand post pillar cut and fill mining is generally the mining method chosen. Post pillars provide ground support to allow selective mining of wide stopes. Backfill is placed to within 1.8 m of the back and the entrance crosscut will be back slashed to provide access to the next cut.

Design for drifts and crosscuts in post pillar stopes is 7 m wide x 5 m high, with 7 m x 7 m post pillars.

Ground control used in post pillar cut and fill mining is 2.2 m resin rebar in the back outside of intersections, 3.6 m resin rebar in the back in intersections, 2.2 m resin rebar in the walls and pillars to within 1.8 m of the sill and welded wire mesh.

Post pillar mining is sequenced to retreat mining towards the access to reduce risk of encountering poor ground conditions. Typical mining sequences are shown for two different mining areas in Figure 16-3 as a section view and Figure 16-4 as a plan view.

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FIGURE 16-3: TYPICAL POST PILLAR CUT AND FILL MINING CROSS SECTION

FIGURE 16-4: TYPICAL POST PILLAR CUT AND FILL MINING PLAN VIEW

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16.4.1.3 Drift and Fill Mining

Where cut and fill mining is planned and economics do warrant consolidated fill, drift and fill mining is generally the mining method chosen. Filling the initial ore drifts with consolidated fill allows successive mining of ore drifts adjacent to the filled areas without leaving pillars allowing complete recovery of the resource. Figure 16-5 shows a two lift sequence of drift and fill mining.

Design for drifts in drift and fill stopes is planned 7 m wide x 5 m high.

Ground control used in drift and fill mining is 2.2 m resin rebar in the back and walls to within 1.8m of the sill and welded wire mesh.

FIGURE 16-5: TYPICAL DRIFT AND FILL MINING CROSS SECTION

16.4.2 Longhole Open Stope Mining

Where the footwall of the ore is steeper than 35°, longhole open stope mining is generally the mining method chosen. The advantages of longhole mining over cut and fill methods include:

• Longhole open stoping is a non-entry method. This reduces our employee’s exposure to risk on a per tonne basis.

• Longhole open stoping is a bulk mining method. The amount of ground support and labour per tonne is reduced lowering the overall cost per tonne compared to cut and fill mining methods.

Longhole stoping areas are evaluated to determine the optimum sill to sill vertical interval. In the shallower dipping areas of the ore this interval is typically 15 m, and in steeper dipping areas the interval is planned up to 25 m. Optimizing the sublevel interval allows minimal dilution and optimal recovery. Stope widths are also evaluated by area typically being less than 25 m and more than 15 m wide. Hole diameter currently used is 7.6 cm, but may be adjusted based on ore fragmentation and cost.

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Depending on economics and availability, consolidated or unconsolidated fill will be used. When using unconsolidated fill a nominal 5 m pillar is established between the stopes to contain the fill.

16.4.2.1 Transverse Longhole Open Stope Mining

The ore is undercut at the top and bottom of the block from cross-cuts off the footwall drift, providing access for drilling and mucking, as shown in Figure 16-6. Drilling is down the hole with a top hammer longhole drill. Stope sequence is reviewed for each area and typically progresses center out, bottom up. In future areas a primary-secondary sequence may be feasible.

FIGURE 16-6: TYPICAL ISOMETRIC VIEW – TRANSVERSE LONGHOLE OPEN STOPING

16.4.2.2 Longitudinal Retreat Longhole Open Stope Mining

The ore is undercut at the top and bottom of the block with sill drifts off the main access drift, providing access for drilling and mucking. Drilling is down the hole with a top hammer longhole drill. Stope sequence is reviewed for each area typically bottom up and retreating along the length of the sill to an access point at the end of the lense or from both ends of the sill toward a central access point, see Figure 16-7.

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FIGURE 16-7: TYPICAL LONG SECTION – LONGITUDINAL RETREAT LONGHOLE OPEN STOPING

16.4.2.3 Uppers Retreat Longhole Open Stope Mining

The ore is undercut at the bottom of the block with sill drifts off the main access drift, providing access for drilling and mucking. Drilling is up the hole with a top hammer longhole drill. Stope sequence is reviewed for each area typically top down and retreating along the length of the sill to an access point at the end of the lens or from both ends of the sill toward a central access point. This method is generally used to recover sill pillars and requires consolidated fill, see Figure 16-8.

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FIGURE 16-8: TYPICAL LONG SECTION – UPPERS RETREAT LONGHOLE OPEN STOPING

16.5 Backfill

All stopes at Lalor mine are backfilled to maintain long term stability and to provide a floor to work from for subsequent mining. Backfill is either:

a)	Unconsolidated waste rock backfill (URF)
b)	Consolidated backfill

a.	Cemented waste rock backfill (CRF)
b.	Paste backfill

URF is used in stopes where pillar or wall confinement is not required and the value of the adjacent pillars does not warrant the added expenditure of consolidated backfill.

Consolidated backfill currently consists of cemented waste rock backfill and is planned to be primarily as paste backfill after commissioning of the paste plant in the first quarter of 2018. Where economically feasible consolidated backfill is used by adding cement to waste rock using a spray bar and placing it in stopes with LHDs or when paste is available using the underground distribution system to transport paste (via gravity) directly to the stope. Consolidated backfill is required to maintain long term stability and allow future recovery of sill pillars.

The majority of consolidated backfill will be paste. Paste backfill is an engineered product comprised of mill tailings and a binder (3-5% cement by weight) mixed with water to provide a thickened paste that is delivered by borehole and pipes to stopes. Hudbay has experience with the design and operation of a paste backfill system, currently in use at the Flin Flon concentrator and 777 Mine. Paste backfill has advantages over unconsolidated fill such as slurried mill tailings or loose waste rock as follows:

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a)	Confines pillars in post pillar cut and fill stopes to increase pillar stability.
b)	Flows to the hanging wall to seal off previous cut and fill cuts, improving ventilation control and limiting the potential for hanging wall failure. By comparison, unconsolidated waste rock backfill typically rills to approximately 50°.
c)	Cures to a solid product. This allows mining up to the paste backfill walls in adjacent longhole stopes, and creates a good mucking floor in all stopes. This also eliminates the possibility of the build up of hydraulic head in stopes and potential flows of re-liquefied unconsolidated tails.

16.6 Ore Handling

Ore is mucked by LHD, loaded into underground haul trucks and hauled to one of the two ore passes that feed the shaft. Ore is dumped onto a grizzly at 910 m level for sizing to less than 0.55 m by a rockbreaker and grizzly. A 40 m raise and bin below the grizzly provides approximately 1,200 tonnes of coarse ore storage. A chute at the bottom of the raise at 955 m level feeds ore to a conveyor that loads a measuring flask with approximately 14 tonnes of ore. Ore is then skipped to surface by two 16 tonne capacity Bottom Dump skips in balance. The ore enters the headframe chute from the skips and is deposited into the surface ore bin or to the exterior concrete bunker via gravity. From the surface bin or bunker, ore is truck hauled to a primary crusher at the Chisel North mine site, crushed, and then trucked to the Stall concentrator to process. Opportunities to increase ore handling capacity and installation of additional shaft ore passes are currently being reviewed. However, based on an internal review the current ore handling system has the capacity to move 4,500 tpd.

16.7 Surface Infrastructure

The Lalor mine surface infrastructure includes the following:

a)	3.5 km mine site access corridor, which includes the mine haul road, 25 kV overhead power line and covered/heat traced process water and mine discharge water pipelines.
b)	Onsite services distribution. This includes pole and buried electrical services and switchgear and buried freshwater and mine discharge water pipelines.
c)	Hoist house, including: hoist foundations, mine service hoist, and production hoist.
d)	Production shaft collar and foundations, including ventilation plenum.
e)	Ore bunker, ore bin and head frame.
f)	Warehouse/shop complex.

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g)	Propane Storage.
h)	Haulage truck route.
i)	Mine water handling systems, including onsite pump house, mine discharge water tanks and pumps and freshwater tanks and pumps. Offsite booster pump station for freshwater and mine discharge water.
j)	Offsite permanent substation with two (2) 115-25 kV 24 MVA transformers.
k)	341 person change house complex that also houses staff offices.
l)	Permanent ventilation installations, including offsite exhaust fan and onsite mine downcast air heater and fans.

Additional information regarding mine infrastructure is included in Section 18. Refer to Figure 16-9 for the Lalor mine site general arrangement drawing.

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FIGURE 16-9: SITE GENERAL ARRANGEMENT

16.8 Geotechnical Design

Initial geotechnical design was completed by Stantec Engineering in 2009 and 2010. The orebody is flat lying beginning at approximately 600m from surface and extending to a depth of approximately 1,100 m. It trends between 320° to 340° azimuth and dips between 30° and 45° to the north with a lateral extent of about 900 m north to south and 700 m east to west. A number of distinct stacked mineralized zones have been interpreted. The geotechnical design was based on a preliminary lithology of the Lalor mine orebody as shown in Table 16-1.

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TABLE 16-1: LATERAL JUMBO DEVELOPMENT

Domain/Zone	Rock/Mineralization Type
Hanging wall Rock	Basaltic wacke, crystal wacke/fragmental polymodal fragmenta. Cordierite + anthophyllite, mafic tuffs
Zone 10	Solid to near solid sulphides and minor disseminated sulphides
Footwall to Zone 10	Sericite + kyanite + pyrite schist, staurolite + garnet + biotite gneiss
Hangingwall Zone 20	Altered dacite, amphibolitized mafic volcanic, calc. Silicate gneiss, diorite
Zone 20	Near to solid sulphides
Footwall Zone 20	Chlorite schist, altered dacite, chlorite sericitic talc schist, quartz-biotite-amphibolite gneiss
Hanging wall Zone 30	Mineralized chlorite schist, quartz chlorite schist
Zone 30	Disseminated to near/solid sulphides
Footwall Rock	Rhyolite fragmental, quartz biotite schist and chlorite schist

The comments below refer to both short term and long term requirements for Lalor mine geotechnical data acquisition, including training requirements.

16.8.1 Geotechnical Requirements

Geotechnical Logging

a)	To ensure that rock mass quality is logged and assessed to industry standard, Lalor has established diamond drill hole (DH) core logging by trained technical personnel.
b)	Lalor core logging format includes geology descriptions and geotechnical data collection (SG, UCS, E, ν, RQD, Q’, Jn, Jr, Ja) and is located in the same summary sheet for ease of data analysis and geotechnical modeling.

Geology and Structures

a)	Geotechnical data from all the know zones is continuously collected following standard procedures. This is done by the Lalor geology department and data is shared with ground control.
b)	All geological core logging is examined to identify and characterize any major structural features, including major fault and/or shear zones that may intersect zones of the mineralization.

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c)	Representative samples from various geotechnical domains, including hanging wall, various ore zones and footwall, are collected to test rock properties, including Unconfined Compressive Strength (UCS), Young Modulus and Poisson Ratio.
d)	A geotechnical database has been established from drill core logging data and geotechnical core sample laboratory tests.
e)	Footwall lithology is interpreted to identify the location and extent of the weak zones where future internal ramp and other major infrastructure may be proposed.
f)	The interpretation of the various lithological units in and around various zones of the orebody, i.e. the major geotechnical domains, and the major structural interpretation, i.e. major faults and shear zones, is generated/mapped in geology and mine planning applications for mine design and ground control use.

Rock Mass Classification Plots and Rock Properties Determination

a)	Rock mass classification data (RQD) is generated in 3D space, similar to the ore grade distribution from all diamond drill holes.
b)	Adequate core samples are collected for all the ore zones to ensure we have sufficient geotechnical data from hanging wall, orebody and footwall to further test rock properties.

16.8.2 Rock Mechanics Numerical Modelling

Numerical modeling is being conducted to help understand mining ground conditions and assist ground control decision making using available numerical modeling software on site (e.g. Examine 2D, ExamineTab). With mining progressing and stress interactions becoming more complicated, 3D numerical modeling may be considered.

16.8.3 Rock Mechanics Instrumentation

Ground monitoring instruments are currently being installed for ground movement monitoring. More advanced ground control instrumentation will be considered in the future.

16.9 Support Systems

Except when using cut and fill mining methods, all other drifts have arched backs for optimized shape and safety.

Ground support is broken down into primary support and secondary support.

Primary support refers to reinforcement of the rockmass immediately following excavation (first pass) to ensure safe working conditions before taking the next round. Primary support is typically undertaken with resin grouted rebar and #6 gauge galvanized welded wire mesh.

Secondary support is additional support applied after the installation of primary support to provide further support in large spans, long term infrastructure excavations and structurally controlled areas where wedge failures may be a concern. Secondary support is installed at a later stage (second pass) and typically is a batch process. Examples of secondary support are rebar, cable bolts, strandlok bolts, inflatable rock anchors, split sets and shotcrete.

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Lalor mine ground support standards are summarized in the following sections.

16.9.1 Primary Support

Standard drift, <7 m wide: 2.2 m long #6 (for jackleg/stopper installations) or #7 (for bolter installations) resin rebar on a 1.2 m x 1.2 m square pattern. Rebar and screen are extended down the walls to within 1.8 m of the sill. Screen overlaps at least 2 squares during installation. FS-33 or FS-39 Friction Stabilizers may be used to pin screen to back and walls but not as primary ground support. Generally, this can be applied to a drift span up to 7.0 m if no major geological structures are encountered. Special design is needed for drift spans larger than 7.0 m or when major geological structure is present.

Intersections, 7 m to 10.8 m wide: ground support uses the same support as for standard drifts except in the back where 3.6 m long #7 resin rebar on a 1.2 m x 1.2 m square pattern is installed.

16.9.2 Secondary Support

Secondary ground support is installed when excavation spans are larger than 10.8 m, major unfavorable ground conditions or rock structures are present, and/or after a site ground condition evaluation indicates it is required. Secondary ground support uses heavy duty longer bolts, such as cement grouted cable bolts. Typically, single cable bolts on a 1.8 m x 1.8 m pattern for long term excavations, or high strength inflatable rock anchors for temporary or short term excavations. The minimum bolt length should be equal to one-third of the final drift span.

16.9.3 Developed in Two Passes – Mechanized Cut and Fill or Infrastructure Headings

Headings wider than 7.0 m are developed using two methods, single pass and double pass. Ground support will depend upon the development method used to excavate the opening.

First pass development ground support is as described in Sections 16.9.1 and 16.9.2 above.

On completion of the first pass and before starting the second pass, secondary ground support is installed. The type of secondary ground support can be longer resin rebar, cable bolts or shotcrete. It should be noted that cablebolts and shotcrete must be allowed to cure for at least 24 hours before any blasting within 30 m. The second pass is supported similarly to the first pass or the area is designated for remote operations only.

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16.9.4 Sill Pillar Support

Sill pillars between mining blocks are occasionally formed. Depending on ore geometry, strength of the mineralization and depth of mining, poor ground conditions are anticipated in the last three cuts (15 m) or last longhole stope of each block, which will then be underneath excavated stopes. Drifts driven under excavated stopes require cablebolting when the sill pillar thickness is less than twice the width of the drift. The minimum lengths of cablebolts are to be equal to the width of the drift. Some shotcreting may also be required if conditions warrant.

All stopes will be backfilled. The fill type and mechanism of placement is important (refer to the following):

a)	Where paste fill is chosen, filling tight will require the use of sacrificial pipes (pours in long and small drifts) with burst disks, or the use of blasting cord to cut the pipe to stage pour locations. Tight fill is defined as 70% fill contact with the back.
b)	Where CRF is chosen, tight fill may be placed. However, the composition of the rock fill must be tightly controlled, as in many cases the mix can be too sloppy to tight fill with a steep rill.
c)	Where URF is chosen, fill may be placed and pushed to within 0.5 m of the back or if the mining plan calls for mining directly above the area to be filled a larger void may be left.

As additional geotechnical data becomes available ground support designs are reviewed. In addition, calibrated two or three dimensional numerical modeling is undertaken to identify problematic areas. Control measures, including enhanced ground support, are installed to reduce the risk of future ground instability if needed.

16.10	Underground Development
16.10.1	Lateral Development

Drifts and ramps will be developed by dedicated mine development crews using the following equipment:

a)	Drilling – Atlas Copco M2D Jumbos equipped with 4.88 m feeds and AC1838HD rockdrills. Development advance is nominally 4.0 m per round.
b)	Bolting – 3 methods are used: 1) Maclean Bolters (MEM-928) equipped with a scissor deck and bolting boom, 2) Atlas Copco Boltec MC equipped with remote arm bolting boom and 3) Maclean Scissor Decks (MEM 977) as a platform for jacklegs and stopers. Bolters have rod adding systems to allow cablebolt and testhole drilling. The units are equipped with AC1638HD, AC50 Monobear or AC1435 rockdrills and screen handling features.
c)	Mucking – LHDs, primarily Atlas Copco ST-14 and ST-1530.
d)	Trucking – Atlas Copco haul trucks primarily MT-42 and MT-6020.

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e)	Development Loading – Rounds will be charged with explosives using a Maclean Explosives Loader (MEM-AC3) equipped to load both ANFO (~1360 kg capacity) and emulsion (~450 kg capacity).
f)	Production Loading – Holes will be charged with emulsion (~1,800 kg capacity) using a Marcotte (M40) toe loader.

Lateral development required by year to mine the LOM reserves is shown in Table 16-2.

TABLE 16-2: LATERAL JUMBO DEVELOPMENT

Year	Sustaining Capital (m)	Mine Operating Waste (m)	Mine Operating Ore (m)	Total (m)
2017	2,918	2,471	4,526	9,915
2018	2,925	1,552	5,933	10,410
2019	1,785	2,120	6,506	10,411
2020	1,837	1,888	6,271	9,996
2021	1,369	1,830	6,672	9,871
2022	1,000	1,815	5,629	8,444
2023	77	2,303	4,788	7,168
2024		2,896	3,925	6,821
2025		1,571	3,816	5,387
2026		1,394	2,175	3,569
2027		71	414	485
Total	11,911	19,911	50,655	82,477

16.10.2 Vertical Development

Vertical development or raising is required at Lalor mine for the following reasons: 1) extension of the ventilation system, 2) secondary egress, 3) main orepass for the upper portion of the mine above the 910 m level rock breakers, 4) temporary transfer ore passes to optimize stope extraction and 5) slot raises for longhole stopes.

When a new raise is planned, three options are considered: raise boring, Alimak and drilled drop raising. Typically raises less than 35 m are designed as conventional drop raises and drilled with the production drill on site. Longer raises are evaluated to choose between raise boring and Alimak considering the conditions, end use and cost of both methods. Vertical development required by year to mine the LOM reserves is shown in Table 16-3.

TABLE 16-3: VERTICAL DEVELOPMENT

Year	Ventilation Raises (m)	Ore Handling Raises (m)	Total (m)
2017	465	438	903
2018	411	253	664

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Year	Ventilation Raises (m)	Ore Handling Raises (m)	Total (m)
2019	654	178	832
2020	682	184	866
2021	548	137	685
2022	350	100	450
2023	344	48	392
2024	240		240
2025	80		80
2026
2027
Total	3,774	1,338	5,112

16.11 Diamond Drilling

Diamond drilling is completed by specialist contractor. Ore delineation holes are typically drilled from one zone to the other, and less than 200 m in length. Stope definition holes are drilled from undercut drifts through the contact to determine mucking sill elevations and grade planning (which could vary sequence in some cases). The diamond drill is equipped with universal rotation to allow test holes to be drilled in all azimuths and dips.

16.12 Drainage System

Lalor mine is a relatively dry underground mine with no significant hydrological concerns. The drainage for Lalor mine, Photo Lake mine and Chisel North mine is interconnected. Mine water reports to the water treatment plant at Chisel Lake where it is treated and released. The main collection areas feeding the water treatment plant are the 140 m level Pump Station at Photo Lake mine via the main ramp, surface portal and Photo Lake Pump House; the 840 m level Sump via the Lalor mine Ventilation Raise and Lalor mine Lift station; and the 955 m level main pumps via the Lalor mine Production Shaft and Lalor mine Lift Station. The Chisel North mine water is collected and that water is pumped to the 410 m level sump where it continues to the 955 m level main pumps at Lalor mine. All water within the mine is collected in intermediary collection sumps and proceeds to the main collection areas via drain lines, drain holes or drainage ditches.

16.13 Mining Operations

Typical development crew equipment consists of a two-boom electric hydraulic jumbo, and a mechanical bolter sized to excavate all lateral development (typical sizes include: 5.0 m x 5.0 m, 6.0 m x 5.0 m and 7.0 m x 5.0 m). The crew also uses LHDs, scissor lifts and backhoes for face preparation and extending services.

Production includes an ore handling system capable of removing 4,500 tpd from underground to surface.

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Current production is approximately 3,000 to 3,500 tpd. At steady state by 2018, Lalor mine will produce 4,500 tonnes per day and be approximately 35% development based and 65% longhole. Development based production methods will produce approximately 4.2 ore rounds (1,600 tonnes) per day. Cut and fill mining areas are assumed to be in the ore producing portion of the mining cycle 75% of the time and in the backfill portion of the mining cycle or otherwise unavailable for mining 25% of the time. Longhole mining based production will produce approximately 2,900 tpd.

16.13.1 Mine Equipment

Lalor mine is a ramp and shaft accessible mine with production and development done by rubber tired underground mining equipment. The mine equipment fleet required to achieve 4,500 tpd is shown in Table 16-4.

TABLE 16-4: MINE EQUIPMENT

Description	Fleet
Underground Trucks 65 tonne	4
Underground Trucks 42 tonne	4
LHD 8yd	5
LHD 10yd	5
Two Boom Jumbo	4
Bolters (includes require for cable bolting)	8
Longhole Drills	3
Powder Trucks	3
Scissor Lift Trucks	8
Grader	1
Boom truck	2
Shotcrete Sprayer	1
Trans-mixers	2
Personnel Carriers Toyota	26
Miscellaneous Underground (Minecats, forklifts, etc.)	19
Miscellaneous Surface (Loader, forklift, pickups, etc.)	22
Total Mobile Equipment.	117
Ventilation Fans – Surface fans (250HP – 2500HP)	6
Ventilation Fans – U/G fans (50HP – 400HP)	54
U/G Submersible Pumps 100HP	7
U/G Submersible Pumps 60HP	1
U/G Submersible Pumps 50HP	1
U/G Submersible Pumps 40HP	13
U/G Submersible Pumps 34HP	1
U/G Submersible Pumps 20HP	7
U/G Submersible Pumps <20HP	5
Portable Refuge Stations	3
Shotcrete Machine - Wet Mix	1
Grout Pump c/w Mixer	3
Portable Welder	1
Total Stationary Equipment	103

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An allowance for replacement equipment has been included in the mine plan. As part of the mobile equipment fleet management plan, major mobile equipment will be replaced at approximately 15,000 operating hours.

16.13.2 Production Schedules

The LOM production schedule, shown in Table 16-5, is currently set to ramp-up to 4,500 tonnes per day by 2018 and continues at that rate to the end of 2021 when a ramp-down begins to 3,000 tonnes per day in 2026. Deswik software was used to assist with the LOM planning and generate the basis for the production schedule. The geologic block model was imported to the software, where a stope optimizer algorithm was applied to create economical mining shapes. These mining shapes were then linked to the development drifts and sequenced individually by their respective locations and geometrical limits. Mine resources and rates, realized through historical data, were applied and levelled through activity priority labels. From this output, adjustments were made to further balance resources and scheduling to create an improved plan.

TABLE 16-5: LOM PRODUCTION SCHEDULE

Year	Tonnes	Au (g/t)	Ag (g/t)	Cu (%)	Zn (%)
2017	1,278,282	1.67	22.68	0.59	7.52
2018	1,616,285	2.13	24.37	0.52	5.71
2019	1,620,000	1.86	21.43	0.48	5.62
2020	1,603,652	2.79	28.43	0.79	4.61
2021	1,620,000	2.86	26.39	0.92	4.83
2022	1,473,657	3.16	26.72	0.95	5.72
2023	1,267,267	3.21	29.87	0.89	5.72
2024	1,212,738	3.14	28.35	0.60	4.49
2025	1,212,739	2.89	27.35	0.66	3.41
2026	1,022,918	2.78	32.83	0.49	3.55
2027	304,098	1.83	23.93	0.37	2.68
Total	14,231,636	2.61	26.50	0.69	5.12

Lalor mine will produce a total of 1,312,563 tonnes of zinc concentrate and 404,864 tonnes of copper concentrate in milling the ore from the LOM production plan, as shown in Table 16-6. The LOM contained metal in concentrate is shown in Table 16-6.

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TABLE 16-6: LOM CONCENTRATE PRODUCTION BY YEAR

	Zinc Concentrate		Copper Concentrate
Year	Tonnes	Zn (%)	Tonnes	Au (g/t)	Ag (g/t)	Cu (%)
2017	176,396	51.0	30,158	42.2	499.1	21.0
2018	166,124	51.0	33,298	55.3	552.6	21.0
2019	163,716	51.0	30,863	54.5	541.8	21.0
2020	130,580	51.0	53,179	48.7	492.7	21.0
2021	138,843	51.0	63,025	45.4	448.7	21.0
2022	151,843	51.0	58,903	49.2	446.9	21.0
2023	130,488	51.0	47,567	51.5	483.5	21.0
2024	99,888	51.0	29,786	70.7	549.8	21.0
2025	74,437	51.0	33,629	62.3	514.7	21.0
2026	65,591	51.0	20,064	74.6	649.6	21.0
2027	14,658	51.0	4,394	70.8	653.4	21.0
Total	1,312,563	51.0	404,864	53.4	502.8	21.0

TABLE 16-7: LOM CONTAINED METAL IN CONCENTRATE

Year	Zn (tonnes)	Cu (tonnes)	Au (oz)	Ag (oz)
2017	89,962	6,333	40,917	483,928
2018	84,723	6,993	59,202	591,589
2019	83,495	6,481	54,079	537,611
2020	66,596	11,168	83,265	842,391
2021	70,810	13,235	91,994	909,201
2022	77,440	12,370	93,174	846,328
2023	66,549	9,989	78,760	739,424
2024	50,943	6,255	67,705	526,512
2025	37,963	7,062	67,359	556,492
2026	33,451	4,213	48,122	419,039
2027	7,476	923	10,002	92,306
Total	669,408	85,022	694,578	6,544,821

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16.13.3 Mine Ventilation

The Chisel North mine ventilation system in sequence with the Lalor mine Downcast Raise, provide 400,000 cfm down the Lalor mine Access Ramp, with 150,000 cfm exhausting to surface via the Chisel North mine Ramp. An additional 555,000 cfm is downcast via the Lalor mine Production Shaft for a total of 955,000 cfm exhausting up the Main Exhaust Shaft. In the summer total volume of air increases slightly. Three heaters heat mine air in the winter: the 36M BTU Chisel North Mine Heater, the 30M BTU Lalor Mine Ramp Heater and an 80M BTU heater at the production shaft.

To mine reserves below the 1025 m level there are plans for modifications to the original ventilation system to provide at least 200,000 cfm from the Chisel North mine Ramp System via existing levels and connect 1025 m level and the 865 m level North exploration drift with a 170 m (2.4 m x 2.4 m) Alimak raise. With approximately 150,000 cfm coming down the 1025 Ramp, the mine expects to have 350,000 cfm at its disposal to ventilate future mining horizons below 1025 m level over the LOM. Fresh air is distributed to different areas of the mine via a series of ventilation raises and cross cuts that are developed off the main ramp. Currently 200,000 cfm ventilates the upper section of 10 Lense, 120,000 cfm ventilates the mining of the upper sections of all other levels. Individual mining faces continue to be ventilated using 75 HP to 250 HP fans and 1.2 m ventilation duct. Several larger fans (3 – 300 HP fans) will be needed to provide air to the bottom of the mine as development continues.

With the increase in mining rate from 3,000 to 4,500 tpd several new areas are being brought into production. As the footprint of the mine expands, the ventilation system will also require expansion to allow fresh air to be delivered to active mining areas.

Refer to Figure 16-10 for a ventilation longsection of Lalor mine.

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	Form 43-101F1 Technical Report

FIGURE 16-10: LALOR MINE VENTILATION PLAN

16.13.4 Mine Power

Grid electricity is supplied by Manitoba Hydro, the provincial power utility. Manitoba Hydro’s 115 kV power line terminates at the Chisel North Mine site, approximately 3.5 road km from the Lalor mine site. This feeds power to the Hudbay owned main distribution substation consisting of two (2) 115-25 kV 24 MVA transformers. Substation is completely equipped with an E-house complete with 4 GE Powervac circuit breakers and Tie breaker. From there power is routed as follows:

1. Breaker 52-C1 – Provides power to the following areas

a) The Chisel North mine site substation provides power to the:

a.	Surface ore crusher
b.	Photo Lake pump house
c.	Office buildings
d.	4160V power to the upper part of the Chisel North mine ramp

b) The Lalor mine waste water treatment plant area located 8 km from Lalor mine shaft. The area is equipped with:

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a.	7.5 MVA Interstate 25-6.9 kV transformer that feeds the waste water treatment plant and Chisel Lake fresh water pump house.
b.	500 KVA 25-0.6 kV transformer for power line technician’s workshop.

c) Chisel North mine downcast fan substation. This substation is equipped with

a. 5.5 MVA 25-4.16 kV transformer that provides power to;

i.	600 HP 4160V Alphair Model 10150 AMF550 downcast fan that supply 350,000 cfm heated mine air to the Chisel North mine underground workings.
ii.	4.16-0.6 kV Portable mine power centers in the lower part of the old Chisel North mine.

b. 5 MVA 25-13.8 kV transformer that provides power to the upper part of the Lalor mine.

d) Lalor mine ramp downcast fan site 750 KVA 25-0.6 kV Transformer for the Lalor mine ramp 400 HP Alphair 8400 VAX 3150 Jetstream downcast fan. Fan supply 240,000 cfm heated air to the upper part of the Lalor mine ramp.

2. Breaker 52-L1 – Lalor CCT1 – Provides power to the following areas

a)	Lalor mine site pump house and water treatment plant 750 KVA 25-0.6 kV Transformer
b)	Lalor mine exhaust fans 6 MVA 25-4.16 kV substation providing power to two (2) Howden 2500 HP exhaust fans. Fans are VFD controlled utilizing Rockwell Powerflex 7000 drives.
c)	Hoist house electrical room Bus 1. Bus 1 further distributes power to the following areas:

a. Mine underground Feeder A feeds

I. 910 m level Shaft Station 750 KVA 25-0.6 kV transformer for

i.	Rock breakers
ii.	Settling cones

II. 955 m level 3 MVA 25-4.16 kV transformer for 2 X 1250 HP 10 Stage Mather and Platt dewatering pumps.

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	Form 43-101F1 Technical Report

III. 835 m level 7.5 MVA 25-13.8 kV Underground distribution transformer. – Currently under construction.

b.	Mill Feeder A (Future)
c.	Davey Markham production hoist 3MVA 25-0.6 kV transformer for VFD drives.
d.	Production Hoist 500 KVA 25-0.6 kV MCC.
e.	Hoist house PDC 2.5 MVA 25-0.6 kV transformer. PDC provides to 3 X GA250 1477 cfm, 120 psi Atlas Copco Compressors and Surface Warehouse/Mechanical shop.
f.	Davey Markham – Service hoist 750 KVA 25-0.6 kV MCC
g.	Davey Markham – Auxiliary hoist

Note: Hoist House electrical room switchgear arrangement is equipped with a 1200A Tie breaker between Bus 1 and 2.

3. Breaker 52-L2 – Lalor CCT2 – Provides power to the following areas:

a) Hoist house electrical room Bus 2. Bus 2 further distributes power to the following areas;

a. Mine UG Feeder B feeds

I.	910 m level 7.5 MVA 25-13.8 kV mine underground distribution transformer.
II.	955 m level 750 KVA 26-0.6 kV Conveyor belt and loading pocket arrangement.

b.	Head frame Complex 750 KVA 25-0.6 kV MCC
c.	Mill Feeder B (Future)
d.	Service hoist 2 MVA 25-0.6 kV transformer for VFD Drives.
e.	1.5 MVA 25-0.6 kV transformer for the 2 X Alphair 1015 AMF 5500 Intake fans that supply heated mine air down the Lalor mine shaft.
f.	Lalor mine Office/Dry Unit 4 MVA 25-0.6 kV substation.

Lalor mine pump house and treatment plant is further backed up with a Cummins 350 kW generator that will supply enough power to keep Freshwater, waste water and fire pumps running in case of an emergency.

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The booster pump station located 3.5 km from the Lalor mine shaft is also equipped with a Cummins 350 kW generator that will supply essential power to:

a)	Process water pumps that will supply process water to Lalor mine pump house
b)	Waste water pumps that will pump Lalor mine effluent water to the Chisel Pit for further waste water treatment.

The Lalor mine Auxiliary hoist is also backed up with a Cummins 1000 KVA generator that will supply sufficient power to the auxiliary hoist to move employees up the shaft in the event of an emergency / power failure.

All generators are equipped with automatic transfer switch that will in the event of a power failure transfer to generator power and vice versa.

Lalor mine underground mine electrical distribution to the mine workings consist 13.8 kV that is further stepped down to 600V. This power is used to power up:

a)	Auxiliary ventilation fans,
b)	Auxiliary pumps, and
c)	Mobile electric Jumbo drills and Bolters

Lalor mine underground mine workings currently consists of:

a)	20 X 1 MVA 13.8-0.6 kV mine power centers (aka. portable sleds) and,
b)	16 X 0.75 MVA 13.8-0.6 kV mine power centers

16.14 Workforce

Lalor mine is operated on a continuous cycle. The majority of operations and maintenance personnel work 11.5 hour shifts on a 5-5-4 day cycle or a 7-7 day cycle. Operations support, technical and administrative personnel work 8 hour day shifts, 40 hours per week. The mine is operated under Collective Bargaining Agreements between Hudbay management and local unions.

The mine operations workforce is comprised of Hudbay hourly operations and maintenance personnel as well as salaried supervision, mine administration and technical staff, plus contractor personnel for specialized work and workforce shortages. Personnel will vary year to year. Steady state personnel requirements are shown in Table 16-7.

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	Form 43-101F1 Technical Report

TABLE 16-7: MINE OPERATIONS WORKFORCE

Discipline	Personnel
Direct Operations	180
Supervision and Administration	47
Health and Safety	4
Mine Maintenance	74
Mine Technical	34
Total Lalor mine	339

16.15 Mine Safety and Health

All personnel are required to work under the applicable laws of the Province of Manitoba, Canada. All contractors working on site are required to have an approved health and safety program in place and have on site representation. Hudbay Plant Safety Rules and Regulations are used at Lalor mine operations including, but not limited to:

a)	Positive Attitude Safety System (PASS) safety program
b)	Health monitoring programs (hearing and lung)
c)	Dust monitoring
d)	Ongoing water and environmental monitoring
e)	Personal Protective Equipment (i.e. reflective outerwear, eye protection, hearing protection, respirators)
f)	Task analysis and job procedures

16.15.1 Refuge Stations

Refuge stations are required at Lalor mine as per mine regulations and Hudbay standards and are incorporated into the mine design. Hudbay’s standard refuge station is excavated from rock and requires two ventilation bulkheads, compressed air and a backup oxygen generator, potable water, stretcher kit and first aid supplies, and supplies to seal off the bulkheads.

In new development where it is impractical to excavate a refuge station, portable refuge stations will be used.

16.15.2 Second Egress

Underground mines require a second means of egress. The primary route in and out of the mine is the production shaft equipped with a service cage. The shaft is equipped with a small auxiliary hoist and six person cage. In case of power failure, the auxiliary hoist can be operated by an emergency diesel generator to evacuate personnel from the mine.

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	Form 43-101F1 Technical Report

In the case that the production shaft is not usable, the second egress from the mine is the main ramp to surface at Chisel North mine.

16.16 MINING METHOD OPPORTUNITIES

The Lalor mine is considering different opportunities to improve mining efficiencies:

• Autonomous operation of LHDs are currently being trialed from surface by tele-remote with changes to standard designs to allow isolation of autonomous areas and buffer storage (transfer raises) for in between shift mucking

• A main ore pass from 755 m level to 910 m level is planned for 2017 to reduce trucking time from the upper levels of the mine

• Alternative truck loading systems are being investigated as an alternative to LHD loading, and

• Stoping block design changes are being considered to allow box hole primary mucking and circle route loading of trucks

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	Form 43-101F1 Technical Report

17 RECOVERY METHODS

17.1 Introduction

The Stall concentrator complex is located approximately 16 km east of the Lalor Mine. Conventional crushing, grinding and flotation operations are used to process the ore. The nominal throughput rate will be expanded from the current 3,000 tpd rate to 4,500 tpd and the mill will operate 24 hours per day, 365 days per year, with scheduled downtime for maintenance as required.

The concentrator produces a copper concentrate with gold and silver credits and a zinc concentrate, both are shipped by truck to Flin Flon, from there the copper concentrate is loaded onto rail cars and shipped to third party smelters. Tailings from the flotation circuit will be utilized to produce a cemented paste backfill for use underground. Tailings not required for paste backfill will continue to be pumped to the existing Anderson TIA.

17.2 Stall Concentrator

A simplified block flow diagram for the planned Lalor concentrator is shown in Figure 17-1. Run of mine ore as large as 0.55 m in one dimension is withdrawn from the head frame ore bin by an apron feeder and transported to a crushing plant located at the Chisel North mine, which is ran by a third party contractor.

The contractor crushing plant reduces the ore to a range of 10 to 15 cm and transports the crushed ore to the Stall concentrator coarse ore bins using belly-dump trucks or to a stockpile located at the Stall concentrator using regular dump-trucks. From the stockpile, belly-dump trucks are loaded using a loader and directed to the coarse ore bin.

A stockpile of 36,000 tonnes, equivalent to 8 days production, is required to blend high-grade zinc to ensure a more consistent zinc feed of less than 13%. Ore is reclaimed from the bins via one of two 1220 mm x 1829 mm Syntron vibrating feeders which discharge onto a 1219 mm wide conveyor No 3. This conveyor discharges directly into a 30 x 48 Hewitt Robins jaw crusher 110 kW (150 HP) installed drive and the discharge is combined with the secondary Symons cone discharge to feed a 2440 mm x 6100 mm double deck vibrating screen which has a 39 mm top deck and 19 mm bottom deck.

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	Form 43-101F1 Technical Report

FIGURE 17-1: LALOR CONCENTRATOR SIMPLIFIED BLOCK FLOW DIAGRAM

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	Form 43-101F1 Technical Report

The undersize from the bottom deck is the final product from the crusher circuit which is conveyed to the fine ore bins (FOB) while the oversize of the two decks is combined to feed a 186 kW (250 HP) 7 ft Symons cone crusher. The product out of the Symons at 19 mm is recirculated back to the screen deck in a closed loop.

For the expansion the crusher will retain all of the existing crushing and screening equipment as mentioned above, except for minor modifications (i.e screen openings, minor conveyor upgrades), as shown in Figure 17-2. If the crushing plant falls short of the target production of 4,500 tpd, a standby outdoor portable crusher will be used to make up the shortfall. This portable crusher will produce a particle size of 19 mm and envisions feeding conveyor No. 10 and eventually the FOB.

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FIGURE 17-2: STALL CONCENTRATOR CRUSHING PROCESS FLOW DIAGRAM

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From the FOB, the crushed ore is reclaimed via three 1370 mm wide variable speed belt feeders; two feed the larger Stall rod mill while the single one feeds the smaller Chisel rod mill. The Chisel grinding circuit will continue operating independent of the Stall circuit, providing the ability to continue operating the plant if one of the lines is down for maintenance. It also permits the use of the existing feed bins and feeders without modification.

17.2.1 Chisel Grinding Circuit

The ore is reclaimed from the FOB via a 1370 mm wide variable speed belt and discharges onto the rod mill feed conveyor which feeds into the 2.13 m x 3.05 m rod mill with 150 kW (200 HP) installed. The rod mill discharges into the cyclone feed pump box where it is combined with the ball mill discharge and with dilution water. The variable speed cyclone feed pump discharges into the cyclone pack fitted with two 500 mm cyclones, one operating and one spare. The cyclone underflow returns to the 3.20 m x 3.96 m ball mill with 600 kW (800 HP) installed. The cyclone overflow flows to the combined cyclone overflow pump which feeds the copper flotation feed trash screen. Flotation reagent 3418A is added to the rod mill and lime is added to the ball mill to control the pH of the feed to the flotation circuit.

17.2.2 Stall Grinding Circuit

The ore is reclaimed from the FOB via two 1370 mm wide variable speed belts and discharge onto the rod mill feed conveyor which feeds into the 3.20 m x 4.88 m rod mill with 600 kW (800 HP) installed. The rod mill discharges into the cyclone feed pump box where it is combined with the ball mill discharge and with dilution water. The variable speed cyclone feed pump discharges into the cyclone pack fitted with two 500 mm cyclones, one operating and one spare. The cyclone underflow returns to the 3.81 m x 5.49 m ball mill with 1,200 kW (1,600 HP) installed. The cyclone overflow flows to the two tertiary high frequency each with five decks 1.22 m x 1.52 m with a 150 micron screen cloth. The screen oversize is discharged into the 2.44 m x 3.66 m ball mill with 300 kW (400 HP) installed. Water is added to the screen oversize to optimize the grinding mill density. The screen undersize flows to the combined cyclone overflow pump which feeds the copper flotation feed trash screen. Flotation reagent 3418A is added to the rod mill and lime is added to the ball mills to control the pH of the feed to the flotation circuit.

17.2.3 Flotation Process

The combined primary grinding circuit cyclone overflow and screen undersize are pumped to a bank of new 20 m³ copper rougher flotation cells, which is part of the concentrator expansion, where it is combined with the copper first cleaner tailings, as shown in Figure 17-3. The tailings from the third cell flows by gravity to the existing copper rougher flotation cells.

The existing copper roughers consist of ten Wemco 8.5 m³ cells. The concentrate from the new 20 m³ copper rougher flotation cells is combined with the concentrate from the first four of the existing Wemco rougher flotation cells and is then pumped to the feed of the Woodgrove SFR flotation cells.

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The concentrate from the remaining copper scavenger flotation cells is pumped to the copper regrind mill circuit which consists of a VTM50 mill with 37 W (50 HP) installed. The mill is operated in closed circuit with a cyclopac with 250 mm cyclones. Methyl isobutyl carbinol (MIBC) frother will be added to stabilize the froth. The regrind cyclone overflow is pumped to the two new Woodgrove SFR flotation cells, which is part of the concentrator expansion. The concentrate from these cells, containing 50% of the final copper concentrate is discharged directly to the final copper concentrate pump box. The tailings of the SFR flotation cells are pumped to the existing copper first cleaners where they are combined with the copper second cleaner tailings. The existing first cleaners consist of nine Denver DR24 1.4 m³ cells. The first cleaner tailings are returned to the feed end of the copper rougher scavenger cells and the concentrate progresses to the second cleaners where it is combined with the third cleaner tailings. There is an option to direct the copper first cleaner tailings to the zinc rougher feed conditioner. The second cleaners consist of seven DR24 cells, with the concentrate going to the third cleaners. The third cleaner concentrate, produced from five DR24 flotation cells is combined with the SFR concentrate and pumped to the copper concentrate thickener.

The zinc flotation feed consists of the copper rougher scavenger flotation tailings and the zinc first cleaner tailings. It is first conditioned with copper sulphate, 7279 collector and MIBC frother and is then pumped to the seven existing Wemco 8.5 m³ flotation cells zinc rougher flotation cells. The rougher tailings are pumped to the existing zinc scavengers consisting of five Wemco 8.5 m³ flotation cells. The zinc scavenger tailings and other streams collect in a tailings box and are the final plant tailings. The zinc scavenger concentrate is combined with the zinc rougher concentrate and the zinc second cleaner tailings which are then pumped to the existing fourteen DR24 flotation cells. The first cleaner concentrate is pumped to the second cleaners which consist of the existing ten DR24 flotation cells. The second cleaner concentrate is the final zinc concentrate and is pumped to the zinc concentrate thickener.

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FIGURE 17-3: STALL CONCENTRATOR FLOTATION PROCESS FLOW DIAGRAM

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17.2.4 Concentrate Dewatering

Flocculated copper concentrate will be pumped to thickeners, two 3.4 m existing and one new one 4.6 m, as part of the concentrator expansion, as shown in Figure 17-4. Thickener overflow will be pumped to the copper thickener overflow pumpbox. Underflow, at a target density of 65% solids, will be pumped to an existing agitated stock tank.

Thickened copper concentrate will be further dewatered to approximately 9% moisture on a pressure filter, as part of the concentrator expansion. Filtrate will be recycled to the copper concentrate thickener to prevent the loss of fine solids and reuse the water. Filter cake will be conveyor fed and gravity dropped to the concentrate shed.

Flocculated zinc concentrate will be pumped to a 7.0 m new thickener, as part of the concentrator expansion. The overflow will be recycled to pump No. 108. The underflow, at a target density of 65% solids, will be pumped to an existing agitated stock tank.

Thickened zinc concentrate will be further dewatered to approximately 9% moisture using a pressure filter. Filtrate will be recycled to the zinc concentrate thickener to prevent the loss of fine solids and reuse the water. Filter cake will be gravity fed to the concentrate shed.

There will be no changes to the concentrate shed area. The shed building is a fully enclosed building and contains partitions for separate areas for zinc and copper concentrate storage.

A front end loader is used to separately load the filtered concentrates into trucks for transport to the Hudbay concentrate handling facilities in Flin Flon. Each truck is weighed on a truck scale located in Flin Flon.

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FIGURE 17-4: STALL CONCENTRATOR DEWATERING PROCESS FLOW DIAGRAM

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17.2.5 Tailings

Flotation tailings, from the tailings pump box, can be pumped to the paste plant or to the Anderson TIA, depending on the demand for paste. In the scenario of pumping to the paste plant, the tailings will be sent to a booster station pump box before reaching the paste plant thickener. In the scenario when paste is not required in the mine, the thickened tailings will be diverted at the splitter box to the tailings pump box and pumped to the Anderson TIA.

In terms of utilities, there will be modifications to the compressed air and electrical requirements that are related to the installation of new equipment. Regarding the electrical power distribution; secondary transformation capacity will be added, new power distribution centers will be established, and centralized process control points for new systems. In terms of mill compressed air systems, sufficient, dedicated compressed air capacity will be added to meet the demands of the filter presses.

17.2.6 Water

The Stall Concentrator ore-grade mineral extraction process utilizes two sources of water – fresh water and reclaim water (or recycled water).

• Fresh Water: approximately 25% of the water usage is fresh water withdrawn from Snow Lake. Used in areas such as the worker’s change-rooms/showers, pump gland water, flocculant mixing, On-Stream X-Ray Analyzer and other processes or equipment that requires good quality water with low levels of suspended solids and dissolved compounds. The fresh water pipe-line run is approximately 7 km from Snow Lake to the Stall concentrator. It is fed into a storage tank located in the concentrator and then distributed to the process.

• Reclaim Water: approximately 75% of the water usage is reclaim water withdrawn from the Anderson TIA; also called recycled water, this water is deposited into the Anderson TIA as final process tailings (approximately 30% solids, 70% water). The water is then recycled from the TIA for re-use in the concentrator and utilized for process density control, launder sprays, hosing and other processes or activities that do not require high quality water.

The reclaim water pipe-line run is approximately 5km from the Anderson TIA to the Stall concentrator. It is fed into a storage tank located in the concentrator and then distributed to the process.

The water balance is summarized in Table 17-1.

TABLE 17-1: STALL MILL WATER BALANCE DATA FOR 2016

Month	Operating Hours	Down Time Hours	Reclaim (m3)	Fresh (m3)	Total Water Use (m3)	% Reclaim	% Fresh Water
Jan	588.99	155.01	251236.38	83969.90	335206.28	74.95	25.05

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Month	Operating Hours	Down Time Hours	Reclaim (m3)	Fresh (m3)	Total Water Use (m3)	% Reclaim	% Fresh Water
Feb	604.11	91.89	233411.67	71326.09	304737.76	76.59	23.41
Mar	626.07	117.93	261091.93	83196.46	344288.38	75.84	24.16
Apr	682.75	37.25	256293.89	85732.02	342025.91	74.93	25.07
May	687.34	56.66	260402.83	85917.26	346320.09	75.19	24.81
Jun	544.58	175.42	203135.96	81736.84	284872.80	71.31	28.69
Jul	682.09	61.91	219229.19	84953.77	304182.96	72.07	27.93
Aug	628.7	115.3	218134.34	78646.05	296780.39	73.50	26.50
Sep	671.75	48.25	218331.58	78871.64	297203.22	73.46	26.54
Oct	673.74	70.26	216471.97	80180.81	296652.78	72.97	27.03
Nov	644.43	75.57	252413.37	66937.53	319350.90	79.04	20.96
Dec	582.16	161.84	204013.93	67165.48	271179.41	75.23	24.77
Total	7616.71	1167.29	2794167.02	948633.86	3742800.88	74.65	25.35

17.2.7 Operating Costs

The 2016 cost per tonne milled at Stall concentrator was $23.62 based on milling 1,089,530 tonnes (average 2,985 tpd). The expansion plan at Stall concentrator from the current tonnage of approximately 3,000 tpd to 4,500 tpd in 2018, increases the daily tonnage by 50% and a yearly expenditure increase of 27.5% is envisaged.

The expansion costs per tonne were further dissected into labour, power, operating supplies, maintenance supplies, outside services and G&A. The estimated percentage of the total cost per tonne by allocated area is shown in Table 17-2.

The methodology assumes there will be a cost increase corresponding with the expansion primarily in power, operating and maintenance supplies. However the higher tonnage will offset cost incrementally, resulting in a $20.09 cost per tonne based on 4,500 tpd.

TABLE 17-2: EXPANSION ESTIMATE COST PER TONNE

Allocation	2016 Unit Operating at 3,000 tpd		Tonnes per day increase (%)	Expansion Unit Operating at 4,500 tpd
	$/tonne	% of Cost	50%	$/tonne	% of Cost
Labour	$7.56	32.0%		$5.54	27.6%
Power	$1.42	6.0%		$1.65	8.2%
Operating Supplies	$7.20	30.5%		$7.20	35.9%
Maintenance Supplies	$1.30	5.5%		$1.39	6.9%
Services	$4.72	20.0%		$3.31	16.5%
GandA	$1.42	6.0%		$0.99	4.9%
Total	$23.62	100.0%	50%	$20.09	100.0%

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17.2.8 Expansion Schedule

In terms of schedule of the 4,500 tpd expansion project, it entails primarily engineering starting in the first quarter of 2017, procurement of long lead items in second quarter of 2017, construction phase in third quarter of 2017 and tie-ins/start-up and commissioning in second quarter of 2018.

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18 PROJECT INFRASTRUCTURE

18.1 Lalor Infrastructure

Lalor mine is designed to hoist 6,000 tpd combined ore and waste. Primary access to the mine is a concrete-lined 6.9 m diameter production shaft with a secondary ramp access from the surface through the Chisel North mine. Ore is hoisted to the surface and trucked to the Chisel North site where it is crushed, then hauled to the Stall concentrator for processing into two concentrates (zinc and copper).

Lalor is 16 km by road from the Town of Snow Lake, Manitoba. General area infrastructure includes provincial roads and 115 kV Manitoba Hydro grid power to within four km of Lalor, and Manitoba Telecom land line and cellular phone service. The Town of Snow Lake is a full-service community with available housing, hospital, police, fire department, potable water system, restaurants and stores. The community is serviced by a 914 m gravel airstrip to provide emergency medical evacuation.

Lalor is located 3.5 km from the Hudbay Chisel North mine. Chisel North infrastructure includes a mined out open pit used for waste rock disposal, fresh (process) water sources, pumps and waterlines, 4160V and 550V power, mine discharge water lines, a 2,500 gpm water treatment plant with retention areas, plus mine buildings including offices and a change house.

The permitted Hudbay Anderson TIA, located approximately 12 km from Lalor is used for tailings disposal.

A site drawing of the Lalor access road and services is shown in Figure 18-1.

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	Form 43-101F1 Technical Report

FIGURE 18-1: SITE ACCESS ROAD AND SERVICES

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As of March 2017, offsite infrastructure for the mine operation included:

• 198 person camp in the Town of Snow Lake to house out of town personnel

• 3.5 km gravel access road connecting Provincial Road 395 to the mine site. The road was constructed to Manitoba Class B Feeder road standard.

• Two 24 MVA - 115 kV to 25 kV power substations located on the Chisel North site. These substations provide power for surface and underground mining activities.

• Four km of 25 kV overhead power lines from Chisel North to Lalor

• Process water is pumped 4 km from Chisel Lake through to the booster pump station at Chisel North where it is then pumped the remaining 3.5 km to Lalor

• The Chisel North complex, which is used for the diamond drilling core processing facility, shop to maintain the surface equipment fleet and offices for the project group

• Crushing of the Lalor ore, which is done at the Chisel North site with a maximum total stock pile capacity of 15,000 tonnes

• Booster pump station at Chisel North with holding tanks and pumps for process and discharge water. Equipped with a backup generator (350 kW).

• Two downcast fans, mine heaters (each with a 30,000 US gallon propane tank); the Chisel North downcast (600 hp and 350,000 cfm) and the Lalor ramp downcast (400 hp and 250,000 cfm)

• Discharge water from the Lalor site is pumped 3.5 km to the booster pump station at the Chisel North site where it is pumped the remaining 3.5 km to the settling ponds by the Chisel Pit

• A high density sludge process acidic water treatment plant that can treat up to 2,500 gpm prior to being discharged to the environment

As of March 2017, onsite surface infrastructure includes:

• Office/change house complex with dry space for 341 personnel

• Hoisthouse containing:

o Electrical distribution for the site

o Hoist and communication control room

o Production hoist (Davy Markham, double drum, 4,828 kW)

o Service hoist (Davy Markham, double drum, 2,414 kW)

o Three GA 250 screw compressors (1,477 cfm @120 psi each)

• Headframe which also contains:

o The utility hoist (Davy Markham, single drum, 314 kW)

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o Bin house (capacity 850 tonnes of ore)

o External bunker (capacity 1,000 tonnes of ore)

o Twin 250 hp downcast fans (313,000 cfm each) and mine heater

• Two 30,000 US gallon propane tanks

• Main pump station includes holding tanks (discharge water, process water and potable), PAL water system and pumps for discharge, potable, process and fire water

• Bio Disk Sewage treatment plant (is a natural biological process based on the principle of rotating biological contactor) for up to 38 m³/day

• Fuel tanks and pumps for diesel and gas.

• Two backup generators one for the utility hoist (1,000 kW) and one for the main pump station (350 kW).

• Temporary offices for health safety, training and mine rescue.

• Temporary change house for contractors on site.

• Temporary trailer for onsite contractor.

• Warehouse/shop

• Vent shaft and two exhaust fans (2,500 hp and 575,000 cfm each)

As of March 2017, underground infrastructure includes:

• Main Production shaft 6.9 m diameter concrete lined with five compartments

o Two - 16 tonne skips

o One double deck cage (50 people per deck)

o Counter weight

o Utility hoist (6 person cage)

o Three main shaft stations at 835 m, 910 m and 955 m levels

• Lateral development consists of 6 m x 5 m ramps and level development

• Secondary egress consists of a ramp access from surface to the 810 m level where it joins the rest of the Lalor ramp system. Total approximately distance of 6.0 km.

• Power Distribution consists of:

o 25 kV power lines down the shaft

o 7.5 MVA 25 kV to 13.8 kV transformer to the 910 mL shaft station

o Primary distribution throughout the mine is 13.8 kV with transformers to 600 V for local distribution

• Compressed air and process water are piped throughout the mine from surface

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	Form 43-101F1 Technical Report

• Underground wireless radio communication throughout the mine is provided by a Leakey Feeder system

• Fiber-optic backbone for data and video

• Ore handling system consists of two rock breakers and bins (1,400 tonnes each) on 910 mL feeding chutes and conveyor system on 955 mL supplying the ore to the skips.

• Mobile Maintenance shop is located in the Chisel North underground workings

• Discharge system consists of a series of drain holes and sumps with submersible pumps that feed the top of the two settling cones on 910 mL. The over flow from the cones goes to the main clean water sumps on 955 mL where the water is then pumped to surface by one 1,250 hp 10 stage Mather Platt (740 gpm). There is a second installed spare.

As of March 2017, the schedule for additional permanent infrastructure is:

• Paste plant is scheduled to start construction in 2017 with completion in early 2018

• Ore pass and chute from 910 mL to 755 mL to be operational in the second half of 2017

• A second 7.5 MVA 25 kV to 13.8 kV transformer is to be installed at 835 mL shaft station in 2017

18.2 Lalor Ore Handling Improvements

Based on a review of Lalor’s ore handling circuit by Stantec in early 2017 Hudbay is planning capital improvements in 2017. These improvements will ensure Lalor is able to maintain a steady 4,500 tpd of production through the ore circuit.

Mine personnel have identified that maintenance, particularly repair and replacement of liners in the ore circuit is challenging and to reduce potential hang-ups in the system, as a result the following activities are envisaged:

• Construct muck deadbed in the skip dump in the headframe

• Modifications to transfer car at loading pocket to replace diverter liners with deadbeds

• Install tramp metal grapples at current 910 m level rockbreakers to remove the tramp metal before entering the ore circuit could help to reduce hang-ups, and

• Measuring flask chute angle shallowing to reduce the energy of the muck striking the wall of the skip and to further reducing maintenance

18.3 Paste Plant

The Lalor paste plant project was approved in February 2017 and is critical for the sustainability of the mine production plan. The paste plant will be located northeast of the existing headframe complex and delivery capacity of the paste is 165 tph solids (tails) or 93 m³/hr paste. The paste plant is designed to fill voids left by mining of approximately 4,500 tpd. Taking into account waste generated from development in the LOM and the plan not to hoist waste from underground the combined paste/waste backfilling capacity is approximately 6,000 tpd. The paste plant will be capable of varying the binder content in the paste to provide flexibility in the strength gain of the paste where higher and early strength may be required depending on mining method.

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	Form 43-101F1 Technical Report

Tails that are currently pumped from the Stall concentrator to the Anderson TIA will be diverted to the Anderson booster pump station. Capacity of the pumping station will range from 110 to 130 tph to allow for some variation in the output of tailings from the concentrator. The tailings will be directed into the Anderson TIA when not required for the paste plant.

Two pipelines will be installed between the Anderson booster pump station and the paste plant located at Lalor mine site, approximately a 13 km distance. The tails slurry pipeline is a nominal 14 inch diameter and the return water pipeline is a nominal 10 inch diameter. The main route of this pipeline will be on top of the existing abandoned rail bed (property owned by Hudbay), then along the west side of Provincial Road #395 and finally along the south side of the Lalor mine access road to the paste plant location.

Paste will be delivered underground via one of two – nominal 8 inch diameter, cased boreholes from surface to the 780 mL of the Lalor mine. Only one borehole is required during normal operation, with the second borehole available as a spare in the event of a plug or excessive wear on the primary hole. The boreholes were drilled and cased in 2016.

A network of underground lateral piping and level to level boreholes will transfer the paste from the base of the discharge hopper to the required underground locations. The 780 mL will be the main distribution level to direct the paste to other levels above and below. Underground development is required to extend an existing drift on the 780 mL to intersect the surface boreholes and short cross-cuts on several levels for level to level boreholes. The majority of the underground distribution system will utilize existing drifts or planned future development.

18.4 Stall Concentrator

Hudbay operates the Stall concentrator approximately 16 km from Lalor. The mill is currently operating seven days per week at 3,000 tpd, processing ore from the Lalor mine. The mill has two circuits, with design capacities of 909 tpd and 2,182 tpd.

The concentrator has two flotation circuits producing a zinc concentrate and a copper concentrate. The tailings associated with the Lalor mine are deposited in the Anderson TIA.

Produced copper concentrate is currently hauled by 40 ton trucks to Flin Flon, where the concentrate is loaded onto gondola rail cars for market. Produced zinc concentrate is hauled by 40 ton trucks to Flin Flon and is processed at the Flin Flon Zinc Plant.

As of March 2017, the current Stall concentrator technical specification includes:

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• 2- 350 tonnes coarse ore bins

• 1 Hewitt Robins model W, 30 inch by 48 inch, 150 hp direct V-belt driven jaw crusher

• 1 Symons Standard 7 ft cone, 250 hp direct V-belt driven, direct coupling

• 1 Tyroc double deck screen – 8 ft by 20 ft, 2 inch on top deck, ½ inch x 2½ inch on bottom deck

• 1 Rod mill – AC – 7 ft diameter by 10 ft, 200 hp induction motor

• 1 Ball mill – AC – 10.6 ft diameter by 13ft, 800 hp induction motor

• 1 Rod mill – AC – 10.5 ft diameter by 16ft, 800 hp induction motor

• 1 Ball mill – AC – 12.5 ft diameter by 18ft, 1,600 hp synchronous motor

• 17 - 8.5 m³ Wemco cells

• 1 Regrind mill – AC – 8 ft diameter by 12 ft, 400 hp induction motor

• 22 – Denver cells

18.5 Stall Concentrator Expansion

Engineering work is currently underway at the Stall concentrator as part of the expansion to 4,500tpd with construction slated in third quarter of 2017 and commissioning in third quarter of 2018. The expansion project will address the following areas (Boge & Boge, 2017):

• Crushing – Increase the ore handling capacity for the crusher house conveyors, and change the screen deck media to allow the crushing system to operate to its maximum capacity to produce 4500tpd at a P80 of 16mm from a coarse ore feed F80 of 150mm.

• Grinding – Increase the ore handling capacity of fine ore bin reclaim conveyors and mill feed conveyors. Reconfigure existing mills into two parallel grinding circuits (a smaller two stage (Chisel) and larger three stage (Stall)), and introduce high frequency screening in closed circuit with the third stage Stall ball mill to ensure product sorting by size rather than density to produce a blended flotation feed P80 of 100 µm at 4500tpd.

• Copper Flotation – Expanded capacity at the copper roughers with 3 new 20m3 tank cells, and expanded capacity of copper cleaners with addition of staged flotation reactors. There will be some pump capacity modifications to address the increased volumes through the circuit. Introduction of a regrind mill to reduce the rougher scavenger tails to target a P80 of 50µm prior to the copper cleaners to improve metal recovery.

• Zinc Flotation – The zinc flotation capacities predicted by the mine planned and head grades are similar to the existing operating conditions and no flotation capacity upgrade requirements were identified. Some pump capacity modifications are required to address extraneous process streams.

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	Form 43-101F1 Technical Report

• Thickening – Introduction of a large (8m) zinc thickener to address the existing and future limitations on zinc concentrate production, and add a new copper thickener (or modify the two existing 3.4m thickeners with de-aeration tanks and froth rings) to extend the existing thickeners’ capacity to meet the increased copper concentrate loads.

• Dewatering – Replace the existing disc thickeners with pressure filters to increase concentrate production capacity to match the new throughput. This will improve copper concentrate from 19% moisture to 9% moisture, and will improve zinc moisture from 14% to 8% moisture.

• Electrical Power Distribution – Add secondary transformation capacity, establish new power distribution centers, and centralize process control points for new systems.

• Mill Compressed Air Systems – Add sufficient, dedicated compressed air capacity to meet the significant demands of the filter presses.

• Mill water distribution systems – Increase the mill distribution pump capacity to satisfy the increase water demand.

18.6 Anderson TIA

Anderson TIA is located in the Snow Lake area between the Stall concentrator and Lalor mine. The purpose of the Anderson TIA is environmental management (storage) of mine tailings produced in the Stall concentrator which processes ore from the Lalor operation.

The Anderson TIA has been in use since 1979, when a control dam was built at the east end of Anderson Lake across Anderson Creek. Seasonal discharge of water out of the Anderson TIA occurs during the open-water season (usually May to October). Water quality at the final Anderson TIA discharge point has at all times been in compliance with applicable regulatory requirements. Tailings are deposited subaqueously into the TIA and no treatment, other than retention in the TIA, has ever been required.

Hudbay has submitted a Notice of Alteration to Manitoba Sustainable Development to expand the TIA within the existing limits to accommodate the future tailings produced through the entire Lalor mine operations. The construction of this expansion is anticipated to be complete prior to 2019.

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	Form 43-101F1 Technical Report

19 MARKET STUDIES AND CONTRACTS

Hudbay has a marketing division that is responsible for establishing and maintaining all marketing and sales administrations of concentrates and metals. As well, Hudbay conducts ongoing research of metal prices and sales terms as part of normal business and long range planning process. Contract terms used in the Lalor financial evaluation are based on this research and the author has reviewed these results and they support the assumptions made in this technical report.

Lalor will produce a zinc concentrate and a copper concentrate with gold and silver credits. Zinc concentrates are trucked to Hudbay’s operations in Flin Flon where they are processed into refined zinc and sold to customers in North America. The key long-term assumptions for the sale of Lalor’s zinc metal and zinc concentrate are summarized in Table 19-1. This report assumes zinc concentrate will be processed at the Flin Flon zinc plant from 2017-2021 and after that time Lalor’s zinc concentrate will be sold to third party refineries.

TABLE 19-1: KEY LONG-TERM ZINC METAL AND ZINC CONCENTRATE ASSUMPTIONS

	Units	LT Total / Average
Zinc Concentrate Grade	%	51%
Moisture Content of Zinc Concentrate	% H2O	9%
Zinc Concentrate Base Treatment Charge	US$/ tonne concentrate	$200
Zinc Concentrate Metal Price Basis	US$ / tonne Zinc metal	$2,204.6
Zinc Concentrate Escalator	%	6%
Zinc Concentrate De-escalator	%	3%
Zinc Concentrate Payability	%	85%
Zinc Concentrate Minimum Deduction	%	8%
Zinc Concentrate Freight Cost	C$/wmt	$118
Freight Allowance/Capture	US$/ wmt concentrate	$40
Zinc Metal Premium	US$/lb	$0.07
Zinc Metal Distribution Cost	US$/lb	$0.055

The copper concentrate produced at Lalor is sent to copper smelters in North America by rail. The key assumptions for the sale of Lalor’s copper concentrate are summarized in Table 19-2 below:

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	Form 43-101F1 Technical Report

TABLE 19-2: KEY LONG-TERM COPPER CONCENTRATE ASSUMPTIONS

	Units	LT Total / Average
Copper Grade in Copper Concentrate	% Cu	21%
Moisture Content of Copper Concentrate	% H2O	9%
Copper Concentrate Base Treatment Charge	US$ / dry tonne con	$80
Copper Refining Charge	US $ / lb Cu	$0.08
Silver Refining Charge	US $ / oz Ag	$0.50
Gold Refining Charge	US$ / oz Au	$5.00
Copper Concentrate Freight Cost	C$ / wet tonne con	$213
Copper Payability	%	96.5%
Copper Minimum Deduction	%	1%
Gold Payability	%	96%
Silver Payability	%	90%

Engineering, supply and construction contracts are initiated, managed and administrated by Hudbay’s Manitoba Business Unit. Hudbay follows a standard contracting out process that specifies contractors’ requirements to be eligible to be considered for work. Contractor selection criteria include ability to complete the work within the required time, safety record and programs, price, and proposed alternatives. The Lalor contracts that are in place have rates and charges that are within industry norms.

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	Form 43-101F1 Technical Report

20 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT

20.1 Environmental Studies and Planning

Commencing in 2007, AECOM carried out the environmental baseline investigations needed to conclude an environmental impact assessment for the Lalor project, including all necessary terrestrial and aquatic field studies. Much of the early baseline work was summarized in AECOM’s Lalor Advanced Exploration Project Plan (“Lalor AEP”), which was submitted to and approved by Manitoba Mines Branch and the Lalor Mine Environment Act Licence (EAL) application submitted to and approved by Manitoba Sustainable Development. Baseline work was utilized in the Lalor Paste Plant Notice of Alteration (NoA) which was submitted to the Manitoba Sustainable Development in Q4 2016 and was approved in January 2017. AECOM has also conducted baseline work and studies which are summarized in the Anderson TIA expansion NoA submitted in Q3 2016 to Manitoba Sustainable Development for approval.

Hudbay is currently reviewing improvement plans for existing operations in the Snow Lake area. For each project associated with previously permitted sites, Hudbay will submit a NoA to Manitoba Sustainable Development for review and approval.

Due to the extensive work completed by AECOM and other existing studies completed as part of Environmental Effects Monitoring programs at the various operations in the Snow Lake area, it is contemplated that no additional baseline studies are necessary for potential future improvement projects. There is no present indication that future approvals will not be obtained to meet potential future construction schedules.

20.2 Waste, Tailings Disposal and Water Management

There are no known environmental concerns which could adversely affect Hudbay’s ability to mine ore from Lalor mine. Because of its location in close proximity to the existing facilities in the Snow Lake area, Lalor was able to utilize existing infrastructure, services, and previously disturbed land that is associated with permitted, pre-existing and current mining operations in the Snow Lake area. The Lalor mine and associated projects are designed to minimize the potential impact on the surrounding environment by keeping the footprint of the operations as small as possible and by using existing licensed facilities for the withdrawal of water and disposal of wastes.

The NoA for the expansion of Anderson TIA was prepared by AECOM utilizing geotechnical tailings dam designs from Hudbay’s Engineer of Record; BGC Engineering Inc. As detailed in the Environmental Assessment of the Proposed TIA Expansion submitted as a NoA, the entire volume of tailings from Lalor LOM was to be stored in Anderson TIA (AECOM, 2016). This conceptual design did not discount the volume of tails that could be used for paste backfill at Lalor. In order to de-risk the construction of the proposed dams the project was proposed in 3 separate stages with the first to occur prior to 2019 based on available storage in the TIA.

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The initial stage of expansion (Stage A) was to provide sufficient storage for tailings production volumes such that future dam raises could be planned based on actual production rates. This would be possible as the actual tailings production could be measured against remaining volumes as ongoing bathymetry surveys are a requirement of the existing licence. Compared to the original design criteria, the predicted volume of tailings from Lalor operation is currently much less due to paste plant operation. This overall reduction in required tailings storage will result in adequate storage volumes with only Stage A being constructed. It is anticipated that Stage A will cost in the order of $7M in 2018. Stage A construction is detailed in Figure 20-1, which was submitted as part of the NoA.

20.3 Permitting Requirements

The existing Lalor mine EAL was obtained in the first quarter of 2014 and covers all facilities on the Lalor site, including sewage and mine wastewater treatment facilities and the pipelines which carry freshwater into the site and remove treated wastewater from it. The sources of freshwater and other facilities where treated wastewater are discharged to the environment are existing operating sites which are licensed provincially and regulated under the federal Metal Mines Effluent Regulation.

The main permits required for the Lalor operation are presently valid licenses and permits for the Lalor mine, Stall Concentrator, Anderson TIA, and New Britannia site. Applications for the Anderson TIA expansion NoA have been submitted for approval. Other upgrades and augmentation plans may require the submission of a NoA to an existing licensed operation but no new tailings impoundment area will be required. No federal permits are anticipated.

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	Form 43-101F1 Technical Report

FIGURE 20-1: ANDERSON TIA STAGE 1A CONSTRUCTION PLAN (AECOM, 2016)

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	Form 43-101F1 Technical Report

Presently the New Britannia site inclusive of the Birch Lake Tailings Management Facility (BLTMF) although currently in care and maintenance, has a current EAL and the seasonal discharge from BLTMF is regulated under the MMER. Hudbay is currently in the process of applying for a new water withdrawal licence for this site which is anticipated to be obtained before potential operational needs. Potential future use of the New Britannia site will require the submission of a NoA in order to process material from the Lalor mine.

The approval process and time requirements have been contemplated in regards to overall project milestones. There is no indication that the approvals will not be obtained within the project schedule.

20.4 Mineral Lease and Surface Lease

Prior to commercial production of ore at the Lalor mine, a mineral lease was applied for and obtained from the Manitoba Mines Branch. The mineral lease grants the holder the exclusive right to mine minerals within the lease area.

As the entire Paste Plant is on Hudbay held property and it utilizes existing infrastructure, there are no land tenure concerns or additional leases required. Approval for this project was received in January 2017.

In specific areas associated with proposed pipeline routes and future improvements to the existing Anderson TIA, surface leases will be required. Activities are currently underway to apply for and obtain the required surface leases. There is no indication that theses leases cannot be obtained in the time lines of the expansion project.

20.5 Community Support

The main settlement in the region of the Lalor mine is the Town of Snow Lake, which is an important mining and service centre for the Ecodistrict and surrounding area. Snow Lake has a population of approximately 840 according to the 2006 data from Statistics Canada, with the majority of these residents employed at or supported by the mines located throughout the area. Many other Snow Lake residents are employed in the industries and services that support the region’s mining operations.

Hudbay and AECOM have carried out public consultation, including meetings to inform local communities about the progress of development of the Lalor mine and expansion of Anderson TIA and environmental effects of these projects. Manitoba Sustainable Development has taken these meetings into account in the environmental licensing process.

The projects will continue to provide jobs for both Flin Flon and the Town of Snow Lake during construction of upgrades and continued operation of the mine. The additional feed from the mine will also help ensure the continued employment of Hudbay employees in the Flin Flon and Snow Lake areas. Since the economies of both communities are based on mining, opposition to the projects is seen as unlikely.

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20.6 Aboriginal People and First Nations

Based on Hudbay’s long-term (more than 50 years) mining experience in the Snow Lake region, there is no known current First Nation or Aboriginal hunting, fishing, trapping or other traditional use in the zone of potential influence for the Lalor mine, other current operations, and potential future projects. There is no First Nation Registered Trapline District or Reserve in the area that will be affected by the Lalor operation. Although development on the Mine Site involved a loss of vegetation and habitat for wildlife, the vegetation and habitat type is common throughout the region.

The Mathias Colomb First Nation (“MCCN”), located 125 km northwest of Snow Lake at the community of Pukatawagan, has asserted a right to be consulted in connection with the Lalor operation and expansion of Anderson TIA. Hudbay continues attempts to initiate an information sharing process with MCCN and expects to be able to provide Manitoba regulators with all information necessary to support previous Crown consultation decisions made prior to approval of the Lalor mine EAL. No impact to current operations or delays in project schedule is anticipated.

20.7 Heritage Resources

Operation of the Lalor mine and construction of potential future upgrades will not affect any known site of potential historical, archaeological or cultural significance. Approximately 20 km south of the Lalor operation is Tramping Lake, which is the site of one of Manitoba’s largest known concentrations of aboriginal pictographs. These paintings are thought to have been created 1,500 to 3,000 years ago by the Algonkian-speaking ancestors of the Cree and Ojibway First Nations. Activities associated with the Lalor operation will not have any impact on this historical site.

20.8 Mine Closure Requirements and Plans

The Manitoba Mines and Minerals Act requires a closure plan and financial assurance for any advanced exploration or mining project. Manitoba accepted Closure Plans prepared by SRK in 2005 and financial assurance to cover the cost of closure for all existing infrastructure that will continue to be used during operation of the Lalor mine. Existing facilities which support the Lalor mine include the Chisel North mine, which is connected by an underground ramp to Lalor, Stall Concentrator and Anderson TIA, piping systems associated with milling and tailings deposition, the Chisel Open Pit and the Chisel North water treatment plant.

Prior to commercial production at the Lalor mine, Manitoba approved the Closure Plan for the Lalor AEP and accepted financial assurance in the amount of $1.5 million. The Lalor AEP Closure Plan was prepared in 2010 and approved as part of the AEP application process. As a requirement of the Lalor mine EAL, an updated closure plan was prepared by SRK and submitted for approval in September 2014. The estimated cost of the closure and post-closure activities detailed in the updated Closure Plan is $1.73 million. It is anticipated that the site of the Lalor mine will be substantially returned to its natural state in about five to ten years post closure, after which no monitoring or other measures will be required.

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NoA applications for the Lalor paste plant, expansion of the Anderson TIA, and upgrades to the New Britannia site also will require the submission of updated closure plans and financial assurance. It is expected that as a condition of Manitoba’s approval, closure planning and financial assurance will be required within a year of final construction activities. Allowances for these applications are contemplated items as part of future years budgeting process.

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21 CAPITAL AND OPERATING COSTS

21.1 Introduction

Capital and operating costs are estimated in constant 2017 Canadian dollars.

21.2 Capital Costs

The total development capital required to increase throughput at Lalor to the targeted 4,500 tpd is estimated to be C$117 million, as shown in Table 21-1, which includes approximately an 18% contingency. This capital is expected to be spent during 2017 and 2018.

TABLE 21-1: DEVELOPMENT CAPITAL COST SUMMARY

Development Capital	000 C$
Paste Backfill	67,786
Ore Handling Underground	3,250
Stall Mill Upgrades	45,870
Total Development Capital	116,906

The development capital costs were estimated internally by Hudbay with input from Golder Associates, Stantec Inc. and Boge & Boge Ltd.

The LOM sustaining capital costs are estimated to be C$220 million. The breakdown of the sustaining capital over the next 5 years and for the LOM is shown in Table 21-2.

TABLE 21-2: SUSTAINING CAPITAL COST SUMMARY

Sustaining Capital (000 C$)	2017	2018	2019	2020	2021	2017-LOM
Mine Capital and Development	17,927	18,381	13,442	13,849	10,635	86,771
Normal Capital	3,000	10,000	3,000	3,000	3,000	28,000
Replacement Equipment	11,629	14,739	13,390	11,004	9,776	91,518
Major Installations	3,640	5,924	1,461	982	835	13,803
Total Sustaining Capital	36,196	49,044	31,293	28,835	24,246	220,092

Normal capital includes C$7 million related to the expansion of the Anderson tailings facility and C$21 million related to the Stall concentrator. This sustaining capital estimate has made no consideration for the availability of used equipment from the 777 and Reed mines. When these mines eventually close, some equipment may be available for use at Lalor and may reduce the sustaining capital estimate above. No contingency has been included in the sustaining capital estimate.

Reclamation costs, salvage value and severance costs have not been considered in this report.

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	Form 43-101F1 Technical Report

21.3 Operating Costs

Operating costs were developed by Hudbay based on a bottom-up approach and utilizing budget quotes from local suppliers, Manitoba operations experience, labor costs within the region and actual costs at Lalor.

The mine plus mill unit operating costs are estimated to be C$99.83/tonne mined over the LOM. The addition of cemented rock fill and paste backfill has increased the mine unit costs, but maximizes recovery of the mineral resource and results in lower capitalized costs than prior years due to less underground development. Table 21–3 summarizes the mine plus mill unit operating costs over the next 5 years and for the LOM.

TABLE 21-3: UNIT OPERATING COST SUMMARY

(C$/tonne mined)	2017	2018	2019	2020	2021	2017-LOM
Mine Development	26.53	15.07	14.87	13.64	13.98	17.12
Ore Extraction	20.28	27.19	33.60	35.59	34.75	30.50
Ore Removal	29.87	30.24	28.22	28.20	28.47	30.70
Total Mine Operating Costs	76.67	72.49	76.68	77.43	77.20	78.32
Mill Operating Costs1	22.02	20.14	20.12	20.19	20.12	21.51
Total Mine + Mill Operating Costs	98.70	92.63	96.80	97.63	97.32	99.83

¹ Milling costs include concentrate haulage to Flin Flon

The total C1 cash costs and sustaining cash costs (net of by-product credits) per pound of zinc over the LOM and over the next 5 years are shown in Table 21-4. C1 cash costs include on-site and off-site costs. Sustaining cash costs include C1 costs plus sustaining capital.

TABLE 21-4: CASH COSTS (NET OF BY-PRODUCT CREDITS)

Cash Costs (Net of By-Product Credits1)	Units	Next 5 Years	LOM
C1 Cash Costs	US$ / lb Zn in con	$0.41	$0.37
C1 Cash Costs + Sustaining Capex	US$ / lb Zn in con	$0.57	$0.50

¹ By-product credits are calculated using the following assumptions: copper price per pound - US$2.60 in 2017, US$2.75 in US$3.00 in 2019 to 2020 and long-term; gold price per ounce - US$1,300 in 2017 to 2020 and US$1,260 long-term; silver price ounce - US$18.00 in 2017 to 2020 and long-term; CAD/USD exchange rate - 1.35 in 2017, 1.25 in 2018, 1.20 in 2019, 1.15 in and 1.10 long-term.

Lalor’s annual zinc production (contained zinc in concentrate) and C1 cash costs (net of products) are shown below in Figure 21-1. Over the first 5 years, annual production is expected average 79 thousand tonnes of zinc at an average C1 cash cost of US$0.41/lb. Over the 10.5 LOM, annual production is expected to average 64 thousand tonnes of zinc at an average C1 cost of US$0.37/lb. Lower C1 cash costs from years 2020 to 2023 are a result of mining the copper gold (Zone 27).

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FIGURE 21-1: LALOR ANNUAL ZINC PRODUCTION AND C1 CASH COSTS

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	Form 43-101F1 Technical Report

22 ECONOMIC ANALYSIS

Hudbay is a producing issuer and has excluded information required by Item 22 of Form 43-101F1 as the updated mine plan does not represent a material increase of Hudbay’s current production.

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23 ADJACENT PROPERTIES

The author is not aware of any current relevant work on properties immediately adjacent to the Lalor deposit.

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24 OTHER RELEVANT DATA AND INFORMATION

24.1 Gold Bulk Sample Program

In the fourth quarter of 2016, Hudbay developed 37 drift rounds at Lalor to assess the continuity and variability of non-contact gold mineralization within discrete areas of zones 21 and 25. Approximately 10,000 tonnes of bulk sample material was mined and hauled to surface. The material was primary crushed and processed through a sample tower to collect a representative subsample of each development round. The integrity of the material was maintained at all times through a rigorous chain of custody process. The mined material, stored on surface, is available for milling pending the potential economic viability of refurbishing the New Britannia gold mill in Snow Lake by Hudbay.

The run of mine (ROM) material for the bulk sampling program came from mining in three different areas of the northern parts of gold zones 21 and 25 on 975 m level, 995 m level and 1000 m level. The areas were selected as to provide maximum spatial distribution of the samples with minimal need for additional capital expenditures on underground development.

The bulk sample program was conducted to evaluate the mining potential of the gold zones and to increase the confidence in results obtained by modeling of the gold zones based on diamond drill hole data.

The grade of the mined material was derived by the collection of manageable sized subsamples representative of each individual round for assaying. The sub samples were collected by the use of a contractor crushing and a custom built sample tower.

To ensure the validity of the obtained results QAQC procedures exceeding best industry practices were in place. The QAQC procedures allowed for the ROM material to be traced back to individual LHD buckets from any particular round. Total round separations was assured by the use of a ROM surface bunker, and elaborate cleaning procedures for all equipment used between rounds.

Other parts of the project included:

• Testing feasibility and performance of various grade control procedures and techniques to be used in connection with the potential to mine future non-contact gold resources (diamond drilling down drifts, sampling while mucking and chip sampling)

• Structural studies to develop a better understanding of the controls of grade distribution within the gold zones

The preliminary results indicate that the gold grades from the bulk sample program are as expected with minor variations when compared to those modeled based on diamond drill data. The bulk sample program has increased the confidence and the understating of the gold zones and gold mineralization at Lalor.

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	Form 43-101F1 Technical Report

Following full assessment of the 2016 bulk sample data it is intended to collect additional subsamples in 2017 from other areas of gold zones 21 and 25 as well as from other previously untested gold zones for confirmation purposes.

24.2	Taxes and Royalties
24.2.1	Applicable Tax Rates

The Lalor mine is not directly taxable as Hudbay pays provincial and federal taxes on a legal entity basis. The combined federal and provincial tax rates are assumed to be approximately 27% for the LOM and Hudbay has approximately C$750 million in tax pools that can be used to offset future income taxes for federal and provincial purposes. Hudbay’s mining operations in Manitoba are also subject to the Manitoba Mining Tax. The Manitoba Mining Tax is not applied to a new mining project until the original capital expenditures are recovered.

24.2.2 Royalties

There are no royalties applicable to Lalor.

The author is not aware of any other information that would impact the reported estimate of mineral resources or estimated mineral reserves for the Lalor deposit.

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	Form 43-101F1 Technical Report

25 INTERPRETATION AND CONCLUSIONS

The Lalor mine operation has been mining ore since August 2012. Since then the mine has operated uninterrupted and been in a continuous production ramp-up cycle, achieving the highest annual tonnage of approximately 1.1 million tonnes in 2016, with complementary throughput at the Stall concentrator. The production ramp-up is planned to continue in 2017 to reach a steady state of 4,500 tpd by the first quarter of 2018.

The production increase of 50%, compared to current production, is supported by an underground ore handling circuit capable of 4,500 tpd, transitioning to more bulk mining methods (65% of reserves) with additional mining fronts and design changes to improve mining efficiencies, developing ore passes and transfer raises to reduce truck haulage cycle times from the upper potions of the mine and commissioning of a paste plant backfill plant in the first quarter of 2018. Autonomous operation of a Load Haul Dump loader underground is currently being trialed from surface by tele-remote monitoring with changes to standard designs to allow isolation of autonomous areas and buffer storage for in between shift mucking.

The increase in production to 4,500 tpd at Lalor is complemented by the Stall concentrator expansion to 4,500 tpd, which is currently underway and is expected to be commissioned in the third quarter of 2018.

The mineral resources, as of September 30, 2016, are estimated as base metal lenses or gold zones based on geological and mineralization properties. The Hudbay validation process and third party review confirmed the resource block model is interpolated using industry accepted modelling techniques and classified in accordance with the 2014 CIM Definition Standards – For Mineral Resources and Mineral Reserves.

A mine reconciliation of the mined out areas compared to the ore reported at the concentrator was very close on the precious metals and a slight conservatism of the zinc and copper grades might be evident. This conservatism of the base metals is likely due to over constraining the high grade samples to 20 m as part of the high yield restriction step.

The mineral resources stated with a metal equivalency cut-offs provide for economic extraction of reserves from stated resources.

The mineral reserves, as of January 1, 2017, are based on a LOM plan that generated a mining inventory based on stope geometry parameters with appropriate dilution and recovery factors The conversion of resources to reserves is based on the LOM plan and NSR cut-offs that primarily focussed on capturing base metal resources for processing at the Stall concentrator. The secondary focus was to capture gold zone resources when in contact with or close proximity to base metal resources. In areas where a large separation existed between base metal and gold lenses, mining blocks were evaluated for economic stope mining shapes. When a non-economic shape was generated in a first pass, a second pass was evaluated for only base metal lenses and if an economic shape was generated the gold zone portion was removed. However, due to this larger separation, majority of these isolated gold lenses could have been evaluated independently of the base metal lenses and could potentially provide feed to a gold processing facility. Below approximately the 950 m level no attempt was made to generate an economic stope mining shape for gold zones 25 and 26 as the separation distance became too large. The author’s opinion is that these resources are potentially better suited for a gold processing facility and should be re-evaluated when Hudbay has a better understanding of their New Britannia gold mill and Birch Tailings Impoundment Area in Snow Lake.

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	Form 43-101F1 Technical Report

The author considers that the mineral reserves as classified and reported comply with all disclosure in accordance with requirements and CIM Definitions. The author is not aware of any mining, metallurgical, infrastructure, permitting or other relevant factors that could materially affect the mineral reserve estimate.

The production and compilation of this technical report was supported by the capable and professional management and staff at Hudbay. The supervision, revision and approval of the assembly of this Technical Report is by the QP Robert Carter, P. Eng., Lalor Mine Manager at Hudbay Manitoba Business Unit.

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	Form 43-101F1 Technical Report

26 RECOMMENDATIONS

It is recommended that the following actions be performed:

• Investigate the impact of under assaying of high grade standards at Hudbay’s Flin Flon laboratory and whether this in turn affects the high grade Lalor samples submitted and has potentially led to an underestimations of gold in the resource estimate, since the proportion of samples assayed at the laboratory was approximately 80% of the total samples assayed between 2012 and 2016

• Investigate the high yield restriction parameters of the high grade base metal samples, and consider whether the restriction distance is suitable or were they over constrained, based on the conservatism noted in the mine reconciliation for zinc and copper

• Pursue the option of a temporary paste backfill plant to utilize the boreholes from surface prior to commissioning of the permanent plant in the first quarter of 2018. This option provides assurance to achieve the ramp-up in production and is another source of backfill rather than relying on waste development.

• Due to the approximate 6 month timing offset of the production ramp-up at Lalor to 4,500 tpd and the Stall concentrator expansion to 4,500 tpd, Hudbay should pursue transporting of ore from Lalor to their Flin Flon concentrator for earlier processing.

• Finalize the evaluation of the gold bulk sample program conducted in the fourth quarter of 2016 and since Hudbay owns a sample tower consider collecting additional subsamples from other areas of gold zones 21 and 25 as well as from other previously untested gold zones for confirmation purposes.

• Hudbay owns the New Britannia mill, a gold leach plant on care and maintenance, in Snow Lake. Hudbay should continue to assess the feasibility of processing a portion of the material mined from the gold zone and copper-gold zone at Lalor at the New Britannia mill at a rate of 1,500 tpd starting in 2019. When combined with the processing capacity of the Stall concentrator, this would enable an aggregate throughput rate of up to 6,000 tpd and utilize the full capacity of the Lalor mine shaft.

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	Lalor Mine
	Form 43-101F1 Technical Report

27 REFERENCES

• AECOM, Internal Report to Hudbay: Environmental Assessment of Proposed Expansion of the Anderson Tailings Impoundment Area, Sadiq, S., Ryman, J., Kjartanson, S., 2016.

• AMEC, Evaluation of Twin and Duplicate Samples: The Hyperbolic Method: Discussion, Simon, A., 2004.

• Arik, A. 2002,” Resource Classification Index”, MineSight in the Foreground.

• Bailes, A.H., and Galley, A.G., 1999. Evolution of the Paleoproterozoic Snow Lake arc assemblage and geodynamic setting for associated volcanic-hosted massive sulphide deposits, Flin Flon Belt, Manitoba, Canada. Canadian Journal of Earth Sciences, 36, pp. 1789-1805.

• Bailes, A.H., and Galley, A.G., 2007. Geology of the Chisel-Anderson Lakes Area, Snow Lake, Manitoba (NTS areas 63K16SW and west half of 63J13SE). Manitoba Science, Technology, Energy and Mines, Manitoba Geological Survey, Geoscientific Map MAP2007- 1, scale 1:20,000 plus notes.

• Boge & Boge Ltd, Internal Report to Hudbay: Debottlenecking Concept Study – Stall Concentrator, March 2017.

• Carter, R, Schwartz, T., West, S., Hoover, K., 2012: Pre-Feasibility Study Technical Report on the Lalor Deposit, Snow Lake, Manitoba, Canada Effective Date: March 29, 2012, Hudbay pp. 1-292, filed on sedar.com

• CIM Definition Standards on Mineral Resources and Mineral Reserves, May 10, 2014, website web.cim.org/standards

• David, J., Bailes, A.H., and Machado, N., 1996: Evolution of the Snow Lake portion of the Paleoproterozoic Flin Flon and Kisseynew belts, Trans-Hudson Orogen, Manitoba, Canada. Precambrian Research, 80, pp. 107-124.

• Franklin, J.M., Gibson, H.L., Jonasson, I.R., and Galley, A.G., 2005. Volcanogenic Massive Sulfide Deposits, in Hedenquist, J. W., Thompson, J. F. H., Goldfarb, R. J., and Richards, J. P., eds., Economic Geology 100th Anniversary Volume: Littleton, CO, Society of Economic Geologists, pp. 523-560.

• Froese, E., and Moore, J.M., 1980. Metamorphism in the Snow Lake area, Manitoba; Geological Survey of Canada, Paper 78-27, p. 16.

• Golder Associates, Internal Report to Hudbay: Lalor Mine Paste Backfill System Snow Lake, Manitoba, March 2017.

• Mercier-Langevin, P., 2009. Field Notes and Observations Lalor deposit, Snow Lake April 22-25^th, 2009.

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	Form 43-101F1 Technical Report

• Manitoba, Department of Growth, Enterprise and Trade (GET) websites,

http://www.gov.mb.ca/iem/geo/exp-sup/files/fig1.pdf,
http://www.gov.mb.ca/iem/geo/exp-sup/files/fig6.pdf,
http://web33.gov.mb.ca/mapgallery/mgg-gmm.html [January 31, 2013].

• Ontario Securities Commission website, http://www.osc.gov.on.ca/en/15019.htm

• SGS Vancouver Metallurgy: A Report on the Recovery of Copper, Zinc Gold and Silver From Lalor Samples, July 20, 2009

• SGS Vancouver Metallurgy; An Investigation Into the Recovery of Copper, Lead, Zinc, and Gold From Lalor Samples – Phase II, February 25, 2011

• SRK Consulting, Internal Memo to Hudbay: Lalor Mine Bulk Sample Area Observations and Recommendations, JF Ravenelle, March 2017.

• SRK Consulting, Internal Report to Hudbay: Structural Geology Investigation of Zone 25, Lalor Zinc, Gold, and Copper Mine, Manitoba, JF Ravenelle, July 2016.

• Stantec Consulting Ltd., Internal Report to Hudbay: Lalor Mine, Mine Throughput Review, February 2017.

• Statistics Canada, Census Profile, 2016 Census for Snow Lake, Manitoba

http://www12.statcan.gc.ca/census-recensement/2016/dp-pd/prof/details/Page.cfm

• Syme, E.C., Lucas, S.B., Zwanzig, H.V., Bailes, A.H., Ashton, K.E., and Haidl, F.M., 1998: Geology, NATMAP Shield Margin Project area, Flin Flon Belt, Manitoba–Saskatchewan; Manitoba Energy and Mines, Geological Services, Map A-98-2, scale 1:100 000 and 1:350 000, with accompanying notes, 54 p.

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	Form 43-101F1 Technical Report

28 SIGNATURE PAGE

This Technical Report titled “NI 43-101 Technical Report, Lalor Mine, Snow Lake, Manitoba, Canada”, dated March 30, 2017 and effective as of March 30, 2017 was prepared under the supervision and signed by the following author:

Dated effective this 30th day of March, 2017.

(signed) Robert Carter

________________________
Signature of Qualified Person

Robert Carter, P. Eng.
Lalor Mine Manager, Hudbay Manitoba Business Unit

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29 CERTIFICATES OF QUALIFIED PERSONS

Robert Carter

CERTIFICATE OF QUALIFICATION

Re: Lalor Mine Technical Report, March 30, 2017

I, Robert Carter, P.Eng., of Burlington, Ontario, do hereby certify that:

1.	I am currently employed as Lalor Mine Manager, Hudbay Manitoba Business Unit, with Hudbay Minerals Inc. (the “Issuer”), 25 York Street, Suite 800, Toronto, Ontario, Canada, M5J 2V5
2.	I graduated from University of Manitoba with a Bachelor of Sciences in Geological Engineering in 1997.
3.	I am a member in good standing of the Association of Professional Engineers and Geoscientists of the Province of Manitoba, Registration #21836.
4.	I am a member in good standing of the Association of Professional Engineers of Ontario, Registration #100089189.
5.	I have practiced my profession continuously for over 19 years and have been involved in mineral exploration, mine site engineering and geology, mineral resource and mineral reserve evaluations, and mine operations for base metal deposits and operations in North and South America.
6.	I have read the definition of “qualified person” set out in National Instrument 43-101 (NI 43- 101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purpose of NI 43-101.
7.	I have reviewed and approved and I am responsible for the preparation of this Technical Report titled “NI 43-101 Technical Report, Lalor Mine, Snow Lake, Manitoba, Canada”, dated March 30, 2017 (the “Technical Report”) and effective as of March 30, 2017.
8.	I last visited the property on March 29, 2017. I am directly involved with Lalor mine on a permanent basis because of my role as mine manager and I personally inspect the operation on a routine basis.
9.	As of the date of this certificate, to the best of my knowledge, information and belief, the parts of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

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10.	I am not independent of the Issuer. Since I am an employee of the Issuer, a producing issuer, I fall under subsection 5.3 (3) of NI 43-101 where “a technical report required to be filed by a producing issuer is not required to be prepared by or under the supervision of an independent qualified person.”
11.	I have been involved with the Lalor property, which is the subject of the Technical Report, continuously since discovery in 2007.
12.	I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with the instrument and form.
13.	I consent to the public filing of the Technical Report with any stock exchange, securities commission or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated this 30th day of March, 2017.

Original signed by:

Robert Carter

Robert Carter, P. Eng.

Lalor Mine Manager, Hudbay Manitoba Business Unit

Page 29-2

EX-99.3

ROBERT CARTER

CERTIFICATE OF QUALIFICATION

Re: Lalor Mine Technical Report, March 30, 2017

I, Robert Carter, P.Eng., of Burlington, Ontario, do hereby certify that:

1.	I am currently employed as Lalor Mine Manager with Hudbay Minerals Inc. (the “Issuer”), 25 York Street, Suite 800, Toronto, Ontario, Canada, M5J 2V5
2.	I graduated from University of Manitoba with a Bachelor of Sciences in Geological Engineering in 1997.
3.	I am a member in good standing of the Association of Professional Engineers and Geoscientists of the Province of Manitoba, Registration #21836.
4.	I am a member in good standing of the Association of Professional Engineers of Ontario, Registration #100089189.
5.	I have practiced my profession continuously for over 19 years and have been involved in mineral exploration, mine site engineering and geology, mineral resource and mineral reserve evaluations, and mine operations for base metal deposits and operations in North and South America.
6.	I have read the definition of “qualified person” set out in National Instrument 43-101 (NI 43- 101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purpose of NI 43-101.
7.	I have reviewed and approved and I am responsible for the preparation of this Technical Report titled “NI 43-101 Technical Report, Lalor Mine, Snow Lake, Manitoba, Canada”, dated March 30, 2017 (the “Technical Report”) and effective as of March 30, 2017.
8.	I last visited the property on March 30, 2017. I am directly involved with Lalor mine on a permanent basis because of my role as mine manager and I personally inspect the operation on a routine basis.
9.	As of the date of this certificate, to the best of my knowledge, information and belief, the parts of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
10.	I am not independent of the Issuer. Since I am an employee of the Issuer, a producing issuer, I fall under subsection 5.3 (3) of NI 43-101 where “a technical report required to be filed by a producing issuer is not required to be prepared by or under the supervision of an independent qualified person.”

11.	I have been involved with the Lalor property, which is the subject of the Technical Report, continuously since discovery in 2007.
12.	I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with the instrument and form.
13.	I consent to the public filing of the Technical Report with any stock exchange, securities commission or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated this 30^th day of March, 2017.

Original signed by:

Robert Carter

Robert Carter, P. Eng.

Lalor Mine Manager, Hudbay Manitoba Business Unit

EX-99.4

CONSENT OF QUALIFIED PERSON

March 30, 2017

British Columbia Securities Commission
Alberta Securities Commission
Financial and Consumer Affairs Authority of Saskatchewan
The Manitoba Securities Commission
Ontario Securities Commission
Autorité des marchés financiers
New Brunswick Financial and Consumer Services Commission
Nova Scotia Securities Commission
Office of the Superintendent of Securities, Prince Edward Island
Office of the Superintendent of Securities, Service Newfoundland and Labrador
Office of the Superintendent of Securities, Northwest Territories
Office of the Yukon Superintendent of Securities
Nunavut Securities Office

Dear Sir/Madam:

Re: Technical Report entitled “NI 43-101 Technical Report, Lalor Mine, Snow Lake, Manitoba, Canada” dated effective March 30, 2017

I, Robert Carter, consent to the public filing of the technical report titled “NI 43-101 Technical Report, Lalor Mine, Snow Lake, Manitoba, Canada” dated effective March 30, 2017 (the “Technical Report”) by Hudbay Minerals Inc. (“Hudbay”).

I also consent to the use of any extracts from, or a summary of, the Technical Report in the annual information form for the year ended December 31, 2016, dated March 30, 2017 (the “AIF”) of Hudbay.

I certify that I have read the AIF and that it fairly and accurately represents the information in the Technical Report for which I am responsible.

(signed) Robert Carter

Robert Carter, P. Eng

Lalor Mine Manager, Hudbay Manitoba Business Unit

EX-99.5

NI 43-101 Technical Report

Feasibility Study

Updated Mineral Resource, Mineral Reserve and Financial Estimates

Rosemont Project

Pima County, Arizona, USA

Issue and Effective Date: March 30, 2017

25 York Street, Suite 800

Toronto, Ontario

Canada M5J 2V5

Prepared by:

Cashel Meagher, P.Geo.

Senior Vice President and Chief Operating Officer, Hudbay

Revision 3

CAUTIONARY NOTE REGARDING FORWARD-LOOKING INFORMATION

This Technical Report contains "forward-looking statements" and "forward-looking information" (collectively, "forward-looking information") within the meaning of applicable Canadian and United States securities legislation. All information contained in this Technical Report, other than statements of current and historical fact, is forward-looking information. Often, but not always, forward-looking information can be identified by the use of words such as “plans”, “expects”, “budget”, “guidance”, “scheduled”, “estimates”, “forecasts”, “strategy”, “target”, “intends”, “objective”, “goal”, “understands”, “anticipates” and “believes” (and variations of these or similar words) and statements that certain actions, events or results “may”, “could”, “would”, “should”, “might” “occur” or “be achieved” or “will be taken” (and variations of these or similar expressions). All of the forward-looking information in this Technical Report is qualified by this cautionary note.

Forward-looking information includes, but is not limited to, our objectives, strategies, intentions, expectations, production, cost, capital and exploration expenditure guidance, including the estimated economics of the Rosemont project, future financial and operating performance and prospects, anticipated production at our Rosemont project and processing facilities and events that may affect Hudbay’s operations, anticipated cash flows from operations and related liquidity requirements, the anticipated effect of external factors on revenue, such as commodity prices, estimation of mineral reserves and resources, mine life projections, reclamation costs, economic outlook, government regulation of mining operations, and expectations regarding community relations. Forward-looking information is not, and cannot be, a guarantee of future results or events. Forward-looking information is based on, among other things, opinions, assumptions, estimates and analyses that, while considered reasonable by us at the date the forward-looking information is provided, inherently are subject to significant risks, uncertainties, contingencies and other factors that may cause actual results and events to be materially different from those expressed or implied by the forward-looking information.

The material factors or assumptions that we identified and were applied by us in drawing conclusions or making forecasts or projections set out in the forward-looking information include, but are not limited to:

• the success of mining, processing, exploration and development activities;
• the accuracy of geological, mining and metallurgical estimates;
• anticipated metals prices and the costs of production;
• the supply and demand for metals we produce;
• the supply and availability of concentrate for our processing facilities;
• the supply and availability of third party processing facilities for our concentrate;

• the supply and availability of all forms of energy and fuels at reasonable prices;
• the availability of transportation services at reasonable prices;
• no significant unanticipated operational or technical difficulties;
• the execution of our business and growth strategies, including the success of our strategic investments and initiatives;
• the availability of additional financing, if needed;
• the ability to complete project targets on time and on budget and other events that may affect our ability to develop our projects;
• the timing and receipt of various regulatory, governmental and joint venture partner approvals;
• the availability of personnel for our exploration, development and operational projects and ongoing employee and union relations;
• the ability to secure required land rights to develop the Pampacancha deposit;
• maintaining good relations with the communities in which we operate, including the communities surrounding our Rosemont project;
• no significant unanticipated challenges with stakeholders at our various projects;
• no significant unanticipated events or changes relating to regulatory, environmental, health and safety matters;
• no contests over title to our properties, including as a result of rights or claimed rights of aboriginal peoples;
• the timing and possible outcome of pending litigation and no significant unanticipated litigation;
• certain tax matters, including, but not limited to current tax laws and regulations; and
• no significant and continuing adverse changes in general economic conditions or conditions in the financial markets (including commodity prices and foreign exchange rates).

The risks, uncertainties, contingencies and other factors that may cause actual results to differ materially from those expressed or implied by the forward-looking information may include, but are not limited to, risks generally associated with the mining industry, such as economic factors (including future commodity prices, currency fluctuations, energy prices and general cost escalation), uncertainties related to the development and operation of our projects (including risks associated with the permitting, development and economics of the Rosemont project and related legal challenges), dependence on key personnel and employee and union relations, risks related to political or social unrest or change, risks in respect of aboriginal and community relations, rights and title claims, operational risks and hazards, including unanticipated environmental, industrial and geological events and developments and the inability to insure against all risks, failure of plant, equipment, processes, transportation and other infrastructure to operate as anticipated, compliance with government and environmental regulations, including permitting requirements and anti-bribery legislation, depletion of Hudbay’s reserves, volatile financial markets that may affect our ability to obtain additional financing on acceptable terms, the failure to obtain required approvals or clearances from government authorities on a timely basis, uncertainties related to the geology, continuity, grade and estimates of mineral reserves and resources, and the potential for variations in grade and recovery rates, uncertain costs of reclamation activities, Hudbay’s ability to comply with its pension and other post-retirement obligations, our ability to abide by the covenants in our debt instruments and other material contracts, tax refunds, hedging transactions, as well as the risks discussed under the heading “Risk Factors” in our most recent Annual Information Form and our management’s discussion and analysis of Hudbay for the year ended December 31, 2016.

Should one or more risk, uncertainty, contingency or other factor materialize or should any factor or assumption prove incorrect, actual results could vary materially from those expressed or implied in the forward-looking information. Accordingly, you should not place undue reliance on forward-looking information. We do not assume any obligation to update or revise any forward-looking information after the date of this Technical Report or to explain any material difference between subsequent actual events and any forward-looking information, except as required by applicable law.

 

	Rosemont Project
	Form 43-101F1 Technical Report

TABLE OF CONTENTS

SECTION			PAGE
TABLE OF CONTENTS			i
LIST OF TABLES			v
LIST OF FIGURES			x
LIST OF APPENDICES			xv
1	SUMMARY		1-1
	1.1	Introduction	1-1
	1.2	Property Description and Location	1-2
	1.3	Accessibility, Climate, Local Resources, Infrastructure and Physiography	1-3
	1.4	History	1-4
	1.5	Geological Setting and Mineralization	1-5
	1.6	Deposit Types	1-6
	1.7	Exploration	1-6
	1.8	Drilling	1-7
	1.9	Sample Preparation, Analyses, and Security	1-7
	1.10	Data Verification	1-8
	1.11	Mineral Processing and Metallurgical Testing	1-9
	1.12	Mineral Resource Estimate	1-10
	1.13	Mineral Reserves Estimate	1-15
	1.14	Mining Methods	1-19
	1.15	Recovery Methods	1-25
	1.16	Project Infrastructure	1-25
	1.17	Market Studies and Contracts	1-27
	1.18	Environmental Studies, Permitting and Social or Community Impact	1-28
	1.19	Capital and Operating Cost	1-29
	1.20	Economic Analysis	1-29
	1.21	Adjacent Properties	1-33
	1.22	Other Relevant Data and Information	1-33
	1.23	Conclusions	1-33
	1.24	Recommendations	1-35
2	INTRODUCTION AND TERMS OF REFERENCE		2-1
	2.1	Information Sources	2-1
	2.2	Unit Abbreviations	2-2
	2.3	Name Abbreviations	2-3
3	RELIANCE ON OTHER EXPERTS		3-1
4	PROPERTY DESCRIPTION AND LOCATION		4-1
	4.1	Location	4-1
	4.2	Land Tenure	4-1
5	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		5-1
	5.1	Accessibility	5-1
	5.2	Climate	5-1
	5.3	Local Resources	5-2

Page i

	Rosemont Project
	Form 43-101F1 Technical Report

	5.4	Infrastructure	5-2
	5.5	Physiography	5-3
6	HISTORY		6-1
	6.1	Helvetia-Rosemont Mining District (1875 – 1973)	6-1
	6.2	Anamax Mining Company (1973 - 1985)	6-1
	6.3	ASARCO, Inc. (1988 – 2004)	6-2
	6.4	Augusta Resource Corporation (2005 – 2014)	6-2
	6.5	Hudbay (2014 – Present)	6-3
7	GEOLOGICAL SETTING AND MINERALIZATION		7-1
	7.1	Tectonic and Metallogenic Setting	7-1
	7.2	Regional Geology	7-2
	7.3	District Geology	7-2
	7.4	Chemostratigraphy	7-8
	7.5	Structural Domains	7-8
	7.6	Mineralization	7-9
	7.7	Mineralization Domains	7-11
	7.8	Alteration and Skarn Development	7-12
	7.9	Clay Proxies	7-13
8	DEPOSIT TYPE		8-1
9	EXPLORATION		9-1
10	DRILLING		10-1
	10.1	Banner Mining Company (1961 to 1963)	10-3
	10.2	The Anaconda Mining Co., (1963 to 1986)	10-3
	10.3	ASARCO Mining Co., (1988 to 2004)	10-3
	10.4	Augusta Resource (2005 to 2012)	10-3
	10.5	Hudbay (2014 to 2015)	10-4
11	SAMPLING PREPARATION, ANALYSES, AND SECURITY		11-1
	11.1	Hudbay 2014	11-1
	11.2	Hudbay 2015	11-13
	11.3	Augusta	11-22
	11.4	Historic	11-29
12	DATA VERIFICATION		12-1
	12.1	Drill Hole Database	12-1
13	MINERAL PROCESSING AND METALLURGICAL TESTING		13-1
	13.1	Overview	13-1
	13.2	Historical Metallurgical Testwork Summary	13-2
	13.3	Hudbay Metallurgical Testing Programs	13-6
	13.4	XPS Phase 1	13-6
	13.5	XPS Phase 2	13-9
	13.6	XPS Phase 3	13-12
	13.7	BML Confirmation Testing	13-14
	13.8	BML Production Period Testwork	13-14
	13.9	Concentrate Quality	13-17
	13.10	Tailings Dewatering	13-18
	13.11	Recovery Estimates	13-19
	13.12	Conclusions and Recommendations	13-21

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	Rosemont Project
	Form 43-101F1 Technical Report

	13.13	Discussion and Recommendations	13-22
14	MINERAL RESOURCE ESTIMATE		14-1
	14.1	Key Assumptions of Model	14-1
	14.2	Wireframe Models and Mineralization	14-1
	14.3	Exploratory Data Analysis	14-7
	14.4	Assays	14-7
	14.5	Composites	14-27
	14.6	Variography	14-29
	14.7	Estimation and Interpolation Methods	14-33
	14.8	Tonnage Factor Assignment	14-36
	14.9	Block Model Validation	14-40
	14.10	Visual Inspection	14-40
	14.11	Metal Removed by Capping	14-45
	14.12	Global Bias Checks	14-48
	14.13	Local Bias Checks	14-51
	14.14	Block Model Quality Control	14-64
	14.15	Grade-Tonnage Statistics	14-64
	14.16	Classification of Mineral Resource	14-67
	14.17	Third Party Review	14-69
	14.18	Internal Peer Review	14-69
	14.19	Reasonable Prospects of Economic Extraction	14-69
	14.20	Mineral Resource Statement Inclusive of Mineral Reserve	14-71
	14.21	Sensitivity of the Mineral Resource	14-72
	14.22	Comparison with the 2012 Resource Estimates	14-74
	14.23	Factors That May Affect the Mineral Resource Estimate	14-75
	14.24	Conclusions	14-75
	14.25	Recommendations	14-76
15	MINERAL RESERVES ESTIMATE		15-1
	15.1	Pit Optimization	15-1
	15.2	Mineral Reserves	15-8
16	MINING METHODS		16-1
	16.1	Mine Overview	16-1
	16.2	Mine Phases	16-2
	16.3	Mine Schedule and Production Plan	16-12
	16.4	Mine Facilities	16-27
	16.5	Mine Equipment	16-30
	16.6	Mine Operations	16-34
	16.7	Mine Engineering	16-35
	16.8	Manpower Requirements	16-40
17	RECOVERY METHODS		17-1
	17.1	Introduction	17-1
	17.2	Buildings	17-3
	17.3	Processing Plant	17-3
	17.4	Crushing	17-6
	17.5	Grinding	17-8
	17.6	Copper Flotation	17-10
	17.7	Copper-Molybdenum Separation	17-13
	17.8	Molybdenum Concentrate Thickening, Filtration and Drying	17-15
	17.9	Copper Concentrate Dewatering and Storage	17-15

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	Rosemont Project
	Form 43-101F1 Technical Report

	17.10	Tailings Thickening	17-17
	17.11	Tailings Filtration Plant	17-17
	17.12	Reagents and Consumables	17-19
	17.13	Plant Services	17-21
	17.14	Process Control Strategy	17-22
18	PROJECT INFRASTRUCTURE		18-1
	18.1	Access Roads, Plant Roads and Haul Roads	18-1
	18.2	Power Supply and Distribution	18-2
	18.3	Water Supply and Distribution	18-3
	18.4	Tailings Management	18-5
	18.5	Communications	18-7
19	MARKET STUDIES AND CONTRACTS		19-1
20	ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT		20-1
	20.1	Reclamation and Closure Plan	20-4
21	CAPITAL AND OPERATING COSTS		21-1
	21.1	Introduction	21-1
	21.2	Capital Costs	21-1
	21.3	Operating Costs	21-4
	21.4	Working Capital Costs	21-5
22	ECONOMIC ANALYSIS		22-1
	22.1	Key Model Assumptions	22-1
	22.2	Annual Cash Flow Model	22-5
	22.3	Financial Analysis (100% Project Basis)	22-10
	22.4	Sensitivity Analysis (100% Project Basis)	22-10
	22.5	Project Ownership Impact on Valuation	22-11
23	ADJACENT PROPERTIES		23-1
24	OTHER RELEVANT DATA AND INFORMATION		24-1
	24.1	Project Implementation	24-1
	24.2	Risk Assessments	24-4
25	INTERPRETATION AND CONCLUSIONS		25-1
26	RECOMMENDATIONS		26-1
27	REFERENCES		27-1
28	SIGNATURE PAGE		28-1
29	CERTIFICATES OF QUALIFIED PERSONS		29-1
	CASHEL MEAGHER		29-1

Page iv

	Rosemont Project
	Form 43-101F1 Technical Report

LIST OF TABLES

TITLE	PAGE
Table 1-1: Rosemont Deposit Drilling Summary	1-7
Table 1-2: Resource by Category, Mineralized Zone and NSR Cut-Off (1)(2)(3)(4)(5)(6)(7)(8)(9)(10)	1-13
Table 1-3: Measured and Indicated, Comparison to 2012 Augusta Estimate	1-14
Table 1-4: Inferred, Comparison to 2012 Augusta Estimate	1-14
Table 1-5: Proven and Probable Mineral Reserves in Rosemont Final Pit	1-17
Table 1-6: Rosemont Mineral Exclusive Resource Estimates (1)(2)(3)(4)(5)(6)(7)(8)(9)	1-18
Table 1-7: Proven and Probable, Comparison to 2012 Augusta Reserve Estimate	1-19
Table 1-8: Rosemont Slope Guidance	1-20
Table 1-9: Rosemont Mine Phases Mineral Reserves	1-22
Table 1-10: Mine Production Schedule Criteria	1-22
Table 1-11: Waste Rock Facility Design Criteria	1-24
Table 1-12: DSTF Buttress Rock Storage Design Criteria	1-24
Table 1-13: Copper Concentrate	1-27
Table 1-14: Molybdenum Concentrate	1-28
Table 1-15: Metal Price Assumptions	1-29
Table 1-16: Life of Mine Financial Metrics (100% Project Basis)	1-31
Table 1-17: After-Tax NPV8%, NPV10%, IRR and Payback Sensitivity at Various Flat Copper Prices (100% Basis)	1-32
Table 1-18: Key Financial Metrics Attributable to Hudbay	1-33
Table 2-1: Unit Abbreviations	2-2
Table 2-2: Name Abbreviations	2-3
Table 6-1: Historical Sulfide Mineral Resource (Augusta 2012)	6-3
Table 10-1: Rosemont Deposit Drilling Summary	10-1
Table 11-1: Bureau Veritas Assay Specifications	11-4
Table 11-2: Summary Of QA/QC Samples	11-5
Table 11-3: Oreas Certified Blanks	11-5
Table 11-4: Summary of Blank Performance	11-6
Table 11-5: Oreas Certified Reference Material	11-6
Table 11-6: Summary of CRM Performance	11-7
Table 11-7: Summary of Coarse Duplicate Analysis	11-9
Table 11-8: Summary of QA/QC Samples	11-14
Table 11-9: Summary of Blank Performance	11-15
Table 11-10: Oreas Certified Reference Material	11-15

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	Rosemont Project
	Form 43-101F1 Technical Report

Table 11-11: Summary of CRM Performance	11-17
Table 11-12: Summary of Coarse Duplicate Analysis	11-18
Table 11-13: Assay Specifications – Skyline	11-24
Table 11-14: Summary of Blank Performance at Skyline	11-25
Table 11-15: Standard Reference Materials – Augusta	11-26
Table 11-16: Performance of Standard Reference Materials at Skyline	11-27
Table 11-17: Rosemont Deposit Drilling Summary	11-29
Table 11-18: Comparison of Historical Assay Results and Twin Half-Split Core Samples Analyzed by Augusta at Skyline	11-31
Table 12-1: Downhole Surveys of Historical Drilling	12-2
Table 12-2: Hudbay 2014 Downhole Results	12-2
Table 12-3: Hudbay 2015 Downhole Surveys	12-2
Table 12-4: Drill Hole Assay Ranking	12-4
Table 13-1: Grinding Mill Sizing Parameters	13-3
Table 13-2: Molybdenite Flotation	13-4
Table 13-3: Lithology of Composite Samples	13-5
Table 13-4: 2012 Closed Circuit Flotation Results	13-5
Table 13-5: XPS Phase 1 - Comminution Test Statistics*	13-7
Table 13-6: XPS Phase 2 - Qemscan Analysis	13-10
Table 13-7: XPS Phase 2 - CEC Analysis	13-10
Table 13-8: XPS Phase 2 - Copper Deportment by Mineral Species	13-10
Table 13-9: XPS Phase 2 - Locked Cycle Test Results	13-11
Table 13-10: XPS Phase 3 - Molybdenum Separation Test	13-13
Table 13-11: XPS Phase 3 - Flotation Blends Test Results	13-13
Table 13-12: XPS Phase 3 - Flotation Water Test Results	13-14
Table 13-13: Production Composites	13-15
Table 13-14: BML Mineral Content	13-15
Table 13-15: BML CEC Analysis	13-16
Table 13-16: BML Locked Cycle Test Results	13-17
Table 13-17: Production Recovery Profile	13-20
Table 14-1: Drilling Data By Company	14-1
Table 14-2: Legend of Interpreted Wireframes	14-4
Table 14-3: Samples and Length Analyzed	14-7
Table 14-4: Capping Thresolds by Lithology	14-12
Table 14-5: Assay Statistics for Total Copper by Lithology in Sulfides	14-13
Table 14-6: Assay Statistics for Acid Soluble Copper by Lithology in Sulfides	14-13

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	Rosemont Project
	Form 43-101F1 Technical Report

Table 14-7: Assay Statistics for Molybdenum by Lithology in Sulfides	14-14
Table 14-8: Assay Statistics for silver by Lithology in Sulfides	14-14
Table 14-9: Assay RMA Regression Parameters, Silver Against Copper	14-16
Table 14-10: Drilling Data by company	14-17
Table 14-11: Matrix of Boundary Conditions, Total Copper	14-23
Table 14-12: Matrix of Boundary Conditions, Acid Soluble Copper	14-24
Table 14-13: Matrix of Boundary Conditions, Molybdenum	14-25
Table 14-14: Matrix of Boundary Conditions, Silver	14-26
Table 14-15: Length Weighted Uncapped and Capped 25-foot Composite Statistics, Copper in Sulfides	14-27
Table 14-16: Length Weighted Uncapped and Capped 25-foot Composite Statistics, Acid Soluble Copper in Sulfides	14-27
Table 14-17: Length Weighted Uncapped and Capped 25-foot Composite Statistics, Molybdenum in Sulfides	14-28
Table 14-18: Length Weighted Uncapped and Capped 25-foot Composite Statistics, Silver in sulfides	14-28
Table 14-19: Variogram Models and Rotation Angles	14-32
Table 14-20: Copper and Acid Soluble Copper Grade Model Interpolation Plans	14-34
Table 14-21: Molybdenum and Silver Grade Model Interpolation Plans	14-35
Table 14-22: Measured Compared to Calculated Specific Gravity	14-36
Table 14-23: Specific Gravity Measurements per Lithology and Oxidation State	14-37
Table 14-24: SG Baseline Values per Lithology and Oxidation Level	14-38
Table 14-25: Tonnage Factors by Lithology and Oxidation State	14-39
Table 14-26: Assay, NN, IDw and OK Model, Copper Removed by Capping in Blocks Within the Resource Pit and Above $5.7/Ton NSR	14-46
Table 14-27: Assay, NN, IDW and OK Model, Acid Soluble Copper Removed by Capping in Blocks Within the Resource Pit and Above $5.7/Ton NSR	14-46
Table 14-28: Assay, NN, IDW and OK Model, Molybdenum Removed by Capping in Blocks Within the Resource Pit and Above $5.7/Ton NSR	14-47
Table 14-29: Assay, NN, IDW and OK Model, Silver Removed by Capping in Blocks Within the Resource Pit and Above $5.7/Ton NSR	14-47
Table 14-30: NN, IDW and OK Model Statistics Mean Block Grade Comparisons for Copper in Blocks Within the Resource Pit and Above $5.7/Ton NSR	14-49
Table 14-31: NN, IDW and OK Model Statistics Mean Block Grade Comparisons for Acid Soluble Copper in Blocks Within the Resource Pit and Above $5.7/Ton NSR	14-49
Table 14-32: NN, IDW and OK Model Statistics Mean Block Grade Comparisons for Molybdenum Silver in Blocks Within the Resource Pit and Above $5.7/Ton NSR	14-50
Table 14-33: NN, IDW and OK Model Statistics Mean Block Grade Comparisons for Silver in Blocks Within the Resource Pit and Above $5.7/Ton NSR	14-50

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	Rosemont Project
	Form 43-101F1 Technical Report

Table 14-34: Quality Control Statistics of the Copper interpolation in Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell	14-64
Table 14-35: Grade-Tonnage Statistics, Copper	14-65
Table 14-36: Lerchs-Grossman Cone Inputs	14-70
Table 14-37: Economic Parameters	14-71
Table 14-38: Resource by Category, Mineralized Zone and NSR Cut-Off (1)(2)(3)(4)(5)(6)(7)(8)(9)(10)	14-72
Table 14-39: Measured Resource by Mineralized Zone and Multiple NSR Cut-Offs	14-73
Table 14-40: Indicated Resource by Mineralized Zone and Multiple NSR Cut-Offs	14-73
Table 14-41: Inferred Resource by Mineralized Zone and Multiple NSR Cut-Offs	14-74
Table 14-42: Measured and Indicated, Comparison to 2012 Augusta Estimate	14-74
Table 14-43: Inferred, Comparison to 2012 Augusta Estimate	14-75
Table 15-1: Metallurgical Recoveries Used In Lerchs-Grossman Evaluations	15-2
Table 15-2: Base-Case Lerchs-Grossman Economic Parameters	15-3
Table 15-3: Overall Slope Angles Used in Lerchs-Grossman Analysis	15-5
Table 15-4: Pit Design Parameters	15-8
Table 15-5: Proven and Probable Mineral Reserves in Rosemont Final Pit	15-10
Table 15-6: Proven and Probable Mineral Reserves in Rosemont Final Pit by Ore Type	15-10
Table 15-7: Rosemont Mineral Exclusive Resource Estimates	15-12
Table 15-8: Proven and Probable, Comparison to 2012 Augusta Reserve Estimate	15-13
Table 16-1: Pit Design Parameters	16-2
Table 16-2: Rosemont Slope Guidance	16-3
Table 16-3: Rosemont Mine Phases Mineral Reserves	16-11
Table 16-4: Rosemont Mine Phases, Mineral Reserves by Ore Type	16-11
Table 16-5: Mine Production Schedule Criteria	16-12
Table 16-6: Mill Ramp-Up Schedule	16-12
Table 16-7: Mine Production Schedule – LOM RP16Aug	16-25
Table 16-8: WRSA Design Criteria	16-27
Table 16-9: DSTF Buttress Rock Storage Design Criteria	16-28
Table 16-10: LOM Waste Rock Distribution and Landforming Storage Plan	16-29
Table 16-11: Material Characteristics	16-30
Table 16-12: Major Fleet Requirements for LOM	16-32
Table 16-13: Major Equipment KPI and Productivity	16-33
Table 16-14: Labor Estimation for Rosemont Mine Operations	16-41
Table 16-15: Labor Estimation for Rosemont Process Operations	16-42
Table 16-16: labor Estimation for Rosemont G&A Operations	16-43
Table 17-1: Key Facility Design Criteria	17-3

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	Rosemont Project
	Form 43-101F1 Technical Report

Table 17-2: Key Design Criteria	17-5
Table 17-3: Plant Utilization Summary	17-6
Table 19-1: Copper Concentrate	19-1
Table 19-2: Molybdenum Concentrate	19-1
Table 21-1: INITIAL Capital Cost Summary	21-1
Table 21-2: Sustaining Capital	21-3
Table 21-3: Operating Cost Summary	21-4
Table 21-4: Cash Costs (Net of By-product Credits at Stream Prices)	21-4
Table 22-1: Metal Price Assumptions	22-1
Table 22-2: Mine and Mill Operating Assumptions used in the Financial Model (100% Project Basis)	22-2
Table 22-3: Capital and Operating Cost Assumptions used in the Financial Model (100% Project Basis)	22-3
Table 22-4: Tax Depreciation for Development Capital	22-5
Table 22-5: Annual Cash Flow Model	22-6
Table 22-6: Life Of Mine Financial Metrics (100% Project Basis)	22-10
Table 22-7: After-Tax NPV8%, NPV10% and IRR Sensitivity at Various Flat Copper Prices (100% Basis)	22-11
Table 22-8: Key Financial Metrics Attributable to Hudbay	22-12
Table 24-1: Risk Assessments	24-4

Page ix

	Rosemont Project
	Form 43-101F1 Technical Report

LIST OF FIGURES

TITLE	PAGE
Figure 1-1: Rosemont Copper Project Property Location	1-4
Figure 1-2: Plan View Contours of Selected Lerchs-Grossman Pit Shell (Pit Shell 30)	1-16
Figure 1-3: AA’ Section View of Selected Lerchs-Grossman Pit Shell (Pit Shell 30)	1-16
Figure 1-4: Rosemont Mine Plan Site Layout	1-20
Figure 1-5: Plan View of Rosemont Mine Phases	1-21
Figure 1-6: AA' Section View of Rosemont Mine Phases	1-21
Figure 1-7: Rosemont Mine Schedule, Material Movement	1-23
Figure 1-8: Dry Stack Tailings Facility NS Section View, LOM Buttress by Year	1-24
Figure 1-9: Rosemont Annual Copper Production and C1 Cash Costs	1-30
Figure 1-10: NPV8% Sensitivity (100% Basis)	1-32
Figure 4-1: Property Location of Rosemont Project	4-1
Figure 4-2: Rosemont Property Ownership	4-3
Figure 7-1: Laramide Belt and Associated Porphyry Copper Mineralization (Barra ET AL., 2005)	7-1
Figure 7-2: Santa Rita Mountain Geology (Adapted From Drewes ET AL., 2002)	7-3
Figure 7-3: Rosemont Regional Geology	7-4
Figure 7-4: Rosemont Stratigraphic Column	7-5
Figure 7-5: Rosemont Deposit Geologic – 4,000 Foot Level Plan	7-6
Figure 7-6: Rosemont Deposit Geologic – 11,555,050 Vertical Section	7-7
Figure 7-7: Chemostratigraphy Rosemont Deposit Geology	7-8
Figure 7-8: Rosemont Deposit Geological Model Structural Domains 3D View (Looking North)	7-9
Figure 7-9: Mineralization Domains Section 11,555,500 N	7-12
Figure 7-10: Ore Types Cart Model	7-14
Figure 10-1: Rosemont Deposit Drill Hole Locations By Company	10-2
Figure 11-1: Copper Coarse Duplicate Minimum and Maximum Plot	11-10
Figure 11-2: Molybdenum Coarse Duplicate Minimum and Maximum Plot	11-10
Figure 11-3: Silver Coarse Duplicate Minimum and Maximum Plot	11-11
Figure 11-4: Soluble Copper Coarse Duplicate Minimum and Maximum Plot	11-11
Figure 11-5: XP Plots of Check Assay Data, Comparing Primary Laboratory Bureau Veritas to Secondary Laboratory SGS	11-13
Figure 11-6: Copper Coarse Duplicate Minimum and Maximum Plot	11-19
Figure 11-7: Molybdenum Coarse Duplicate Minimum and Maximum Plot	11-19
Figure 11-8: Silver Coarse Duplicate Minimum and Maximum PLot	11-20
Figure 11-9: Soluble Copper Coarse Duplicate Minimum and Maximum Plot	11-20

Page x

	Rosemont Project
	Form 43-101F1 Technical Report

Figure 11-10: XP Plots of Check Assay Data, Comparing Primary Laboratory Bureau Vertias To Secondary Laboratory SGS	11-21
Figure 11-11: Boxplots of SG Measured By Hudbay and Augusta at Inspectorate and Skyline Laboratories, Respectively	11-23
Figure 11-12: Boxplots of Raw Molybdenum Data and Factored Data Reported by Wet and XRF (A- A = Anaconda-ANamax and Y-Axis in Logarithmic Scale)	11-32
Figure 13-1: % COPPER Rougher Flotation Recovery VS % Acid Soluble / Total Copper	13-9
Figure 13-2: LCT Final Concentrate Fluorine Levels	13-18
Figure 14-1: 3D View of Interpreted Lithology Wireframes, Looking Northwest	14-3
Figure 14-2: 3D View of Interpreted Oxidation Wireframes, Looking Northwest	14-5
Figure 14-3: EW Cross Section of the Ore Types Wireframes	14-6
Figure 14-4: Box Plots of Total Copper Assays in Sulfides	14-8
Figure 14-5: Box Plots of Acid Soluble Copper in Sulfides	14-9
Figure 14-6: Box Plots of Molybdenum Assays in Sulfides	14-10
Figure 14-7: Box Plots of Silver Assays in Sulfides	14-11
Figure 14-8: Scatter Plot of Capped Silver and Capped Copper, All Lithology Domains	14-15
Figure 14-9: Scatter Plot of Capped Molybdenum and Capped Copper, All Lithology Domains	14-16
Figure 14-10: QQ Plot of Original Molybdemum Grade Versus Corrected Molybdenum Grade	14-17
Figure 14-11: Box Plot of Gold in Oxide	14-18
Figure 14-12: Box Plot of Gold in Mix	14-19
Figure 14-13: Box Plot of Gold in Hypogene	14-20
Figure 14-14: Contact Profile, Upper and Lower Earp	14-22
Figure 14-15: Histogram, 25-foot Copper Composites, Horquilla Lithology	14-29
Figure 14-16: Downhole Variogram Copper, Lower Group of Lithologies in Sulfides	14-30
Figure 14-17: Correlogram of the Main Structure of Copper, Lower Group of Lithologies in Sulfides	14-31
Figure 14-18: Correlogram of the Nested Structure of Copper, Lower Group of Lithologies in Sulfides	14-31
Figure 14-19: Scatterplot of Total Copper and Specific Gravity	14-37
Figure 14-20: OK, IDW and NN Specific Gravity Distribution	14-38
Figure 14-21: Vertical E-W Section 11,554,900 Showing OK Model and Composites - Copper	14-41
Figure 14-22: Vertical E-W Section 11,554,900 Showing OK Model and Composites – Acid Soluble Copper Grade	14-42
Figure 14-23: Vertical E-W Section 11,554,900 Showing OK Model and Composites - Molybdenum Grade	14-43
Figure 14-24: Vertical E-W Section 11,554,900 Showing OK Model and Composites – Silver Grade	14-44
Figure 14-25: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Copper Swath Plot by Easting	14-52
Figure 14-26: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Copper Swath Plot by Northing	14-53

Page xi

	Rosemont Project
	Form 43-101F1 Technical Report

Figure 14-27: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Copper Swath Plot by Elevation	14-54
Figure 14-28: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Acid Soluble Copper Swath Plot by Easting	14-55
Figure 14-29: Measured and Indicated Blocks above $5.7/Ton NSR within the resource pit shell, acid soluble Copper Swath Plot by Northing	14-56
Figure 14-30: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Acid Soluble Copper Swath Plot by Elevation	14-57
Figure 14-31: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Molybdenum Swath Plot by Easting	14-58
Figure 14-32: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Molybdenum Swath Plot by Northing	14-59
Figure 14-33: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Molybdenum Swath Plot By Elevation	14-60
Figure 14-34: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Silver Swath Plot by Easting	14-61
Figure 14-35: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Silver Swath Plot by Northing	14-62
Figure 14-36: Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell, Silver Swath Plot by Elevation	14-63
Figure 14-37: NN, IDW and OK Copper Grade-Tonnage Curves, All Lithologies in Measured and Indicated Blocks Above $5.7/Ton NSR Within the Resource Pit Shell	14-66
Figure 14-38: Vertical e-w Section 11,554,600 Showing Resource Classification and Drill Holes	14-68
Figure 15-1: Plan View Contours of Selected Lerchs-Grossman Pit Shell	15-5
Figure 15-2: Rosemont Whittle Results, Revenue Factor Sensitivity	15-6
Figure 15-3: Plan View Contours of Selected Lerchs-Grossman Pit Shell	15-7
Figure 15-4: AA’ Section View of Selected Lerchs-Grossman Pit Shell	15-7
Figure 15-5: Plan View of Rosemont Final Pit and Economic Shell 30	15-11
Figure 15-6: Section view BB’ of Rosemont Final Pit and Economic Shell 30	15-11
Figure 16-1: Rosemont Mine Plan Site Layout	16-2
Figure 16-2: Rosemont Geotechnical Sectors	16-3
Figure 16-3: Plan View of Mining Pit Phase 1	16-4
Figure 16-4: Plan View of Mining Pit Phase 2	16-5
Figure 16-5: Plan View of Mining Pit Phase 3	16-6
Figure 16-6: Plan View of Mining Pit Phase 4	16-7
Figure 16-7: Plan View of Mining Pit Phase 5	16-8
Figure 16-8: Plan View of Mining Pit Phase 6 (Ultimate Pit)	16-9
Figure 16-9: Plan View of Rosemont Mine Phases	16-10
Figure 16-10: AA’ Section View of Rosemont Mine Phases	16-10

Page xii

	Rosemont Project
	Form 43-101F1 Technical Report

Figure 16-11: Mine Plan End of Period Pre-Production	16-14
Figure 16-12: Mine Plan End of Period Year 1	16-15
Figure 16-13: Mine Plan End of Period Year 2	16-15
Figure 16-14: Mine Plan End of Period Year 3	16-16
Figure 16-15: Mine Plan End of Period Year 4	16-16
Figure 16-16: Mine Plan End of Period Year 5	16-17
Figure 16-17: Mine Plan End of Period Year 6	16-17
Figure 16-18: Mine Plan End of Period Year 7	16-18
Figure 16-19: Mine Plan End of Period Year 8	16-18
Figure 16-20: Mine Plan End of Period Year 9	16-19
Figure 16-21: Mine Plan End of Period Year 10	16-19
Figure 16-22: Mine Plan End of Period Year 11	16-20
Figure 16-23: Mine Plan End of Period Year 12	16-20
Figure 16-24: Mine Plan End of Period Year 13	16-21
Figure 16-25: Mine Plan End of Period Year 14	16-21
Figure 16-26: Mine Plan End of Period Year 15	16-22
Figure 16-27: Mine Plan End of Period Year 16	16-22
Figure 16-28: Mine Plan End of Period Year 17	16-23
Figure 16-29: Mine Plan End of Period Year 18	16-23
Figure 16-30: Mine Plan End of Period Year 19	16-24
Figure 16-31: Mine Plan, Final Topography	16-24
Figure 16-32: Rosemont Mine Schedule, Material Movement	16-26
Figure 16-33: Rosemont Mine Schedule, Mill Feed Ore by Lithology	16-26
Figure 16-34: Rosemont Mine Schedule, Mill Feed Ore by Ore Type	16-27
Figure 16-35: DSTF NS Section View, LOM Buttress by Year	16-29
Figure 16-36: Rosemont Geotechnical Sectors	16-37
Figure 16-37: Section AA’ Showing RQD Values in Final Rosemont Pit	16-37
Figure 16-38: Rosemont Final Pit, Lithology in Final Wall	16-38
Figure 16-39: Section BB’ showing Hard Values in Final Rosemont Pit	16-38
Figure 16-40: Section AA’ Showing Hard Values in Final Rosemont Pit	16-40
Figure 17-1: Overall View of Process Plant Looking North	17-2
Figure 17-2: Process Plant Process Flow Diagram	17-5
Figure 17-3: Primary Crusher	17-7
Figure 17-4: Stockpile Fabric Cover	17-8
Figure 17-5: Grinding Building From Stockpile Looking North (Roof and Walls removed)	17-9

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Figure 17-6: Pebble Crushing Looking South From the Grinding Area	17-10
Figure 17-7: Copper Flotation	17-11
Figure 17-8: Copper Regrind Area Looking West	17-12
Figure 17-9: Molybdenum Plant Looking West (Roof and walls Removed)	17-14
Figure 17-10: Copper Concentrate Filtration, Storage and Load out Looking South	17-16
Figure 17-11: Tailing Thickeners Looking East	17-17
Figure 17-12: Tailings Filter Plant (roof and walls removed)	17-18
Figure 17-13: Tailings Shiftable Conveyor/Mobile Tripper	17-19
Figure 17-14: Reagents Area	17-20
Figure 18-1: Plant and Access Road	18-1
Figure 18-2: CEC Approved Utility Corridor for 138Kv Transmission Line	18-3
Figure 18-3: Utility Corridor for Water Line	18-4
Figure 22-1: Rosemont Annual Copper Production and C1 Cash Costs	22-4
Figure 22-2: NPV8% Sensitivity (100% Basis)	22-11
Figure 24-1: Overall Construction Management & HSEC (EPCM)	24-1

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LIST OF APPENDICES

TITLE	PAGE
A1-1 Land Tenure	30-1
A1-2 Rosemont Project Patented Claims	30-2
A1-3 Rosemont Project Unpatented Claims	30-5
A1-4 Rosemont Project Fee Owned (Associated) Lands	30-27
A2-1 Permits and Authorizations	30-28

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1        SUMMARY

The information that follows provides an executive summary of important information contained in this Technical Report.

1.1        Introduction

The author has prepared this Technical Report for Hudbay Minerals Inc. (“Hudbay”) with respect to its Rosemont Project (the “Project”), located in Arizona (the “Property”), issued and effective as of March 30, 2017. The purpose of this Report is to present Hudbay’s estimate of the mineral reserves and mineral resources for the Project based on the current mine plan, the current state of metallurgical testing, operating cost and capital cost estimates.

Hudbay is a Canadian integrated mining company with assets in North and South America principally focused on the discovery, production and marketing of base and precious metals. Hudbay’s objective is to maximize shareholder value through efficient operations, organic growth and accretive acquisitions, while maintaining its financial strength.

Hudbay completed the acquisition of the Project on September 23, 2014 through its acquisition of all issued and outstanding common shares of Augusta Resource Corporation (“Augusta”) pursuant to the take-over bid, which expired July 29, 2014.

Hudbay owns a 92.05% interest in the 132 patented claims and 1,064 unpatented claims that comprise the Project, all of which are duly registered in the name of Hudbay’s wholly-owned subsidiary, Rosemont Copper Company¹; Rosemont Copper Company also has the required surface rights to develop the Project. This Technical Report represents the first technical report filed by Hudbay since its acquisition of Augusta and also updates and supersedes Augusta’s Updated Feasibility Study dated August 28, 2012, prepared by M3 Engineering and Technology Corporation.

This Technical Report provides current estimates of the mineral reserves and mineral resources at the Project and describes the latest resource model, mine plan and the current state of the permitting process, metallurgical testing, operating cost and capital cost estimates. The information presented in this Technical Report relating to the Rosemont deposit, including the estimates of mineral reserves and resources therein, is the result of “feasibility study” level work conducted partly by external contractors and partly internally by Hudbay’s personnel under the overall supervision of, Cashel Meagher, the Qualified Person (the “QP”).

^{____________________________________________________
1} Hudbay’s ownership in the Project is subject to an earn-in agreement and joint venture agreement dated September 16, 2010 between Rosemont Copper Company and United Copper & Moly LLC, pursuant to which UCM has earned a 7.95% interest in the project and may earn up to a 20% joint venture interest.

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	Form 43-101F1 Technical Report

This Technical Report conforms with the 2014 CIM Definition Standards – for Mineral Resources and Mineral Reserves and the requirements in Form 43-101F1 of National Instrument (NI) 43-101, Standards of Disclosure for Mineral Projects.

The QP and Principal Author who supervised the preparation of this Technical Report is Cashel Meagher, P.Geo., Senior Vice President and Chief Operating Officer for Hudbay. Mr. Meagher last visited the property on April 21, 2016 and numerous times prior to this date. The personal site inspections were conducted as part of the mineral resource estimation and technical report process, to become familiar with conditions on the Property and the Project, to observe the geology and mineralization and verify the work completed on the Property. Mr. Meagher has reviewed and approved the 3D block model and determination of mineral resources and mineral reserves of the Project.

As Hudbay is a “producing issuer”, as defined in NI 43-101, this Technical Report is not required to be prepared by or under the supervision of an independent QP.

1.2        Property Description and Location

The Project is located within the historic Helvetia-Rosemont Mining District on the eastern flanks of the Santa Rita Mountain Range, approximately 30 miles southeast of Tucson in Pima County, Arizona. The property consists of a comprehensive land package that includes patented and unpatented mining claims, fee land and grazing leases that cover most of the old mining district. The lands are under a combination of private ownership by Rosemont Copper and Federal ownership. The lands occur within Townships 18 and 19 South, Ranges 15 and 16 East, Gila & Salt River Meridian. The Project’s geographical coordinates are approximately 31º 50’N and 110º 45’W.

Hudbay’s ownership in the Project is subject to an earn-in agreement and joint venture agreement dated September 16, 2010 between Rosemont Copper Company and United Copper & Moly LLC (‘‘UCM’’), pursuant to which UCM has earned a 7.95% interest in the Project and may earn up to a 20% joint venture interest.

Hudbay has all of the surface and mineral rights required to conduct the open pit mining operation, processing and concentrating facilities, storage of tailings, and disposal of waste rock as documented in this Technical Report. The core of the Project mineral resource is contained within the 132 patented mining claims that in total encompass an area of approximately 2,000 acres. Surrounding the patented claims is a contiguous package of 1,064 unpatented mining claims with an aggregate area of more than 16,000 acres.

There is a 3% Net Smelter Return (“NSR”) royalty on all 132 patented claims, 603 of the unpatented claims, and one parcel of fee owned associated land. Pursuant to a precious metals stream agreement with Silver Wheaton Corp. (“Silver Wheaton”) entered into on February 11, 2010, as amended and restated on February 15, 2011, Hudbay will receive deposit payments of $230 million against delivery of 100% of the payable gold and silver from the Project. The deposit will be payable upon the satisfaction of certain conditions precedent, including the receipt of permits for the Project and the commencement of construction. In addition to deposit payments, as gold and silver is delivered to Silver Wheaton, Hudbay will receive cash payments equal to lesser of (i) the market price and (ii) $450 per ounce (for gold) and $3.90 per ounce (for silver), subject to one percent annual escalation after three years. Approximately 50% of the copper concentrate has been contracted under existing commitments that are on benchmark-based terms.

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1.3        Accessibility, Climate, Local Resources, Infrastructure and Physiography

Existing graded dirt roads connect the Project site with State Route 83, which provides easy access to the Project for the communities of Tucson and Benson to the north, and to Sierra Vista, Sonoita, Patagonia and Nogales to the south. The city of Tucson, Arizona provides the nearest major railroad and air transport services to support the Project.

The Project site is located immediately adjacent and west of Arizona State Route 83, approximately eleven miles south of Interstate 10 (I-10). This system of state and interstate highways allows convenient access to the site for all major truck deliveries. The majority of the labour and supplies for construction and operations come from the surrounding areas in Pima, Maricopa, Cochise, and Santa Cruz counties.

The southern Arizona climate is typical of a semi-arid continental desert with hot summers and temperate winters. However, higher elevations in the Project area (4,550 to 5,350 feet AMSL) result in a milder climate than at lower elevations across the region. Summer daily high temperatures are above 90°F with significant cooling at night. Winter in the Project area is typically drier with mild daytime temperatures and overnight temperatures that are typically above freezing.

The average annual precipitation in the Project area is estimated between 16 and 18 inches with more than half of the annual precipitation occurring during the monsoon season from July through September. Rainfall has minor effects on a mining operation, which is considered to be 365 days per year.

The Project is located within the northern portion of the Santa Rita Mountains that form the western edge of the Mexican Highland section of the Basin and Range Physiographic Province characterized by high mountain ranges adjacent to alluvial filled basins. Vegetation in the Project area reflects the climate with the lower slopes of the Santa Rita Mountains dominated by mesquite and grasses while the higher elevations, receiving greater rainfall, support an open cover of oak, pine, juniper and cypress trees.

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FIGURE 1-1: ROSEMONT COPPER PROJECT PROPERTY LOCATION

1.4        History

The first recorded mining activity in the Helvetia-Rosemont mining district occurred in 1875 and the mining district was officially established in 1878. Production from mines on both sides of the Santa Rita ridgeline supported the construction and operation of two smelters. Copper production from the district ceased in 1951 after production of about 227,300 tons of ore.

By the late 1950s, the Banner Mining Company (“Banner”) had acquired most of the claims in the area and had drilled the discovery hole into the Rosemont deposit. In 1963, Anaconda Mining Co. acquired options to lease the Banner holdings. Over the next ten years, they carried out an extensive drilling program on both sides of the ridgeline.

In 1973, the Anaconda Mining Co. and Amax Inc. formed a 50/50 partnership to form the Anamax Mining Co. (“Anamax”) and in 1985, Anamax ceased operations and liquidated their assets.

ASARCO Inc. (“Asarco”) purchased the patented and unpatented mining claims from Anamax’s real estate interests in August 1988 and renewed exploration and engineering studies. Asarco expanded the core of the mineral deposit in 1995 by patenting 347 acres, the last of the available claims for the orebody. In 1999, Grupo Mexico acquired the Helvetia-Rosemont property through a merger with Asarco and in 2004 Grupo Mexico sold the property to a Tucson real estate developer.

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In April 2005, Augusta purchased the Property from Triangle Ventures LLC and initiated a series of extensive drill programs on the property. A Technical Report issued by Augusta in 2012 estimated mineral reserves of 667.2 million tons at an average grade of 0.44% copper, 0.015% molybdenum and 0.12 ounces per ton of silver based on $4.90 per ton net smelter return cut-off using metal prices of 2.50/lb. copper, $15.00/lb. molybdenum and $20.00/oz. silver.

Note that Hudbay has treated Augusta’s publicly disclosed estimated mineral reserves and resources as a historical estimate under NI 43-101 and not as current mineral reserves and resources, as a qualified person has not performed sufficient work for Hudbay to classify the 2012 estimate for the Project’s mineral reserves or resources as current mineral reserves or mineral resources.

Following its acquisition of Augusta, Hudbay acquired all of the issued and outstanding common shares of Augusta pursuant to a take-over bid, which expired July 29, 2014, and a subsequent acquisition transaction, which closed on September 23, 2014. Hudbay’s ownership in the Project is subject to an earn-in agreement with UCM, pursuant to which UCM has earned a 7.95% interest in the Project and may earn up to a 20% interest. A joint venture agreement between Hudbay’s subsidiary, Rosemont Copper Company, and UCM governs the parties’ respective rights and obligations with respect to the Project.

Hudbay completed a 43-hole, 92,909 feet (28,319 m) drill program from September to December 2014 and a 46-hole, 75,164 feet (22,910 m) drill program from August to November 2015 in further efforts to better understand the geological setting and mineralization of the deposit and to collect additional metallurgical and geotechnical information.

1.5        Geological Setting and Mineralization

The Laramide belt is a major porphyry province that extends for approximately 621 miles (1,000 km) from Arizona to Sinaloa, Mexico. It hosts a number of world-class deposits including the Rosemont deposit. The northern block of the Santa Rita Mountains, where the Rosemont deposit lies, is dominated by Precambrian granite, with some dismembered slices of Paleozoic and Mesozoic sediments on the eastern and northern sides.

Paleozoic sedimentary carbonate units are the predominant host rocks for the copper mineralization. Structurally overlying these predominantly carbonate units at Rosemont are Mesozoic clastic units, including conglomerates, sandstones, and siltstones. These clastic upper sequences have andesitic flows and host mineralization. Quartz monzonite and quartz latite sill-shaped porphyries intruded both sequences and are associated with the porphyry/skarn mineralization.

Post-mineral features partially delimit the defined resource, dividing the deposit into major structural blocks with contrasting intensities and types of mineralization. The north-trending, steeply-dipping Backbone Fault juxtaposes marginally mineralized Precambrian granodiorite and Lower Paleozoic quartzite and limestone to the west against a block of younger, well-mineralized Paleozoic limestone units to the east.

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The Rosemont deposit consists of copper-molybdenum-silver-gold mineralization primarily hosted in skarn that formed in the Paleozoic rocks as a result of the intrusion of quartz latite to quartz monzonite porphyry intrusions. Bornite-chalcopyrite-molybdenite mineralization occurs as veinlets and disseminations in the skarn.

Three mineralization domains (oxide, mixed and sulfide) were defined based on the soluble to total copper ratio (ASCu/TCu) collected in the Augusta (2005 to 2012) and Hudbay (2014 and 2015) drilling programs. The oxidation and mixed mineralization occurs mainly above a low angle fault defining the contact between the Palozoic and Mesozoic rocks as chrysocolla, copper carbonates and supergene chalcocite.

1.6        Deposit Types

As mentioned above, the Rosemont deposit consists of copper-molybdenum-silver-gold mineralization primarily hosted in skarn, genetically, it is a style of porphyry copper deposit, although intrusive rocks are volumetrically minor within the resource area. The skarns are formed as the result of thermal and metasomatic alteration of Paleozoic carbonate and to a lesser extent Mesozoic clastic rocks. Near surface weathering has resulted in the oxidation of the sulfides in the overlying Mesozoic units however, oxidation also occurs in the underlying Paleozoic carbonates.

1.7        Exploration

Prospecting began in the Rosemont and Helvetia Mining Districts in the mid-1800s and by 1875 copper production was first recorded, which continued sporadically until 1951. By the late 1950s, exploration drilling had discovered the Rosemont deposit. A succession of major mining companies subsequently conducted exploratory drilling of the Rosemont deposit and the nearby Broadtop Butte, Peach Elgin and Copper World mineralized areas.

Augusta acquired the Rosemont property in 2005 and performed infill drilling of the Rosemont deposit along with exploration geophysical surveys. A Titan 24 induced polarization/resistivity (DCIP) survey over the Rosemont deposit, performed in 2011, discovered significant chargeability anomalies, which were partially tested. These anomalies appear to define mineralization and certain unmineralized lithologic units. A regional scale airborne magnetics survey was also completed in 2008.

Two infill drilling campaigns were completed by Hudbay in and beneath the Rosemont deposit in the fall of both 2014 and 2015. In addition to chemical assaying, magnetic susceptibility and conductivity measurements were taken. A single test-line of DCIP data was collected over the Rosemont deposit using the DIAS Geophysical in April 2015 for comparison to the previously completed Titan 24 survey.

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Hudbay analyzed all samples of the 2014 and 2015 drilling programs with ICP multi-element geochemistry. This new geochemical data set was used to classify rocks according to chemical indexes in a ternary diagram defined by siliciclatic, limestone and dolomitic vertices. The lithogeochemical groups honour the deposit stratigraphy and geochemical attributes and proved to be a useful tool for geological modeling and vectoring.

A mapping and geochemical sampling program was completed in the latter half of 2015 on the Rosemont property to reassess the interpretation of the regional geology and deposit setting. This was followed by a structural interpretation using both surface and drill core measurements to aid in the geotechnical evaluation of the Project.

1.8        Drilling

Extensive drilling has been conducted at the Rosemont deposit by several successive property owners. The most recent drilling was done by Hudbay, with prior drilling campaigns completed by Banner, Anaconda Mining Co., Anamax and Asarco and Augusta. Table 1-1 summarizes the drill holes used to estimate the current mineral resource estimate, with regional exploration holes excluded. The drillholes are approximately 200 feet apart over the core of the deposit.

TABLE 1-1: ROSEMONT DEPOSIT DRILLING SUMMARY

Company	Time Period	Drill Holes
		Number	Feet
Banner Mining	1950s to 1963	3	4,300
Anaconda Mining	1963 to 1973	113	136,838
Anamax	1973 to 1986	52	54,350
ASARCO	1988 to 2004	11	14,695
Augusta	2005 to 2012	87	132,525
Hudbay	2014 to 2015	90	168,286
Total		355	510,780

The recent Hudbay drilling went deeper by approximately 300 feet on average than the Augusta drilling and almost twice as deep as the Anaconda and Anamax drilling program. This most recent drilling has helped to confirm the size and quality of the deposit estimated by previous owners and to also establish its continuation at depth resulting in an improved definition of the optimum open pit design.

1.9        Sample Preparation, Analyses, and Security

During the Hudbay 2014 and 2015 drill programs, the samples were transported to the Inspectorate America Corporation (“Inspectorate”) preparation facility at Sparks, Nevada, USA. Once the samples were pulverized, a 150 g subsample pulp was collected and air-freighted to Bureau Veritas Commodities Canada Ltd., in Vancouver, Canada, for analysis. A total of 18,361 drill core samples in 2014 and 14,868 samples in 2015 were analyzed for copper, molybdenum and silver, through a multi-element (45 elements) determination by Inductively Coupled Plasma Mass Spectrometry after 4-acid digestion. A total of 1,677 samples in 89 drill holes were collected for specific gravity determinations by a standard water displacement method at the Inspectorate preparation facility.

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As part of Hudbay’s quality control and quality assurance (“QA/QC”) program, QA/QC samples were systematically introduced in the sample stream to assess adequate sub-sampling procedures, potential cross-contamination, precision, and accuracy. A total of 1,000 representative pulp samples (5.4%) from 2014 drilling and 742 representative pulp samples (5.0%) from 2015 drilling were selected and re-analyzed at the SGS Canada Inc., laboratory in Vancouver.

The core samples from the Augusta drilling programs from 2005 to 2012 were transported to Skyline Assayers and Laboratories (Skyline), in Tucson, Arizona, USA for preparation and analysis. In total, 21,197 samples were analyzed for total copper and 16,619 samples for molybdenum. Total copper and molybdenum were dissolved using a hot 3-acid digestion at 482°F and subsequently analyzed by AAS and ICP-OES, respectively. The lower detection limits for molybdenum are high relative to the average molybdenum grade of the Rosemont deposit. Silver was determined in 15,334 samples, which were digested using an aqua regia leach in 0.25 g subsample pulp and analyzed by AAS. A total of 391 drill core samples across the Rosemont deposit were measured for specific gravity at Skyline.

Augusta conducted its own internal QA/QC program to independently evaluate the quality of the assays reported by Skyline. Standards and blanks were systematically inserted in the sample stream. Duplicates were not periodically inserted.

Prior to Hudbay and Augusta, significant diamond drilling, drill core sampling, and assaying programs were executed by several property owners. Records are not available that detail the sampling and security protocols used by these property owners. There are no available QA/QC records for sample preparation and assaying methodologies for Banner, Anaconda, and Anamax. Copper, molybdenum, silver, and soluble copper were analyzed by Anaconda and Anamax at their in-house laboratories. Silver was regularly analyzed by Anamax, but not commonly assayed by Banner and Anaconda. Asarco assayed drill core samples for total copper, molybdenum, and acid soluble copper (“ASCu”) at Skyline laboratory.

1.10        Data Verification

Hudbay built an entirely new drill hole database from all pre-Hudbay drilling and assaying information. Orix Geoscience Inc. was employed to digitally enter collar, downhole surveys and assay information from scanned drill logs and assay certificates for all holes drilled prior to ownership of the property by Augusta.

The infill drilling conducted by Hudbay and Augusta together with re-assaying of historical holes have closely replicated previous drilling campaign results confirming that the historical data can be used with a sufficient level of confidence for resource and reserve estimation. A bias was identified in the historical molybdenum assays and the data was corrected.

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The author’s opinion is that the data verification is adequate for the purposes used in the Technical Report.

1.11        Mineral Processing and Metallurgical Testing

The earliest reported testwork on Rosemont ores comprising preliminary grinding and flotation tests was completed by Anamax in 1974. This early work was followed by a larger testwork campaign by Augusta in 2006 and 2007 to support the preparation of a feasibility study and technical report. Further testwork was then completed by Augusta between 2008 and 2012 to support engineering design and updates to the original technical report.

Historical metallurgical testwork programs were undertaken by Mountain State R&D International (MSRDI), SGS and G&T Metallurgical Services, with dewatering and rheology testing undertaken by Pocock, Outotec and FLSmidth. In 2014, Hudbay engaged XPS Consulting & Testwork Services (XPS) to undertake mineral characterization and metallurgical testwork. Base Met Laboratory (“BML”) was engaged in late 2015 to provide confirmation testwork of the XPS testwork and additional process optimization.

The testwork investigated key geo-metallurgical variables such as copper oxide content, swelling clays, magnesium clays and ore hardness. The copper oxide content, as measured by the acid soluble procedure, is a good indicator of the recoverable copper content of the ore. Clay content varies considerably in type and quantity throughout the oxide, transition and sulfide mineralization. Ore hardness varies from soft to very hard; testing results, together with geomet proxy modelling, were utilized to calculate hardness in the resource model.

Production period composites, together with the geo-metallurgical samples, underwent flotation testing for process engineering design as well as a recovery estimator for mine planning and the financial model.

Through the course of all the mineral processing and metallurgical testing, no deleterious elements were found to have a negative impact on plant performance or on the marketable value of the copper and molybdenum concentrates to be produced at the Project.

Based on the body of testwork that exists, including both the historical testwork, and the testing programs completed by Hudbay since the acquisition of the Project, forecasts of recovery, concentrate grade and quality, as well as characteristics of the resultant tailing product have been developed. The following summarizes long range mine plan (“LOM”) average recoveries expected.

Concentrate	Average LOM recoveries
Copper (Cu)	80.4%
Molybdenum (Mo)	53.4%

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Concentrate	Average LOM recoveries
Silver (Ag)	74.4%
Gold (Au)	65.1%

1.12        Mineral Resource Estimate

Hudbay prepared a 3D block model of the Rosemont deposit. The 3D block model and determination of the mineral resources were reviewed and approved by Cashel Meagher, P.Geo., Senior Vice President and Chief Operating Officer for Hudbay and QP of this Technical Report.

1.12.1        Wireframe Models and Mineralization

The Rosemont deposit trends approximately along an azimuth of N020° with a general dip of 50° to the east. The Backbone Fault forms the footwall contact along the entire length of the deposit. Geologically, Rosemont is a skarn deposit. The deposit is continuous along a strike length of 4,000 feet in a north-south direction, 3,000 feet in an east-west direction and goes to a maximum vertical depth of approximately 2,500 feet.

Three sets of structures were recognized, a north-northeast trending set, an east-west trending set and a gently east dipping set. The structures locally offset mineralization but some also appear to control mineralization, especially the oxidation. Wireframes were constructed for each lithological unit, oxidation level and fault structure.

1.12.2        Exploratory Data Analysis

A statistical analysis (basic statistics, histograms, box plots, contact plots, regressions) for total copper, acid-soluble copper, molybdenum, silver and sample length was performed on all the assays and composites to ensure mineralized domains were understood and that bias was not introduced during the data preparation stage.

1.12.3        Variography

Experimental variograms were calculated for total copper, acid-soluble copper, molybdenum and silver from the 25-foot capped composites. Directional and down-the-hole correlograms were fitted. The down-the-hole models were used to select the nugget used in subsequent modelling of directional correlograms. The total copper variograms show very low to moderate nugget effects with ranges of correlation generally varying between 340 to 2,000 ft, with the majority of the variability occurring within the first 200 to 300 feet in all directions.

1.12.4        Estimation and Interpolation Methods

The block model consists of regular blocks (50 feet along strike x 50 feet across strike x 50 feet vertically). The block size was chosen such that geological contacts are reasonably well reflected and to support a large-scale open pit mining scenario.

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The interpolation plan was completed on the uncapped and capped composites via ordinary kriging (“OK”) interpolation method using three passes with increasing search distances.

The first interpolation pass was restricted to a minimum of nine composites, a maximum of 12 composites (with a maximum of three composites per hole) and quadrant declustering. The second interpolation pass was restricted to a minimum of six composites, a maximum of 12 composites (with a maximum of three composites per hole) and quadrant declustering. Finally, the third interpolation pass was restricted to a minimum of four composites, a maximum of 12 composites (with a maximum of three composites per hole) without quadrant declustering.

1.12.5        Block Model Validation

The Rosemont block model was validated to ensure appropriate honouring of the input data and to verify the absence of bias by the following methods:

• Visual inspection of the OK block model grades in plan and section views in comparison to composites grades
• Assessment of the quantity of metal removed via the grade capping methodology
• Comparison between the different interpolation methods, including nearest neighbour, inverse distance squared and OK
• Swath plot comparisons of the estimation methods
• Review of block model OK quality control parameters
• Review of grade tonnage curves and statistics for each estimation method
• Third party and internal peer review

1.12.6        Classification of Mineral Resource

The resource category classification relies on the relative difference between the kriged grade and the composites grades, the number of composite used, the closest and farthest distance between the composites used and the centre of the blocks.

A smoothing algorithm was applied to remove isolated blocks of measured category blocks within areas of mostly indicated category or isolated indicated blocks within areas of mostly measured category blocks. Proportions of measured and indicated category blocks were not changed significantly by this process.

1.12.7        Reasonable Prospects of Economic Extraction

The component of the mineralization within the block model that meets the requirements for reasonable prospects of economic extraction was based on the application of a Lerchs-Grossman (“LG”) cone pit algorithm. The mineral resources are therefore contained within a computer-generated open pit geometry.

The following assumptions were applied to the determination of the mineral resources:

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• Economic benefit was applied to measured, indicated and inferred classified material within the resource cone.
• The LG optimization was conducted on a NSR value that reflects for each block in the resource model, the copper (“Cu”), molybdenum (“Mo”) and silver (“Ag”) grades, mill recoveries, contained metal in concentrate, deductions and payable metal values, metal prices, freight costs, smelting and refining charges and royalty charges.
• The pit shell selected to report mineral resources was based on a revenue factor of 1.0 (break-even value) using the following metal prices: $3.15/lb. copper, $11.00/lb. molybdenum, and $18.00/oz. silver.
• A constant 45-degree pit slope was used for the resource estimate.
• No haulage increment or bench discounting was applied to the resource estimate.

1.12.8        Mineral Resource Statement inclusive of the Mineral Reserve

Mineral resources for the Rosemont deposit were classified under the 2014 CIM Definition Standards for Mineral Resources and Mineral Reserves by application of a NSR calculation that reflects the combined benefit of producing copper, molybdenum and silver in addition to mine operating, processing and off-site costs. The cut-off used for resource reporting is based on a reasonable estimate of the investment required to construct and sustain a viable operating complex.

The mineral resources, classified as Measured, Indicated and Inferred and prior to any conversion to mineral reserves, inclusive of the portion of the mineral resources that was converted to mineral reserves, are summarized in Table 1-2.

Mineral resources that are not mineral reserves do not have demonstrated economic viability. Due to the uncertainty that may be associated with Inferred mineral resources it cannot be assumed that all or any part of Inferred resources will be upgraded to an Indicated or Measured Resource.

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TABLE 1-2: RESOURCE BY CATEGORY, MINERALIZED ZONE AND NSR CUT-OFF
^{(1)(2)(3)(4)(5)(6)(7)(8)(9)(10)}

Notes:

1.	The above mineral resources include mineral reserves.
2.	Domains were modelled in 3D to separate mineralized rock types from surrounding waste rock. The domains were based on core logging, structural and geochemical data.
3.	Raw drill hole assays were composited to 25-foot lengths broken at lithology boundaries.
4.	Capping of high grades was considered necessary and was completed for each domain on assays prior to compositing.
5.	Block grades for copper, molybdenum and silver were estimated from the composites using OK interpolation into 50 ft x 50 ft x 50 ft blocks coded by domain.
6.	Tonnage factors were interpolated by lithology and mineralized zone. Tonnage factors are based on 2,066 measurements collected by Hudbay and previous operators.
7.	Blocks were classified as Measured, Indicated or Inferred in accordance with CIM Definition Standards 2014.
8.	Mineral resources are constrained within a computer generated pit using the LG algorithm. Metal prices of $3.15/lb copper, $11.00/lb molybdenum and $18.00/troy oz silver. Metallurgical recoveries of 85% copper, 60% molybdenum and 75% silver were applied to sulfide material. Metallurgical recoveries of 40% copper, 30% molybdenum and 40% silver were applied to mixed material. A metallurgical recovery of 65% for copper was applied to oxide material. NSR was calculated for every model block and is an estimate of recovered economic value of copper, molybdenum, and silver combined. Cut-off grades were set in terms of NSR based on current estimates of process recoveries, total process and G&A operating costs of $5.70/ton for oxide, mixed and sulfide material.
9.	The oxide resource will be processed in the mill via flotation
10.	Totals may not add up correctly due to rounding.

The reporting of the mineral resource by NSR within the LG pit shell reflects the combined benefit of producing copper, molybdenum and silver as per the following equations based on mineralized type, in addition to mine operating and processing costs:

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The copper equivalency is calculated, using metal contributions, for each block using the following formula:

CuEq = Copper + (Contribution of Molybdenum) + (Contribution of Silver)

Since molybdenum and silver are not considered in oxide material, the copper equivalency value equals the copper value.

1.12.9        Comparison with the 2012 Resource Estimate

A review and comparison of 2017 Hudbay mineral resource and 2012 Augusta Resource mineral resource was completed. The results of Measured, Indicated and Inferred are summarized in Table 1-3 and Table 1-4.

TABLE 1-3: MEASURED AND INDICATED, COMPARISON TO 2012 AUGUSTA ESTIMATE

TABLE 1-4: INFERRED, COMPARISON TO 2012 AUGUSTA ESTIMATE

The 2017 measured and indicated resource estimates constitute a 29% increase in tonnage with copper grades 8% lower to those estimated in 2012 by Augusta. Also, the molybdenum grade is 17% lower than reported in 2012 for the sulfide mineralized material while the 2016 oxide tonnage and grade have more than doubled. These differences result mainly from the reinterpretation of the oxide blanket surface. Molybdenum grades are also lower as a result of factoring historical molybdenum assays. The reduction of tonnage in the Inferred category is related to the infill drilling completed by Hudbay in 2014 and 2015 which resulted in the reclassification of the 2012 inferred and indicated resources to indicated and measured categories in 2017.

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1.13        Mineral Reserves Estimate

The Mineral Reserves estimate for the Project are based in a LOM which uses the block model compiled under Section 14, Mineral Resource Estimates, with an economic value calculation per block (NSR in $/ton) and mining, processing, and engineering detail parameters. The Mineral Reserves estimate for the Project has been prepared by Hudbay senior mine engineer experts under the supervision of the QP. The mineral reserve economics are described in Section 22, Economic Analysis, of this Technical Report.

This Mineral Reserves estimate has been determined and reported in accordance with NI 43-101 and the classifications adopted by CIM Council in 2014. NI 43-101 defines a Mineral Reserve as “the economically mineable part of measured and indicated mineral resources”.

Mine design and reserves estimation for the Rosemont pit use the NSR block model, which consists of an NSR value calculation for each block in the block model, taking into account grade mill recoveries (Cu, Mo and Ag), contained metal in concentrate, deductions and payable metal values, metal prices, freight costs, smelting and refining charges and royalty charges. These parameters were applied to the block model to form the basis of the reserve estimate.

The LG analyses were conducted for the purpose of reporting reserves. The selected pit shell corresponds to a revenue factor of 0.8 (Pit shell 30), which represents metal prices 20% lower than the base case (revenue factor of 1.0 - pit 40). It was selected as the basis for the ultimate pit design, and is approximately 10% smaller than the base economic pit. This pit has better economic indicators in comparison with other pits in terms of free discounted cash flow and total revenue, stripping ratio and capital costs. All LG analyses were restricted to prevent the pit shells from crossing the topographic ridge immediately west of the deposit. This was done due to a permit commitment.

The selected LG pit shell 30 is shown in plan view in Figure 1-2 and in cross-section in Figure 1-3.

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FIGURE 1-2: PLAN VIEW CONTOURS OF SELECTED LERCHS-GROSSMAN PIT SHELL (PIT SHELL 30)

FIGURE 1-3: AA’ SECTION VIEW OF SELECTED LERCHS-GROSSMAN PIT SHELL (PIT SHELL 30)

The Rosemont mineral reserve estimate is based on measured and indicated resources. Therefore, the potential exists for Inferred Mineral Resources within the ultimate pit to be included and reported as waste, as they currently do not meet the economic and mining requirements to be categorized as

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Mineral Reserves. It cannot be assumed that all or any part of Inferred Mineral Resources will ever be upgraded to a higher category.

The selective mining unit (“SMU”) dimension in the resource block model is 50x50x50 ft. The interpolated metal grade is averaged for the entire block. When the Project commences operations, ore feed will be delineated by implementing a detailed blasthole sampling program. Drill blast patterns will be smaller, 30 ft to 30 ft, than the resource block dimensions, thereby providing better definition than from the resource model. This new definition will be provided by a new block model built by assays from blasthole projects, dynamic or short-range block model, which is a common practice in Hudbay operations.

1.13.1        Mineral Resources and Mineral Reserves Statement

Proven and probable mineral reserves within the designed final pit total are 592 million tons grading 0.45% Cu, 0.012% Mo and 0.13 oz Ag/ton. There are 1.25 billion tons of waste material (including pre-stripping material), resulting in a stripping ratio of 2.1:1 (tons of waste per ton of ore). Total material in the pit is 1.84 billion tons. Contained metal in Proven and Probable mineral reserves is estimated at 5.30 billion pounds of copper, 142 million pounds of molybdenum and 79 million ounces of silver. Proven and Probable mineral reserves for the Rosemont deposit are summarized in Table 1-5.

TABLE 1-5: PROVEN AND PROBABLE MINERAL RESERVES IN ROSEMONT FINAL PIT

	Short Tons	TCu %	SCu %	ASCu %	Mo %	Ag opt	NSR $/ton	CuEq %	Tonnes	Ag (g/t)
Proven	469,708,117	0.48	0.43	0.05	0.012	0.14	22.11	0.56	426,112,017	4.96
Probable	122,324,813	0.31	0.28	0.03	0.010	0.09	14.66	0.38	110,971,199	3.09
Total	592,032,930	0.45	0.40	0.05	0.012	0.13	20.57	0.53	537,083,216	4.58

Notes:

1.	TCu % corresponds to the total copper grade.
2.	SCu % grade corresponds to the sulfide copper in the Ore. As per formula SCU = TCU – ASCu
3.	ASCu % grade corresponds to the soluble copper.
4.	CuEq% is calculated based on metal prices of $3.15/lb Cu, $11.00/lb Mo and $18.00/oz Ag.

Economics of the Mineral Reserves were demonstrated by the mine plan’s financial analysis, documented in Section 22 of this Technical Report, which confirmed a 15.6 percent after-tax internal rate of return, based on a copper price of $3.00/lb, silver price of $18.00/oz and molybdenum price of $11.00/lb.

Table 1-6 presents the mineral resource estimates exclusive of the Mineral Reserve estimate, i.e. the mineral resources located inside the resource pit shell but outside of the reserve pit design. The mineral reserve estimate represents the portion of the mineral resource estimates with potential for economic extraction after the current mineral reserves estimate has been mined and processed.

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TABLE 1-6: ROSEMONT MINERAL EXCLUSIVE RESOURCE ESTIMATES ^{(1)(2)(3)(4)(5)(6)(7)(8)(9)}

	Short Tons	NSR Cut-off	CuEq %	TCu %	Mo (%)	Ag opt	Tonnes	Ag (g/t)
Measured
Oxide	54,000,000	> = $5.70	0.41	0.41			49,000,000
Mixed	5,000,000	> = $5.70	0.45	0.41	0.01	0.047	4,500,000	1.63
Hypogene	118,700,000	> = $5.70	0.44	0.36	0.01	0.117	107,700,000	4.01
Summary	177,700,000		0.43	0.38	0.01	0.079	161,200,000	2.72
Indicated
Oxide	18,600,000	> = $5.70	0.27	0.27			16,900,000
Mixed	2,600,000	> = $5.70	0.36	0.34	0.01	0.037	2,400,000	1.27
Hypogene	392,000,000	> = $5.70	0.31	0.25	0.01	0.080	355,600,000	2.73
Summary	413,200,000		0.31	0.25	0.01	0.076	374,900,000	2.60
Measured + Indicated
Oxide	72,700,000	> = $5.70	0.38	0.38			66,000,000
Mixed	7,600,000	> = $5.70	0.42	0.38	0.01	0.044	6,900,000	1.50
Hypogene	510,700,000	> = $5.70	0.34	0.27	0.01	0.088	463,300,000	3.03
Summary	591,000,000		0.35	0.29	0.01	0.077	536,200,000	2.64
Inferred
Oxide	3,500,000	> = $5.70	0.33	0.33			3,200,000
Mixed	1,300,000	> = $5.70	0.47	0.45	0.00	0.019	1,200,000	0.66
Hypogene	63,900,000	> = $5.70	0.35	0.29	0.01	0.049	58,000,000	1.69
Summary	68,700,000		0.35	0.30	0.01	0.046	62,400,000	1.58

Notes:

1.	Domains were modelled in 3D to separate mineralized rock types from surrounding waste rock. The domains were based on core logging, structural and geochemical data.
2.	Raw drill hole assays were composited to 25-foot lengths broken at lithology boundaries.
3.	Capping of high grades was considered necessary and was completed for each domain on assays prior to compositing.
4.	Block grades for copper, molybdenum and silver were estimated from the composites using OK interpolation into 50 ft x 50 ft x 50 ft blocks coded by domain.
5.	Tonnage factors were interpolated by lithology and mineralized zone. Tonnage factors are based on 2,066 measurements collected by Hudbay and previous operators.
6.	Blocks were classified as Measured, Indicated or Inferred in accordance with CIM Definition Standards 2014.
7.	Mineral resources are constrained within a computer generated pit using the LG algorithm. Metal prices of $3.15/lb copper, $11.00/lb molybdenum and $18.00/troy oz silver. Metallurgical recoveries of 85% copper, 60% molybdenum and 75% silver were applied to sulfide material. Metallurgical recoveries of 40% copper, 30% molybdenum and 40% silver were applied to mixed material. A metallurgical recovery of 65% for copper was applied to oxide material. NSR was calculated for every model block and is an estimate of recovered economic value of copper, molybdenum, and silver combined. Cut-off grades were set in terms of NSR based on current estimates of process recoveries, total process and G&A operating costs of $5.70/ton for oxide, mixed and sulfide material.
8.	The oxide resource will be processed in the mill via flotation.
9.	Totals may not add up correctly due to rounding.

1.13.2        Comparison with the 2012 Mineral Reserves

A review and comparison of the 2017 Hudbay mineral resource and 2012 Augusta mineral reserves was completed. The results in Table 1-7 of proven and probable reserves show that Hudbay reports a tonnage 11% lower; with copper grades 2% higher, molybdenum grades 17% lower and silver grades 11% higher compared to those estimated in 2012.

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TABLE 1-7: PROVEN AND PROBABLE, COMPARISON TO 2012 AUGUSTA RESERVE ESTIMATE

Category	Hudbay Reserves 2017				Augusta Reserves 2012 Model
	Tons	Cu (%)	Mo (%)	Ag (opt)	Tons	TCu (%)	Mo (%)	Ag (opt)
Proven	469,708,117	0.48	0.012	0.14	308,075,000	0.46	0.015	0.12
Probable	122,324,813	0.31	0.010	0.09	359,131,000	0.42	0.014	0.12
TOTAL	592,032,930	0.45	0.012	0.13	667,206,000	0.44	0.014	0.12

The changes between the 2012 and 2017 mineral reserve estimates can be mostly attributed to a revision of the mining, processing and general & administration cost assumptions resulting in a marginally higher cut-off in 2017.

1.14        Mining Methods

The Rosemont deposit is a high-tonnage, skarn-hosted, porphyry-intruded, copper-molybdenum deposit located in close proximity to the surface. The Project will be a traditional open pit shovel/truck operation. It consists of open pit mining and flotation of sulfide minerals to produce commercial grade concentrates of copper and molybdenum. Payable silver and gold will report to the copper concentrate.

The proposed pit operations will be conducted from 50-foot-high benches using large-scale mine equipment, including: 10-5/8-inch-diameter rotary blast hole drills, 60 yd³ class electric mining shovels, 46 yd³ class hydraulic shovels, 25 yd³ front-end loaders, and 260-ton off-highway haul trucks.

The Rosemont final pit will measure approximately 6,000 feet east to west, 6,000 feet north to south, and will have a total depth of approximately 2,900 feet down to 3,100 feet (AMSL). There is one primary waste rock storage area (“WRSA”), which is located 1,200 feet southeast of the Rosemont final pit. The processing facility is located approximately 1,000 feet east of the final pit, while the dry stacking tailings facility (“DSTF”) is located 1,500 feet southeast of the Rosemont pit. The final pit and facilities can be seen in Figure 1-4.

The mine production plan contains 592 million tons of ore and approximately 1.25 billion tons of waste, yielding a life of mine waste to ore stripping ratio of 2.1 to 1 (including pre-stripping material). The mine has a 19-year life, with ore to be delivered to the processing plant at a throughput ramping up to 90,000 tons per day (tpd).

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FIGURE 1-4: ROSEMONT MINE PLAN SITE LAYOUT

1.14.1        Mine Phases

The mine phases and ultimate pit for the Project are designed for large-scale mining equipment (specifically, 60 yd³ class electric shovels and 260-ton haulage trucks) and is derived from the selected LG pit shells described in the previous section. The design process included smoothing pit walls, eliminating or rounding significant noses and notches that may affect slope stability, and providing access to working faces by developing internal ramps (including a dual ramp for the final pit).

For the pit design, the targeted minimum mining width is 320 ft. and honored the wall slope design provided by Call and Nicholas, Inc. (“CNI”) and Hudbay. Table 1-8 lists the configuration of the recommended pit slope configuration for each sector.

TABLE 1-8: ROSEMONT SLOPE GUIDANCE

Geotechnical Sector	Bench Height, ft.	Bench Face Angle°	Inter-Ramp Slope Angle°	Catch BBench, ft.	Overall Slope Angle°
1	100	70	50	48	42
2	100	65	46	50	40
3	100	65	48	44	45
4	100	65	48	44	45
5	50	65	46	25	43
6	50	65	44	29	41
7	50	55	33	42	31
8	50	55	33	42	31

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Total ore reserves in the final pit are estimated to be 592 million tons. Approximately 55 million tons of medium and low grade oxide, mixed and sulfide ore will be stockpiled. This material will be reclaimed and processed during operations.

Final configuration of mine phases in plan view is presented in Figure 1-5 and in cross section in Figure 1-6. Mineral reserves for the Rosemont deposit by mine phase are summarized in Table 1-9.

FIGURE 1-5: PLAN VIEW OF ROSEMONT MINE PHASES

FIGURE 1-6: AA' SECTION VIEW OF ROSEMONT MINE PHASES

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TABLE 1-9: ROSEMONT MINE PHASES MINERAL RESERVES

	Ore M Tons	TCu %	SCu %	ASCu %	Mo %	Ag opt	NSR $/ton	CuEq %	Waste M Tons	Total M Tons	S.R.
PH01	84.8	0.49	0.43	0.06	0.011	0.16	21.80	0.57	190.3	275.1	2.24
PH02	88.3	0.43	0.38	0.05	0.010	0.15	19.77	0.51	115.6	203.9	1.31
PH03	74.8	0.50	0.45	0.04	0.012	0.15	23.18	0.58	177.9	252.7	2.38
PH04	63.5	0.53	0.50	0.03	0.014	0.13	25.26	0.62	182.5	246.0	2.87
PH05	59.4	0.47	0.44	0.03	0.014	0.12	22.65	0.56	150.3	209.8	2.53
PH06	221.2	0.39	0.34	0.05	0.012	0.12	17.64	0.46	431.9	653.1	1.95
Total	592.0	0.45	0.40	0.05	0.012	0.13	20.57	0.53	1,248.6	1,840.6	2.11

Notes:

1.	TCu % corresponds to the total copper grade.
2.	SCu % grade corresponds to the sulfide copper in the Ore. As per formula SCU = TCU – ASCu
3.	ASCu % grade corresponds to the soluble copper.
4.	CuEq% is calculated based on metal prices of $3.15/lb Cu, $11.00/lb Mo and $18.00/oz Ag.

1.14.2        Mine Schedule and Production Plan

The operating and scheduling criteria used to develop the mining sequence plans are summarized in Table 1-10 below.

TABLE 1-10: MINE PRODUCTION SCHEDULE CRITERIA

Parameter	Value
Annual Ore Production Base Rate	32,850,000 tons
Daily Ore Production Base Rate	90,000 tons
Operating Hours per Shift	12
Operating Shifts per Day	2
Operating Days per Week	7
Scheduled Operating Days per Year	365
Number of Mine Crews	4

Pit operations and mine maintenance will be scheduled around the clock. Allowances for down time and weather delays have been included in the mine equipment and manpower estimations.

A mill ramp up period for concentrator start-up has been considered. Provisions are included to reach a full and steady production (throughput) by the end of the sixth month of operation. This assumption is based on the actual ramp-up achieved by Hudbay in 2016 at the Constancia Project in Peru.

An elevated cut-off grade strategy has been implemented to bring forward a slightly higher-grade ore from the pit to the early part of the ore production schedule. Delivering higher-grade ore to the mill in the early years will improve the net present value and internal rate of return of the Project. NSR values were calculated for each block in the resource model to represent the net Cu, Mo, and Ag metal values. The pit reserves were estimated at a cut-off with an NSR value of $6.00/ton. This is the minimum value of mineralized material that will cover the processing and G&A costs and is therefore reserved for mill feed. Priority plant feed will consist of high grade material (NSR above $12.00/ton) . The medium and low grade material (NSR between $6.00/ton and $12.00/ton) will be fed as needed to make up any immediate ore short-fall, but the bulk of this material will be stockpiled.

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The stripping analysis determined a minimum preproduction stripping requirement of approximately 94 million tons of waste. Approximately 11 million tons of ore will also be mined and stockpiled during this period.

A mine life of approximately 19 years is projected by this development plan. Peak mining rates of 367,000 tpd of total material will be realized in year 1 through year 11. Average mining rates during years 12-14 will be 180,000 tpd of total material, and will then be reduced to an average of 105,000 tpd from years 15 – 17 as the strip ratio drops.

The estimated mine production schedule, in terms of annual movement of material, is summarized in Figure 1-7.

FIGURE 1-7: ROSEMONT MINE SCHEDULE, MATERIAL MOVEMENT

1.14.3        Waste Rock Storage Area (WRSA)

Overburden and other waste rock encountered during the course of mining will be placed into a WRSA located to the south and southeast of the planned open pit and within the permitted landform area (i.e., combined WRSA and DSTF). The design criteria for the WRSA and associated haul roads are summarized in Table 1-11 below. The general mine site layout is shown in Figure 1-4.

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TABLE 1-11: WASTE ROCK FACILITY DESIGN CRITERIA

Description	Criteria
Angle of Repose	37°
Average Tonnage Factor (with swell)	16.02 ft3/ton
Overall Slope Angle	3.5H:1V
Total Height, ft	600
Lift, ft	100
Haul Road, ft	120
Max Elevation, ft (AMSL)	5,700

1.14.4        Dry StackTailings Facility (DSTF) Buttress

The DSTF is north of the WRSA area and east-northeast of the pitt. The DSTF is the repository where processed ore tailings will be placed behind large containmen nt buttresses constructed from mine waste rock. The design criteria for the DSTF and associated haul roads are summarized in Table 1-12 below. The general mine site layout is shown in Figure 1-4 and a N-S cross section view of the DSTF buttress by year is shown in Figure 1-8 below.

TABLE 1-12: DSTF BUTTRESS ROCK STORAGE DESIGN CRITERIA

Description	Criteria
Angle of Repose	37°
Average Tonnage Factor (with swell)	16.02 ft3/ton
Overall Slope Angle	3.5H:1V
Total Height, ft	700
Haul Road, ft	120
Max Elevation, ft (AMSL)	5,490

FIGURE 1-8: DRY STACK TAILINGS FACILITY NS SECTION VIEW, LOM BUTTRESS BY YEAR

1.14.5        Mine Equipment

The proposed pit operations will be conducted from 50-foot-high benches using large-scale mine equipment, including: 10-5/8-inch-diameter rotary blast hole drills, 60 yd³ class electric mining shovels, 46 yd³ class hydraulic shovels, 25 yd³ front-en nd loaders, and 260 ton off-highway haul trucks.

The mine will operate two shifts per day, 12 hours per shift for 365 days a year. No significant weather delays are expected and the mine will not be shut down for holidays. Crew work schedule will consist of a standard four crew rotation.

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A summary of fleet requirements by time period for major mine equipment is shown in Table 16-12. This represents the equipment necessary to perform the following mine tasks:

• Mine Site clearing and topsoil salvage and stockpiling.
• Construction of the main haul roads.
• Production drilling.
• Loading and hauling of sulfide ore to the primary crusher (located on the east side of the pit), and waste rock to the WRSA and DSTF buttresses.
• Maintain mine haulage and access roads.
• Maintain WRSA, DSTF and berms, and allow re-grading of slopes and final surfaces for concurrent reclamation.
• Control dust.

1.14.6        Mine Manpower Requirements

Mine supervision, technical staff, mine maintenance, workshop personnel and equipment operator requirements over the life of the mine is based on the mine plan. During the Pre-Production period, direct (workshop and operators) and indirect (staff, supervision and technicians) requirements are 337, building up to a peak of 459 in year 7.

Mine staff manpower employees and salaries were developed for Mine Administration, Mine Geology, Mine Operations, and Mine Maintenance. Salaries were a composite of information provided by Hudbay which was calibrated against local mine salaries. Salary information includes wages, burden and bonus for staff employees.

1.15        Recovery Methods

The Rosemont process plant is a conventional copper-molybdenum concentrator and its process design is typical of concentrators treating low sulfur copper porphyry-skarn style ores. The process involves crushing, grinding, flotation, concentrate dewatering, molybdenum separation and tailings dewatering.

The process plant design is based on a combination of metallurgical testwork, Project production plan, and in-house information, and is modelled after the Constancia Copper Project design with changes made where necessary to address differences. With minor modifications, the process plant is designed to treat an average of 90,000 tons/d (32.8 million tons/y).

1.16        Project Infrastructure

The Project Infrastructure consists of access and plant roads, electric power supply and distribution, water supply and distribution, voice and data communication, and DSTF, and other ancillary facilities.

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1.16.1        Access and Plant Roads

Access and plant roads consist of an access road into the plant from State Highway 83, in-plant roads, haul roads and a perimeter road around the toe of the WRSA and DSTF. The plant and access roads are shown in Figure 18-1.

1.16.2        Power Supply and Distribution

An agreement between Tucson Electric Power (“TEP”), Trico Electric, and Hudbay will be realize to provide the electrical power supply, estimated to be approximately 183 MVA, for the Project. A proposed switchyard (Toro Switchyard) will tap into the existing TEP 138 kV transmission line that extends from the South Substation to the Green Valley Substation. A 13.2 -mile-long proposed 138 kV transmission line originates at the Toro Switchyard and terminates on private property to the Rosemont substation as shown in Figure 18-2.

1.16.3        Water Supply and Distribution

The fresh water requirement for the Rosemont facilities is approximately 6,000 acre-feet per year. The water supply source identified for the Project is groundwater from the Santa Cruz basin, which lies west of the Project and the Santa Rita Mountains.

There are 4 pump stations located strategically to pump the necessary water to the storage tank located at the mine site. Water from the storage tank will be provided for the following systems:

• potable water system,
• fresh water system,
• process water system, and
• fire water system.

1.16.4        Tailings Management

The Rosemont tailings dry stack is designed as a low hazard facility with fully drained waste rock placed as buttressing material. The slope stability analyses performed on the outer slope indicate the dry tailings stack operations can be constructed with stable 3H:1V inter-bench slopes and an overall stable slope of approximately 3.5H:1V. The design was developed based on hydrological and geotechnical studies that included review of regional climate data, drilling and testing programs, and laboratory characterization of subsurface and tailings samples.

An initial starter buttress around the tailings facility will be constructed with waste rock. Concurrent tailings and waste rock placement in the buttress will occur throughout the life of the tailings facility.

1.16.5        Communication

The proposed approach is to integrate data networking and telecommunication systems into a common infrastructure to meet the requirements for accounting, purchasing, maintenance, and general office business as well as specialized requirements for control systems. Mobile radios will

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also be used by the mine and plant operation personnel for daily control and communications while outside the offices.

A security system has been incorporated into the plant network. Using a dedicated video server and monitors, I/P cameras utilizing power over ethernet connections will be plugged into dedicated switches.

1.17        Market Studies and Contracts

Hudbay has a marketing division that is responsible for establishing and maintaining all marketing and sales administrations of concentrates and metals. Rosemont copper concentrates are expected to be a clean, high grade concentrate containing small gold and silver by-product credits which will be suitable as a feedstock for smelters globally. Approximately 50% of the copper concentrate production has been contracted under long term sales contracts.

Table 1-13 below summarizes the key assumptions for the sale of Rosemont’s copper concentrate.

TABLE 1-13: COPPER CONCENTRATE

	Units	LOM Total / Average
Copper Concentrate Base Treatment Charge	$ / dry short ton con	$73
Copper Refining Charge	$ / lb Cu	$0.08
Silver Refining Charge	$ / oz Ag	$0.50
Copper Concentrate Transport & Freight	$ / wet short ton con	$127
LOM Copper Grade in Copper Concentrate	% Total Cu	34.3%
Moisture Content of Copper Concentrate	% H2O	8.0%

No deleterious elements are expected to be produced in quantities which would result in material selling penalties.

Pursuant to a precious metals stream agreement with Silver Wheaton entered into on February 11, 2010, as amended and restated on February 15, 2011, Hudbay will receive deposit payments of $230 million against delivery of100% of the payable gold and silver from the Project . The deposit will be payable upon the satisfaction of certain conditions precedent, including the receipt of permits for the Project and the commencement of construction. In addition to deposit payments, as gold and silver is delivered to Silver Wheaton, Hudbay will receive cash payments equal to lesser of (i) the market price and (ii) $450 per ounce (for gold) and $3.90 per ounce (for silver), subject to one percent annual escalation after three years.

Rosemont is expected to produce a marketable 45% molybdenum concentrate. Table 1-14 below summarizes the key assumptions for the sale of Rosemont’s molybdenum concentrate.

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TABLE 1-14: MOLYBDENUM CONCENTRATE

	Units	LOM Total / Average
Molybdenum Concentrate Base Treatment Charge	$ / lb Mo	$1.50
Molybdenum Concentrate Transport & Freight	$ / wet short ton con	$124
LOM Molybdenum Grade in Molybdenum Concentrate	% Mo	45.0%
Moisture Content of Molybdenum Concentrate	% H2O	8.0%

1.18        Environmental Studies, Permitting and Social or Community Impact

Permitting status for the Project is well advanced and has continued to progress since July 2007. The final approvals required include the Final Record of Decision (“ROD”) from the U.S. Forest Service (“USFS”) and the 404 Permit from the U.S. Army Corps of Engineers (“USACE”). These final Federal permits are currently in the review process.

Since 2013 when the Final Environmental Impact Statement (“EIS”) and Draft ROD were issued, the USFS has finalized two Supplemental Information Reports (“SIRs”) and a Supplemental Biological Assessment (“SBA”)and completed a consultation with the U.S. Fish and Wildlife Services (“USFWS”) culminating in an Amended Final Biological and Conference Opinion (“BO”) in April 2016. The SIRs determined that nothing disclosed to date would indicate that the information in the SIR falls outside the information disclosed in the EIS. The BO determined that none of the endangered species were jeopardized by the Project.

The USFS is expected to issue its ROD once the USACE is clear on the decision it will make for the Project. This will allow the USFS to include additional analysis into their record and review it against the disclosed impacts in the EIS if the USACE determines it is necessary to make adjustments to their portion of the Project, mitigation, or evaluations. Once the ROD is issued, Hudbay will submit the Mine Plan of Operations (“MPO”) to the USFS for their approval. This approval is expected to take up to six months, and once the MPO is approved, site access is granted.

The USACE is evaluating the overall project record and a mitigation package that provides mitigation for impacts to the ephemeral channels on the Project site. This mitigation incorporates the restoration of a floodplain that was impacted by agriculture; mitigation for two sites impacted by grazing, poor roadway maintenance, and other activities; as well as preservation of sites near to the Project site. Once the USACE evaluation is complete, a decision will be made by the USACE on permit issuance, terms and conditions and appropriate financial assurance will be negotiated.

At this time, the State of Arizona Permits and Approvals dealing with the environment have been issued for the Project, and all permits remain in force and are current. The Project continues to comply with permit terms and conditions. No additional environmental permits are necessary to begin construction of the facilities, and only minor environmental permits (e.g., septic system permits, water system approvals and registration) will be needed during construction.

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The Project refinements included in this document were specifically designed and evaluated to fall within the envelope included in the EIS review and as such are not expected to cause concern by the agency. State of Arizona permits that were issued based on early designs will be amended to include designs included in the EIS. Such amendments are customary in the state of Arizona.

1.19        Capital and Operating Cost

Initial project capital costs are estimated to be $1,921 million including 15% contingency on all items. The LOM sustaining capital costs are estimated to be $387 million excluding capitalized stripping and $1,168 million including capitalized stripping. The capital cost estimate is considered to be a Class 3 estimate as defined by AACE Recommended Practice 47R-11 for the mining and mineral process industry.

The average LOM operating costs (mining, milling and G&A) are estimated to be $9.24/ton milled (before deducting capitalized stripping) and $7.92/ton milled (after deducting capitalized stripping). Refer to Section 21 for greater capital and operating cost detail.

Over the first 10 years, C1 cash costs (net of by-product credits at stream prices) are estimated to average $1.40 per pound of copper before deducting capitalized stripping and $1.14 per pound of copper after deducting capitalized stripping. LOM C1 cash costs are estimated to be $1.47 per pound of copper before deducting capitalized stripping and $1.29 per pound of copper after deducting capitalized stripping. Including royalties and sustaining capital, sustaining cash costs are estimated to be $1.59 per pound of copper over the first 10 years and average $1.65 over the LOM.

1.20        Economic Analysis

The economic viability of the Project has been evaluated using the metal prices outlined in Table 1-15. The metal prices used in the economic analysis are based on a blend of consensus metal price forecasts from over 30 well-known financial institutions and Wood Mackenzie.

TABLE 1-15: METAL PRICE ASSUMPTIONS

Metal	Units	Price
Spot Copper	$/lb	$3.00
Spot Molybdenum	$/lb	$11.00
Spot Silver	$/oz	$18.00
Streamed Silver1	$/oz	$3.90

1. Subject to a 1% annual inflation adjustment

The terms of the existing precious metals streaming agreement with Silver Wheaton were included in the analysis. Silver Wheaton will make upfront cash payments totalling $230 million to fund initial development capital in exchange for 100% of the silver and gold production from Rosemont. Silver Wheaton will make ongoing payments of $3.90 per ounce of silver and $450 per ounce of gold subject to a 1% inflation adjustment starting on the third anniversary of production.

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Although gold is not part of the current reserve estimate, metallurgical testing has demonstrated economic concentrations of gold in copper concentrate as outlined in Section 13. Over the LOM, approximately 309 thousand ounces of gold are expected to be recovered in copper concentrate (although the financial impact has not been included).

At the effective realized prices including the impact of the stream, the revenue breakdown at Rosemont is approximately 92% copper, 6% molybdenum, and 2% silver.

Rosemont’s annual copper production (contained copper in concentrate) and C1 cash costs (net of by-products at stream prices after deducting capitalized stripping) are shown below in Figure 1-9. Over the first 10 years, annual production is expected to average 140 thousand tons of copper at an average C1 cash cost of $1.14/lb. Over the 19-year LOM, annual production is expected to average 112 thousand tons of copper at an average C1 cash cost of $1.29/lb.

FIGURE 1-9: ROSEMONT ANNUAL COPPER PRODUCTION AND C1 CASH COSTS

Rosemont (on a 100% basis) has an unlevered after-tax NPV8% of $769 million and a 15.5% after-tax IRR using a copper price of $3.00/lb as summarized in Table 1-16. The Project NPV and IRR are calculated using end of period quarterly discounting in the quarter immediately before development capital is spent.

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TABLE 1-16: LIFE OF MINE FINANCIAL METRICS (100% PROJECT BASIS)

Metric	Units	LOM Total
Gross Revenue (Stream Prices)	$M	$13,377
TCRCs	$M	($1,837)
On-Site Operating Costs (After Deducting Capitalized Stripping)	$M	($4,691)
Royalties	$M	($368)
Operating Margin	$M	$6,480
Development Capital	$M	($1,921)
Stream Upfront Payment	$M	$230
Sustaining Capital (excludes capitalized stripping)	$M	($387)
Capitalized Stripping	$M	($781)
Pre-Tax Cash Flow	$M	$3,622
Cash Income Taxes	$M	($718)
After-Tax Free Cash Flow	$M	$2,903
After-Tax NPV8%	$M	$769
After-Tax NPV10%	$M	$496
After-Tax IRR	%	15.5%
After-Tax Payback Period	Years	5.2

The NPV8% (100% Project basis) was sensitized based on percentage changes in various input assumptions above or below the base case. Each input assumption change was assumed to occur independently from changes in other inputs. The sensitivity analysis is summarized in Figure 1-10. The Project is most sensitive to the copper price, followed by initial capital costs, on-site operating costs, and the molybdenum price.

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FIGURE 1-10: NPV8% SENSITIVITY (100% BASIS)

Table 1-17 below reports the after-tax NPV8%, NPV10%, IRR and Payback of the Project at various flat copper prices assuming all other inputs remain constant.

TABLE 1-17: AFTER-TAX NPV8%, NPV10%, IRR AND PAYBACK SENSITIVITY AT VARIOUS FLAT COPPER PRICES (100% BASIS)

	Flat Copper Price ($/lb)
	$2.50	$2.75	$3.00	$3.25	$3.50
After-Tax NPV8% ($M)	$45	$412	$769	$1,115	$1,448
After-Tax NPV10% ($M)	($122)	$192	$496	$792	$1,076
After-Tax IRR (%)	8.5%	12.2%	15.5%	18.5%	21.2%
After-Tax Payback (years)	6.9	5.9	5.2	4.4	4.3

The existing Joint Venture Agreement requires cash payments from UCM totaling $106 million to the Joint Venture (“JV”) in order for UCM to complete its earn-in for 20% ownership of the Project. The payments will be made on an installment basis to fund the initial development capital and payments will commence once certain milestones are achieved. The NPV attributable to Hudbay is improved beyond 80% of the standalone project NPV due to the JV payments, and the IRR attributable to Hudbay is improved beyond the standalone project IRR as a result of the reduced time period between development capital spending and positive project cash flow. Table 1-18 shows the adjusted key financial metrics attributable to Hudbay.

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TABLE 1-18: KEY FINANCIAL METRICS ATTRIBUTABLE TO HUDBAY

Metric	Units	LOM Total
Development Capital (100% Basis)	$M	$1,921
Stream Upfront Payment	$M	($230)
Joint Venture Earn-in Payment	$M	($106)
JV Share of Remaining Capital (20%)	$M	($317)
JV Loan Repayment to Hudbay1	$M	($20)
Hudbay's Share of Development Capital	$M	$1,248
After-Tax NPV8% to Hudbay	$M	$719
After-Tax NPV10% to Hudbay	$M	$499
After-Tax IRR to Hudbay	%	17.7%
After-Tax Payback Period to Hudbay	Years	4.9

1. Hudbay is funding the JV’s share of project expenditures until receipt of material permits and approximately $20M in principal and accrued interest is due to Hudbay

1.21        Adjacent Properties

There is no material information concerning mineral properties immediately adjacent to the Project.

1.22        Other Relevant Data and Information

The author is not aware of any other information that would impact the reported estimate of mineral resources for the Project.

A draft feasibility study was completed for the Project which included information on the basis of design, infrastructure, design strategies, Project Execution Strategy, risks assessments and recommendations. The EPCM team has also completed a draft construction execution plan.

The Project has undergone various risk assessments and workshops during the years and continues to hold quarterly risk assessment workshops.

1.23        Conclusions

The purpose of this Technical Report is to present Hudbay’s estimate of the mineral reserves and mineral resources for the Project based on the current mine plan, the current state of metallurgical testing, operating cost and capital cost estimates. The results of “feasibility study” level work conducted partly by external contractors and partly internally by Hudbay, completion of the drill program and bench-marking against other mines including Hudbay’s Constancia mine has resulted in the following fundamental conclusions:

• The Rosemont deposit consists of copper-molybdenum-silver-gold mineralization primarily hosted in skarn formed on a chemical/siliciclastic sedimentary sequence after the intrusion of Laramide quartz monzonite porphyry intrusions.

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• A new geological model was built based on chemostratigraphy and lithogeochemistry from 33,000 samples covering the full footage of the 2014 and 2015 Hudbay drilling programs. Geochemical based geological model reduces uncertainties in the formational, lithological and alteration logging. The updated geological model incorporated a revised structural framework based on a surface mapping and downhole structural review.
• The resource economic parameters utilized are slightly different than those supporting the mineral reserve statement. The cost and price inputs for the mineral resource economic parameters are considered an approximation and were used to test the economic viability of the resource. Although these costs and prices differ from the ones used for the mineral reserves, it is the opinion of the QP that changing the resource parameters would not materially change the output of the reserve.
• A proven and probable reserve of 592 million tons has been identified and its economic viability demonstrated. An additional 591 million tons of measured and indicated resources and 69 million tons of inferred resources have been identified and outlined as having further potential for economic extraction once the mineral reserves have been extracted.
• Mining equipment performance and maintenance requirements and costs are benchmarked from Constancia’s actual operating information and are robust.
• Metallurgical testwork has confirmed that the Rosemont ores respond well to proven and widely used sulfide mineral processing techniques.
• The tailing properties have been sufficiently characterized as well as the dewatering performance of vendor equipment over the life of the operation to satisfy the estimated number, type and size of tailing filters for this Project. To be conservative, expansion space has been allocated for additional filtering equipment, to the extent that it may be necessary.
• The Rosemont Process Plant design has been modelled after the operating Constancia processing plant’s flow sheet and is sufficiently robust to routinely achieve key production targets such as throughput, recovery, and concentrate grade as stated in the production schedule based on the metallurgical testing conducted.
• Flexibility exists within the mine plan to optimise plant recovery and performance through the management of feed types including clays, oxides and hardness.
• The Project is one of many large projects scheduled to be constructed. The author believes that schedule slip will be the principal pressure on cost should the Project experience construction delays. In the opinion of the author, the contingency allotted to construction capital cost should mitigate most cost risk. At the time of publication of this Technical Report, committed and spent dollars would raise this contingency to approximately 15% of the required project capital as estimated in the draft Detailed Feasibility Study (dDFS).

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• Related to the capital costs, M3 Engineering and Amec Foster Wheeler (acting under a joint venture agreement) and Ausenco each completed a value engineering phase and independently produced capital estimates for the Project which were within 5% of each other. Since then, Ausenco has also completed a feasibility quality capital estimate, benchmarked their findings with Constancia (and other similar projects), and engaged third party construction contractors and various consulting firms to provide additional input into the estimate. Hudbay has also completed an independent third party review of the feasibility study estimate.
• The net present value of the project is most sensitive to the price of copper. The resulting project NPV8% ($769 million) and IRR (15.5%) utilizing the current Hudbay long term view on metal prices, TCRC’s and other economic assumptions, in the opinion of the author, support the declaration of mineral reserves as outlined in the CIM guidelines.
• The project execution plan is modelled after the Constancia delivery method. Many key personnel who developed the Constancia mine are members of the Project who will be involved in the development of the Project and therefore considered a robust project execution plan.

The Project is uniquely located in a copper mining jurisdiction that has sustained economic copper production for close to 140 years. Since it is located approximately 30 miles from Tucson it is expected to have a significant impact on employment and economic gain for the region. The proposed mining, processing, and logistics plan provides a step forward in innovation and sustainability. The dry stack tailings deposition proposed would be among the largest in size and address industry and stakeholder concerns regarding the use of water and the stability of tailings impoundment facilities. The proposed design and operating practice that will be applied in respect of the Project is expected to set a new standard by which other large mining projects are judged with respect to their impact on stakeholders, the ecology and the environment.

In recognition of the scarcity of world economic copper reserves in an environment of ever increasing consumption of the metal, Hudbay has carefully considered the ecological, environmental, and ethical extraction methods to be applied to the Project in an effort to set it apart from others in the world. The Project is located in a first world leading nation, where extraction and production is governed by laws with due process and human rights fundamental to the consumer.

This Technical Report also concludes that the estimated mineral reserves and mineral resources for the Project conform to the requirements of 2014 CIM Definition Standards – for Mineral Resources and Mineral Reserves and requirements in Form 43-101F1 of NI 43-101, Standards of Disclosure.

1.24        Recommendations

The Author recommends the following:

• The author recommends that Hudbay further investigate the cause(s) of the differences in average molybdenum grade of the historical assays. Hudbay should also evaluate the application of non-linear interpolation or wireframing methods in the minor geological units.

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• A drill hole twinning program of the pre-Augusta drilling to bolster the confidence of the re- assaying campaign as conducted by Augusta.
• Further investigate change-of-support correction and alternative approaches to resource classification taking into account the high production rate. This should be performed to ensure that the resource classification properly reflects the reduced risk when a large volume is mined and delivered to the mill on a quarterly and annual basis.
• It is recommended by the author that 5% of the samples should be sent for check assay in future drill hole campaigns.
• Future drilling campaigns should consider the by-product credit contribution of gold noted contained in copper concentrate produced through testing and increase the confidence in geological continuity such that it can be included in the reserve statement.
• Metallurgical test-work has confirmed that marketable copper concentrates can be produced. Mine and mill production sequencing and planning will require care to manage clay content that can adversely affect flotation and tailings filtration.
• A geometallurgical program is recommended as a component of the operating plan to further refine, monitor and optimise the mine to mill performance. It is concluded that fluorine can be readily rejected from copper concentrate, and further study is recommended to develop understanding of ore conditions and indicators that trigger elevated fluorine content in concentrate.
• Ongoing optimization of pit slope designs should be conducted during operation and should be based on observed conditions accounting for more detailed mapping of local alteration, jointing and faulting. These observations and compilation can then provide the basis for revised slope geometry and pushback (mining phase) configurations that have the potential to increase mineral reserves and reduce overall stripping requirements.

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2        INTRODUCTION AND TERMS OF REFERENCE

The Principal Author has prepared this Feasibility-Level Technical Report for Hudbay Minerals Inc. (“Hudbay”) on the Project, located approximately 30 miles (50 km) southeast of Tucson, in Pima County, Arizona. This Technical Report conforms with the 2014 CIM Definition Standards for Mineral Resources and Mineral Reserves and requirements in Form 43-101F1 of NI 43-101, Standards of Disclosure for Mineral Projects.

Hudbay is a Canadian integrated mining company with assets in North and South America principally focused on the discovery, production and marketing of base and precious metals. Hudbay’s objective is to maximize shareholder value through efficient operations, organic growth and accretive acquisitions, while maintaining its financial strength.

Hudbay acquired all of the issued and outstanding common shares of Augusta Resource Corporation (“Augusta”) pursuant to take-over bid, which expired July 29, 2014, and a subsequent acquisition transaction, which closed on September 23, 2014.

This Technical Report describes the latest resource model and mine plan and the current state of metallurgical testing, operating cost and capital cost estimates. The information presented in this Technical Report is the result of “feasibility study” level work conducted partly by external contractors and partly by internal Hudbay personnel under the overall supervision of the Qualified Person (“QP”).

The QP who supervised the preparation of this Technical Report is Cashel Meagher, P.Geo, Senior Vice President & Chief Operating Officer for Hudbay. Mr. Meagher last visited the property on April 21, 2016 and numerous times prior to this date. The personal site inspections were conducted as part of the 2014-2015 diamond drilling program to become familiar with conditions on the property, to observe the geology and mineralization and to verify the work completed on the Property. Mr. Meagher has also reviewed and conducted sufficient confirmatory work to act as QP for the reporting of the mineral resource and mineral reserve estimates for the Project.

2.1        Information Sources

Information used to support this Technical Report was based on current primary-source data when available, previous technical reports on the property and from the reports and documents listed in Section 27, References. Notable information reviewed and relied upon by the QP was as follows:

• 3D Block Model – Hudbay prepared a 3D block model of the Rosemont deposit. The 3D block model and determination of the mineral resources at the Rosemont deposit were performed by internal Hudbay employees, following Hudbay procedures and were reviewed and approved by Cashel Meagher, Chief Operating Officer for Hudbay and Qualified Person of this Technical Report.

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• Lerchs-Grossman Analyses – A Lerchs-Grossman (“LG”) cone algorithm was applied by Hudbay to the block model to establish the component of the deposit that has a “reasonable prospect of economic extraction”.

Additional sources of information that the QP relied upon are described in Section 3 of this Technical Report.

2.2         Unit Abbreviations

The units of measure in this report are a combination of US standard units and metric units. Unless stated otherwise, all dollar amounts (“$”) are in United States dollars. Unit abbreviations used in this report are noted below:

TABLE 2-1: UNIT ABBREVIATIONS

Abbreviation	Description
$	United States dollar
°C	degree Celsius
°F	degree Fahrenheit
%	percent
μm	microns
cm	centimetres
ft	feet
ft2	Square feet
g	gram
g/mt	grams per (metric) tonne
HP	horsepower
km	Kilometre
kV	Kilovolt
kW	kilowatt
kWh	Kilowatt-hour
m2	Square meter
m3	Cubic meter
mm	Millimetres
tonne	metric tonne
Mt	Million (short) tons
ppb	parts per billion
ppm	parts per million
t, ton	short ton
w/w	Weight per weight

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2.3        Name Abbreviations

Abbreviations of company names and terms used in the report are as shown in Table 2-2.

TABLE 2-2: NAME ABBREVIATIONS

Abbreviation	Description	Abbreviation	Description
3D	Three-Dimensional	ID2	Inverse Distance Squared
AAS	Atomic Absorption Spectrometry	Inspectorate	Inspectorate America Corporation
ADWR	Arizona Department of Water Resources	LAF	Low Angle Fault
		LG	Lerchs-Grossman
Ag	Silver	MPO	Mine Plan of Operations
AMSL	Above mean seal level	Mo	Molybdenum
ASCu	Acid soluble copper	NaHS	Sodium Hydrosulfide
Augusta	Augusta Resource Corporation	NEPA	National Environmental Policy Act
AV	Average	NI	National Instrument
BADCT	Best Available Demonstrate Control Technology	NN	Nearest Neighbour
		NQ	HQ drill core size 1.875 inches or 47.6 mm diameter
BQ	BQ drill core size 1.43 inches or 36.4mm	NSR	Net Smelter Return
Bureau Veritas	Bureau Veritas Commodities Canada Ltd.	OK	Ordinary Kriging
CBV	Certified best value	OSA	On-Stream Analyser
CEC	Cation Exchange Capacity	OREAS	Ore Research and Exploration
CIM	Canadian Institute of Mining, Metallurgy and Petroleum	PQ	PQ drill core size 3.3 inch or 83 mm diameter
Cu	Copper	QA/QC	Quality Assurance and Quality Control
Cu-Mo	Copper-molybdenum	QUEMSCAN	Quantitative Evaluation of Minerals by
CRM	Certified reference materials		Scanning electron microscopy
CV	Coefficient of Variance	R2	Coefficient of Determination
DGM	Discrete Gaussian Model	RE	Absolute relative error
DSTF	Dry Stack Tailings Facility or Tailings Management Facility (TMF)	RMA	Reduced-to-Major-Axis regression
		ROD	Record of Decision
EDX	Energy Dispersive X-ray	RQD	Rock Quality Designation
EGL	Equivalent Grinding Length	RSE	Relative standard error of the kriged
EIS	Environmental Impact Statement		estimate
EPMA	Electron Probe Micro-analysis	RSD	Relative standard deviations
FEL	Front End Loader	SAG	Semi-Autogenous Grinding
FileMaker	FileMaker Inc.	SABC	SAG and Ball Mill and Crushing Comminution Circuit
GMD	Gearless Motor Drive
GT	Grade-tonnage	SBA	Supplemental Biological Assessment
H2SO4	Sulfuric acid	SCu	Copper in sulfides
HCT	High Compression Thickeners	SD	Standard deviation
HQ	HQ drill core size 2.50 inches or 63.5	SEM	Scanning Electron Microscopy
	mm diameter	SFR	Staged Flotation Reactor
Hudbay	Collectively all Hudbay Minerals Inc. subsidiaries and business groups	SG	Specific Gravity
		SGS	SGS Canada Inc.
ICP	Inductively Coupled Plasma	SIR	Supplemental Information Report
ICP-MS	Inductively Coupled Plasma Mass Spectrometry	Skyline	Skyline Assayers & Laboratories
ICP-OES	Inductively Coupled Plasma Optical Emission Spectroscopy	SMU	Selective mining unit
		SRM	Standard reference materials

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Abbreviation	Description	Abbreviation	Description
TCu	Total copper	USFS	U.S. Forest Service
TEP	Tucson Electric Power Company	USFWS	U.S. Fish and Wildlife Service
TIA	Tucson International Airport	USACE	U.S. Army Corps of Engineers
TRICO	TRICO Electric Cooperative Inc.	XPS	XPS Consulting & Testwork Services
UCM	United Copper & Moly LLC	XRD	X-Ray Diffraction

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3        RELIANCE ON OTHER EXPERTS

Standard professional procedures were followed in preparing the contents of this Technical Report. Data used in this report has been verified where possible and the author has no reason to believe that the data was not collected in a professional manner and no information has been withheld that would affect the conclusions made herein.

Hudbay has retained a number of contractors/consultants to prepare technical and cost information to support this Technical Report. All the information used from this work in the current Technical Report has been duly verified and validated by the author.

The information, conclusions, opinions, and estimates contained herein are based on:

• Information available to Hudbay at the time of preparation of this Technical Report,
• Assumptions, conditions, and qualifications as set forth in this Technical Report, and
• For purposes of this Technical Report, the author has relied on title and property ownership information based on select records of the U.S. Bureau of Land Management and the Assessor’s Office in Pima County, Arizona.
• The author has also relied on tax information provided by Hudbay’s tax department, Social and Environmental information as provided by appropriate Arizona based personnel from Hudbay, marketing information from Hudbay’s marketing group and forward price forecasts from Hudbay’s treasury department.

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4        PROPERTY DESCRIPTION AND LOCATION

4.1        Location

The Project is located within the historic Helvetia-Rosemont Mining District that dates back to the 1800’s. The deposit lies on the eastern flanks of the Santa Rita Mountain range approximately 30 miles (50 km) southeast of Tucson, in Pima County, Arizona off of State Route 83 (see Figure 4-1). The core land position includes patented and unpatented mining claims, fee land and grazing leases that cover most of the old Mining District. The lands are under a combination of private ownership by Rosemont and Federal ownership. The lands occur within Townships 18 and 19 South, Ranges 15 and 16 East, Gila & Salt River Meridian. The Project geographical coordinates are approximately 31º 50’N and 110º 45’W.

FIGURE 4-1: PROPERTY LOCATION OF ROSEMONT PROJECT

4.2        Land Tenure

Hudbay acquired all of the issued and outstanding common shares of Augusta Resource Corporation pursuant to take-over bid, which expired July 29, 2014, and a subsequent acquisition transaction, which closed on September 23, 2014. Hudbay’s ownership in the Project is subject to an earn-in agreement with United Copper & Moly LLC (‘‘UCM’’), pursuant to which UCM has earned a 7.95% interest in the Project and may earn up to a 20% interest subject to cash payments from UCM totaling $106 million to the Joint Venture. A joint venture agreement between Hudbay’s subsidiary, Rosemont Copper Company, and UCM governs the parties’ respective rights and obligations with respect to the Project.

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Hudbay continues to maintain the property in good standing which consists of a combination of fee land, patented and unpatented lode, mill site mining claims, and rights-of-way from the Arizona State Land Department. Taken together, the land position is sufficient to allow access to an open pit mining operation, processing and concentrating facilities, storage of tailings, disposal of waste rock and a utility corridor to bring water and power to site. The Federal lands covered by unpatented mining claims are accessible under the provisions of the Mining Law of 1872, subject to approval from the U.S. Forest Service (“USFS”) after the completion of an Environmental Impact Statement (“EIS”) as per the National Environmental Policy Act (“NEPA”) process.

The core of the Project mineral resource is contained within the 132 patented mining claims that in total encompass an area of approximately 2,000 acres (809 hectares) as shown in Figure 4-2. Surrounding the patented claims is a contiguous package of 1,064 unpatented mining claims with an aggregate area of more than 16,000 acres (6,475 hectares). Associated with the mining claims are 38 parcels of fee (private) land consisting of approximately 2,300 acres (931 hectares) (the Associated Fee Lands). The area covered by the patented claims, unpatented claims and Associated Fee Lands totals approximately 20,300 acres (8,215 hectares). A listing of the patented claims, unpatented claims and Associated Fee Lands is provided in Appendix A1-1 & A1- 2.

Rosemont has also acquired 62 parcels of fee (private) land and one parcel of leased land that are more distal from the Project area that are planned for: (1) various infrastructure purposes including, well fields, pump stations, utilities and ranch operation; and (2) for environmental mitigation and conservation purposes (together, the Distal Fee Lands). The Distal Fee Lands constitute an additional approximately 3,700 acres (1,497 hectares).

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FIGURE 4-2: ROSEMONT PROPERTY OWNERSHIP

The patented mining claims are considered to be private lands that provide the owner with both surface and mineral rights. The patented mining claim block, including the core of the mineral resource, is monumented in the field by surveyed brass caps on short pipes cemented into the ground. The fee lands are located by legal description recorded at the Pima County Recorder’s Office. The patented claims and Associated Fee Lands are subject to annual property taxes amounting to a total of approximately $8,800.

Mineral Rights on USFS and Bureau of Land Management (“BLM”) lands have been reserved to Rosemont Copper Company, via the unpatented claims that surround the patented claims. Wooden posts and stone cairns mark the unpatented claim corners, end lines and discovery monuments, all of which have been surveyed. The unpatented claims are maintained through the payment of annual maintenance fees of $155.00 per claim, for a total of approximately $165,000 per year, payable to the BLM.

The rights-of-way over State Land are all non-exclusive but grant Rosemont the rights to construct certain utility infrastructure connecting the well field and power supply to the site. Two of these rights-of-way have a term of ten years while the other four have a term of fifty years. These rights-of-way across State Land are not shown in Figure 4-2, but generally run northwest from the project site towards the Town of Sahuarita.

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There is a 3% NSR royalty on all 132 patented claims, 603 of the unpatented claims, and one parcel of the Associated Fee Lands with an area of approximately 180 acres. In the original royalty deeds, a 1.5% NSR is reserved to each of (1) Dennis Lauderbach et. ux. and (2) Pioneer Trust Company of Arizona, as Trustee under Trust No. 11778. These royalties have since been assigned and Rosemont is in the process of verifying current ownership.

A precious metals stream agreement with Silver Wheaton Corp. for 100% of payable gold and silver from the Project was entered into on February 11, 2010. Under the agreement, Hudbay will receive payments equal to lesser of either the market price or $450 per ounce for gold and $3.90 per ounce for silver, subject to 1% annual escalation after three years.

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5       ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, 
         INFRASTRUCTURE AND PHYSIOGRAPHY

5.1        Accessibility

The Project is easily accessible to the communities of Tucson and Benson to the north and Sierra Vista, Sonoita, Patagonia and Nogales to the south by way of State Route 83.

Existing graded dirt roads connect the property with State Route 83 which include Forest Service (“FS”) roads 4062 into the Hidden Valley complex, FS4064 into Rosemont Camp and FS231 transecting the property north to south. FS4051 and FS4059 provide good access into and around the Project area.

The city of Tucson, Arizona provides the nearest major railroad and air transport services to the Project and is approximately 30 miles southeast of Tucson in Pima County.

5.2        Climate

The southern Arizona climate is typical of a semi-arid continental desert with hot summers and temperate winters. The Project area is at the north end of the Santa Rita Mountain Range at elevations between 4,550 and 5,350 feet (1,387 and 1,631 meters) above mean sea level (“AMSL”). The higher elevation in the Project area results in a milder climate than at the lower elevations across the region.

Summer daily high temperatures are above 90°F (32°C) with significant cooling at night. Winter in the Project area is typically drier with mild daytime temperatures and overnight temperatures that are typically above freezing. Winter can have occasional low intensity rainstorm and light snowfall patterns that can last for several days.

The average annual precipitation in the Project area is estimated between 16 and 18 inches (41 and 46 cm) based on historical data from eight meteorological stations within a 30 mile (50 km) radius of the Project area. More than half of the annual precipitation occurs during the monsoon season from July through September. The monsoon season is characterized by afternoon thunderstorms that are typically of short duration, but with high-intensity rainfall that has minor effects on a mining operation, which is considered to be 365 days per year. The lowest precipitation months are April through June.

A meteorological station was installed at the property in April, 2006. The station is located near the center of the deposit at an elevation of 5,350 feet (1,631 m) AMSL. The station monitors site-specific weather data including temperature, precipitation, wind speed, and wind direction. Pan evaporation was added to this station in mid-2008. This station was decommissioned and a new station was located near the core shed on private property in 2015. Data from the weather station is automatically recorded and downloaded monthly by site personnel.

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5.3        Local Resources

The largest city near the Project area is Tucson with a population of over 520,000 based on the 2014 census. Tucson is also the county seat for Pima County with a population of approximately one million, which encompasses the Tucson Metropolitan Area.

Arizona produces 65% of the copper in the USA² and Tucson is a mining industry hub in the state with nine operating copper mines within a 125 mile (200 km) radius. The cultural and educational facilities provided in the Tucson Metropolitan Area attract experienced technical staff into the area. Tucson is home to a well-established base of contractors and service providers to the mining industry.

5.4        Infrastructure

The Project site is located immediately adjacent to and west of Arizona State Route 83 (South Sonoita Highway), approximately 11 miles (18 km) south of Interstate 10 (“I-10”). This system of state and interstate highways allows convenient access to the site for all major truck deliveries. The majority of the labour and supplies for construction and operations can come from the surrounding areas in Pima, Cochise and Santa Cruz Counties.

The Union Pacific mainline east-west railroad route passes through Tucson, Arizona and generally follows I-10. The Port of Tucson has rail access from the Union Pacific mainline consisting of a two mile (3.2 km) siding complimented by an additional 3,000 foot (914 m) siding.

The Tucson International Airport (“TIA”) is located approximately 30 miles (50 km) from the Project site and in close proximity to Interstate highways I-10 and I-19. TIA provides international air passenger and air freight services to businesses in the area with seven airlines currently providing nonstop service to 15 destinations with connections worldwide.

The power supply to the Project falls within the Tucson Electric Power Company (“TEP”) and TRICO Electric Cooperative Inc. (“TRICO”) service territories. Presently there exists a 13.8 kV TEP distribution line that routes through the Project site and power from this distribution line was used to service the related activities for the 2014 and 2015 drill program.

Geographically, the area east of the deposit that includes the majority of the mineral resource is in the TEP service territory, while the area west of the deposit falls within the TRICO service territory. Since most of the estimated electrical load for a mining and process operations would be located in the TEP service territory, TEP will be the electrical utility service provider for the entire facility. A joint venture business arrangement is expected to be established between TEP and TRICO to compensate both service providers in accordance with the Arizona Corporation Commission review and approval.

^{___________________________________________
2} Arizona Mining Association economic impact 2014 (azmining.com)

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The most viable source of water supply for the Project is from groundwater in the upper Santa Cruz basin aquifer. The Project currently holds a permit granted by the Arizona Department of Water Resources (“ADWR”) in 2009 to pump 6,000 acre-feet of water annually for 20 years that meet mining and processing operations requirements. In addition, there are bedrock and/or shallow alluvium aquifers on or near the Project area that supplied water for the 2014 and 2015 drill program; however, they are considered to be insufficient as a primary source of water supply for a mining operation.

The Project consists of sufficient area of land to the east of the deposit, which is suitable for mining and processing operations, waste rock storage area (“WRSA”) and dry stack tailings facility (“DSTF”) for a deposit of this size.

5.5        Physiography

The Project is located within the northern portion of the Santa Rita Mountains that form the western edge of the Mexican Highland section of the Basin and Range Physiographic Province of the southwest United States (Wardrop, 2005). The Basin and Range physiographic province is characterized by high mountain ranges adjacent to alluvial filled basins. The property occupies flat to mountainous topography in the northeastern and northwestern flanks of the Santa Rita Mountains.

Vegetation in the Project area reflects the climate with the lower slopes of the Santa Rita Mountains dominated by mesquite and grasses. The higher elevations, receiving greater rainfall, support an open cover of oak, pine, juniper and cypress trees.

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6        HISTORY

The early history and production from the Property has been described in Anzalone (1995), M3 (2012) and Briggs (2014) from which the following summarization is taken. Hudbay considers the mineral reserve and resource estimates referred to in this chapter (including the estimates prepared by Augusta) to be historical in nature since no work was done by a qualified person to verify such estimates and such estimates should not be relied upon.

6.1        Helvetia-Rosemont Mining District (1875 – 1973)

The first recorded mining activity in the Helvetia-Rosemont mining district occurred in 1875. The Helvetia-Rosemont mining district was officially established in 1878. Production from mines on both sides of the Santa Rita ridgeline supported the construction and operation of the Columbia Smelter in Helvetia and the Rosemont Smelter in Old Rosemont. Copper production from the district ceased in 1951 after production of about 227,300 tons of ore containing 17,290,000 pounds of copper, 1,097,980 pounds of zinc and 180,760 ounces of silver.

By the late 1950s, the Banner Mining Company (Banner) had acquired most of the claims in the area and had drilled the discovery hole into the Rosemont deposit. In 1963, the Anaconda Mining Co. acquired options to lease the Banner holdings and over the next ten years they carried out an extensive drilling program on both sides of the mountain for a total of 136,838 feet (41,708 m) from 113 drill holes. The exploration program demonstrated that a large scale porphyry/skarn existed at Rosemont. Regional exploration included targets at Broadtop Butte and Peach-Elgin prospects. In 1964, Anaconda produced a historical resource estimate for the Peach-Elgin deposit located in the Helvetia District. Based on assays from 67 churn and diamond drill holes, the estimate identified 14 million tons of sulfide material averaging 0.78% copper and 10 million tons of oxide material averaging 0.72% copper.

6.2        Anamax Mining Company (1973 - 1985)

In 1973, Anaconda Mining Co. and Amax Inc. formed a 50/50 partnership to form the Anamax Mining Co. In 1977, following years of drilling and evaluation, the Anamax joint venture commissioned the mining consulting firm of Pincock, Allen & Holt, Inc. to estimate a resource for the Rosemont deposit. Their historical resource estimate of about 445 million tons of sulfide mineralization averaged 0.54% copper using a cut-off grade of 0.20% copper. In addition to the sulfide material, 69 million tons of oxide mineralization averaging 0.45% copper was estimated. Subsequent engineering designed a pit based on 40,000 tons/day production rate for a mine life of 20 years.

In 1979, Anamax carried out a resource estimate for the Broadtop Butte deposit located about a mile north of the Rosemont deposit. Based on assays from 18 widely spaced diamond drill holes, a historical estimate identified 9 million tons averaging 0.77% copper and 0.037% molybdenum. In 1985, Anamax ceased operations and liquidated their assets. Today, most of the Anaconda/Anamax

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core is currently stored at Hidden Valley core storage facility at Project site. Hudbay considers the estimate done by Anamax to be historical in nature since no work has been done by a Hudbay QP to verify the estimate and the estimate should not be relied upon.

6.3        ASARCO, Inc. (1988 – 2004)

Asarco purchased the patented and unpatented mining claims in the Helvetia-Rosemont mining district from real estate interests in August 1988 and renewed exploration of the Peach-Elgin and initiated engineering studies on Rosemont. In 1995, Asarco succeeded in acquiring patents on 21 mining claims in the Rosemont area just prior to the moratorium placed on patented mining claims in 1996.

In 1999, Grupo Mexico acquired the Helvetia-Rosemont property through a merger with Asarco. During the 16 years of ownership by Asarco and Grupo Mexico, 11 diamond drill holes were completed for a total of 14,695 feet (4,479 m) at Rosemont. Asarco estimated historical reserves of 294,834,000 tons at 0.673% copper based on a mine production schedule with a strip ratio of 3.7:1. The Asarco drill core is currently stored at the Hudbay core storage facility on site. In 2004, Grupo Mexico sold the Rosemont property to a Tucson developer.

6.4        Augusta Resource Corporation (2005 – 2014)

In April 2005, Augusta purchased the property from Triangle Ventures LLC. Between mid-2005 and January 2007, Augusta drilled 55 diamond drill holes for a total of 96,129 feet (29,300 meters) in order to bring the resource estimate at Rosemont into compliance with NI 43-101 standards. The program was designed to better define the geology, distribution of copper mineralization as well as gather geotechnical data required to design a pit. In June 2006, Washington Group Int. completed a preliminary assessment and economic evaluation of the Project. Augusta submitted a mine plan of operations (“MPO”) to the USFS in July 2006. It was deemed incomplete and in 2007, Augusta resubmitted the MPO. Following a positive feasibility study conducted by M3 Engineering in August 2007 the Forest Service accepted a revised MPO in March 2008 marking the start of the formal EIS process mandated by the National Environmental Protection Act (“NEPA”) of 1969.

Over the next several years, Augusta continued to evaluate the mineral potential at Rosemont and refine the economics of developing this resource. A 20-hole diamond drilling program (17,522 feet or 5,341 meters) was conducted from December 2007 through July 2008. This was followed by a twelve-hole diamond drilling program (18,874 feet or 5,753 meters), which was completed in February 2012. A Technical Report issued by Augusta in 2012 estimated mineral reserves of 667.2 million tons at an average grade of 0.44% copper, 0.015% molybdenum and 0.12 ounces per ton of silver based on $4.90 per ton NSR cut-off using metal prices of $2.50/lb copper, $15.00/lb molybdenum and $20.00/oz silver. Augusta’s mineral resource estimate, shown in Table 6-1, is inclusive to their mineral reserves, stated above.

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Hudbay is treating Augusta’s publicly disclosed estimated mineral reserves and resources as a historical estimate under NI 43-101 and not as current mineral reserves and resources, as a Qualified Person has not done sufficient work for Hudbay to classify Rosemont’s mineral reserves or resources as current mineral reserves or mineral resources.

TABLE 6-1: HISTORICAL SULFIDE MINERAL RESOURCE (AUGUSTA 2012)

Category	Tons (millions)	Cu (%)	Mo (%)	Ag (oz/ton)
Measured	334.619	0.440	0.015	0.124
Indicated	534.735	0.373	0.014	0.105
Inferred	128.488	0.397	0.013	0.104

6.5        Hudbay (2014 – Present)

Hudbay completed a 43-hole, 92,909 feet (28,319 meters) drill program from September to December 2014 and a 46-hole, 75,164 feet (22,910 meters) drill program from August to November 2015. These drilling programs were completed in further efforts to gain a better understanding of the geological setting and mineralization of the deposit and to collect additional metallurgical and geotechnical information.

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7       GEOLOGICAL SETTING AND MINERALIZATION

7.1        Tectonic and Metallogenic Setting

Mesozoic subduction and associated magmatism and tectonism in the southwestern United States and northern Mexico, generated extensive and relevant porphyry copper mineralization (Figure 7-1). Compressional tectonism during the Mesozoic and early Cenozoic Laramide Orogeny caused folding and thrusting, accompanied by extensive calc-alkaline magmatism (Barra et al., 2005). The Laramide belt is a major porphyry province that extends for approximately 600 miles (1,000 km) from Arizona to Sinaloa, Mexico, and includes the Rosemont deposit and hosts a number of other world class deposits (e.g. Morenci, Resolution, and Cananea).

FIGURE 7-1: LARAMIDE BELT AND ASSOCIATED PORPHYRY COPPER MINERALIZATION (BARRA ET AL., 2005)

Tertiary extensional tectonism followed the Laramide Orogeny, accompanied by voluminous felsic volcanism (Barra et al., 2005). Steeply-to shallowly-dipping normal faults became active during this time, including rotational listric faulting. At Rosemont, it appears that tertiary faulting has significantly segmented the original deposit, juxtaposing mineralized and unmineralized rocks. The extensional tectonics culminated in the large-scale block faulting that produced the present basin and range geomorphology that is typical throughout southern Arizona (Maher, 2008).

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7.2        Regional Geology

The geology of the Santa Rita Mountains has recently been reviewed by Rasmussen et. al. (2012) identifying two main blocks (Figure 7-2). The northern block, where the Rosemont deposit lies, is dominated by Precambrian granite (brown on the map), with some slices of Paleozoic and Mesozoic sediments on the eastern and northern sides (blue and green on the map). This block includes small stocks and dikes of quartz monzonite or quartz latite porphyry that are related to porphyry copper and skarn mineralization.

7.3        District Geology

The Precambrian meta-sedimentary and intrusive rocks in the Rosemont area form the regional basement beneath a Paleozoic carbonate and siliciclastic sedimentary sequence (Figure 7-3 and Figure 7-4). Paleozoic sedimentary carbonate units are the predominant host rocks for the copper mineralization. Structurally overlying these predominantly carbonate units at Rosemont are Mesozoic clastic units, including conglomerates, sandstones, and siltstones. These clastic upper sequences have andesitic intercalations and also host mineralization. Quartz monzonite and quartz latite sill-shaped porphyries intruded both sequences and are associated with the porphyry/skarn mineralization. Tertiary conglomerates locally lay over the Mesozoic sedimentary units in late fault grabens. The Rosemont stratigraphy is summarized in Figure 7-4 and the geological configuration of the deposit is shown on a level section in Figure 7-5 and in a vertical section in Figure 7-6.

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FIGURE 7-2: SANTA RITA MOUNTAIN GEOLOGY (ADAPTED FROM DREWES ET AL., 2002)

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FIGURE 7-3: ROSEMONT REGIONAL GEOLOGY

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FIGURE 7-4: ROSEMONT STRATIGRAPHIC COLUMN

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FIGURE 7-5: ROSEMONT DEPOSIT GEOLOGIC – 4,000 FOOT LEVEL PLAN

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FIGURE 7-6: ROSEMONT DEPOSIT GEOLOGIC – 11,555,050 VERTICAL SECTION

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7.4        Chemostratigraphy

Hudbay 2014 and 2015 drilling programs included complete inductively coupled plasma multi-element assays (4 acid digestion) for every sample. The new data set (> 33,000 samples) was used to classify the different stratigraphic units according to their geochemical affinities. The original formations were grouped into equivalent chemostratigraphic units that reflect chemical changes induced by mixing of siliciclastic, dolomitic, and calcareous sediments as well as a hydrothermal component. The chemostratigraphic groups honour both the deposit stratigraphy and geochemical attributes and ultimately reflect the mineralogy (Figure 7-7). The geological model built implicitly in Leapfrog is based mainly on the downhole chemostratigraphy of the holes drilled between 2014 and 2015 (90 holes). The density of the chemostratigraphy data is higher in the pit area, where the new model was exclusively based on chemostratigraphy. In zones that are distal to the recent drilling (e.g. Backbone footwall domain), lithology (logged) data was incorporated into the model.

FIGURE 7-7: CHEMOSTRATIGRAPHY ROSEMONT DEPOSIT GEOLOGY

7.5        Structural Domains

The updated geological model incorporated a revised structural framework based on a surface and downhole structural review. The temporal and special relations between the main fault surfaces define 4 structural domains: Backbone Footwall, Lower Plate, Upper Plate and Graben Block (Figure 7-8).

The north trending, steeply dipping Backbone Fault juxtaposes Precambrian granodiorite and Lower Paleozoic quartzite and limestone marginally mineralized to the west (Backbone Footwall block) against a block of an homoclinical sequence of younger mineralized metamorphosed sedimentary units to the east (Lower Plate). A series of subparallel, anastomosing, curviplanar faults that generally strike north and dip steeply within the Lower Plate define a zone along the Backbone Fault strike.

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The Low Angle Faults are a series of shallowly east-dipping faults that are comprised of one major fault and a series of steep to shallow splay structures. The main Low Angle Fault forms the nonconformable contact between the Upper (siliciclastics and volcanics) and Lower Plate rocks.

The southeast-dipping Graben (60-65°) fault terminates mineralization continuity to the southeast and east. This fault postdates mineralization and has been interpreted as an extensional normal fault.

The east-west striking faults, including the Weigle Faults are a series of steeply-dipping, anastomosing structures that are oriented oblique to bedding and rock contacts. Locally, the most well-known fault is the Weigle Fault Zone, which displaces the Precambrian, Paleozoic, and Mesozoic rocks and generates a deepening of the oxide front.

FIGURE 7-8: ROSEMONT DEPOSIT GEOLOGICAL MODEL STRUCTURAL DOMAINS 3D VIEW (LOOKING NORTH)

7.6        Mineralization

Drilling to date at Rosemont has defined a mineral resource approximately 4,000 feet (1,200 meters) in diameter that extends to a depth of approximately 2,500 feet (750 meters) below the surface. The main fault systems partially delimit the defined resource, dividing the deposit into major structural blocks with contrasting intensities and types of mineralization. The north-trending, steeply dipping Backbone Fault juxtaposes marginally mineralized Precambrian granodiorite and Lower Paleozoic quartzite and limestone to the west (Back Bone Footwall Block) against a block of younger, well-mineralized Paleozoic limestone units to the east (Lower Plate).

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Most of the copper sulfide resource is contained in the eastern hanging wall of the Backbone Fault. Structurally overlying the sulfide resource is a block of Mesozoic sedimentary and volcanic rocks (Upper Plate) that contains lower grade copper mineralization (predominantly as oxides). These two blocks are separated by the shallowly dipping Low Angle Fault (“LAF”). Other post-mineral features include a deep, gravel-filled Tertiary paleo valley on the south side of the deposit and a significant thickness of Cretaceous and Tertiary volcaniclastic material to the northeast of the deposit.

Sulfide mineralization on the east side of the Backbone Fault and below the LAF is hosted in an east-dipping package of Paleozoic-age sedimentary rocks that includes the Escabrosa Limestone, Horquilla Limestone, Earp Formation, Colina Limestone, and Epitaph Formation.

Relatively minor mineralization occurs in the other Paleozoic units. To the south, the mineralization in this block appears to weaken and eventually die out. To the north, mineralization appears to narrow but continues under cover amid complex faulting (Weigles Fault system). To the east, mineralization is covered by an increasingly thick block of Mesozoic sediments due to normal faulting (east block down) along the graben fault.

The Mesozoic rocks of the structural block above the LAF consist predominantly of arkosic siltstones, sandstones, and conglomerate. Subordinate andesite flows or sills within the Arkose range from a few tens of feet to several hundred feet thick. Also, structurally wedged into the Upper Plate block at the base of the Arkose is the Glance Conglomerate, a limestone-cobble conglomerate, and some occurrences of relatively fresh Paleozoic formations.

New QUEMSCAN® data from 107 composite samples (averaging 30 feet (9.1 meters) of core each) collected from Augusta and Hudbay drilled core provides a preliminary mineralogical characterization of the Rosemont deposit. In bulk terms, the total sulfide volume content of the non-oxidized mineralized skarn is less than 2.7% . Pyrite and chalcopyrite comprise approximately 25% and 35% of the total sulfides content, respectively; along with bornite (20%) and chalcocite (12%). The ratio of these main sulfide minerals is variable through the stratigraphy of the deposit owing to competing, over-printing pulses of mineralization and possible supergene effects. Mineralization in the Horquilla formation is richer in bornite and chalcocite (40% and 35% of total sulfides, respectively) and lower in pyrite and chalcopyrite (5% and 15% respectively) compared to the other mineralized units. Molybdenite is a minor phase but appears to be distributed throughout the skarn and in peripheral portions of the deposit. Gold and silver are present in small amounts across the deposit and are thought to be contained in the primary sulfide minerals. Chalcocite, covellite, native copper and a suite of other secondary copper oxide and carbonate minerals are found in fault and fracture zones in the skarn.

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7.7        Mineralization Domains

Three mineralization domains (oxide, mixed and sulfide) were defined based on the soluble to total copper ratio (“ASCu/Tcu”) collected in the Augusta (2005-2012) and Hudbay (2014 and 2015) drilling programs. The Augusta analytical protocol included soluble copper assays only in zones where copper oxides were observed in core. Hudbay’s 2014 and 2015 programs included soluble copper assays for all samples regardless of the dominant logged mineralogy.

For the domains definition the ASCu/TCu ratio was modelled in Leapfrog using samples with TCu > 0.05% . Two ASCu/TCu ratio shells were interpolated (spheroidal indicator interpolant) for values of > 0.3 and > 0.5. The remaining part of the deposit constitutes the sulfide domain.

The oxidized zones including the 0.3 – 0.5 ASCu/TCu (Mixed) and the > 0.5 ASCu/TCu (Oxide) in part are bounded by a continuous blanket with a gentle east-dipping attitude. The blanket is defined by a sharp decrease in ASCu/TCu ratios that coincides with the LAF.

Other irregular oxidized zones are located in the hanging wall of the Backbone Fault with special development at the intersection with the Weigles Butte Fault, where a bulbous body of mixed mineralization projects deeply down-dip into the Horquilla Formation.

Some significant differences were noted between the current and previous domaining outcomes (Figure 7-9). Augusta domain modelling was built on down hole coding based on the soluble copper data selectively collected in some holes and visual examination of core. The new modelling is based exclusively on analytical data spread throughout the mineralized zones, including the hole core length unbiased acid soluble data collected by Hudbay. This new approach allows refining the shape and continuity of the domains within the upper and lower plate. In the upper plate the oxide blanket topography is better resolved and in the lower plate the continuity of the mixed and oxide zones associated with faults and fractures is better constrained.

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FIGURE 7-9: MINERALIZATION DOMAINS SECTION 11,555,500 N

7.8        Alteration and Skarn Development

The Rosemont deposit consists of copper-molybdenum-silver-gold mineralization primarily hosted in skarn that formed in the Paleozoic rocks as a result of the intrusion of quartz latite to quartz monzonite porphyry intrusions. Bornite-chalcopyrite-molybdenite mineralization occurs as veinlets and disseminations in the skarn.

Garnet-diopside-wollastonite skarn, which formed in impure limestone, is the most important skarn type volumetrically. Diopside-serpentine skarn which formed in dolomitic rocks is less significant. Marble was developed in the most purest carbonate rocks, while the more siliceous, silty rocks were converted to hornfels. Both marble and hornfels are relatively poor hosts to mineralization. The main skarn minerals are accompanied by quartz, potassium feldspar, amphibole, magnetite, epidote, chlorite and clay minerals. Quartz latite to quartz monzonite intrusive rocks host strong quartz-sericite-pyrite alteration with minor mineralization. Where the mineralized package of Paleozoic rocks and quartz-latite intrusive outcrop on the western side of the deposit, near surface weathering and oxidation has produced disseminated and fracture-controlled copper oxide minerals.

The Mesozoic and lesser Paleozoic rocks above the LAF are propylitically altered to an assemblage including epidote, chlorite, calcite, and pyrite. Copper mineralization is irregularly developed. The rocks are commonly deeply weathered and limonitic. The original chalcopyrite is typically oxidized to chrysocolla, copper wad and copper carbonates. Supergene chalcocite is locally present.

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7.9        Clay Proxies

Geochemical proxies were developed using mineralogy data paired to multi-element geochemistry. This allowed populating the whole length of the core drilled in 2014 and 2015 with relevant clay estimates, and modelling the distribution of both Mg and swelling clays within the deposit.

Magnesium-rich clays (“Mg-clays”), defined as the combined percentage of talc + serpentine, were analyzed by XRD and QEMSCAN as part of the metallurgical work. A total of 431 samples were analyzed by XRD and 107 samples were analyzed by QEMSCAN.

Swelling clays (“Sw-clays”) were analyzed using cation exchange capacity (“CEC”) for a total of 431 samples also for metallurgical work. Sw-clays were collectively determined as a group in which individual swelling clay types were not discriminated.

For Mg-clays, the QEMSCAN data was used as a training data set to fit a multi-linear regression (MLR) using the multi-element geochemical analysis as input variables. QEMSCAN was preferred, given that it has better detection limits than XRD mineralogy. The XRD data has approximately 83% of observations below the detection limit for Mg-clays and, accordingly, it is not suitable to fit the models. For Sw-clays, the CECF proxies data was used as training data set to fit a MLR model using the multi-element geochemical analysis as input variables.

Higher magnesium clay content is generally associated with a dolomitic protore (e.g. Epithaph) and higher swelling clays with the siliciclastic component of the chemical sedimentary rocks (e.g. Earp).

7.9.1        Ore Types

A 6 node classification and regression tree (“CART”) was used to classify ore types using 107 measurements of total copper rougher recoveries (“RCu %”) as the response variable. The bond work index (“BWI”), sag power index (“SPI”), Sw-clay, Mg-clay; talc + serpentine, and the ratio of soluble copper relative to total copper (pctCuox = Soluble Cu/TCu), a proxy for oxidation of ore minerals, were used as predictor variables in the CART model. The results of the regression tree indicate that the most important variables in hierarchical order are pctCuox, Sw-clay, and Mg-clay.

The CART model suggests that the Rosemont deposit can be subdivided in at least 6 ore types (Figure 7-10). Ore types 1 and 2 are considered clean material. Ore types 3 and 4 are clay-rich material; in which ore type 4 is Mg-clay rich. Ore types 5 and 6 are highly oxidized ores including mixed material (Ore 5) and oxide (Ore 6) material (Figure 7-10).

Once the ore types were established, the geochemical data of these groups were used to create proxies for ore types of the entire deposit. Mineralogical data collected for metallurgical purposes was used to train geochemical data. Mg-clays were analyzed by QEMSCAN (107 samples). Sw-clays were analyzed using CEC for a total of 431 samples. Sw-clays were collectively determined as a group in which; individual Sw-clay types were not discriminated.

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Mixed and oxide materials (Ore 5a and Ore 5b, respectively) were modeled using the pctCuox. Geochemical proxies for both Sw-Clays and Mg-clays were developed. For Mg-clays, the QEMSCAN data was used as a training data set to fit a multi-linear regression (“MLR”) using the multi-element geochemical analysis as input variables. For Sw-clays, the CEC data was used as training data set to fit a MLR model using the multi-element geochemical analysis as input variables.

FIGURE 7-10: ORE TYPES CART MODEL

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8        DEPOSIT TYPE

The Rosemont deposit consists of copper-molybdenum-silver-gold mineralization primarily hosted in skarn that formed in the Paleozoic rocks as a result of the intrusion of quartz latite to quartz monzonite porphyry intrusions. Genetically, skarns form part of the suite of deposit styles associated with porphyry copper centers, although intrusive rocks are volumetrically minor within the resource area. The skarns were formed as the result of thermal and metasomatic alteration of Paleozoic carbonate and to a lesser extent Mesozoic clastic rocks. Near surface weathering has resulted in the oxidation of the sulfides in the overlying Mesozoic units.

Mineralization is mostly in the form of primary (hypogene) copper, molybdenum and silver bearing sulfides, found in stockwork veinlets and disseminated in the altered host rock. Some oxidized copper mineralization is also present in the upper portion of the deposit. The oxidized mineralization is primarily hosted in Mesozoic rocks, but is also found in Paleozoic rocks on the west side of the deposit and deeper along some faults. The oxidized mineralization occurs as mixed copper oxide and copper carbonate minerals. Locally, enrichment of supergene chalcocite and associated secondary mineralization are found in and beneath the oxidized mineralization.

The Twin Buttes Mine, operated by Anaconda and later by Cyprus, was developed on a deposit with a number of geologic similarities, located approximately 20 miles (32 kilometers) to the west of Rosemont. The Twin Buttes mine was in production from 1969 to 1994. In addition, the Asarco Mission Mine, also located about 20 miles (32 kilometers) to the west of Rosemont, has some common geologic characteristics.

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9        EXPLORATION

Prospecting began in the Rosemont and Helvetia Mining Districts in the mid-1800s and by 1875 copper production was first recorded, which continued sporadically until 1951. By the late 1950s, exploration drilling had discovered the Rosemont deposit. A succession of major mining companies subsequently conducted exploratory drilling of the Rosemont deposit and the nearby Broadtop Butte, Peach Elgin and Copper World mineralized areas.

Augusta acquired the Rosemont property in 2005 and performed infill drilling at the Rosemont deposit along with exploration geophysical surveys. A Titan 24 induced polarization/resistivity (“DCIP”) survey over the Rosemont deposit, performed in 2011, discovered significant chargeability anomalies which are partially-tested. These anomalies appear to define mineralization and also certain unmineralized lithologic units. A regional scale airborne magnetics survey was also completed in 2008.

Two infill drilling campaigns were completed by Hudbay in and beneath the Rosemont deposit in the fall of both 2014 and 2015. In addition to chemical assaying, magnetic susceptibility and conductivity measurements were taken using the Terraplus’ KT-10 & KT-20 instruments at approximately every 10-feet (3 meters) intervals of recovered core from the drilling program. The magnetic susceptibility data has been used from both drilling programs as a constraint for a 3D inversion of the deposit with an interpretation in progress. A single test-line of DCIP data was collected over the Rosemont deposit using the DIAS Geophysical (3D Survey/Mapping) in April 2015 for comparison to the previously completed Titan 24 survey.

Hudbay analyzed all samples of the 2014 and 2015 drilling programs with ICP multi-element geochemistry. This new geochemical data set was used to classify rocks according to chemical indexes in a ternary diagram defined by Siliciclatic, Limestone and Dolomitic vertices. The lithogeochemical groups honour the deposit stratigraphy and geochemical attributes and proved to be a useful tool for geological modeling and vectoring.

A mapping and geochemical sampling program was completed in the latter half of 2015 on the Rosemont property to reassess the interpretation of the regional geology and deposit setting. This was followed by a structural interpretation using both surface and drill core measurements to aid in the geotechnical evaluation of the Project.

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10      DRILLING

Extensive drilling has been conducted at the Rosemont deposit by several successive property owners. The most recent drilling was by Hudbay, with prior drilling campaigns completed by Banner, Anaconda Mining Co., Anamax and Asarco and Augusta. Table 10-1 summarizes the drill holes used to estimate the current mineral resource estimate, with regional exploration holes excluded.

TABLE 10-1: ROSEMONT DEPOSIT DRILLING SUMMARY

Company	Time Period	Drill Holes
		Number	Feet	Meters
Banner Mining	1950s to 1963	3	4,300	1,311
Anaconda Mining	1963 to 1973	113	136,838	41,708
Anamax	1973 to 1986	52	54,350	16,566
ASARCO	1988 to 2004	11	14,695	4,479
Augusta	2005 to 2012	87	132,525	40,394
Hudbay	2014 to 2015	90	168,286	51,294
Total		355	510,780	155,686

The drill holes in the database were all drilled using diamond drilling (coring) methods. In some cases, the top portion of the older holes were drilled using a rock bit to set the collar or by rotary drilling methods and then switching to core drilling before intercepting mineralization. A map showing the location of the drill holes by company is provided in Figure 10-1 along with an outline of the mineral resource pit shell limits for the Rosemont deposit. Exploration holes drilled using rotary or older “churn” drill holes were excluded from the resource database.

In all of the drilling campaigns, efforts were consistently made to obtain representative samples by drilling either H-size (2.5 inch or 63.5 mm diameter) or N-size (1.9 inch or 47.6 mm diameter) core.

Core recoveries within the ore zone for the Hudbay and Augusta drilling programs average 96% and core recoveries within the pit elsewhere average 89%, lending confidence that quality samples were obtained including for oxidised intervals. Generally, drill programs were on east-west grid lines spaced approximately 200 feet (61 meters) apart.

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FIGURE 10-1: ROSEMONT DEPOSIT DRILL HOLE LOCATIONS BY COMPANY

The majority of the Anaconda Mining Co., Anamax and Asarco drill core is available and was systematically re-logged by Augusta personnel to be geologically consistent with their drilling from 2005 to 2012. In addition, with re-logging, they completed some re-sampling for geochemical analyses.

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10.1        Banner Mining Company (1961 to 1963)

The first significant core drilling campaign on the Property was by the Banner, beginning in about 1961. Banner completed primarily shallow diamond drill holes, many of which were subsequently deepened by Anaconda Mining Co. Three drill holes included in the resource database were shallow holes initially collared by Banner and were significantly deepened during subsequent drilling programs conducted by Anaconda Mining Co.,. These holes have a combined length of 4,300 feet (1,311 meters).

10.2        The Anaconda Mining Co., (1963 to 1986)

Anaconda acquired Banner Rosemont Holdings around 1963 and conducted exploration at the Rosemont deposit and in adjacent mineralized areas. Between the years of 1963 and 1973, they completed 113 diamond drill holes at Rosemont for a total of 136,838 feet (41,708 meters). These holes were primarily drilled vertically. Down-hole and collar surveys completed by company surveyors were conducted during drilling or immediately following drill hole completion for selected holes. Anaconda drilled approximately 85% of the larger N-size core and 15% of the smaller B-size core (1.4 inch or 36.4 mm diameter). Overall core recovery was more than 85%.

Exploration subsequently transferred to Anamax Mining Co., (an Anaconda Mining Co., and Amax Inc., joint venture) around 1973, which continued extensive diamond drilling and analytical work until 1986. Anamax completed 52 core holes for a total of 54,350 feet (16,566 meters). These holes were almost exclusively drilled as angle holes inclined -45° to -55° to the west, approximately perpendicular to the east-dipping, Paleozoic, metasedimentary host rocks. Down-hole and collar surveys by company surveyors were conducted during drilling or immediately following drill hole completion for the majority of the holes. Anamax drilled approximately 80% N-sized core and 20% B-sized core, with an overall core holes recovery of more than 88%.

10.3        ASARCO Mining Co., (1988 to 2004)

Asarco acquired the Rosemont property in 1988 and conducted exploration until 2004, completing 11 vertical drill holes for a total of 14,695 feet (4,479 meters) in the deposit area (a 12th hole was drilled to the east of the deposit and is not part of the Project’s database). Data was available from eight of the Asarco core holes in the Rosemont deposit area and were incorporated into Hudbay’s resource estimate. Down-hole survey data, if taken, were not available for the Asarco holes. Drill hole collars were surveyed by company surveyors. The size of core collected by Asarco was predominantly N-sized. Core recovery information was not available but re-logging by Augusta personnel indicated it to be of similar quality to that of other drilling campaigns.

10.4        Augusta Resource (2005 to 2012)

Augusta optioned the Rosemont property in 2005 and conducted diamond drilling in several campaigns, from 2005 to 2012. In total, Augusta completed 87 core holes for a total of 132,525 feet (40,394 meters). Of these, 60 holes were drilled for the purposes of delineating the deposit and providing infill information, while six were exploration holes outside of the planned pit area, but close enough to be a part of the resource database. The remaining 21 core holes support geotechnical (13) or metallurgical (8) studies. Augusta holes were usually collared by rock-bitted through overburden, and then drilled with larger HQ-sized core as deeply as possible and finished with NQ-sized core if a reduction in core size was required due to ground conditions.

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Most of the holes were oriented vertically, although a few of the holes were inclined to intercept targets from reasonably accessible drill pad locations. All drill holes were down-hole surveyed using a Reflex EZ-Shot survey instrument which measures inclination/dip and azimuth direction, with measurements generally taken every 100 feet (30 meters) down the hole during 2008 and every 200 or 500 feet (61 or 152 meters) down the hole during 2005, 2006 and 2011 to 2012 drill campaigns. The initial drill hole collar locations were surveyed by Putt Surveying of Tucson, Arizona, while all later drilling locations were measured and certified by Darling Environmental & Surveying of Tucson, Arizona.

10.5        Hudbay (2014 to 2015)

Shortly after acquiring the Project, Hudbay initiated a 44 core hole drill program in September 2014 and completed 93,122 feet (28,384 meters) of diamond drilling by December 2014. The Phase I drill program was conducted entirely within the Rosemont resource, on patented claims and was designed to gain an initial understanding of the geological setting and mineralization, provide infill drilling density along with metallurgical, geochemical and geophysical data.

Diamond drilling was primarily HQ-sized core as deeply as possible and finished with NQ-sized core, if a reduction in core size was required due to ground conditions. If ground conditions warranted, drill holes were collared in larger PQ size (3.3 inch or 83 mm diameter) and reduced to HQ as ground conditions improved. Drilled length and respective recoveries were PQ 4,326 feet (1,319 meters) with 83.5% recovery, HQ 85,583 feet (26,086 meters) with 95.9% recovery, and NQ 3,213 feet (979 meters) with 92.8% recovery (statistics include HB-2119 that was abandoned due to poor ground conditions after drilling approximately 200 feet (60 meters).

Forty-three of the drill holes were orientated vertically, with one inclined in order to intercept a target area from an accessible drill pad location. Down hole surveying was conducted on 200 feet (61 meters) intervals with either a Multishot Reflex or a Surface Recording Gyro Survey instrument, both instruments measured inclination/dip and azimuth direction. Collar locations were surveyed and certified by Darling Environmental & Surveying of Tucson, Arizona

From August to November 2015, Hudbay completed a 46 core hole, 75,164 feet (22,910 meters) diamond drill program. The Phase II drill program was conducted entirely within the Rosemont resource, on patented claims and was designed to gain a further understanding of the geological setting and mineralization, provide infill drilling density along with metallurgical, geotechnical, geochemical and geophysical data.

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Diamond drilling was primarily HQ-sized core as deeply as possible and finished with NQ-sized core, if a reduction in core size was required due to ground conditions. If ground conditions warranted, drill holes were collared in larger PQ size and reduced to HQ as ground conditions improved. Twenty-two of the drill holes were oriented vertically, with 24 inclined drill holes. Eight holes were inclined for drilling oriented core utilizing the Reflex ACT III instrument to gather geotechnical structural data, and 16 holes were inclined in order to intercept a target area from an accessible drill pad location. Down hole surveying was conducted on 200 feet (61 m) intervals with either a Multishot Reflex or a Surface Recording Gyro Survey instrument, both instruments measured inclination/dip and azimuth direction. Collar locations were surveyed and certified by Darling Environmental & Surveying of Tucson, Arizona.

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11      SAMPLING PREPARATION, ANALYSES, AND SECURITY

11.1        Hudbay 2014

11.1.1        Core Logging

The drilling contractors thoroughly cleaned the drill core retrieved from the core tube before piecing all the segments together in the core boxes. Footage marker blocks were inserted in the core boxes after each run to indicate the relative down-hole depth. Core boxes were labelled with the hole name, box number and from - to footage measurements before securely closing the box with a tightly fitted lid. Core boxes were delivered to the core processing areas of either Rosemont Camp or Hidden Valley Ranch by the drilling contractors, and neatly stacked on top of pallets. Private 24-hour per day security guards administered by Securitas Inc., controlled site access and oversaw sample security at each camp and drill site.

Core boxes were loaded onto conveyor racks by the geotechnicians and geologists for logging. Prior to measuring the core recovery parameters and Rock Quality Data (“RQD”), visual checks were performed for incorrect placement and orientation of core fragments. Any discrepancies caused by mislabeled or misplaced footage tags were resolved by consulting the drilling contractors. The drill core was marked with cut lines designed to provide the most representative split.

Standard parameters for core recovery and RQD for each drill run were measured by either the trained geotechnicians or geologists and recorded on tablets. The RQD program was administrated and monitored by the consulting engineering firm CNI. All core logging was completed by experienced contract geologists. At the start of each drilling campaign, all geologists were provided with three days of training on the rock types, alterations, mineralization and structures found on the property.

All drill holes were logged using tablets networked to a FileMaker Inc. (FileMaker) database hosted on a laptop using a local hotspot network. Drill core was divided into sub intervals based on the rock types observed by the geologists. A local formation name was assigned to each interval if a positive identification was made. Each interval was further described for alteration, mineralization, and oxidation state of the primary sulfides. Any significant veins found were also logged along with identifiable structures.

11.1.2        Sample Selection

All core samples for assaying were assigned by the core logging geologist. The typical sample interval was about 5 feet (1.5 meters) while being mindful of lithological contacts. Geologists were responsible for filling out two of the three paper tags for each sample in the sampling book with the hole name and sample interval. Sample tag numbers along with the sampled intervals were also entered into the core logging database. For core samples, two of the three tags were stapled into the core box at the starting point of each sample, one to remain there, and the second to accompany the sampled split in the sample bag. Lines were drawn on the core using a permanent marker to indicate the beginning and end of each sample. For QA/QC samples consisting of duplicates (two coarse rejects and analyses to be generated from the same interval), two sets of sample tags were stapled into the core box, and a double line was drawn on the core. For other QA/QC samples (standards and blanks), a single sample tag was stapled into the core box indicating the QA/QC sample’s relative position in the sequence and the sample type.

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11.1.3        Core Photographs

Completely logged core boxes with the sample intervals marked and sample tags inserted were passed on to geotechnicians who photographed all core boxes individually or in groups of two using a digital camera (Nikon Coolpix L830) mounted to a tripod in natural light. The hole name, box numbers and the from and to intervals were written on a white board, which was photographed with the core boxes along with a color patch and a scale for reference. All photos taken were loaded on to a laptop computer dedicated to core photos and reviewed by the on-site database manager or the lead geotechnician. Any photos deemed unacceptable were retaken by the geotechnicians. All core photos were renamed by the site database manager based on hole name, box numbers and intervals and uploaded to the Google Drive.

11.1.4        Core Cutting

Prior to cutting core, the database manager printed a sample list for each drill hole that included the sample identification number, hole name, sample type and the start and end footage of each sample. This list was used by the geotechnicians to label sample bags. At the core cutting station, a bucket was lined with the correctly labelled sample bag and the corresponding core box was placed on to the work table next to the core saw. The core cutters separated one of the two sample tags stapled in the core box at the start of the sample and placed them in the sample bag. For coarse duplicate samples, the second QA/QC sample tag was also placed in the same bag. Core was cut along the cut line drawn by the geologist and the right half of the sample was placed in the sample bag and the left was placed back in the core box. In gouge and rubble intervals, an aluminium sampling scoop was used to separate the gouge into two halves in the core boxes and the right half was scooped into the sample bag. Completed sample bags were closed using the bag draw strings and secured at the neck using two zip ties. These bags were moved to a dry storage area away from the core cutting saws and stacked in orderly rows. All saws and sampling buckets were rinsed with water after cutting each sample to prevent cross contamination.

11.1.5        Sample Dispatching

Samples were dispatched using the dispatching module in the core logging database. Samples were dispatched typically in batches of 100 samples from the same drill hole. A requisition form was automatically created that listed the range of sample numbers, job order number, requested analytical codes and any special instructions. A corresponding sample list for each requisition was also created. The requisition form and the sample list were emailed to the preparation laboratory prior to sample shipment. QA/QC samples including Blanks and Standards were prepared by the database manager prior to sample shipment. On the day of sample shipment, sample bags were cross-checked with the sample requisition list before packing them into large sacks placed on wooden pallets. These sacks were secured to the wooden pallets using shrink wrap and strap cables before being loaded on the truck.

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11.1.6        Sample Preparation

Drill core samples were picked up at the core processing facilities and transported via UPS to Inspectorate America Corporation’s (“Inspectorate”) preparation facility at Sparks, Nevada, USA. Samples were weighed upon arrival, dried at 60°C, and crushed in jaw crushers to ≥70% passing through 10 mesh (2 mm). The entire crushed sample was homogenized, riffle split, and a 1,000 g subsample was pulverized to ≥85% passing through 200 mesh (75 μm) using Essa standard steel grinding bowls. Jaw crushers, preparation pans, and grinding bowls were cleaned by brush and compressed air between samples. Cleaning with a quartz wash was conducted between jobs and between highly mineralized samples.

Once samples were pulverized a 150 g subsample pulp was collected and air freighted to Bureau Veritas Commodities Canada Ltd., (“Bureau Veritas”) in Vancouver, Canada, for analysis. The remaining 850 g master pulps and the coarse rejects were stored temporarily at the Inspectorate laboratory and then moved to a storage facility in Tucson.

Bureau Veritas is independent from Hudbay and has a quality system that is compliant with the International Standards Organization (“ISO”) 9001 Model for Quality Assurance and ISO/IEC 17025 General Requirements for the Competence of Testing and Calibration Laboratories.

The sample preparation, analysis, and security procedures are considered industry standard, adequate, and acceptable.

11.1.7        Bulk Density

A total of 922 samples in 43 drill holes were collected for specific gravity determinations at the Inspectorate preparation facility. The drill core samples, 4 to 6 inches (10 to 15 cm) long, were taken every 100 feet (30 meters) from half-split core.

At the laboratory, samples were dried at 105°C overnight and then allowed to cool to room temperature. The initial weight of the sample is determined using a top loading balance and recorded. Balances are calibrated using a 10 g, 50 g and 250 g calibration weight. The sample is immersed in a pan containing molten paraffin, then immediately removed from the molten paraffin and shaken a few times to remove excess wax while hardening. The wax coated sample is reweighed using the top loader and the weight is recorded. The standard water displacement method is then used to calculate the specific gravity.

The method is industry standard and suitable for bulk density determinations.

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11.1.8        Assay Methodology

A total of 18,361 drill core samples were analyzed for 45 elements by Inductively Coupled Plasma Mass Spectrometry (ICP-MS) after 4-acid digestion (Method MA200). The specifics of the analyses for copper, molybdenum and silver are given in Table 11-1. Samples with concentrations above the upper limit of detection and those with copper ≥ 8,000 ppm and molybdenum ≥ 3,200 ppm were systematically re-assayed by high-grade 4-acid digestion, Method MA370, and ICP-OES. The 8,000 and 3,200 ppm copper and molybdenum grade thresholds, respectively, were selected to maintain accuracy at grade levels within 20% of the upper limit of detection in Method MA200.

TABLE 11-1: BUREAU VERITAS ASSAY SPECIFICATIONS

Element	Cu	Cu over limits	Cu Soluble	Mo	Mo over limits	Ag	Au
Unit	ppm	%	%	ppm	%	ppm	ppm
Lower Detection Limit	0.1	0.001	0.001	0.1	0.001	0.1	0.005
Upper Detection Limit	10,000	-	10	4,000	-	200	10
Digestion	4 acids	4 acids	Sulfuric acid at 5%	4 acids	4 acids	4 acids	Fire assay
Instrumental Finish	ICP-MS	ICP-OES	AAS	ICP-MS	ICP-OES	ICP-MS	AAS
Method Code	MA200	MA370	GC921	MA200	MA370	MA200	FA430

To investigate the oxidation of primary copper sulfides, all drill core samples were analyzed for acid soluble copper (“ASCu”) Method GC921 with an acid leach at room temperature using 5% sulfuric acid (H₂SO₄). Samples were agitated in a mechanical shaker for one hour, and then made up to volume with demineralised water. The solution was filtered and analyzed by Atomic Absorption Spectroscopy (“AAS”).

Gold was analyzed in all samples from seven selected drill holes across the Rosemont deposit. A total 3,155 samples were analyzed for gold; which represents approximately 18% of Hudbay’s 2014 drilling and sampling program. Gold was determined by lead-collection fire assay fusion, for total sample decomposition, and AAS instrumental finish (“Method FA430”). Fire assays were performed on 30 g subsample pulps to circumvent potential problems due to nugget effect.

As part of Hudbay’s QA/QC program, QA/QC samples, shown in Table 11-2, were systematically introduced in the sample stream to assess adequate sub-sampling procedures, potential cross-contamination, precision, and accuracy. On average, the sampling program included 6% certified reference materials (“CRM”), 6% certified blanks, and 6% coarse duplicates. Blanks and CRMs were prepared by Ore Research and Exploration (“OREAS”) in Australia. All QA/QC samples were analyzed following the same analytical procedures as those used for the drill core samples.

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TABLE 11-2: SUMMARY OF QA/QC SAMPLES

Category	No. of samples	Relative Frequency1	Type	Elements
OREAS 501b	214	1.2%	CRM	Cu-Mo-Ag-Au
OREAS 502b	211	1.1%	CRM	Cu-Mo-Ag-Au
OREAS 503b	210	1.1%	CRM	Cu-Mo-Ag-Au
OREAS 504b	198	1.1%	CRM	Cu-Mo-Ag-Au
OREAS 902	55	0.3%	CRM	Cu-Mo-Ag- Cu Soluble
OREAS 930	179	1.0%	CRM	Cu-Mo-Ag
OREAS 22d	534	3.0%	Certified Blank	Cu-Mo-Ag-Au
OREAS 26a	554	3.0%	Certified Blank	Cu-Mo
Duplicates	1,086	6.0%	Coarse Duplicate	Cu-Mo-Ag
Duplicates	189	5.9%2	Coarse Duplicate	Cu-Mo-Ag-Au

¹Frequencies estimated relative to a sampling program comprising 18,361 samples
²Frequencies estimated relative to 3,155 samples analyzed for gold

11.1.9        Blanks

Certified OREAS blanks, shown in Table 11-3, were inserted into the sample stream approximately one every twenty samples to monitor potential cross-contamination.

TABLE 11-3: OREAS CERTIFIED BLANKS

	Cu	Mo	Ag	Au
Unit	ppm	ppm	ppm	ppb
OREAS 22d	9.23	2.36	<0.1	<1
OREAS 26a	50	1.50	-	-
Certified Method	4 acids	4 acids	4 acids	Fire assay

Fine and coarse blanks were systematically inserted at the same rate for a total of 1,088 blanks representing 5.9% of the sampling program. OREAS 22d is a certified fine blank prepared from quartz sand. OREAS 26a is a certified coarse blank sourced from fresh and non-mineralized olivine basalt.

Fine blank OREAS 22d and coarse blank OREAS 26a were used to assess potential contamination during assaying and sample preparation, respectively. These blanks contain low trace level concentrations of copper, molybdenum, silver, and gold. Blank failure due to potential contamination issues is documented when the blank values exceed five times the lower limit of detection. For those blanks with concentration levels above the lower detection limits the failure thresholds are set to values that exceed the certified best value (“CBV”) plus three standard deviations. A summary of the blank performance is shown in Table 11-4.

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TABLE 11-4: SUMMARY OF BLANK PERFORMANCE

Fine Blank OREAS 22d
Element	No. of Blanks	Failed Blanks	Failure Rate	Maximum Contamination of CBV	Average Contamination of CBV
Cu	534	32	6.0%	7.4 ppm	1.7 ppm
Mo	534	12	2.2%	4.0 ppm	1.2 ppm
Ag	534	1	0.2%	3.2 ppm	3.2 ppm
Au	89	0	0.0%	0	0
Coarse Blank OREAS 26a
Element	No. of Blanks	Failed Blanks	Failure Rate	Maximum Contamination of CBV	Average Contamination of CBV
Cu	554	39	7.0%	294 ppm	40 ppm
Mo	554	8	1.4%	69 ppm	19 ppm
Ag	554	1	0.2%	0.1 ppm	0.1 ppm
Au	98	0	0%	0	0

A total of 1,088 blank samples were systematically inserted along with the drill core samples and analyzed at Bureau Veritas. Contamination with copper and gold was insignificant. There are very few isolated cases of contamination at high-grade levels for silver and molybdenum. However, the overall blank failure rates are very low ranging from 0 to 7%.

The performance of blanks indicates no significant issues with contamination and therefore it is concluded that the results are acceptable and adequate for the resource estimation.

11.1.10        Standards

OREAS certified reference materials, as shown in Table 11-5, were inserted one every twenty samples. In total, 1,067 CRMs were analyzed for a total insertion rate of 5.8% .

TABLE 11-5: OREAS CERTIFIED REFERENCE MATERIAL

Unit	Cu %	Mo ppm	Ag ppm	Au ppm	Cu Soluble %
OREAS 501b	0.260	99	0.778	0.248	-
OREAS 502b	0.773	238	2.090	0.495	-
OREAS 503b	0.531	319	1.540	0.695	-
OREAS 504b	1.110	499	3.070	1.610	-
OREAS 902	0.301	12.2	0.343	-	0.111
OREAS 930	2.520	<1.5	9.0	-	-
Certified Method	4 acids	4 acids	4 acids	Fire assay	Sulfuric acid at 5%

OREAS 501b to 504b represent a blend of porphyry copper-gold mineralization, barren gangue, and minor quantities of copper and molybdenum concentrate. Copper and gold mineralization occurs as stockwork quartz veins and disseminations associated with potassic alteration. Primary copper sulfides include bornite and chalcopyrite.

OREAS 902 was prepared from oxidized copper ore hosted in dolomitic, carbonaceous, and argillaceous sedimentary rocks. Copper oxides consist primarily of malachite, cuprite, chrysocolla, and chalcocite. Chalcopyrite is the primary copper sulfide.

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OREAS 930 was prepared from a copper orebody hosted in carbonaceous siltstones and mudstones. Sulfides include chalcopyrite, bornite, pyrrhotite, pyrite, sphalerite, galena, and cubanite.

More than 30 and up to 214 samples were analyzed per CRM, as summarized in Table 11-6, which provide sufficient information to set acceptance criteria relative to the average (“AV”) and standard deviation (“SD”) of the actual assay values of the CRMs. However, if the absolute relative bias is >10% the acceptance criteria is set relative to the CBV and standard deviation recommended by the CRM certificates.

TABLE 11-6: SUMMARY OF CRM PERFORMANCE

Total Cu (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 501b	214	0	0.0%	2,600	2,581	-0.7%
OREAS 502b	211	0	0.0%	7,730	7,532	-2.6%
OREAS 503b	210	0	0.0%	5,310	5,241	-1.3%
OREAS 504b	198	2	1.0%	11,100	11,000	-0.9%
OREAS 902	55	1	1.8%	3,010	3,028	+0.6%
OREAS 930	179	4	2.2%	25,200	25,480	+1.1%
Mo (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 501b	214	2	0.9%	99	96	-3.0%
OREAS 502b	211	0	0.0%	238	230	-3.4%
OREAS 503b	210	2	1.0%	319	310	-2.8%
OREAS 504b	198	1	0.5%	499	487	-2.4%
OREAS 902	55	2	3.6%	12.2	11.9	-2.5%
Ag (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 501b	214	4	1.9%	0.778	0.78	+0.3%
OREAS 502b	211	2	0.9%	2.09	2.17	+3.8%
OREAS 503b	210	0	0.0%	1.54	1.61	+4.5%
OREAS 504b	198	3	1.5%	3.07	3.28	+6.8%
OREAS 902	55	9	16.4%	0.343	0.38	+10.8%
OREAS 930	179	25	14.0%	9.00	9.98	+10.9%
Au (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 501b	37	0	0%	0.248	0.252	+1.6%
OREAS 502b	35	0	0%	0.495	0.496	+0.2%
OREAS 503b	32	0	0%	0.695	0.697	+0.3%
OREAS 504b	34	0	0%	1.610	1.596	-0.9%
Soluble Copper (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 902	55	0	0%	1,110	1,140	+2.7%

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Accordingly, CRM assayed values within AV±2SD and isolated values between AV±2SD and AV±3SD were accepted. In contrast, two consecutive assayed values between AV±2SD and AV±3SD and all values outside the AV±3SD were rejected and triggered re-analysis.

To evaluate the accuracy of assaying using CRMs, the relative analytical bias was calculated after excluding the outlier values located outside the AV±3SD:

Bias (%) = 100*[(AVeo/CBV)-1]

AVeo represents the average of actual assay values after excluding outliers. The analytical bias was assessed according to the following ranges: good between 0 and ±5%, reasonable between ±5% and ±10%, and unacceptable for values ±10%.

The analytical bias of CRMs for total copper, molybdenum, soluble copper, and gold was good with values in the range between -3.4% and +2.7% (Table 11-6).

The analytical bias of silver in OREAS 501b to 504b (0.8 to 3 ppm Ag) was good to reasonable, with values between +0.3% and +6.8% . However, silver in OREAS 902 and 930 displayed a larger analytical bias, approximately +11% (Table 11-6).

The large bias of silver in OREAS 902 is attributed to the low silver content and large variance (0.343 ppm, SD = 0.043) of this CRM. The certified silver content of OREAS 902 is between three and four times the lower limit of detection (0.1 ppm), and the relative standard deviation (12%) of this CRM is larger than the analytical bias of +11%. Good performance is expected at values at least 10 times the lower limit of detection. Therefore, the performance of OREAS 902 is considered reasonable.

The certified best value for silver of OREAS 930 (9.00 ppm, SD = 1.09) is more than 10 times the lower detection limit. This CRM displayed a bias of 10.9% which is considered unacceptable (Table 11-6). However, the bias is lower than the relative standard deviation of this CRM which is 12%. To further investigate this issue, pulp duplicate re-analysis were resubmitted to Bureau Veritas laboratory for all failed OREAS 930 including eight drill core pulp samples centered on the failed CRMs. In total, 21 failed OREAS 930 and 135 drill core pulp samples with silver between 0.1 and 116 ppm were re-analyzed. After re-assaying, 86% of CRMs yielded silver values within the certified best value. Despite the poor reproducibility of OREAS 930, 90% of the re-assayed drill core pulps in the sample stream of OREAS 930 yielded similar results (silver 01.-117 ppm) evaluated for an absolute relative difference between pulp pairs equal to or smaller than 10%.

The CRM analysis indicates that the analytical accuracy for total copper, molybdenum, and soluble copper is of good quality for the resource estimation. The cause of the poor performance of silver in OREAS 930 is attributed to the large variability (SD = 1.09) of this CRM. It is also noted that OREAS 930 includes minerals such as sphalerite, galena, and cubanite, which are not found in significant quantities at Rosemont. Given the analysis discussed above, and good performance of silver in OREAS 501b to 504b, it is concluded that the accuracy for silver is also adequate for resource estimation.

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11.1.11      Duplicates

Coarse duplicates, approximately one in every twenty samples, were requested to Bureau Veritas laboratory in order to monitor sub-sampling precision. Accordingly, after crushing to 10 mesh (2 mm), a 1,000 g coarse duplicate sub-sample was riffle split and pulverized to ≥85% passing through 200 mesh (75 μm). The duplicate sample was analyzed immediately after its paired sample. A total of 1,086 coarse duplicate samples were inserted for a total rate of 6%. Quarter-core twin sample duplicates and pulp duplicates were not analyzed during Hudbay’s 2014 drilling program.

Coarse duplicates were reviewed using the hyperbolic method (Table 11-2) developed by AMEC (Simón, 2004). Minimum and maximum element concentrations of the sample pairs are plotted in the y and x axis, respectively. In the Minimum-Maximum diagrams, all samples plot along and above the x = y line and the failure boundary is given by the equation y²=m²x²+b². The coarse duplicates were evaluated using a failure boundary that asymptotically approaches the line with slope “m” corresponding to a 15% absolute relative error (“RE”). The RE is calculated as the absolute value of the pair difference divided by the pair average and expressed in percentage. An acceptable level of sub-sampling variance is achieved when the failure rate does not exceed 10% of all sample pairs.

The failure rates of the duplicate pairs for total copper, molybdenum, silver, gold and soluble copper range between 4% to 6% based on the hyperbolic method for an absolute relative error of 15% (Table 11-7, Figure 11-1 to Figure 11-4). It is concluded that the sub-sampling procedures were adequate for all metals used in the resource model.

TABLE 11-7: SUMMARY OF COARSE DUPLICATE ANALYSIS

Element	No. of Samples	No. of Failures	Failure Rate	Accepted Absolute RE
Cu	1,086	58	5.3%	15%
Mo	1,086	66	6.1%	15%
Ag	1,086	66	6.1%	15%
Au	189	8	4.2%	15%
Soluble Cu	1,086	65	6.0%	15%

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FIGURE 11-1: COPPER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

FIGURE 11-2: MOLYBDENUM COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

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FIGURE 11-3: SILVER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

FIGURE 11-4: SOLUBLE COPPER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

11.1.12      Check Assaying

A total of 1,000 representative pulp samples (5.5%) were selected and re-analyzed at SGS Canada Inc. (“SGS”) laboratory in Vancouver. This laboratory is independent from Hudbay and has a quality system that is compliant with the International Standards Organization (“ISO”) 9001 Model for Quality Assurance and ISO/IEC 17025 General Requirements for the Competence of Testing and Calibration Laboratories. Only samples with ≥500 ppm copper were submitted for re-analysis at the secondary laboratory.

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CRMs, certified blanks, and pulp duplicates were inserted along with the check samples following the same protocols used for monitoring Bureau Veritas. However, pulp duplicates, rather than coarse duplicates, were submitted to SGS. Duplicates and CRMs indicate that SGS achieved good levels of precision and accuracy. The overall bias deduced from the CRMs was -2.3% for copper, -4.5% for molybdenum, -2% for silver, and +2.1% for soluble copper. The analysis of blanks identified a few cases of economically insignificant copper contamination with average contamination of <30 ppm copper. It is concluded that the assay results from SGS are of good quality to evaluate the performance of Bureau Veritas.

A Reduced-to-Major-Axis regression (“RMA”) was used to evaluate the check samples (Kermack and Haldane, 1950). The RMA regression calculates an unbiased fit for values that are independent from each other. The coefficient of determination (R²) is used to assess the variance explained by the linear relationship between the pairs. The bias, expressed as a percent, is calculated as Bias (%) = 1-RMAS in which RMAS is the slope of the RMA regression.

There is a good fit for copper (R² = 0.996), silver (R² = 0.976), molybdenum (R² = 0.993), soluble copper (R² = 0.970), and gold (R² = 0.987) . The slope of the RMA regression for all metals ranges between 0.93 and 1.01 and all intercepts are below the practical limit of detection and approximate zero, as shown in Figure 11-5, where the black line represents the y = x line and the red dash line represents the RMA regression line.

The overall analytical bias of Bureau Veritas relative to SGS is +1.2% for copper, -1.0% for silver, +2.7% for molybdenum, +6.8% for soluble copper and -2.0% for gold. The overall bias estimated by the RMA regression analysis indicates that the accuracy achieved by Bureau Veritas for copper, molybdenum, silver, soluble copper is of good quality for resource estimation.

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FIGURE 11-5: XP PLOTS OF CHECK ASSAY DATA, COMPARING PRIMARY LABORATORY BUREAU VERITAS TO SECONDARY LABORATORY SGS

11.2        Hudbay 2015

For consistency, the 2015 drilling campaign followed identical core logging, sample selection, core cutting, sample dispatching, and sample preparation protocols as those followed in 2014 as previously described in detail. The only notable exception in 2015 was the closure of Hidden Valley camp as all logging related activities took place at Rosemont camp which was expanded to accommodate up to 8 core logging geologists at a given time.

11.2.1        Bulk Density

A total of 755 samples in 46 drill holes were collected for specific gravity determinations at the Inspectorate preparation facility. The drill core samples, 4 to 6 inches (10 to 15 cm) long, were taken every 100 feet (30 meters) from half-split core and wax-coated following similar procedures to those used during the 2014 drilling program as previously described in detail.

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11.2.2        Assay Methodology

A total of 46 drill holes were sampled from top to bottom and assayed following the same digestion and analytical procedures used during 2014 and described in detail on Table 11-1, Section 11.1.

During 2015, 14,844 drill core samples were analyzed for 45 elements, including copper, molybdenum, and silver by ICP-MS after 4-acid digestion (“Method MA200”). All samples were also analyzed for ASCu (“Method GC921”) with a 5% sulfuric acid (H₂SO₄) leach at room temperature. For gold assays, 5 drill holes were sampled from top to bottom with a total of 1,957 samples, 13% of Hudbay’s 2015 program, assayed by lead-collection fire assay fusion (Table 11-1).

QA/QC protocols, duplicates, CRMs, and certified blanks, were similar to those used during the 2014 drilling program. However, OREAS 930 was replaced by OREAS 931 (Table 11-8). The QA/QC samples were systematically introduced to assess adequate sub-sampling procedures, potential cross-contamination, precision, and accuracy. The sampling program included 6% CRM, 6% certified blanks, and 6% coarse duplicates. Blanks and CRMs were prepared by OREAS in Australia. All QA/QC samples were analyzed following the same analytical procedures as those used for the drill core samples.

TABLE 11-8: SUMMARY OF QA/QC SAMPLES

	No. of samples	Relative Frequency1	Type	Elements
OREAS 501b	160	1.1%	CRM	Cu-Mo-Ag-Au
OREAS 502b	166	1.1%	CRM	Cu-Mo-Ag-Au
OREAS 503b	160	1.1%	CRM	Cu-Mo-Ag-Au
OREAS 504b	158	1.1%	CRM	Cu-Mo-Ag-Au
OREAS 902	91	0.6%	CRM	Cu-Mo-Ag- Cu Soluble
OREAS 931	159	1.1%	CRM	Cu-Mo-Ag
OREAS 22d	436	2.9%	Certified Blank	Cu-Mo-Ag-Au
OREAS 26a	438	2.9%	Certified Blank	Cu-Mo
Duplicates	870	5.8%	Coarse Duplicate	Cu-Mo-Ag
Duplicates	115	5.9%2	Coarse Duplicate	Cu-Mo-Ag-Au

¹Frequencies estimated relative to a sampling program comprising 14,868 samples
²Frequencies estimated relative to 1,957 samples analyzed for gold

11.2.3        Blanks

Fine and coarse blanks were systematically inserted at the same rate for a total of 874 blanks representing 6% of the sampling program. OREAS 22d is a certified fine blank prepared from quartz sand. OREAS 26a is a certified coarse blank sourced from fresh and non-mineralized olivine basalt. The certified values for OREAS blanks are shown on Table 11-3.

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OREAS 22d and OREAS 26a contain low trace level concentrations of copper, molybdenum, silver, and gold. Blank failure due to potential contamination issues is documented when the blank values exceed five times the lower limit of detection. For those blanks with concentration levels above the lower detection limits the failure thresholds are set to values that exceed the certified best value (CBV) plus three standard deviations. A summary of the blank performance is shown in Table 11-9.

TABLE 11-9: SUMMARY OF BLANK PERFORMANCE

Fine Blank OREAS 22d
Element	No. of Blanks	Failed Blanks	Failure Rate	Maximum Contamination of CBV	Average Contamination of CBV
Cu	436	17	3.9%	45 ppm	7 ppm
Mo	436	9	2.1%	1 ppm	1 ppm
Ag	436	0	0	0	0
Au	57	0	0	0	0
Coarse Blank OREAS 26a
Element	No. of Blanks	Failed Blanks	Failure Rate	Maximum Contamination of CBV	Average Contamination of CBV
Cu	438	13	3.0%	40 ppm	10 ppm
Mo	438	1	0.2%	12 ppm	12 ppm
Ag	438	0	0	0	0
Au	57	0	0	0	0

Contamination with copper, silver, and gold was insignificant (Table 11-9). There are very few cases of contamination at high-grade levels (12 ppm) of molybdenum. However, the overall blank failure rates are <3%.

The performance of blanks indicates no significant issues with contamination; therefore, it is concluded that the results are acceptable and adequate for the resource estimation.

11.2.4        Standards

OREAS certified reference materials (Table 11-10) were inserted one every twenty samples. In total, 894 CRMs were analyzed for a total insertion rate of 6.0% .

TABLE 11-10: OREAS CERTIFIED REFERENCE MATERIAL

Unit	Cu %	Mo Ppm	Ag ppm	Au ppm	Cu Soluble %
OREAS 501b	0.260	99	0.778	0.248	-
OREAS 502b	0.773	238	2.090	0.495	-
OREAS 503b	0.531	319	1.540	0.695	-
OREAS 504b	1.110	499	3.070	1.610	-
OREAS 902	0.301	12.2	0.343	-	0.111
OREAS 931	3.82	-	14.04	-	-
Certified Method	4 acids	4 acids	4 acids	Fire assay	Sulfuric acid at 5%

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The geological matrices for OREAS 501b to 504b and OREAS 902 are described in detail in Section 11. OREAS 931 is composed of a geological matrix, similar to OREAS 930, prepared from a copper ore body hosted in carbonaceous siltstones and mudstones mineralized with chalcopyrite, bornite, pyrrhotite, pyrite, sphalerite, galena, and cubanite.

Between 90 and 170 samples were analyzed per CRM (Table 11-8), which is sufficient information to set acceptance criteria relative to the average (“AV”) and standard deviation (“SD”) of the actual assay values of the CRMs. However, if the absolute relative bias is >10% the acceptance criteria is set relative to the CBV and standard deviation recommended by the CRM certificates.

The analytical bias for copper, molybdenum, and gold was good ranging between -2.5% and +1.1% (Table 11-11). The bias for soluble copper was reasonable at +6.3% .

The analytical bias of silver in OREAS 501b to 504b (0.8 to 3 ppm Ag) was good to reasonable with values between -1.1% and +7.7% . However, silver in OREAS 902 and 931 displayed larger analytical bias (+14.1% to +16%).

The large bias of silver in OREAS 902 and OREAS 931 is attributed to the low silver content of OREAS 902 and large variability of silver in these standards with relative standard deviations of 13% and 14%, respectively. The average silver value measured for these CRMs from the assays is within 1 and 2 SD of the CBV, respectively. Thus, the performance of silver in OREAS 902 and OREAS 931 is considered reasonable (Table 11-11).

Pulp duplicate re-analyses were requested to Bureau Veritas laboratory for all failed silver assays for OREAS 902 and OREAS 931 including six drill core pulp samples centred on the failed CRMs. In total, 21 failed CRMs and 126 drill core pulp samples with silver between 0.1 and 36.6 ppm were reanalyzed. After re-assaying, 90% of CRMs yielded silver values within the CBV and 92% of the re-assayed drill core pulps in the sample stream yielded similar results (0.1 - 27.8 ppm silver) evaluated for an absolute relative difference between pulp pairs equal to or smaller than 10%.

The CRM analysis indicates that the analytical accuracy for total copper, molybdenum, soluble copper, silver, and gold is of good quality for the resource estimation.

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TABLE 11-11: SUMMARY OF CRM PERFORMANCE

Total Cu (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 501b	160	2	1.3%	2,600	2,589	-0.4%
OREAS 502b	166	3	1.8%	7,730	7,539	-2.5%
OREAS 503b	160	1	1.3%	5,310	5,233	-1.4%
OREAS 504b	158	1	0.6%	11,100	10,971	-1.2%
OREAS 902	91	2	2.2%	3,010	3,003	-0.2%
OREAS 931	159	1	0.6%	38,200	38,488	+0.8%
Mo (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 501b	160	1	0.6%	99	97	-1.7%
OREAS 502b	166	1	0.6%	238	236	-0.7%
OREAS 503b	160	2	1.2%	319	313	-1.9%
OREAS 504b	158	1	0.6%	499	491	-1.6%
OREAS 902	91	0	0%	12.2	11.9	-2.4%
Ag (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 501b	160	2	1.3%	0.778	0.77	-1.1%
OREAS 502b	166	3	1.8%	2.09	2.16	+3.3%
OREAS 503b	160	2	1.3%	1.54	1.59	+3.5%
OREAS 504b	158	0	0%	3.07	3.31	+7.7%
OREAS 902	91	14	15.4%	0.343	0.40	+16.0%
OREAS 931	159	13	8.2%	14.04	16.01	+14.1%
Au (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 501b	23	0	0%	0.248	0.251	+1.1%
OREAS 502b	22	0	0%	0.495	0.493	-0.4%
OREAS 503b	20	0	0%	0.695	0.693	-0.2%
OREAS 504b	21	0	0%	1.610	1.600	-0.7%
Soluble Copper (ppm)
Standard	No. of Samples	No. of Failures	Failure Rate	CRM Value (ppm)	Assay Average	Relative Bias
OREAS 902	91	1	1.1%	1,110	1,180	+6.3%

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11.2.5        Duplicates

Coarse duplicates, approximately one in every twenty samples, were requested to Bureau Veritas laboratory in order to monitor sub-sampling precision (Section 11.1.11) . Accordingly, after crushing to 10 mesh (2 mm), a 1,000 g coarse duplicate sub-sample was riffle split and pulverized to ≥85% passing through 200 mesh (75 μm). The duplicate sample was analyzed immediately after its paired sample.

A total of 870 coarse duplicate samples were inserted for a total rate of 6%. Quarter-core twin sample duplicates and pulp duplicates were not analyzed during Hudbay’s 2015 drilling program. The coarse duplicates were evaluated using the hyperbolic method developed by AMEC (Simón, 2004) and explained in detail on Section 11.1.11.

TABLE 11-12: SUMMARY OF COARSE DUPLICATE ANALYSIS

Element	No. of Samples	No. of Failures	Failure Rate	Accepted Absolute RE
Cu	870	71	8.2%	20%
Mo	870	72	8.3%	20%
Ag	870	25	2.9%	20%
Au	115	0	0%	20%
Soluble Cu	870	69	7.9%	20%

The results from the coarse duplicate analysis are presented on Table 11-12 and illustrated on Figure 11-6 to Figure 11-9. During 2015, an acceptable level of sub-sampling variance was achieved with, failure rates between 0% and 8.3%, for sample pairs evaluated for a maximum absolute relative error of 20% (Table 11-12). It is noteworthy that the sub-sampling variance achieved during 2015 was larger than the variance indicated by the coarse duplicates analysis during Hudbay’s 2014 drilling program (Table 11-7). During 2014, an acceptable level of sub-sampling variance was achieved for an absolute relative error of 15%.

Despite the larger sub-sampling variance observed during the 2015 program, the duplicate failure rate for a maximum RE of 20% is acceptable for all metals used in the resource model.

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FIGURE 11-6: COPPER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

FIGURE 11-7: MOLYBDENUM COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

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FIGURE 11-8: SILVER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

FIGURE 11-9: SOLUBLE COPPER COARSE DUPLICATE MINIMUM AND MAXIMUM PLOT

11.2.6        Check Assaying

A total of 742 representative pulp samples (5%) were selected and re-analyzed at SGS Canada Inc. (“SGS”) laboratory in Vancouver. Only samples with ≥1000 ppm copper were submitted for re-analysis at the secondary laboratory.

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CRMs, certified blanks, and pulp duplicates were inserted along with the check samples following the same protocols used for monitoring Bureau Veritas. However, pulp duplicates, rather than coarse duplicates, were submitted to SGS. Duplicates and CRMs indicate that SGS achieved good levels of precision and accuracy. The overall bias deduced from the CRMs was +2.7% for copper, -5.8% for molybdenum, +14% for silver, and +2% for soluble copper. The large bias of silver is explained by the large variance of silver in OREAS 931. The analysis of blanks indicates that there was no economically significant contamination. It is concluded that the assay results from SGS are of good quality to evaluate the performance of Bureau Veritas.

A Reduced-to-Major-Axis regression (“RMA”) was used to evaluate the check samples (Kermack and Haldane, 1950). The RMA regression calculates an unbiased fit for values that are independent from each other (Section 11.1.12) .

FIGURE 11-10: XP PLOTS OF CHECK ASSAY DATA, COMPARING PRIMARY LABORATORY BUREAU VERTIAS TO SECONDARY LABORATORY SGS

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There is a good fit for copper (R² = 0.999), silver (R² = 0.979), molybdenum (R² = 0.990), soluble copper (R² = 0.989), and gold (R² = 0.977) . The slope of the RMA regression for all metals ranges between 0.92 and 1.02 and all intercepts are below the practical limit of detection and approximate zero, as shown in Figure 11-5, where the black line represents the y = x line and the red dash line represents the RMA regression line.

The overall analytical bias of Bureau Veritas relative to SGS is -2.17% for copper, +3.32% for silver, +1.66% for molybdenum, +8.25% for soluble copper and +2.55% for gold. The overall bias estimated by the RMA regression analysis indicates that the accuracy achieved by Bureau Veritas for copper, molybdenum, silver, soluble copper, and gold is of good quality for resource estimation.

11.3        Augusta

A detailed description of sample preparation procedures and data verification processes conducted by Augusta are provided in several reports including two NI 43-101 technical reports prepared by M3 Engineering & Technology Corporation (2009, 2012). However, Hudbay conducted its own technical review and verification of the data; the results of which are summarized in this section.

11.3.1        Sample Preparation

Overall, the documented protocols for handling diamond drill core, data security, drill core sampling, and sample custody by Augusta are acceptable and industry standard. All core drilled by Augusta was systematically logged for RQD, lithology, alteration, mineralization, and structures. Logging was paper-based and later recorded in spreadsheets. Geologists marked the logged core for cutting with diamond rock saws and sampling was conducted on half-split core. Samples were collected at fixed 5 feet (1.5 meters) intervals, assigned a unique sample number, and securely sealed in sample bags. The sample intervals were shortened to accommodate smaller zones with abrupt changes in copper and molybdenum mineralization, but typically the sampled intervals straddle geological boundaries.

11.3.2        Bulk Density

Augusta analyzed a total of 391 drill core samples across the Rosemont deposit for their specific gravity (“SG”) at Skyline Assayers & Laboratories (“Skyline”), Tucson, Arizona, USA. Skyline followed a protocol based on the differential weight of the sample in air and water. No paraffin coating was applied.

The SG results obtained by Augusta compare well with those measured by Hudbay at Inspectorate during its 2014 and 2015 drilling programs, as shown in Figure 11-11. The global average (2.74) and median (2.70) SG values of 391 samples measured at Skyline are comparable with the average (2.69) and median (2.66) of 954 samples analyzed by Hudbay during 2014, and 755 samples analyzed in 2015 (average= 2.72, median= 2.69) . It is concluded that the SG reported by Augusta are of good quality and appropriate for resource estimation.

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FIGURE 11-11: BOXPLOTS OF SG MEASURED BY HUDBAY AND AUGUSTA AT INSPECTORATE AND SKYLINE LABORATORIES, RESPECTIVELY

11.3.3        Assay Methodology

Augusta assayed drill core samples at Skyline’s laboratory in Tucson, Arizona. Drill core samples were dried before being crushed using jaw crushers to produce a coarse fraction with ≥70% passing through 10 mesh (2 mm). The entire crushed sample was homogenized, riffle split, and a 300 to 400 g subsample split was pulverized to pass ≥95% through 150 mesh (105 μm) using Essa standard steel grinding bowls. Jaw crushers, preparation pans, and grinding bowls were cleaned with compressed air between samples. Coarse rejects and pulps were returned to Augusta.

Table 11-13 summarizes the assay methodologies and instrumental finishes conducted by Skyline.

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TABLE 11-13: ASSAY SPECIFICATIONS – SKYLINE

Element	Cu	Cu Soluble	Mo	Mo	Ag	Ag	Au
Lower Detection Limit	100 ppm	100 ppm	50 ppm	10 ppm	0.4 pm	0.1 ppm	0.005ppm
Upper Detection Limit	10%	5%	1%	1%	100 ppm	100 ppm	3 ppm
Digestion	3 acids HCl- HNO3- HClO4	Sulfuric acid at 10% H2SO4- Na2SO3	3 acids HCl- HNO3- HClO4	3 acids HCl- HNO3- HClO4	Aqua Regia HCl-HNO3-	Aqua Regia HCl- HNO3-	Fire assay
Instrumental Finish	AAS	AAS	ICP-OES	ICP-OES	AAS	AAS	AAS
Method Code	MEA	Cu-AS	MEA	MEA	FA-O8	FA-08	FA-01
Time Period	2005-2012	2005-2012	2005-2006	2006-2012	2005	2005-2012	2005-2012

In total 21,197 samples were analyzed for total copper and 16,619 samples for molybdenum. Total copper and molybdenum were dissolved using a hot 3-acid digestion at 250°C and subsequently analyzed by AAS and ICP-OES, respectively. The lower detection limits for molybdenum are high relative to the average molybdenum grade of the Rosemont deposit (Table 11-13).

A total of 9,030 samples were analyzed for soluble copper using an acid leach at 10% sulfuric acid with sodium sulfite. The acid leach was conducted for an hour at room temperature and the solution was analyzed by AAS.

Silver, analyzed in 15,334 samples, was digested using an aqua regia leach in 0.25 g subsample pulp and analyzed by AAS. Two different lower limits of detection, 0.4 and 0.1 ppm, were used in 2005. The 0.4 ppm detection limit is high relative to the average silver grade of the deposit. However, the lower limit of detection was improved in the following years (Table 11-13).

A total of 4,932 samples were analyzed for gold by fire assay with an AAS finish.

Augusta conducted its own internal QA/QC program to independently evaluate the quality of the assays reported by Skyline. Standards and blanks were systematically inserted in the sample stream. Duplicates were not periodically inserted. The QA/QC program was initially provided by Geochemist, Kenneth A. Lovstrom (deceased). After 2006 the QA/QC program was managed by Geochemist, Shea Clark Smith of Minerals Exploration & Environmental Geochemistry.

Skyline is a certified laboratory accredited in accordance with the recognized International Standard ISO/IEC 17025:2005 General Requirements for the Competence of Testing and Calibration Laboratories. The sample preparation, analysis, and security procedures followed by Skyline are considered industry standard.

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11.3.4        Blanks

In order to track potential contamination processes Augusta inserted non-certified blanks in the sample stream for an average insertion rate of 2%. Coarse barren marble and fine quartz sand were used as blanks in early drill programs through 2007, after which the marble blank was no longer used. The marble blank was used after high grade samples as a cleaner and to test for cross contamination, and this blank was excluded from statistical analyses because its contained metal content is unknown. The distinction between these two blanks was not documented by Augusta in the database and the results are evaluated as a combined single blank.

The assays results for copper, molybdenum, silver, and gold indicate that the blanks are barren relative to the metals of economic interest and appropriate to assess contamination. Blank failure due to potential contamination issues is triggered when the blank values exceed five times the lower limit of detection. A few cases of contamination at higher grade levels are documented for silver and molybdenum. However, all metals of economic interest have very low failure rates ranging from 0 to 6%, indicating that contamination is not a significant problem in the samples analyzed at Skyline as shown in Table 11-14.

TABLE 11-14: SUMMARY OF BLANK PERFORMANCE AT SKYLINE

Element	Count	Failed Blanks	Failure Rate (%)	Maximum Contamination	Average Contamination
Cu	553	5	1.0%	690 ppm	310 ppm
Ag	552	33	6.0%	7.6 ppm	0.9 ppm
Mo	440	7	1.6%	120 ppm	40 ppm
Au	123	0	0.0%	0	0

11.3.5        Standards

Augusta used 14 standard reference materials (“SRM”) inserted in the sample stream with an average insertion rate of 4.3% . The insertion rate is appropriate to assess the accuracy of the data.

Standards KM5, GRS3, GRS4, OC43, and OC48 were developed by Mr. Lovstrom. The R-series standards were prepared at MEG Labs in Carson City, Nevada. M3 Engineering & Technology Corporation (2012) has indicated that the MEG SRMs were prepared from mineralized rock collected from the Rosemont deposit with best values (“BV”) determined following a round robin program from a minimum of 25 samples analyzed by at least 5 different laboratories. The certificates of the SRMs used by Augusta are not available and details of the digestion protocols, sample matrix, and analytical finish are unknown.

Table 11-15 provides a summary of best values, standard deviations, and relative standard deviations (“RSD”) for the SRMs extracted from internal reports by Augusta. There are no records in the database indicating the use of SRM R4A. Therefore, the analysis presented here is based on the

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13 remaining SRMs. Table 11-16 summarizes the analytical performance of the SRMs used by Augusta.

TABLE 11-15: STANDARD REFERENCE MATERIALS – AUGUSTA

Total Cu (%)
SRM	Best Value	SD	RSD
KM5	0.99	0.02	1.5%
GRS3	1.23	0.03	2.0%
GRS4	2.02	0.02	1.0%
R1	0.47	0.02	3.2%
R2	0.72	0.02	2.8%
R4A	1.43	0.02	1.4%
R4B	0.57	0.02	3.5%
R4C	0.39	0.02	3.8%
R4D	0.30	0.02	6.7%
R4E	0.22	0.01	4.5%
R4F	0.14	0.01	7.1%
R4G	0.07	0.01	14.3%
Mo (%)
SRM	Best Value	SD	RSD
OC43	0.035	0.001	2.9%
OC48	0.078	0.004	5.1%
R1	0.025	0.003	10.0%
R2	0.017	0.002	11.8%
R4A	0.032	0.002	6.2%
R4B	0.030	0.002	6.7%
R4C	0.033	0.002	6.1%
R4D	0.018	0.002	8.3%
R4E	0.011	0.001	9.1%
R4F	0.010	0.001	10.0%
R4G	0.016	0.001	6.2%
Ag (ppm)
SRM	Best Value	SD	RSD
R1	5.1	0.52	10.20%
R2	7.1	0.74	10.44%
R4A	7.0	0.84	11.96%
R4B	3.9	0.49	12.60%
R4C	3.1	0.58	18.89%
R4D	2.4	0.67	28.51%
R4E	1.7	0.64	36.99%
R4F	1.4	0.66	46.81%
R4G	1.2	0.80	66.81%

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TABLE 11-16: PERFORMANCE OF STANDARD REFERENCE MATERIALS AT SKYLINE

Total Cu (%)
SRM	No. of Samples	No. of Failures	Failure Rate (%)	SRM Best Value	Skyline Average	Relative Bias
KM5	59	0	0.0%	0.99	1.01	+2.0%
GRS3	18	0	0.0%	1.23	1.12	-8.9%
GRS4	20	1	5.0%	2.02	1.90	-5.9%
R1	417	15	3.6%	0.47	0.47	0.0%
R2	233	2	0.9%	0.72	0.71	-1.4%
R4B	33	1	3.0%	0.57	0.57	0.0%
R4C	151	1	0.7%	0.39	0.40	+2.6%
R4D	74	2	2.7%	0.30	0.30	0.0%
R4E	90	1	1.1%	0.22	0.21	-4.5%
R4F	93	3	3.2%	0.14	0.14	0.0%
R4G	23	1	4.3%	0.07	0.07	0.0%
Total	1,211
Mo (%)
SRM	No. of Samples	No. of Failures	Failure Rate (%)	SRM Best Value	Skyline Average	Relative Bias
OC43	22	0	0.0%	0.035	0.034	-2.9%
OC48	21	1	4.8%	0.078	0.073	-6.4%
R1	233	6	2.6%	0.025	0.025	0.0%
R2	225	1	0.4%	0.017	0.018	+5.9%
R4B	33	0	0.0%	0.030	0.029	-3.3%
R4C	141	1	0.7%	0.033	0.031	-6.1%
R4D	51	1	2.0%	0.018	0.018	0.0%
R4E	81	2	2.5%	0.011	0.009	-18.2%
R4F	74	2	2.7%	0.010	0.009	-10.0%
R4G	23	1	4.3%	0.016	0.014	-12.5%
Total	904
Ag (ppm)
SRM	No. of Samples	No. of Failures	Failure Rate (%)	SRM Best Value	Skyline Average	Relative Bias
R1	233	14	6.0%	5.1	5.1	0.0%
R2	225	2	0.9%	7.1	7.0	-1.4%
R4B	25	1	4.0%	3.9	3.8	-2.6%
R4C	124	6	4.8%	3.1	2.5	-19.4%
R4D	51	1	2.0%	2.4	2.0	-16.7%
R4E	41	0	0.0%	1.7	1.3	-23.5%
R4F	122	1	0.8%	1.4	1.0	-28.6%
R4G	21	0	0.0%	1.2	0.5	-58.3%
Total	842

The analytical accuracy for copper was good to reasonable for all SRMs analyzed at Skyline with relative bias ranging from -9% to +3% (Table 11-16). Two SRMs (R4E and R4G) displayed poor accuracy for molybdenum with negative bias of less than -10%. However, 90% of the SRMs measured for molybdenum displayed relative bias between -10% and +6% indicating good to reasonable accuracies for molybdenum SRMs at grade levels of ≥ 100 ppm.

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The results of the standards analyzed for silver indicate good to reasonable accuracies at grade levels ≥3 ppm. However, all SRMs with recommended best values <3 ppm silver displayed unacceptable negative bias between -58% and -19% (R4C to R4G; Table 11-16). The poor performance of these silver standards is attributed to several factors including imprecise characterization of the standards and silver grades close to the lower detection limit. For instance, the reported RSDs for standards R4C to R4G range from 19 to 67% indicating poor precision in the determination of the recommended best values (Table 11-15). The large RSDs diminish the value of these standards as reference materials.

It is concluded that the analytical accuracy achieved by Skyline for all metals of economic interest is good to reasonable and therefore adequate for resource estimation.

11.3.6        Duplicates

Augusta did not insert duplicates periodically within the sample stream. On average, only 0.2% coarse duplicate (<50 samples) samples were analyzed. The insertion of duplicates is significantly below recommended rates of 2% to 6%. There are insufficient duplicate samples to correctly evaluate batch reproducibility.

11.3.7        Check Assays

Augusta resubmitted sample pulps to Skyline for check assays. The significance of the check samples to independently estimate the confidence of Skyline is diminished given that a secondary laboratory was not used. Augusta submitted an average of 1.5% of samples for re-analysis of copper, molybdenum, silver, and soluble copper. The check assay rate is lower than a recommended rate of 4% to 5%.

A RMA was used to evaluate the check samples. The coefficient of determination (R²) is used to assess the variance explained by the linear relationship between the pairs and is calculated as Bias (%) = 1-RMAS in which RMAS is the slope of the RMA regression.

Augusta re-assayed 373 samples for copper, 326 samples for silver, 326 samples for molybdenum, and 203 samples for soluble copper. The RMA regression for the sample pairs was calculated after removing extreme outliers that clearly represent switched samples. In total, there were three outliers for copper, four outliers for silver, two outliers for molybdenum, and one outlier for soluble copper.

There is a good fit for copper (R² = 0.999), silver (R² = 0.992), molybdenum (R² = 0.995), and soluble copper (R² = 0.968) . The slope of the RMA regression for all metals ranges between 0.98 and 1.04 and all the intercepts approximate zero and are below the lower limit of detection for each element. The overall bias of the primary analysis relative to the check assays is +1.0% for copper, +2.0% for silver, +1.6% for molybdenum, and -3.7% for soluble copper. However, on a per sample basis, >10% relative error (“RE”) is observed for silver grades ≤ 3 ppm and molybdenum grades ≤ 50 ppm. This is expected given the use of relatively high lower limits of detection for these metals (Table 11-13).

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The overall bias estimated by the RMA regression analysis indicates that the global accuracy for copper, molybdenum, silver, and soluble copper is good for resource estimation.

11.4        Historic

Table 11-17 provides historical sample preparation procedures and data verification processes conducted by several property owners prior to 2005.

TABLE 11-17: ROSEMONT DEPOSIT DRILLING SUMMARY

Company	Time Period	No. of Drill Holes	Metres Drilled
Banner	1950s to 1963	3	1,311
Anaconda	1963 – 1973	113	41,708
Anamax	1973 – 1986	52	16,566
ASARCO	1988 – 2004	11	4,479
Total		179	64,008

11.4.1        Sample Preparation

For over 50 years, significant diamond drilling, drill core sampling, and assaying programs were executed by several property owners preceding Hudbay and Augusta. Records are not available with details of sampling and security protocols used by these property owners.

Banner, Anaconda, and Anamax used similar methodologies for drill core logging and sampling. In general, lithology, alteration, structures, and mineralization were documented on paper logs. Drill core was half-split using mechanical splitters and sampled for assaying. Mineralized intervals were entirely sampled with sample length ranging from 1 to 5 feet (0.3 to 1.5 meters). Poorly mineralized intervals were sampled every 20 to 30 feet (6 to 10 meters) along 5 feet (1.5 meters) intervals.

Asarco logged drill core following the same methodology used by Banner, Anaconda, and Anamax. All geological information was captured on paper logs. The length of drill core samples was variable and subjected to the criteria of core logging geologists. Typical sampling intervals were approximately 10 feet (3 meters) in length.

Augusta data verification program included re-logging of the majority of available drill core mostly from Anaconda, Anamax, and Asarco following the procedures described in Section 11.3.1. The information was collected on paper logs and typed in spreadsheets. Augusta confirmed that historical drill core recoveries were better than 85% (M3 Engineering & Technology Corporation, 2009). Augusta also sampled intervals of historical drill core to fill-in missing analytical information and conducted a re-sampling program of 10 historical drill holes for data verification purposes.

11.4.2        Quality Control Evaluation

There are no available QA/QC records for sample preparation and assaying methodologies for Banner, Anaconda, and Anamax. These previous property owners regularly analyzed drill core

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samples for total copper and molybdenum. Silver was regularly analyzed by Anamax, but not commonly assayed by Banner and Anaconda. Visible oxidized zones were analyzed by soluble copper, but molybdenum was commonly not analyzed for oxide zone samples.

Copper, molybdenum, silver, and soluble copper were analyzed at Anaconda and Anamax in-house laboratories. The only existing record of digestion methodologies and analytical instruments used by these laboratories was provided by former Anaconda Chief Chemist, Mr. Dale Wood, in phone interviews on November 28, 2005, and January 21, 2006 (M3 Engineering & Technology Corporation 2009, 2012). Accordingly, copper and molybdenum were preliminary screened using x-ray fluorescence (“XRF”). Samples with > 0.2% Cu and > 200 pm Mo were then selected for wet chemical analysis. There is no documentation on the methods used to analyze silver and soluble copper.

Sample pulps for XRF analysis were placed on Mylar® film or prepared by adding cellulite and pressing it into a ring. There is no documentation on the type of XRF instrument, XRF technique, and internal laboratory QA/QC protocols.

Wet chemical analysis was a hot 3-acid digestion with hydrochloric, nitric, and perchloric acid, with a few drops of hydrofluoric acid. Copper and molybdenum were analyzed by colorimetry following phenolthylanaline titration for copper and iodine titration for molybdenum. There are no records on the model of colorimeters and internal laboratory QA/QC protocols.

11.4.3        ASARCO

Asarco assayed drill core samples for total copper, molybdenum, and ASCu at Skyline; which is a certified laboratory accredited in accordance with the recognized International Standard ISO/IEC 17025:2005 General Requirements for the Competence of Testing and Calibration Laboratories. However, there are no records of the QA/QC practices followed by Asarco to independently monitor the quality performance of Skyline. Descriptions of digestion methodologies and analytical instruments are not available for the assay results conducted by Asarco.

11.4.4        Augusta Re-Sampling Program

Augusta collected twin samples in 10 historical drill holes to verify the assay results reported by historical drilling and sampling programs. The remaining half-split core was sampled entirely along intervals similar to the original intervals sampled by previous property owners. Nine of these drill holes were drilled by Anaconda and Anamax and one (AH-4) by Asarco. Anaconda and Anamax used their in-house laboratories for assaying, whereas Asarco used Skyline. Copper and molybdenum were re-analyzed in all holes and only three drill holes were reanalyzed for silver. The twin sample re-assays were conducted at Skyline.

Duplicate analysis shows poor reproducibility of each individual Anaconda and Anamax assay relative to their twin sample analyzed by Augusta at Skyline. The poor reproducibility is attributed to factors including sample heterogeneity and poor analytical precision and accuracy of the Wet and XRF methods used by the Anaconda and Anamax laboratories.

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However, on a drill hole by drill hole comparison, the average historical copper assays display differences of less than 10% relative to the assays at Skyline, as shown in Table 11-18. Overall, the slightly lower copper content is deemed to result from the deterioration of these historical core samples over time (i.e. oxidation).

TABLE 11-18: COMPARISON OF HISTORICAL ASSAY RESULTS AND TWIN HALF-SPLIT CORE SAMPLES ANALYZED BY AUGUSTA AT SKYLINE

Copper
Drill Hole	Cu (%) Historic	Cu (%) Re- Analyzed	No. of Samples	Bias (%)	Historic Laboratory
A-804	0.45	0.43	315	+4.7%	Anaconda-Anamax
A-813	0.51	0.51	239	0.0%	Anaconda-Anamax
A-821	0.57	0.53	302	+7.5%	Anaconda-Anamax
A-834	0.48	0.45	282	+6.7%	Anaconda-Anamax
A-858	0.35	0.34	199	+2.9%	Anaconda-Anamax
1485	0.43	0.39	239	+10.3%	Anaconda-Anamax
1508	0.94	0.90	207	+4.4%	Anaconda-Anamax
1916	0.39	0.39	256	0.0%	Anaconda-Anamax
1917	0.24	0.25	63	-4.0%	Anaconda-Anamax
AH-4	0.37	0.38	255	-2.6%	Skyline
Molybdenum
Drill Hole	Mo (ppm) Historic	Mo (ppm) Re-Analyzed	No. of Samples	Bias (%)	Historic Laboratory
A-804	70	50	244	+40.0%	Anaconda-Anamax
A-813	330	280	178	+17.9%	Anaconda-Anamax
A-821	380	290	230	+31.0%	Anaconda-Anamax
A-834	180	180	244	0.0%	Anaconda-Anamax
A-858	180	160	199	+12.5%	Anaconda-Anamax
1485	170	60	239	+183.3%	Anaconda-Anamax
1508	240	230	207	+4.3%	Anaconda-Anamax
1916	260	160	257	+62.5%	Anaconda-Anamax
1917	230	130	65	+76.9%	Anaconda-Anamax
AH-4	90	90	226	0.0%	Skyline
Silver
Drill Hole	Ag (ppm) Historic	Ag (ppm) Re-Analyzed	No. of Samples	Bias (%)	Historic Laboratory
A-804	10.8	8.0	208	+35.0%	Anaconda-Anamax
A-813	10.6	5.9		+79.7%	Anaconda-Anamax
A-821	10.2	4.6		+121.7%	Anaconda-Anamax

Molybdenum reported by the Anaconda and Anamax laboratories (Wet and XRF) show significant positive bias of up to 183% on a drillhole by drillhole basis, relative to the twin samples analyzed at Skyline in Table 11-18. The average historical molybdenum over 2,136 samples is 195 ppm versus an average of 145 ppm in the twin samples, as shown in Figure 11-12. Relative to the twin samples, molybdenum reported as wet assays is 20% higher whereas XRF values are 130% higher. To circumvent problems related to the strong positive bias of the Anamax and Anaconda data, molybdenum grades reported by wet assays were multiplied by 0.85 and those reported by XRF by 0.45. After factoring, the average molybdenum in 2,136 assays by Wet and XRF is 147 ppm and compares well with an average of 145 ppm molybdenum in the twin samples, as shown in Figure 11-12. The factored molybdenum was used for resource estimation to minimize the impact of the large positive bias in the historical data.

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FIGURE 11-12: BOXPLOTS OF RAW MOLYBDENUM DATA AND FACTORED DATA REPORTED BY WET AND XRF (A-A = ANACONDA-ANAMAX AND Y-AXIS IN LOGARITHMIC SCALE)

Silver reported by Anamax and Anaconda in-house laboratory also display a high bias of over 35% relative to the twin samples. However, due to the very small population of re-analyzed silver from only three holes and the Anaconda and Anamax representing approximately 12% of the entire dataset, the author decided not to impact the silver values until further investigation is complete.

Overall, the copper and molybdenum assays by Asarco (drill hole AH-4) compare well with the twin samples with very low bias. No silver was reported by Asarco.

In 2011, Augusta compared the historical drilling data to its more recent drilling results. Even though there were no twin holes drilled on the Property, six metallurgical holes located 13 to 29 feet from the historical holes were used for comparison purposes.

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For most comparisons, copper grades show only minor variability with an average copper grade of 0.65% Cu for Augusta 0.65% Cu and 0.63% Cu for the historical data. Augusta concluded that the grade difference was linked to the natural variability of the skarn mineralization and the spacing between the Augusta and historical drilling.

More information can be found in the Rosemont feasibility study published by Augusta in 2012 and available on SEDAR website (http://www.sedar.com).

In the opinion of the author, the results from the re-assay program of Augusta and the comparison of metallurgical holes with closely located historical holes validated the use of historical copper and silver assays for resource estimation while Hudbay continues to perform confirmatory drilling. Thus far, this confirmation of drilling has confirmed resource quality and permitted the expansion of resource tonnage down the plunge of the deposit. The Molybdenum grade shows an over-estimation bias of approximately 15% in the historical drilling and this data was corrected prior to being used for resource estimation.

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12       DATA VERIFICATION

12.1        Drill Hole Database

Hudbay built an entirely new drill hole database from all pre-Hudbay drilling and assaying information. Orix Geoscience Inc. was employed to digitally enter collar, downhole surveys and assay information from scanned drill logs and assay certificates for all holes drilled prior to Augusta.

The following subsections describe the process Hudbay used to build a completely new database of the drilling, assay values and the steps taken to verify the information. All pre-Augusta (prior to 2005) drill holes will hereby be referred to as “historical drill holes”.

The author’s opinion is that the data verification is adequate for the purposes used in the Technical Report.

12.1.1        Drill Hole Collars

Drill hole collar coordinates of historical drilling were reported in a local Anaconda grid system. The coordinates were converted to NAD83 UTM Zone 12N using MapInfo software by Augusta in 2005. The conversion was based on a best fit transformation using drill hole collars and corners from patented claim boundaries (approximately 6,000 points in total). This conversion is verified by plotting the converted coordinates against a drill hole collar compilation map prepared by Anamax Mining Co., in 1979 and the results are within acceptable margin of error (+/- 5 feet). Further verification was provided by Richard Darling, who located and surveyed 12 of these historical holes in UTM coordinates in 2006 and 2008.

Augusta drill hole collars were surveyed by J. Edmonson in 2006 and Darling Geomatics from 2006 to 2012 using a Trimble GLONASS (“Trimble”) sub-centimeter survey grade GPS. Darling Geomatics were also employed for surveying Hudbay drill hole collars using the same Trimble unit. All coordinates are reported in UTM Zone 12, NAD 83 horizontal datum in International Feet and NAVD 88 vertical datum in International Feet.

12.1.2        Downhole Surveys

Downhole survey files exist for 25 of the 181 historical drill holes, as shown in Table 12-1. The majority of the downhole surveys were conducted by Mollen-Hauer Surveying Company using a gyroscope that measured the drift angle and azimuth. The readings were generally recorded every 100 feet (30 meters). From the record sheets it cannot be determined if the azimuth recorded was adjusted for magnetic declaration, hence no further adjustments were made to these readings. However, of the 181 historical drill holes, 136 were drilled vertically.

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TABLE 12-1: DOWNHOLE SURVEYS OF HISTORICAL DRILLING

Company	Surveyed by	Instrument	Number of Holes	Year
Anaconda Mining	Eastco	Single Shot	4	1966
Anamax	Anamax	Gyroscope	1	1974
Anamax	Parsons Surveying Company	Gyroscope	3	1974
Anamax	Mollen-Hauer Surveying Company	Gyroscope	17	1974-1983

Downhole surveys were conducted on all Augusta drill holes using a Reflex EZ-Shot which measures the dip and azimuth. The surveys were measured at 500 feet (152 meters) intervals by the drilling contractors of Layne-Christensen and Boart Longyear.

In 2014, Hudbay completed downhole surveys on either 200 feet (61 meters) intervals using a Reflex EZ Shot tool or on 50 feet (15 meters) intervals using a gyroscopic tool for their drill hole program, as shown in Table 12-2. Except for the single shot surveys measured using a Reflex EZ-Shot, all down hole data was digitally imported from the instrument into the database. Data was further assessed for reliability based on corresponding magnetic readings, subsequently discarding any readings above the threshold magnetic field.

In 2015, Reflex EZ shot tool was used to survey angled holes every 200 feet (61 meters) while drilling, and the gyroscopic tool was used to survey the holes every 50 feet (15 meters) at the conclusion of drilling (see Table 12-3). Single shot measurements were used to monitor the dip of the drill holes during its progress, while the gyroscopic readings are treated as the official downhole surveys with the exception of two holes for which no gyro survey was conducted. All data was digitally imported into the database from the instrument output files.

TABLE 12-2: HUDBAY 2014 DOWNHOLE RESULTS

Surveyed by	Instrument	Number of Holes
Major Drilling	Reflex EZ-Shot	5
Layne-Christensen / Major Drilling	Reflex EZ-Trac	14
IDS Directional Surveys	Televiewer	3
Southwest Geophysics	GyroTracer Directional	21

TABLE 12-3: HUDBAY 2015 DOWNHOLE SURVEYS

Surveyed by	Instrument	Number of Holes
Layne-Christensen /National Drilling	Reflex EZ-Shot	26
Southwest Geophysics	GyroTracer Directional	44

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12.1.3        Historical and Augusta Assay Information

Hudbay acquired a compiled drill hole database from Augusta in 2014. In order to verify the data, Hudbay undertook the task of re-creating the historical assay database from the original paper certificates. The services of Orix Geosciences were employed to retrieve drill hole name, sample number, start and end depth of sample, assay values, and analytical methods from scanned copies of the historical paper certificates. Assay values were entered as reported on the paper logs including the lower than the detection limit values in the original reported units.

The newly compiled database was rechecked against the paper copies by Hudbay personnel to identify and fix any data entry errors and typos. Reoccurrences of sample identification for 590 samples from various sampling campaigns are present in the database, hence it was ruled out as the primary field for each sample. A unique key combining the drill hole name, along with start and end depth of the sample was created to adequately identify and index all of the samples in the database. Each assay field passed through several validation queries that flagged records outside of expected range, missing values and potential mismatch of characters.

In an effort to improve the density of analyses where core was only partially analyzed, Augusta performed a re-sampling program in conjunction with re-logging historic drill holes. Augusta also completely re-analyzed 10 historic drill holes as a validation of the quality of the historic analyses. In this process they collected over 1,800 samples (9,334 feet) of core. The re-analysis was completed at Skyline laboratories in Tucson.

Re-assay data was appended into the new Hudbay database as a separate column if the hole number matched perfectly with down-the-hole depth intervals. There were 10,056 samples that match this criterion with total footage of about 43,200 feet (13,200 meters). For approximately 1,000 samples, where the down-the-hole depth intervals did not match the original intervals Hudbay employed a weighted average method to import the re-assay information.

Augusta drilled 81 holes from 2005 to 2012 that were sampled and assayed. Laboratory assay certificates for all these holes were provided digitally by Skyline. Hudbay imported these digital certificates directly into their database.

For each original sample interval, every element assayed was ranked using the following criteria shown in Table 12-4. A separate field for each element was populated using the defined ranking with historical information given preference over re-assay, and for copper and molybdenum ranking Wet over XRF analysis. In instances where the original data is missing or reported as “Nil”, the next best ranked value was chosen. An associated data source field for every ranked assay documents the origin of the assay value in the database. Less than detection limits are reported as half the limit and assay unit measurements for the re-assay program were converted to historical units. All fields without any data to report are represented by a “-1” to distinguish them from a missing value.

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TABLE 12-4: DRILL HOLE ASSAY RANKING

Metal	Historical Information		Re-assay at Skyline
	Wet	XRAY
Copper	1	2	3
Molybdenum	1	2	3
Copper Soluble	1		2
Silver	1		2
Lead	1		2
Zinc	1		2
Gold	N/A		1

Augusta measured 391 SG samples from both the historical and their drill programs. This point data was merged into the assay database by matching corresponding sample interval and down-the-hole depth.

12.1.4        Hudbay Assay Information

For the 2014 drilling campaign, Hudbay enlisted the services of an independent consultant to build a custom core logging database using the FileMaker database platform. This database was further enhanced and tested by Hudbay personnel. As the logging and sampling was completed at two different sites (Rosemont Camp and Hidden Valley Camp), separate clones of the main database were created and dispatched onto two laptop computers dedicated to data capture. Core logging and sampling was recorded on to a single database at each camp with up to seven geologists logging core simultaneously using tablets.

The second drilling campaign in 2015 eliminated the usage of Hidden Valley Camp and all core logging activities were centralized at Rosemont Camp. All logging was completed using tablets synced with a centralized FileMaker database hosted on a laptop.

Quality assurance protocols built into the core logging database prevented the loggers from duplicating sample numbers and entering out of range values for sample intervals. Sample types were restricted to a down-drop list in order to prevent several variations of a sample type. For quick data input and to prevent data entry errors, the sampling module was designed to predict the sample number, the interval length and the down-the-hole depths based on previous entered values, while allowing the users to edit and adjust the predicted values at their discretion. The sampling module was improved further in 2015 by pre-assigning sample numbers and sample types for each core logging geologist. This prevented incidents of incorrectly assigning sample types and typos in the sample numbers.

The sampling module portion of the FileMaker database generated color coded reports for each logged drill hole that listed hole name, depth interval, sample number and type of sample. These reports were visually examined by the lead geotechnician prior to core cutting and any sample overlaps or skipped samples were brought to the attention of the on-site database manager. Any discrepancies were fixed immediately in the database and the reports were reprinted.

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The sample dispatch module of the FileMaker database was used to create assay requisition forms to submit samples to the assay laboratory, as well as export the samples included in each requisition to a PDF and CSV file. To maintain consistency, sampling dispatching was handled by the on-site database manager.

In order to minimize data loss, each of the databases was regularly saved. The FileMaker database created a backup of the data three hours on the host laptop. The database manager also saved a copy on a flash drive at the end of the day which was uploaded to a Google Drive. The updated copy of the database was merged into a centralized database at Hudbay’s Toronto office and backed up on a different server on a daily basis.

The database manager in Toronto reviewed all the samples and logging from the previous day and communicated any edits or discrepancies that required amendments by the core loggers. Maintenance and updates to the field databases were carried out on bi-weekly basis.

Digital copies of the assay certificates received from the assay laboratory were imported into the main FileMaker database in Toronto by the database manager using custom-built importers. All fields were imported as they appeared on the certificates without substituting values for codes or special characters. All element fields in the database are named appropriately to include the element name, analytical method and units of measurement. Attribute fields which include hole name, down hole depths, sample type and requisition identification were populated using lookup functions for each sample number. A sample tracking module was created to track all jobs submitted to the assay laboratory along with analytical methods, number of samples and the date samples were sent and received.

SG samples were entered into the logging database by the geologist. The SG measured results were later imported into the assay table by matching the depth of the sample and assay interval.

An Open Database Connectivity was established between FileMaker database and Excel to import sample interval table and the sample assay table. A copy of this Excel workbook was placed in a secure location on the Hudbay server that was accessible to a small group of approved Hudbay personnel. These Excel files were imported into several different software (Minesight, Target for Geosoft, Leapfrog etc.) that allowed further validation of the data.

Overall quality of samples taken, recorded and submitted to the laboratory was excellent with few data entry errors that were identified and corrected. Of the 21,647 samples submitted from the Phase I campaign only one sample was lost. No samples were reportedly lost from the 17,485 samples submitted during the Phase II campaign. Importing the assay certificates directly into the database virtually eliminated any data entry errors.

In the author’s opinion, the drill hole and assay database is acceptable for resource estimation.

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12.1.5        Data Security

The historical assay database and the Hudbay assay database are administered by the database manager with working copies kept on the local drive of a secure computer and backups placed on a secure location on a Hudbay server. Any requests for edits to the database are made to the database manager who updates all the copies. All paper copies of the historical assay certificates and logs are available on the Hudbay’s internal Sharepoint website with restricted access.

Moving forward, all of the Rosemont historical and current data will be migrated to the AcQuire platform that provides robust data security and long-term data storage solutions.

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13     MINERAL PROCESSING AND METALLURGICAL TESTING

13.1        Overview

Recorded metallurgical testwork on Rosemont ores comprises work beginning as early 1974 by Anamax Mining Company. Following Augusta’s acquisition of the Rosemont group of properties in 2005, Augusta continued the work and concluded it with the publication of NI 43-101 technical reports, the first in 2007, followed by another dated August 28, 2012. These reports were authored by M3 Engineering & Technology QPs.

Following its acquisition of Rosemont in 3Q2014, Hudbay completed two drilling programs (the first commencing late 2014 and the second in late 2015) and initiated a series of phased metallurgical testing programs, each designed to advance its understanding of the deposit and metallurgical performance in response to treatment.

A mine planning effort was also initiated, beginning with an effort that utilized the two phases of drilling and associated metallurgical testwork programs that were conducted through 2014 and 2015. In early 2016, an updated block model was developed, and the mine plan was subsequently updated in mid-2016 to reflect the changes in the understanding of the deposit. The new mine plan drove some changes to pit phasing and mining sequences (see additional detail about the updated model and new mine plan in Sections 14 through 16).

The principal objectives of the phased metallurgical testing programs were to:

• Enhance the understanding of the Rosemont resource
• Confirm the quality of the prior metallurgical testwork
• Identify downstream processing methods, forecast recoveries and quality of final products
• Evaluate characteristics of tailing products
• Derive a required ore processing flowsheet and size process equipment

Three composite samples were prepared for metallurgical testing in 2015, and among these were a set of three samples that corresponded to ore that was projected (under the earlier mine plan) to report from the mine during the first five years of operation, the second five years, and a third sample for the balance of operations.

As reporting from the first phase of metallurgical testwork programs became available, a revised set of composites was prepared to further enhance the understanding of the orebody, more particularly as it related to metallurgical performance in the presence of clays, as well as process equipment sizing and selection (principally flotation and dewatering equipment). For this phase of the testwork, a subset of these composites was selected again on the basis of when the ore was scheduled to report from the mine (again, under the earlier draft of the mine plan), this time with date ranges chosen as production years 1 through 3, 4 through 7, and after year 7. Again, due to timing all of these composites were chosen using the production schedule from the original mine plan.

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The period descriptors for these several composites may not be strictly accurate under the revised mine plan developed by Hudbay in 2016 since testing has shown copper recovery to be strongly correlated to the ratio of sulfide copper to total copper, regardless of ore type or location in the deposit. The balance of this report section will discuss:

• Historical metallurgical testwork programs
• Hudbay’s phased metallurgical testwork program
• Principal conclusions and recommendations that result from this body of work

13.2        Historical Metallurgical Testwork Summary

The information presented in the following historical summary was important to Hudbay’s early understanding of Rosemont mineralogy and Augusta’s strategies for liberating metals of interest from the deposit. It was this early review and investigation of the prior work that drove the definition of the subsequent drilling and metallurgical testing program implemented immediately following Hudbay’s acquisition of Rosemont. Subsequent drilling, sampling and metallurgical testing programs discussed in later paragraphs of this Section will provide an appreciation of the evolution of Hudbay’s understanding of the deposit and final conclusions with respect to processing strategies, flowsheet development and forecasts of recoveries and product quality.

13.2.1        Early Work

The earliest reported testwork on Rosemont ores comprising preliminary grinding and flotation tests was completed by Anamax Mining Company in 1974. This early work was followed by a larger testwork campaign by Augusta in 2006-2007 to support the preparation of a feasibility study and technical report. Further testwork was then completed by Augusta between 2008 and 2013 to support engineering design and updates to the original technical report. The description of the Augusta work is provided in this Section and was part of the “Rosemont Copper Project, NI 43-101 Technical Report”, dated August 28, 2012.

Historical metallurgical testwork programs were undertaken by Mountain State R&D International (“MSRDI”), SGS and G&T Metallurgical Services, with dewatering and rheology testing undertaken by Pocock, Outotec and FLSmidth. Early attempts to characterize the deposit were difficult due to the large differences in mineralogy and high degree of variability within the major lithologies. The testwork programs had previously isolated and tested different lithologies and period composites without successfully correlating metallurgical performance with specific ore types

The balance of this Section summarizes the results of the previous metallurgical test programs, those conducted prior to Hudbay’s acquisition of Rosemont. These are more fully described in Augusta’s technical report titled “Rosemont Copper Project, NI 43-101 Technical Report, Updated Feasibility Study” dated August 28, 2012.

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13.2.2        Basis

The earliest existing records of metallurgical testing are from the period 1974 to 1975, at which time grinding and flotation tests were performed. In the first half of 2006, Augusta initiated test-work to provide a better understanding of the metallurgy of the Rosemont mineralization and establish criteria for the design of a process facility.

The program tested both composites and individual variability samples and are considered to be fairly representative of the variety of ore conditions within the deposit.

13.2.3        Mineralization & Ore Types

The ore contains three main copper sulfide minerals (in order of relative abundance): chalcopyrite, bornite, and chalcocite/covellite. The deposit was described as having three major and several minor lithological units, within which the various types of sulfide mineralization occur:

• Horquilla
• Earp
• Colina
• Other including Epitaph and Escabrosa

Two samples of ground Horquilla sulfide material were examined by detailed mineralogical modal analysis. The result of this analysis indicated a large difference in copper mineralogy within the Horquilla rock type and association of silver and gold with the copper sulfide minerals. Molybdenite,

^MoS2^{, was the only molybdenum mineral identified.}

The copper oxide minerals identified as primarily chrysocolla, tenorite, malachite, and azurite. Oxide resources are distributed in three major rock units as follows:

• Arkose
• Porphyry – Quartz Monzonite (“QMP”) or Quartz Latite (“QLP”)
• Andesite

13.2.4        Comminution Work

Grinding mill sizing parameters were provided to mill manufacturers for use in their mill sizing methods. The mill sizing parameters are shown in Table 13-1.

TABLE 13-1: GRINDING MILL SIZING PARAMETERS

Parameter	Value
CWi	4.90
RWi	12.40
BWi	11.40
Tonnage	3,400 tph
SAG Mill Feed Size	150,000µ
Transfer Size	3,000µ
Ball Mill Product Size	105µ

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13.2.5        Flotation Testwork

Flotation test-work was performed during the years 1974 to 1975 and 2006 to 2008. The tests included bench-scale rougher-scavenger and cleaner tests, rougher variability tests, and rougher cleaner optimization tests. Based on the test results the flotation conditions were indicated to be as follows:

• Primary grinding to P80 = 105μ
• Rougher flotation pH= 9.7 to 10.8
• AP-238 and AX-343 collectors
• Regrind to P80 = 74μ
• One stage of cleaner flotation

The result of the variability tests indicated that the grind size has an effect on both copper recovery and rougher concentrate grade. The mineralogical modal analyses indicated that the chalcopyrite liberates at a coarser size, between 150 and 75μ, than bornite and chalcocite. The molybdenite began to liberate from the gangue between 150 and 75μ, but remained locked to a significant degree with gangue to about 22μ.

13.2.6        Molybdenum Testwork

During 2008, flotation tests were conducted at MSRDI on composite samples of five individual rock lithology samples and one composite sample representing the material expected to be processed during the first three years of process plant operation. The test program was designed to examine the process of producing molybdenite concentrate.

The bulk (copper-molybdenite) flotation concentrate from mineralized Horquilla produced a molybdenite concentrate grading 52.7% molybdenum with a 93% molybdenum recovery from bulk concentrate. The results of testing the other samples indicated lower molybdenite concentrate grades and with variable molybdenite recovery from the bulk concentrate with the procedure used. The results of the testing are presented in Table 13-2.

TABLE 13-2: MOLYBDENITE FLOTATION

Molybdenite Flotation
Sample	Concentrate Assay %			Recovery % Mo
	Cu	Mo	Insol
Horquilla	0.44	52.7	1.8	93.0
Colina	0.70	26.5	16.9	96.5
Earp	0.50	42.8	6.5	93.0
Epitaph	0.30	39.3	17.5	55.7
Escabrosa	0.50	27.9	25.8	84.8
1 – 3 Year Composite	0.06	41.6	13.5	96.5

13.2.7        2012 Metallurgical Test Program

In 2012, a metallurgical test program was designed by Augusta to prepare composite samples representing four periods of expected mine production and test them by bench scale test procedures. The test procedures followed the treatment methods proposed for the proposed process plant. The metallurgical test composite samples were prepared from half-core from six holes drilled in late 2011. The half-core drill segments were selected so that the composite samples were representative of grade, lithology, and spatial characteristics of material predicted to be produced during the mine operating periods of years 1 through 3, years 4 through 7, years 8 through 12, and years 13 through 21. The composition of the composite samples by lithology is shown in Table 13-3.

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TABLE 13-3: LITHOLOGY OF COMPOSITE SAMPLES

Lithology	Composite Samples Representing Expected Mine Production Years
	1 through 3	4 through 7	8 through 12	13 through 21
Epitaph			10%	16%
Colina		11%	17%	25%
Earp	16%	28%	23%	16%
Horquilla	84%	61%	50%	43%

The result of closed circuit flotation tests are summarized in Table 13-4:

TABLE 13-4: 2012 CLOSED CIRCUIT FLOTATION RESULTS

Period	Copper Recovery	Molybdenum Recovery	Final Bulk Concentrate Grade
			Copper	Molybdenum	Silver
Yr 1-3	87.9%	62%	41%	1.02%	502ppm
Yr 4-7	81.2%	2.5%	44%	0.047%	NA
Yr 8-12	92%	84%	28%	1.22%	NA
* Yr 13-21 1st composite	75.8%	31.1%	36%	0.56%	NA
* Yr 13-21 2nd composite	91.4%	66%	37%	0.84%	NA

* Note: Core submitted for the years 13 to 21 composite were found to have a higher content of oxidized material than was expected to be mined during that phase of mining and resulted in low overall metal recovery. For that reason, a second composite was assembled containing a lesser percentage of oxidized material and the flotation tests were repeated.

The anomalous value obtained for the molybdenum recovery (years 4-7) was checked by re-testing the same composite sample, resulting in improved rougher concentrate molybdenum recovery of 40% to 60%. Previous results from testing samples containing the Colina mineralization indicated that lower molybdenum recovery was to be expected, however the cause was not specifically known at the time.

Additional analysis of the concentrate produced in the testwork (years 8-12) showed that the concentrate contained low amounts of contaminants such as arsenic (<80 g/ton) and mercury (0.8 g/ton) and contained both gold (1.91 g/ton) and silver (294 g/ton).

13.2.8        Principal Observations

The significant conclusions that can be drawn from the prior metallurgical testing program are as follows:

• Rosemont ore responds well to proven and widely used mineral separation techniques.

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• Concentrate grade vs. recovery and flotation feed grind size vs. recovery relationships were established and follow expected trends.
• Early process plant design criteria were developed on the basis of this work.

13.3        Hudbay Metallurgical Testing Programs

Following the acquisition of the Project in the third quarter of 2014, Hudbay undertook a series of additional drilling, sampling and metallurgical testwork programs. Drilling programs were undertaken in late 2014 and 2015, and are discussed in greater detail in Section 10 of this Report.

In 2014, Hudbay engaged XPS Consulting & Testwork Services (“XPS”) to undertake mineral characterization and metallurgical testwork. Base Met Laboratory (“BML”) was engaged in late 2015 to provide confirmation testwork of the XPS testwork and additional process optimization.

The Hudbay metallurgical testing programs are separated into the following phases:

• XPS Phase 1: Variability Test Program
• XPS Phase 2: Geometallurgical Variables
• XPS Phase 3: Copper / Moly Separation, flotation response to clays
• BML Confirmation Testing
• Production Period Testwork

13.4        XPS Phase 1

In late 2014, Hudbay initiated a variability test program (XPS Phase 1) that would improve the understanding of the mineral and lithological data and help define geo-metallurgical characteristics. The objective was to improve the correlation between mineralogy/geology and metallurgical variability observed in prior metallurgical testwork conducted by others.

The data presented in this section is taken from the 2015 SGS report on grindability characteristics of samples provided by XPS.

All of the samples in this program were analyzed by Inductively Coupled Plasma (ICP), X-Ray Diffraction (“XRD”), Quantitative Evaluation of Minerals by Scanning electron microscopy (“QEMSCAN”), Cation Exchange Capacity (“CEC”) and Near-Infrared (“NIR”). ICP measured elemental assays while XRD measured mineral composition. CEC and NIR analysis determined the clay content and other alteration minerals content.

A total of 140 samples (Met1A, Met1B and Met2 samples) were sent to SGS Canada in Lakefield, Ontario for comminution testing, specifically for the JK drop-weight test, the SPI® test and the Bond ball mill grindability test at a closing size of 150 mesh (106 µm). JK drop-weight tests require coarser material and hence this testing was only possible on the 33 new full HQ core samples (Met2).

Statistics from the comminution test results are summarized in Table 13-5.

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TABLE 13-5: XPS PHASE 1 - COMMINUTION TEST STATISTICS*

Statistic	DWT			SPI (min)	BWi (kWh/tonne)
	Relative Density*	A x b	ta
Average	2.84	46.9	0.54	94.6	13.0
Standard Deviation	0.17	17.1	0.31	54.4	2.5
Minimum*	2.56	94.2	1.49	24.9	8.2
Median	2.83	45.6	0.49	82.1	13.0
75th Percentile	2.94	37.0	0.39	117	14.4
90th Percentile	3.06	25.1	0.22	151	16.0
Maximum*	3.28	18.6	0.14	401	21.7

* Reference: 5a-SGS-14816-001 FINAL Rpt Apr 17 2015.pdf

13.4.1        General Observations

Hudbay’s technical services team initiated an effort to map geochemical characteristics of the various ore types, intending to utilize this data as predictors of recovery on the basis of ore type, indicators of clay distribution in the orebody, and other proxies. While this work has produced results, a stronger indicator of recovery is the sulfide component of total copper in the sample. Hudbay will continue to develop the geochemical database with the intent to leverage its value during subsequent project execution and operational phases.

Results from the XPS Phase 1 sample characterization and testwork program suggested the following:

• There was significant variability within the major lithologies with respect to copper oxide, clay content and ore hardness.
• All of the samples tested contained measurable amounts of swelling and/or magnesium- bearing clays, indicating widespread clay presence in the deposit and across all ore types.
• Swelling clay, as identified by its CEC content, varied widely from 4% to over 30%, averaging about 8.5%.
• Magnesium-bearing clays, including serpentine and talc, varied from 0% to over 35%, with an average of 1.8%.

Sample analysis results also showed that copper oxide varies widely, from 0% to 90%, averaging about 5.4% for these samples.

13.4.2        Comminution Results

Both JK drop-weight test (“DWT”) and Bond ball mill work index (“BWI”) results ranged from very soft to very hard while SAG Power Index (“SPI”) test results ranged from soft to very hard. A 75^th percentile DWT Axb of 37.0 was determined based on 33 samples, while a 75^th percentile BWI of 14.4 kWh/ton was determined based on 140 samples. A 75^th percentile parameters were chosen as the basis for design of the comminution circuit.

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13.4.3        Flotation Results

The XPS Phase 1 flotation program consisted of rougher kinetic flotation tests for 107 samples (Met1A and Met1B). These tests followed a standardized flotation schedule drawn from previous work, employing the following conditions:

• PAX 50 g/tonne (0.1 lb/ton).
• Fuel Oil 30 g/tonne (0.06 lb/ton).
• MIBC 30 g/tonne (0.06 lb/ton).
• Lime to pH 9.5.
• Target grind P80 105 μm.

There was no attempt in Phase 1 to optimize flotation conditions. Grind determinations were recorded for each test. Lab results report the following average (rougher feed) grind sizes by sample type:

Sample	P80, μm	Std Deviation, μm
Met1A	94	21
Met1B	107	31

Clay content, and specifically swelling (CEC) clay content, appears to have an impact on flotation recovery. However, the impact is not consistent across all of the tests. Samples with swelling clay content above about 12% typically yielded lower rougher recoveries.

Rougher flotation results showed a high level of variability, however, a strong correlation was determined between oxide copper (defined as the ratio of acid soluble copper to total copper) and rougher flotation recovery as shown in Figure 13-1.

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FIGURE 13-1: % COPPER ROUGHER FLOTATION RECOVERY VS % ACID SOLUBLE / TOTAL COPPER

13.5        XPS Phase 2

The principal objective of the XPS Phase 2 testwork was to investigate the key geo-metallurgical variables identified in Phase 1, which are:

• Copper oxide content
• Swelling clays
• Magnesium clays
• Ore hardness

A series of composites were prepared, three according to production period criteria (production years 1 through 5, 6 through 10, and 11 through LOM), and four geometallurgical subtype composites and were tested under the Phase 2 testwork program as follows:

• Base 1: Main sulfide ore, Production Years 1 – 5
• Base 2: Main sulfide ore, Production Years 6 – 10
• Base 3: Main sulfide ore, Production Years > 10
• Sub 4: ~25% copper oxide ore
• Sub 5A: Swelling clay (CEC) rich ore >10%
• Sub 5B: Magnesium clay (Serpentine and talc) rich ore >10%
• Sub 6: Hard sulfide ore with BWi >14
• Note: “Production years” above refers to the preliminary, not final 2016 mine plan

The results of this phase of testwork are reported in the Geometallurgical Program Technical Review (XPS, 2015b).

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13.5.1        Mineralogy

Phase 2 composites were submitted for mineralogical characterization using XRD with Rietveld Refinement, CEC analysis and QEMSCAN along with Electron Probe Micro-analysis (“EPMA”) to validate phase compositions and copper deportments.

XRD analysis was completed as a cross-check to the QEMSCAN analyses and was found to correlate well. A summary of the QEMSCAN data is presented in Table 13-6. As QEMSCAN is limited in its ability to isolate and quantify both fine clays and swelling species, CEC analysis was used to define the swelling-clay content. These results are presented in Table 13-7.

TABLE 13-6: XPS PHASE 2 - QEMSCAN ANALYSIS

Mineral	Base 1	Base 2	Base 3	Sub 4	Sub 5A	Sub 5B	Sub 6
Serpentine Talc Clays	2.2	3.4	6.4	1.3	2.6	18.7	0.9
Muscovite	0.5	0.9	1.0	0.8	3.4	0.5	2.4
Biotite	0.7	1.1	1.4	1.3	4.8	1.9	3.5
Chlorite	1.7	2.6	1.8	2.7	4.0	3.0	2.0
Quartz	23.3	15.4	7.1	25.5	24.5	0.3	19.1
K-Feldspar	7.0	8.4	3.1	9.2	13.6	0.4	21.7
Garnet	24.2	16.5	14.6	21.7	8.5	5.8	11.0
Calcite	17.9	26.9	39.5	23.3	14.6	40.5	6.4
Pyrite	0.4	0.5	0.2	0.1	0.3	0.7	0.3
Chalcopyrite	0.4	0.9	1.0	0.3	0.7	1.0	0.7
Bornite	0.2	0.2	0.1	0.1	0.1	0.3	0.3
Chalcocite/Cov.	0.1	0.2	0.1	0.2	0.3	0	0.1
Cu Oxide-other	0.1	0.3	0.1	0.8	0.1	0.1	0.1
Other	21.3	22.7	23.6	12.7	22.4	26.8	31.5
Total	100.0	100.0	100.0	100.0	100.0	100.0	100.0

TABLE 13-7: XPS PHASE 2 - CEC ANALYSIS

Test	Base 1	Base 2	Base 3	Sub 4	Sub 5A	Sub 5B	Sub 6
CEC	7.1	9.5	5.9	6.3	10.6	6.1	8.3

Based on a combination of the EPMA mineral composition data and the QEMSCAN modal abundance data, the elemental deportment by mineral can be calculated for copper as is summarized in Table 13-8.

TABLE 13-8: XPS PHASE 2 - COPPER DEPORTMENT BY MINERAL SPECIES

Minerals	Base 1	Base 2	Base 3	Sub 4	Sub 5A	Sub 5B	Sub 6
Chalcopyrite	33.4	49.1	67.2	24.8	39.4	58.1	48.1
Bornite	31.5	18.2	16.2	7.6	11.1	32.7	28.9
Chalcocite	20.5	22.0	8.3	32.9	33.1	3.1	18.5
Covellite	6.1	1.3	0.1	1.3	4.8	0.0	0.1
Other Sulfides	0.4	0.3	0.2	0.4	4.3	0.3	0.1
Total Sulfide Copper	92.0	90.9	92.1	67.1	92.8	94.3	95.6
Chrysocolla	0.6	2.1	0.9	3.9	0.1	0.0	0.1
Cu Chlorite	3.6	3.6	3.3	5.3	5.1	4.1	3.0
Goethite	0.8	0.8	0.3	3.2	0.5	0.4	0.3
Cu Oxide Other	2.8	2.3	3.4	20.0	1.5	0.9	1.0

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Minerals	Base 1	Base 2	Base 3	Sub 4	Sub 5A	Sub 5B	Sub 6
Other	0.1	0.2	0.0	0.1	0.0	0.1	0.0
Total Oxide Copper	8.0	9.1	7.9	32.9	7.2	5.7	4.4
TOTAL	100	100	100	100	100	100	100
Cu-Oxide by Assay	5.1	7.7	7.1	24.7	3.8	5.7	3.0

In general terms, testwork conducted on the composited samples provide the following as indicators of variability in the orebody:

• Copper oxide content is variable and continues at depth.
• Copper deportment to chalcopyrite increases with depth while chalcocite and bornite contents are reduced.
• Base 2 composite has relatively high molybdenum levels when compared to others.
• Base 2 composite has elevated CEC clay content relative to Base 1 and Base 3.
• Serpentine and talc content increases with depth (Base 3 has three times the Mg clay content of Base 1)
• Fe and Si contents decrease with depth, while Ca and Mg contents increase.
• Calcite content increases from Base 1 to Base 3.

13.5.2        Results - Phase 2 Flotation Testwork

XPS conducted over 104 flotation tests to investigate the effect of primary grind size, reagents, pH modifiers, dispersants and rougher and cleaner pulp densities in parallel with locked cycle testing. A primary grind size of 140 μm was selected for subsequent flotation testing on all composites.

Locked cycle tests were undertaken on Base 1, Base 2, Sub 4 and Sub 5A. The locked cycle flotation test results are given in Table 13-9.

TABLE 13-9: XPS PHASE 2 - LOCKED CYCLE TEST RESULTS

Sample	Test	Stream	Mass Percent	Grade - Percent		Distribution - Percent
				Cu	Mo	Cu	Mo
Base 1	Float 094 Cycles 3-5	Feed	100	0.58	0.012	100	100
		Cu Concentrate	1.5	31.2	0.451	82.7	59.7
		Cu Cleaner Tail	4.8	0.69	0.041	5.7	16.8
		Cu Rougher Tail	93.7	0.07	0.003	11.6	23.6
Base 2	Float 114 Cycles 4-6	Feed	100	0.59	0.022	100	100
		Cu Concentrate	1.7	22.1	0.394	64.0	30.1
		Cu Cleaner Tail	11.0	0.88	0.070	16.5	34.7
		Cu Rougher Tail	87.3	0.13	0.009	19.6	35.2
Base 2	Float 145 Cycles 4-6	Feed	100	0.59	0.022	100	100
		Cu Concentrate	1.6	28.2	0.624	74.7	44.6
		Cu Cleaner Tail	10.4	0.30	0.037	5.3	17.4
		Cu Rougher Tail	88.0	0.14	0.010	20.0	38.0
Sub 4	Float 104 Cycles 4-6	Feed	100	0.54	0.005	100	100
		Cu Concentrate	1.0	31.9	0.198	61.3	43.3
		Cu Cleaner Tail	3.1	0.75	0.009	4.4	5.7
		Cu Rougher Tail	95.8	0.19	0.003	34.3	51.0
Sub 5A	Float 103 Cycles 4-6	Feed	100	0.46	0.010	100	100
		Cu Concentrate	1.7	22.6	0.327	81.6	54.4

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Sample	Test	Stream	Mass Percent	Grade - Percent		Distribution - Percent
				Cu	Mo	Cu	Mo
		Cu Cleaner Tail	6.9	0.44	0.033	6.6	22.6
		Cu Rougher Tail	91.5	0.06	0.003	11.8	22.9

13.6        XPS Phase 3

The XPS Phase 3 of work focused on the separation of copper and molybdenum, investigating the flotation response of blended high-clay ore mixtures and the effect of calcium rich water on flotation behavior.

13.6.1        Copper-Molybdenum Separation Testwork

Copper-Molybdenum separation testwork was of necessity constrained by the limited sample available, a consequence of the relatively small quantity of Cu/Mo concentrate available for testwork. Subsequent paragraphs will describe general conclusions that may be drawn from the copper-moly testwork that has been completed, and suggests some additional work that may be required during project execution and operational phases of the Project.

Stored cleaner 1 concentrate from XPS Phase 2 tailings generation work was upgraded to produce a copper-molybdenum bulk concentrate suitable for copper-molybdenum separation. Application of three additional stages of cleaning raised the copper grade to 35.8%, 40.6% and 41.4% copper, respectively.

The copper cleaner 4 concentrate, or copper molybdenum bulk concentrate, was split for assays, mineralogical analysis, a scoping separation test and the remainder used for the copper molybdenum demonstration test.

To depress the copper minerals, the pH of the bulk concentrate was elevated to pH 12 with lime and then treated with sodium hydrosulfide (NaHS). Unfortunately for this phase of the testwork, there was not enough molybdenum rougher concentrate to do more than a single stage of molybdenum cleaning.

13.6.2        Observations & Discussion

While the results above suggest that additional stages of cleaning may be required to produce an acceptable bulk copper concentrate, subsequent analysis has demonstrated that acceptable CuMo concentrate grades can be achieved by applying staged flotation reactor (“SFR”) flotation technology in the bulk cleaner circuit flowsheet. SFR flotation cells have been proven to achieve recovery rates equivalent to or better than conventional flotation cells while realizing exceptionally high upgrade ratios. This performance is accomplished though separating the three phases of flotation (particle collection, bubble disengagement, and froth recovery) into different zones such that each phase can be optimized independent of the others. To further enhance upgrading performance, under-froth dilution/wash water can be applied in the froth recovery unit to significantly improve gangue rejection from the concentrate product. Accordingly, the incorporation of SFR flotation technology in the flotation circuit design supersedes the need for additional stages of cleaning.

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13.6.3        Results – Phase 3 Copper – Moly Separation

The results from this test are shown in Table 13-10.

TABLE 13-10: XPS PHASE 3 - MOLYBDENUM SEPARATION TEST

	Mass %	Grades			Recovery
		Cu %	Mo %	MgO %	Cu %	Mo %	MgO %
Mo Cleaner Concentrate	0.4	3.45	30.7	5.38	0.0	38.5	7.8
Mo Rougher Concentrate	1.5	19.4	18.0	4.48	0.7	79.6	22.8
Mo Rougher Tails	98.5	41.2	0.07	0.23	99.3	20.4	77.2
Cleaner 4 Concentrate (Recalculated)	100.0	40.9	0.34	0.29	-	-	-

Copper was depressed producing a molybdenum concentrate of 30.7% molybdenum upgraded from 18% and 0.34% molybdenum in the molybdenum rougher concentrate and bulk concentrate respectively. This corresponds to 79.6% molybdenum recovery from the bulk concentrate to the molybdenum rougher concentrate and 38.5% molybdenum recovery to the molybdenum cleaner. Additional cleaning stages were not pursued in this program due to the small sample size available.

13.6.4        Flotation of Composite Blends and Water Testwork

Blends were made up of clean ore (“Sub 6”), swelling-clay ore (“Sub 5A”) and magnesium-clay ore (“Sub 5B”) to determine if the blends behaved as the sum of the components. Additionally, one test was conducted with calcium saturated water. This water was used in all stages of grinding and flotation for Float 158, which was otherwise identical to the fresh water equivalent test in Float 150. The results are summarized in Table 13-11 and Table 13-12.

TABLE 13-11: XPS PHASE 3 - FLOTATION BLENDS TEST RESULTS

Test	Sub 6	Sub 5A	Sub 5B	Rougher		Cleaner 1, 2		Overall Circuit
				R%Cu	%Cu	R%Cu	%Cu	R%Cu	%Cu
Float 105	100	-	-	92.0	16.6	93.9	29.3	86	29.3
Float 093	-	100	-	90.4	3.7	77.3	22.1	70	22.1
Float 121	-	-	100	75.8	1.7	74.4	24.3	56	24.3
Float 150	75	25	-	88.4	8.7	94.1	28.7	83	28.7
Prediction	75	25	-	91.6	9.4	90.2	27.6	83	27.6
Float 149	75	-	25	88.8	4.7	70.5	24.5	63	24.5
Prediction	75	-	25	87.8	5.7	89.5	28.2	79	28.2
Float 151	75	12.5	12.5	86.7	6.2	70.7	27.3	61	27.3
Prediction	75	12.5	12.5	89.7	7.1	89.9	27.9	81	27.9
Float 152	50	25	25	85.4	3.5	75.2	25.6	64	25.6
Prediction	50	25	25	87.3	4.4	85.6	26.4	75	26.4

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TABLE 13-12: XPS PHASE 3 - FLOTATION WATER TEST RESULTS

	Sub 6	Sub 5A	Sub 5B	Rougher		Cleaner 1, 2		Overall Circuit
				R%Cu	%Cu	R%Cu	%Cu	R%Cu	%Cu
Float 150	75	25	-	88.4	8.7	94.1	28.7	83	28.7
Float 158*	75	25	-	91.0	10.6	79.7	29.9	73	29.9

* High calcium water test.

The rougher results were close but generally slightly less than the mathematical weighted sum of the end member floats. However, cleaner results departed from predictions for the magnesium-clay blends as well as for the calcium-saturated float. Cleaner tests were conducted at low densities (ranging from 4 to 18% w/w solids) leaving reagent strategy, dispersants and physical set-up as possible mitigation factors.

13.7        BML Confirmation Testing

BML conducted a testwork program at their Kamloops laboratory in late 2015. The main objective was to confirm the flotation process parameters developed by XPS during the Phase 1 and Phase 2 testwork and to replicate the locked cycle results with the Base 1 and Base 2 samples. Results of this testwork are summarized in BML report BL065.

The test program was completed using two period composites, Base 1 and Base 2. The main difference between these replicate tests and previous XPS tests was the use of traditional manual froth scraping and shorter flotation times. More specifically, tests conducted at BML were performed with constant froth levels and a technician manually recovering froth, while the XPS program used a mechanical froth paddle system with level manipulation to achieve desired mass recoveries.

Further, BML locked cycle tests used three stages of cleaning and returned the cleaner scavenger concentrate and cleaner 2 tail to cleaner 1 rather than regrind as these streams should already be sufficiently reground.

The basic parameters of the XPS flowsheet were confirmed for Base 1, however, BML’s Base 2 LCT grade-recovery performance was different, producing a higher copper concentrate grade of 34.5% (versus 28.2%) and lower copper recovery of 63.8% (versus 74.7%) . Analysis of these results suggests that they are essentially different points on the same grade-recovery curve.

13.8        BML Production Period Testwork

In early 2016, production period samples, based on the 131-million tons/y mine plan developed for Hudbay by Independent Mining Consultants (“IMC”) in November 2015, were bench tested for additional metallurgical and project engineering data. The purpose of this program was to add detail to the prior metallurgical testwork results in key areas in order to improve confidence in recovery and concentrate quality forecasts, as well as for sizing downstream process equipment, most notably the tailing dewatering and filtration equipment necessary to accomplish the drystack tailing deposition strategy. Results of this testing program are recorded in BML report BL076.

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A program of batch and locked cycle testing was performed using the previously developed flowsheet and basic reagent scheme. In addition, SFR rougher pilot testing was done in parallel to the bench testing. Due to wall effects of the small SFR pilot units, the results of the pilot plant rougher tests fell short of bench test (conventional) results, thus SFR cells were not incorporated in the rougher flotation circuit design.

As noted previously, these “period composites” are not strictly accurate when referenced to the new 2016 Mine Plan; regardless, given the strong correlation of copper recovery to the sulfide copper component of the ore, moderate shifting of period composites in terms of production phase are not expected to materially impact the recovery conclusions.

TABLE 13-13: PRODUCTION COMPOSITES

Composite	Years	Grade		Acid Soluble Cu
		%Cu	%Mo	ASCu/TCu, %
Period	1 1 – 3	0.54	0.013	9
Period	2 4 – 6	0.48	0.013	13
Period	3 7 - LOM	0.72	0.018	13

13.8.1        Production Period Mineralogy

The feed sample mineral content was analyzed by SEM-EDX and XRD and the results presented in Table 13-14.

TABLE 13-14: BML MINERAL CONTENT

	SEM-EDX			XRD
Mineral	Period 1	Period 2	Period 3	Period 1	Period 2	Period 3
Chalcopyrite	0.39	0.44	0.59
Tetrahedrite	0.01	0.01	0.02
Bornite	0.21	0.29	0.61
Chalcocite/Covellite	0.07	0.13	0.29
Other Copper	0.01	0.00	0.05
Sphalerite	0.04	0.05	0.07
Arsenopyrite	0.00	0.01	0.00
Pyrite	0.26	0.28	0.60
Molybdenite	0.00	0.00	0.02
Fluorite	0.1	0.1	0.1
Apatite	0.1	0.2	0.1
Carbonates	7.3	16	16	11	21	20
Oxides	0.4	0.5	0.5
Quartz	25	12	12	18	9	10
Feldspars	16	18	15	18	17	17
Mica	1.5	2.1	1.8	0.0	2.5	1.5
Pyroxene	22	23	24	20	19	20
Clays	0.2	05	0.4	7.0	7.2	8.4
Olivine	0.0	0.1	0.1
Talc	0.1	0.7	0.3

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	SEM-EDX			XRD
Mineral	Period 1	Period 2	Period 3	Period 1	Period 2	Period 3
Epidote	3.0	1.8	2.3	1.2	0.0	0.0
Chlorite	0.9	2.4	3.3	2.3	0.0	3.8
Garnet	19	10	16	22	13	16
Amphibole	1.0	1.7	1.6	0.0	1.6	1.8
Serpentine	0.0	4.0	2.0	0.0	4.0	0.8
Other minerals	2.0	5.5	2.6	1.2	6.3	1.6
Total	100.0	100.0	100.0	100.0	100.0	100.0

The chemical, mineralogical and clay analyses were performed using the same laboratories as the previous Phase 2 test program. The swelling clay content as determined by CEC is given in Table 13-15.

TABLE 13-15: BML CEC ANALYSIS

Test	Period 1	Period 2	Period 3
CEC	7.0	7.2	8.4

The composite samples were dominated by pyroxene, quartz, feldspars, garnet and carbonates. Previous testwork indicated several minerals that may affect flotation performance and include the serpentine group minerals, swelling clays (CEC), talc and fluorine bearing minerals (apatite, fluorite and micas). Period 1 had the lowest levels of these minerals while Period 2 had the highest.

As determined by SEM-EDX, sulfide content ranged from 1 to 2 percent of the feed, and were for the most part copper minerals with low levels of pyrite and trace levels of sphalerite, tetrahedrite and arsenopyrite.

The previously developed flowsheet and reagent scheme were used as a basis in the production period test program. The primary grind size was 140µm, and small changes in reagent additions and trial of gangue dispersants were investigated. Preliminary rougher and cleaner tests were performed prior to executing the locked cycle tests.

13.8.2        Results of Production Period Testwork

Rougher flotation tests indicated the addition of sodium hexametaphosphate (“SHMP”), a dispersant/chelating agent improved control of the non-sulfide gangue. The performance of Period composites 2 and 3 improved copper recovery by 8 and 3 units respectively. The batch cleaner tests showed shorter regrind time and reduced copper losses from cleaner tailings stream.

Based on BML results, the locked cycle tests averaged 97% recovery of the sulfide copper in the roughers and 93.7% copper recovery in the cleaners. Concentrate grade varied from 32% for Period 1, 34% for Period 2 and 38% for Period 3. The locked cycle test results for the Period composites together with the XPS replicate testing done previously at BML is given in Table 13-16.

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	Form 43-101F1 Technical Report

TABLE 13-16: BML LOCKED CYCLE TEST RESULTS

Sample	Test	Stream	Mass Percent	Grade - Percent		Distribution - Percent
				Cu	Mo	Cu	Mo
Base 1	Float 011 Cycles D+E	Feed	100	0.58	0.012	100	100
		Cu Concentrate	1.2	36.5	0.298	78.5	32.2
		Cu Cleaner Tail	7.1	0.79	0.046	9.7	28.0
		Cu Rougher Tail	91.7	0.08	0.005	11.9	39.8
Base 2	Float 012 Cycles D+E	Feed	100	0.61	0.022	100	100
		Cu Concentrate	1.2	32.7	0.586	64.0	32.0
		Cu Cleaner Tail	12.6	0.44	0.029	9.1	16.9
		Cu Rougher Tail	86.2	0.19	0.013	26.9	51.0
Base 2	Float 014 Cycles D+E	Feed	100	0.60	0.022	100	100
		Cu Concentrate	1.1	34.5	0.558	63.8	28.0
		Cu Cleaner Tail	10.3	0.56	0.035	9.6	16.1
		Cu Rougher Tail	88.5	0.18	0.014	26.6	55.8
Period 1	Float Avg. 20,23,26 Cycles D+E	Feed	100	0.53	0.011	100	100
		Cu Concentrate	1.5	31.9	0.620	90.4	82.7
		Cu Cleaner Tail	3.7	0.45	0.013	3.1	4.3
		Cu Rougher Tail	94.8	0.037	0.002	5.9	13.0
Period 2	Float Avg. 21,24,27 Cycles D+E	Feed	100	0.48	0.011	100	100
		Cu Concentrate	1.05	33.9	0.497	73.1	48.8
		Cu Cleaner Tail	4.3	1.03	8.7	6.6	9.5
		Cu Rougher Tail	94.6	0.093	0.005	18.2	41.6
Period 3	Float Avg. 22,25,28 Cycles D+E	Feed	100	0.73	0.016	100	100
		Cu Concentrate	1.45	37.6	0.625	75.0	57.0
		Cu Cleaner Tail	4.1	1.24	0.024	7.1	6.4
		Cu Rougher Tail	94.4	0.14	0.006	18.0	36.6

13.9        Concentrate Quality

Fluorine is a deleterious element identified in Rosemont copper concentrates. Fluorine levels in copper concentrate above 350-400 ppm typically incur a penalty, with concentrates often rejected by smelters at fluorine levels greater than 900-1,000 ppm.

To mitigate the risk of copper concentrate rejection in the event of higher than normal fluorine levels, the maximum design fluorine grade in concentrate is 800 ppm. Cleaner flotation tests show that upgrading the concentrate rejects entrained fluorine. Available fluorine assays from locked cycle test of 8 concentrate samples are summarized in Figure 13-2. Four of these tests were carried out with only two stages of cleaning; the results suggest that additional cleaning tends to improve copper grade and decrease fluorine levels.

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	Form 43-101F1 Technical Report

FIGURE 13-2: LCT FINAL CONCENTRATE FLUORINE LEVELS

The LCT results show that fluorine can be managed to acceptable levels (less than 800 ppm) with the use of two stage cleaning and froth washing. Good results were achieved with the Base 1 (Year 1-5) composite. Fluorine levels tended to be higher for later period composites.

Testing of concentrates also indicated that other common contaminants, such as arsenic and mercury, were found in concentrations sufficiently low to not warrant concern.

Given the presence of secondary copper sulfides and the need to produce a higher grade concentrate to reject fluorine, final concentrate grades may vary y from 30-38% Cu. Nominal concentrate grades of 32% Cu for Years 1-5 and 33% Cu for Years 6-LOM were selected for equipment sizing. Operating assumptions included in the mill production model utilize concentrate grades of 32% Cu for operating years 1 through 3, 34% in year 4, and 35% in year 5 and beyond on the basis of expected increase in requirement to reject fluorine from concentrate.

13.10        Tailings Dewatering

Tailings samples were generated by XPS to be tested by Andritz, Bilfinger, FLSmidth (FLS), Outotec and Pocock for water separation and recovery from tailing prior to deposition in the DSTF. As expected, clay content and size distribution has a significant effect on tailings dewatering. The samples with lower clay content (Base 1 and Sub 4) generally achieved the highest thickener underflow densities. As expected, specific high clay samples (Sub 5A and Sub 5B) achieved lower densities across most tests.

On average, the high compression thickener tests achieved underflow densities 3%–4% higher than the high rate thickening tests. Generally, high rate thickeners could be expected to achieve an underflow density of 65% for lower clay content ore, while high compression thickeners could be expected to achieve an underflow density up to 65% even for higher clay content ore (Sub 5A and Sub 5B).

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	Form 43-101F1 Technical Report

A solids loading rate of 0.6 ton/m²/h was sufficient to achieve the target underflow density and samples achieved an overflow clarity lower than 200 ppm based on dynamic thickening testwork.

Tailings Proctor compaction testwork indicated the maximum moisture criteria to achieve compaction of 15.2% (equivalent to a dry-weight-basis moisture of 18%). The target moisture for tailings filtration was therefore deemed to be 15%.

A key outcome from the filtration testwork was that membrane filters can achieve lower moisture content at higher machine throughputs compared to the chamber, or recessed plate, filter press. The 15% moisture target was generally achieved after one minute with membrane filters. Increasing feed pressure and air blowing times generally improved the results.

The tailings material produced during the Period composite testing was sent for filtration and thickening tests. Together with the previous filtration results, the laboratory filtration rates were scaled to industrial sized filter criteria (pressure, mechanical time, filtering area, etc.) to determine the number of filters required in the engineering design. Filter sizing and counts considered the anticipated clay content as a component of mill feed according to the mine plan, and as determined by the resource clay proxy model.

13.11        Recovery Estimates

On the basis of the body of testwork that exists, including both the historical testwork, and the testing programs completed by Hudbay since the acquisition of Rosemont in 2014, forecasts of recovery, concentrate grade and quality, as well as characteristics of the resultant tailing product have been developed. The following paragraphs summarize the best estimate of these criteria.

13.11.1        Copper (Cu)

The results from the XPS Phase 1 and 2 and BML Replicate and Period testing as well as prior locked cycle tests (LCTs) confirmed there is a strong relationship between copper recovery and the content of oxide copper in the feed, as determined by acid soluble assay. Overall, rougher flotation recoveries of the copper sulfide component of the feed averaged 96.5%, and cleaner flotation recoveries averaged 93% of the copper sulfide component. The results from the various test programs were consolidated and modelled and resulted in the following equation to forecast recovery of copper as a function of total and ASCu in the feed:

Copper Recovery (%) = (1-ASCu:TCu) x 90

Overall copper recovery corresponding to the 2016 mine plan presented in this report is summarized in TABLE 13-17, expressed as percent of TCu:

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	Form 43-101F1 Technical Report

TABLE 13-17: PRODUCTION RECOVERY PROFILE

	Production Years
	1−5	6−10	11−15	16−LOM	1−LOM
Head Grade %Tcu	0.53%	0.53%	0.39%	0.28%	0.45%
Ratio %AsCu	10.0%	8.6%	12.0%	15.2%	10.6%
Copper Recovery	81%	82%	79%	71%	80%

Note: Based on Mine Plan RP16AUG Mine Plan

13.11.2        Molybdenum (Mo)

The ability to fully characterize molybdenum recoveries has been hampered because of limited sample availability on which to conduct testing. Nevertheless, a reasonable effort has been made to forecast molybdenum recoveries and molybdenum concentrate quality.

The limited XPS and BML copper-molybdenum separation testing provides the basis for the viability of producing separate molybdenum concentrate. In the production year composites, the Mo recovery into the Cu/Mo concentrate ranged as follows:

• Period 1 Composites: 83%
• Period 2 Composites: 49%
• Period 3 Composites: 62.8%

The Cu/Mo separation testing into Mo rougher concentrate reported Mo recoveries of 94% with 92% of the Cu reporting to the tails or Cu concentrate. A Mo separation circuit recovery factor of 90% was applied to the period sample bulk recoveries to estimate overall recovery of Molybdenum as follows;

Production Period	Mo Recovery
Period 1 (Yrs. 1-3)	74.4%
Period 2 (Yrs. 4-6)	43.9%
Period 3 (Yrs. 7-LOM)	51.3%
LOM Average	53.4%

In accordance with the RP16Aug Mine Plan and corresponding mill production schedule, the calculated Mo recovery figures were adjusted slightly to account for expected additional losses during the initial production ramp-up period.

Mo rougher concentrate grades were low (15%-22%) and additional testing is recommended, especially to improve performance in the presence of talc.

13.11.3        SILVER (Ag)

The head samples in XPS and BML testing ranged from 4.5 -9.0 grams per ton, averaging approximately 6 grams per ton in the feed ore. The Ag assays in the copper concentrates ranged from 260 to 460 grams per ton averaging 323 grams per ton. There were no assays performed on the tails, the Ag recovery is estimated based on feed and concentrate mass and assays.

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	Form 43-101F1 Technical Report

The 2016 production schedule developed by Hudbay processes ore with an average Ag content of 0.16 troy ounces/ton or 5.4 g/ton, similar to the test composite samples. Given an estimated Ag grade of the Cu concentrate of 323 g/ton (average test composite assays) the average recovery is:

Average Ag Recovery LOM = 74.4%

13.11.4        GOLD (Au)

Similarly to the silver, gold recovery can be estimated based using ore feed and concentrate grades as there are no tails assays.

The head samples in the XPS and BML testing ranged from 0.03 -0.05 grams per ton, averaging approximately 0.04 grams per ton in the ore feed. The Au assays in the copper concentrates ranged from 1.1 to 2.8 g/ton, averaging 2.1 g/ton. Similar to the Ag recovery, the gold recovery estimate is based on feed and concentrate assays;

Average Au Recovery LOM = 65.1%

However, gold had not been systematically assayed in all the drill holes and is therefore not part of the 2016 resource estimate. Nevertheless, 66 drill holes were assayed for gold, representing 17% of all the assaying conducted on the property. A geostatistical analysis was performed on the drill holes with gold results and has shown good similitude between the gold grade assayed in the drill holes and the gold grade assayed in the metallurgical tests heads.

13.12        Conclusions and Recommendations

The following paragraphs summarize, on the basis of the preceding narrative describing both the historical metallurgical testing programs, as well as the programs completed by Hudbay in the time since acquiring the Project. These recommendations will serve as the basis for the production phase recovery criteria that will drive inputs into the economic model for the Project.

Principal conclusions:

1)	Despite the work of Augusta and the extensive work of Hudbay on matters of lithology and ore type, as well as to associate recoveries by ore type, the overwhelming evidence of the testwork, both past and present is that recovery is driven primarily by the component of soluble copper in the ore sample. When the grade of the sample is discounted by the amount of oxide ore in the sample, recovery of the remaining copper in sulfide minerals is 90%.
2)	While lithology appears not to control to any significant degree the recovery of sulfide copper, it does appear to influence molybdenum recovery and more importantly grade of molybdenum concentrate.
3)	Contaminants (Fluorine) may affect concentrate quality, but subsequent testwork, completed late in the period of time that Hudbay has been evaluating this project suggests that fluorine can be controlled in copper concentrate through flotation equipment selection and operational strategies in a way that minimizes concerns with respect to concentrate quality over the life of the mine.

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4)

Clay in the ore with high concentrations was shown to have a detrimental effect on flotation performance under standard conditions, however, altering conditions (reducing pulp density, dispersing agent addition) was shown to counteract the negative effects. On average, clay content in the ore is expected to remain below concentrations that could result in reduced metal recovery, and when elevated clay contents are encountered, mitigating operational strategies in the process and blending strategies in the mine can be invoked if necessary to offset any negative effects.

5)

The tailing properties have been sufficiently characterized as well as the dewatering performance of vendor equipment over the life of the operation to satisfy the estimated number, type and size of tailing filters for this Project. To be conservative, expansion space has been allocated for additional filtering equipment, to the extent that it may be necessary.

13.13      Discussion and Recommendations

While the production period composites used in the latter stages of the testwork campaign are no longer an exact match for the operating years they were intended to represent due to recent changes to the mine plan, the test results are nonetheless representative of changes in ore conditions as mining progresses deeper into the deposit.

It is recommended that efforts continue in seeking to utilize the extensive geometallurgical database to identify trends and metallurgical indicators that can inform and optimize production plans.

There is ample information in the database regarding serpentine group minerals, but not specifically for talc. There remains some uncertainty for Molybdenum production with respect to the occasional presence of talc in test samples and its potential to interfere with the production of saleable Mo concentrate. It is recommended to study the occurrence of talc in the deposit to better understand the potential effects on Mo production.

Although it has been concluded that fluorine can be readily rejected from copper concentrate, further study is recommended to develop understanding of ore conditions and indicators that trigger elevated fluorine content in concentrate.

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	Form 43-101F1 Technical Report

14     MINERAL RESOURCE ESTIMATE

Hudbay prepared a 3D block model of the Rosemont deposit using MineSight® version 11.00 -3, an industry standard commercial software that specializes in geologic modelling and mine planning. A Lerchs-Grossman (“LG”) cone algorithm was applied to the block model to establish the component of the deposit that has a “reasonable prospect of economic extraction”. The 3D block model and determination of the mineral resources at the Rosemont deposit were performed by Hudbay personnel following Hudbay procedures. The work was reviewed and approved by Cashel Meagher, P.Geo., Chief Operating Officer for Hudbay, Qualified Person and author of this Technical Report.

14.1        Key Assumptions of Model

As shown in Table 14-1, there are 356 drill holes totalling approximately 510,951 feet within the Rosemont database used to support the mineral resource estimation.

TABLE 14-1: DRILLING DATA BY COMPANY

Company	Time Period	Number of Drill holes	Total Length (in feet)
Banner (Anaconda)	1950 - 1963	3	4,300
Anaconda	1963 - 1971	113	136,838
Anamax	1973 -1983	52	54,350
Asarco	1988 - 1992	11	14,695
Augusta	2005 - 2012	87	132,483
Hudbay	2014	44	93,122
Hudbay	2015	46	75,164
Summary		356	510,951

The drill hole database was provided in Microsoft Excel® format with a cut-off date for mineral resource estimate purposes of January 19^th, 2016. The files were imported as collar, downhole survey and assay data into MineSight® Version 11.00 -3.

The topographic surface is based on two LIDAR surveys performed in 2006 (10-feet contours) and 2008 (2-feet contours) by Cooper Aerial. Drill hole collars were compared to the topographic surface and only minor differences (98% of < 5 feet) in elevation between drill hole collars and the surveyed topography were found and corrected.

14.2        Wireframe Models and Mineralization

The Rosemont deposit trends approximately along an azimuth of N020° with a general dip of 50° to the east. The Backbone Fault forms the footwall contact along the entire length of the Rosemont deposit. Geologically, Rosemont is a skarn deposit. Higher grade mineralization correlates with the Horquilla, Earp (upper and lower) and, Epitaph lithologies and also with the intensity of skarn alteration. The Rosemont deposit is continuous along a strike length of 4,000 feet (1,200 m) in north-south direction, 3,000 feet (900 m) in an east-west direction and to a vertical depth of approximately 2,500 feet (750 m).

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Three sets of structures were recognized, a north-northeast trending set, an east-west trending set and a gently east dipping set. The structures locally offset mineralization but some also appear to control mineralization, especially the oxidation.

Interpretations of the lithology, oxidation state, fault structures and ore types were built with Leapfrog® version 3.01 using the structural information, core angles and geochemical proxies developed by Hudbay. Figure 14-1 presents the Rosemont lithologies. Refer to Table 14-2 for the lithology legend. The details regarding geochemical proxy models developed by Hudbay to model the different lithologies and ore types can be found in section seven of this Technical Report.

Oxidation levels of the deposit have been modelled using the acid soluble copper to total copper ratio, where a ratio of ≥ 50% is defined as oxide, ≥30 and <50% as mixed and <30% as sulfides as shown in Figure 14-2.

As mentioned in section seven of this Technical Report, six different ore types, based on their level of oxidation, swelling and magnesium clays content, are found at Rosemont. Figure 14-3 presents the Rosemont ore types.

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FIGURE 14-1: 3D VIEW OF INTERPRETED LITHOLOGY WIREFRAMES, LOOKING NORTHWEST

Note: Lithology colour legend is approximately the same as that shown in Table 14-2. Resource pit is represented by the dashed black line. The resource pit is not indicative of the mine plan.

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TABLE 14-2: LEGEND OF INTERPRETED WIREFRAMES

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FIGURE 14-2: 3D VIEW OF INTERPRETED OXIDATION WIREFRAMES, LOOKING NORTHWEST

Note: Oxide in blue, Mixed in orange and Sulfides and un-altered in green. Resource pit is represented by the dashed black line. The resourc e pit is not indicative of the mine plan .

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FIGURE 14-3: EW CROSS SECTION OF THE ORE TYPES WIREFRAMES

Note: Resource pit is represented by the black line. The resource pit is not indicative of the mine plan.

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14.3        Exploratory Data Analysis

Exploratory data analysis (“EDA”) comprised basic statistical evaluation of the assays and composites for total copper, acid-soluble copper, molybdenum, silver and sample length.

14.4        Assays

The Table 14-3 presents the number of samples collected and total length anaylzed by element.

TABLE 14-3: SAMPLES AND LENGTH ANALYZED

14.4.1        Box Plots

Box plots of the basic statistics for TCu, ASCu, molybdenum (Mo) and silver (Ag), for each lithology within the sulfide portion of the deposit are displayed in Figure 14-4 to Figure 14-7.

These box plots confirm that most of the mineralization of economic interest in sulfides will occur in the four main units of the lower plate group with some high copper grade also occurring in the Andesite, Arkose and QMP units. The minor units exhibit higher skewness in the copper sample statistics. Molybdenum and silver statistics display a high skewness in all the lithological units.

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FIGURE 14-4: BOX PLOTS OF TOTAL COPPER ASSAYS IN SULFIDES

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FIGURE 14-5: BOX PLOTS OF ACID SOLUBLE COPPER IN SULFIDES

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FIGURE 14-6: BOX PLOTS OF MOLYBDENUM ASSAYS IN SULFIDES

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FIGURE 14-7: BOX PLOTS OF SILVER ASSAYS IN SULFIDES

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14.4.2        Grade Capping

Since most of the lithological units show a high skewness in the statistical distribution of the metal grade, length weighted, log-scaled probability plots and deciles analysis of the assays were used to define grade outliers for TCu, ASCu, molybdenum (Mo) and silver (Ag) within each of the separately evaluated domains. The capping thresholds are shown below in Table 14-4.

TABLE 14-4: CAPPING THRESOLDS BY LITHOLOGY

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14.4.3        Assay Statistics

Exploratory data analysis of assay statistics are summarized by uncapped and capped grades in Table 14-5, Table 14-6, Table 14-7 and Table 14-8. Since the earlier drill programs were mostly focused on copper, not all samples were analyzed for Mo and Ag and the exploratory data analysis shows fewer assays for these metals. There are a total of 9,174 samples with missing molybdenum assays and 21,578 missing silver assays. Capping was completed on the assays prior to compositing.

TABLE 14-5: ASSAY STATISTICS FOR TOTAL COPPER BY LITHOLOGY IN SULFIDES

TABLE 14-6: ASSAY STATISTICS FOR ACID SOLUBLE COPPER BY LITHOLOGY IN SULFIDES

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TABLE 14-7: ASSAY STATISTICS FOR MOLYBDENUM BY LITHOLOGY IN SULFIDES

TABLE 14-8: ASSAY STATISTICS FOR SILVER BY LITHOLOGY IN SULFIDES

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14.4.4        Scatter Plots, Regression Analyses, Grade Adjustments and Analysis of Gold

Exploratory data analysis of assay scatter plots were examined between TCu, ASCu, molybdenum (Mo) and silver (Ag).

14.4.4.1    Silver Adjustment

The scatter plot for silver against total copper is shown in Figure 14-8.

FIGURE 14-8: SCATTER PLOT OF CAPPED SILVER AND CAPPED COPPER, ALL LITHOLOGY DOMAINS

The scatter plot shows a relatively poor correlation (correlation coefficient of 0.58) between total copper and silver when looking at all the data. Better correlations were found when filtering by lithology and oxidation state. A reduction-to-major-axis (“RMA”) regression analyses was performed on silver against total copper for all the lithologies. Missing silver assays were assigned silver grades using the regression formula with copper grades when the correlation coefficient was above 0.7. The regression parameters used in the formula y = mx + b for assays within each lithology domain are tabulated in Table 14-9. Where the correlation coefficients were inferior to 0.7, missing silver values were replaced by zero as a conservative approach.

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TABLE 14-9: ASSAY RMA REGRESSION PARAMETERS, SILVER AGAINST COPPER

Note: Regression parameters shown above using the formula y = mx + b. Slope is given by standard deviation silver / standard deviation of total copper.

14.4.4.2    Molybdenum Adjustment

In regards to the relationship between copper and molybdenum, the scatter plot of molybdenum against copper shows a poor correlation (correlation coefficient of 0.09) as shown in Figure 14-9. Given the poor correlation, no effort was made to calculate RMA equations.

FIGURE 14-9: SCATTER PLOT OF CAPPED MOLYBDENUM AND CAPPED COPPER, ALL LITHOLOGY DOMAINS

A bias in historical molybdenum assays analysed by Wet geochemical and X-ray analytical methods has been observed. As a result of the poor correlation between copper and molybdenum, no regression could be applied to the assays. Instead, two correction factors were applied to the affected historical assays, as described in Section 11 of this Technical Report. The formulas are given below:

Mo (corrected) = Mo (Wet) x 0.85
Mo (corrected) = Mo (X-ray) x 0.45

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Figure 14-10 presents the QQ plot between the original and the corrected molybdenum grade.

FIGURE 14-10: QQ PLOT OF ORIGINAL MOLYBDEMUM GRADE VERSUS CORRECTED MOLYBDENUM GRADE

14.4.4.3    Preliminary Analysis of Gold

Out of the 356 drill holes within the Rosemont database, only 66 drill holes were analyzed for gold, which representing only 17% of all the assaying conducted on the property. Table 14-10 presents the basic statistics of the drill data by company, while Figure 14-11 to Figure 14-13 present the box plots per lithology and oxidation state. Refer to Table 14-2 for the code equivalency.

TABLE 14-10: DRILLING DATA BY COMPANY

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FIGURE 14-11: BOX PLOT OF GOLD IN OXIDE

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FIGURE 14-12: BOX PLOT OF GOLD IN MIX

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FIGURE 14-13: BOX PLOT OF GOLD IN HYPOGENE

In order to evaluate the global gold content of the Rosemont deposit, assay intervals were regularized by compositing drill hole data. The 25 ft intervals (+/- 12.5 feet of threshold) were composited using “honor geology” from the coded drill hole file. Once composited, oxidation levels were coded into the composite file.

For bias assessment purposes, assay intervals were also composited into 50 ft lengths (+/- 25 feet of threshold) using the same methodology. The 50 ft composites were used to estimate nearest neighbor models.

Down-the-hole and directional correlograms of gold were calculated using SAGE® software. However, given the limited number of pairs, the correlograms structures were found to be too erratic to produce meaningful variogram parameters. When directional correlograms were valid, the range continuity of the gold structures extended between 200 ft to 400 ft.

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The preliminary gold grade estimation was completed on the 25ft length composites using Inverse-Distance-Squared (“IDW”) grade interpolation method and three passes with increasing requirements using a composite and block matching system based on the lithology and oxidation codes.

The preliminary gold grade estimation was validated to ensure appropriate honoring of the input data. Nearest-neighbor (NN) from 50 ft composites was used to validate the IDW. Overall, no significant differences were observed.

Even though the gold grade estimation is well-constrained by three-dimensional wireframes representing geologically realistic volumes of mineralization, the confidence level is considered “low” given the lack of data. Considering the fact that only a limited number of gold assay results are available, the gold mineralization for the Rosemont deposit cannot be classified under the 2014 CIM Definition Standards for Mineral Resources.

Nevertheless, the basic statistics of gold in the drill hole database and the interpolated blocks are showing similitude with the amount of gold measured from the head grade of the metallurgical tests (i.e. 0.03 to 0.05 g/ton or 0.0008 to 0.001 once per ton). Additional work should be undertaken to gain a better understanding of the gold distribution within the Rosemont deposit, both for the grade content and its spatial distribution.

14.4.5        Contact Profiles

Exploratory data analysis (“EDA”) of contact plots displaying average grades of Cu, ASCu, Mo and Ag by distance classes on either side of the contact between each lithology domain were created. The contact profiles show that there are sharp (hard), gradual (firm) and no (soft) changes in metal grade across the contacts. An example is shown in Figure 14-14 for the Earp lithologies. A matrix of boundary conditions for sulfide material is shown in Table 14-11, Table 14-12, Table 14-13 and Table 14-14.

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FIGURE 14-14: CONTACT PROFILE, UPPER AND LOWER EARP

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TABLE 14-11: MATRIX OF BOUNDARY CONDITIONS, TOTAL COPPER

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TABLE 14-12: MATRIX OF BOUNDARY CONDITIONS, ACID SOLUBLE COPPER

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TABLE 14-13: MATRIX OF BOUNDARY CONDITIONS, MOLYBDENUM

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TABLE 14-14: MATRIX OF BOUNDARY CONDITIONS, SILVER

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14.5        Composites

In order to normalize the weight of influence for each sample, assay intervals were regularized by compositing drill hole data into 25-feet lengths using lithology boundaries to break composites. The 25-foot intervals (+/- 12.5 feet of threshold) were composited using “honor geology” from the coded drill hole file. For bias assessment purposes, assay intervals were also composited into 50-foot lengths (+/- 25 feet of threshold) using the same methodology. The 50-foot composites were used to estimate nearest neighbor models. Exploratory data analysis of the 25-foot composite statistics for TCu, ASCu, molybdenum (Mo) and silver (Ag) are shown in Table 14-15, Table 14-16, Table 14-17 and Table 14-18.

TABLE 14-15: LENGTH WEIGHTED UNCAPPED AND CAPPED 25-FOOT COMPOSITE STATISTICS, COPPER IN SULFIDES

TABLE 14-16: LENGTH WEIGHTED UNCAPPED AND CAPPED 25-FOOT COMPOSITE STATISTICS, ACID SOLUBLE COPPER IN SULFIDES

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TABLE 14-17: LENGTH WEIGHTED UNCAPPED AND CAPPED 25-FOOT COMPOSITE STATISTICS, MOLYBDENUM IN SULFIDES

TABLE 14-18: LENGTH WEIGHTED UNCAPPED AND CAPPED 25-FOOT COMPOSITE STATISTICS, SILVER IN SULFIDES

The length weighted mean grades of both 25-foot and 50-foot length composites are similar to those of the assays; therefore, providing confidence that the compositing process is working as intended. The appreciable amounts of sulfide mineralization, located within the Horquilla, Earp (lower and upper), and Epitaph lithologies, consist of low to moderate CV values for all metal types. These CV values suggest that no further domaining is warranted and that a linear interpolation method can be used. Linear interpolation was also used for the other lithological units given their minor contribution to the mineralization of economic interest. Applying non-linear interpolation methods and/or revisions of the wireframing criteria will be further investigated for these lithologies in future updates of the resource model.

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Histogram and basic statistics for capped total copper within the Horquilla lithology unit are shown in Figure 14-15.

FIGURE 14-15: HISTOGRAM, 25-FOOT COPPER COMPOSITES, HORQUILLA LITHOLOGY

14.6        Variography

Down-hole and directional correlograms for total copper, acid-soluble copper, molybdenum and silver using three combined groups of lithologies were created using SAGE® software. The Footwall Group of lithologies lies to the west of the Backbone Fault and includes the Granodiorite, Bolsa, Abrigo, Martin and Escabrosa. The Lower Plate group of lithologies lies on the hanging wall of the Backbone Fault and includes Horquilla, Earp (lower and upper) and Epitaph. The Upper Plate group of lithologies lies above the Lower Plate group and includes Scherrer, Glance, Gila, Arkose, Andesite and the QMPs. Due to a limited number of pairs in the oxide and mixed zones, the analysis was conducted on oxidation state only rather than lithology.

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The total copper variograms show very low to moderate nugget effects with values between 2% and 52% of the total variance for the sulfide mineralization. The ranges of correlation generally vary between 340 and 2,000 feet (103 and 609 meters). The downhole variogram for the lower plate group of lithologies is shown in Figure 14-16.

FIGURE 14-16: DOWNHOLE VARIOGRAM COPPER, LOWER GROUP OF LITHOLOGIES IN SULFIDES

A nugget and a nested exponential model were fitted to the experimental correlograms. An example of a variogram showing the anisotropy of the fitted model, together with the three principal directions is shown in Figure 14-17 and Figure 14-18. Correlogram model parameters for TCu, ASCu, molybdenum (Mo) and silver (Ag) are shown in Table 14-19.

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FIGURE 14-17: CORRELOGRAM OF THE MAIN STRUCTURE OF COPPER, LOWER GROUP OF LITHOLOGIES IN SULFIDES

FIGURE 14-18: CORRELOGRAM OF THE NESTED STRUCTURE OF COPPER, LOWER GROUP OF LITHOLOGIES IN SULFIDES

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TABLE 14-19: VARIOGRAM MODELS AND ROTATION ANGLES

Note: Ranges are in feet and search ellipse orientations are given using MEDS rotation convention.

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14.7        Estimation and Interpolation Methods

Lithology solids were used to code assay and composite intervals. The same solids were used to code blocks in the model based on a minimum 50% majority code threshold. Aside from the lithologies in the footwall of the backbone fault which have a limited number of composites, metal grade estimation used a composite and block matching system based on the lithology and oxidation codes. For example, in the case of the Horquilla lithology in the sulfides, only composites coded as Horquilla and sulfides were used to estimate block grades. In the case of swelling and magnesium clays, the ore types solids were used to code the assays, composites and blocks.

The block model consists of regular blocks (50 feet along strike x 50 feet across strike x 50 feet vertically). The block size was chosen such that geological contacts are reasonably well reflected and to support a large-scale open pit mining scenario.

The interpolation plan was completed on the uncapped and capped composites, 25 feet in length, using ordinary kriged (“OK”) grade interpolation method using three passes with increasing search distances.

The composite selection parameters for grade estimation in each domain (minimum, maximum, and maximum number of composites per hole) were selected to minimize bias. Table 14-20 and Table 14-21 show the search distances and search ellipse orientations for the estimation domains.

The first interpolation pass is restricted to a minimum of nine composites, a maximum of 12 composites (with a maximum of three composites per hole) and quadrant declustering. The second interpolation pass is restricted to a minimum of six composites, a maximum of 12 composites (with a maximum of three composites per hole) and quadrant declustering. Finally, the third interpolation pass is restricted to a minimum of four composites, a maximum of 12 composites (with a maximum of three composites per hole) without quadrant declustering.

Since the skarn mineralization and alteration system is driving the copper mineralization and the clays alteration at Rosemont, the copper interpolation plans were used to interpolate the magnesium and swelling clays content in every block using the ore types code matching between the composites and the blocks.

The swelling and magnesium clays were interpolated using the same multi pass system as describe above. At the end of the interpolation runs, some blocks located in small pods of ore type 2, 3, 4, 5A and 5B were left un-interpolated since they did not meet the interpolation requirements for the number of composites. In order to interpolate these isolated blocks, two additional interpolation pass had to be used. One used a minimum of two composites to interpolate the swelling and magnesium clays and the another one used a minimum of one composite.

The ordinary kriging (“OK”) interpolation results were validated against a NN model and an IDW model. The three models show similar values, hence giving confidence in the clays interpolation.

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TABLE 14-20: COPPER AND ACID SOLUBLE COPPER GRADE MODEL INTERPOLATION PLANS

Note: Ranges are in feet and search ellipse orientations are given using MEDS rotation convention.

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TABLE 14-21: MOLYBDENUM AND SILVER GRADE MODEL INTERPOLATION PLANS

Note: Ranges are in feet and search ellipse orientations are given using MEDS rotation convention.

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14.8        Tonnage Factor Assignment

There are a total of 2,486 specific gravity (SG) measurements in the drill hole database, including 2,066 Hudbay measurements with matching full geochemistry analysis. The pre-Hudbay specific gravity measurements were performed using a non-wax sealed immersion technique. The Hudbay measurements were performed by Inspectorate laboratory using a wax-sealed immersion technique to measure the weight of each sample in air and in water.

The lithologies identified at the Rosemont deposit display variable density contrast between the chemical sediments such as the limestones and dolostones, and the siliclastic sediments and crystalline rocks. In order to circumvent the relative low number of specific gravity measurements available, two linear regression models were developed based on the Hudbay specific gravity measurements and the geochemistry data. One linear regression model was fitted for the chemical sediments while the other model was adapted to the siliclastic sediments and crystalline rocks.

Table 14-22 presents the measured specific gravity measurements and the predicted specific measurements by decile and quantile.

TABLE 14-22: MEASURED COMPARED TO CALCULATED SPECIFIC GRAVITY

Missing SG measurements in the Hudbay drill holes were replaced by the calculated SG linear regression models. Table 14-23 shows the number of samples with density measurements and their basic statistics.

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TABLE 14-23: SPECIFIC GRAVITY MEASUREMENTS PER LITHOLOGY AND OXIDATION STATE

Since the skarn mineralization and alteration system was driving the copper mineralization and therefore the density, the copper interpolation plans were used to interpolate the specific density in every block using the lithology and oxidation code matching between the composites and the blocks. Figure 14-19 displays the relationship between the copper grade and the density values.

FIGURE 14-19: SCATTERPLOT OF TOTAL COPPER AND SPECIFIC GRAVITY

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The OK interpolation results were validated against a NN model and an IDW model. The three models show a similar distribution (Figure 14-20), confirming the absence of a global bias. The blocks that did not have a SG value after the interpolation were assigned an average SG value based on lithology and oxidation level (Table 14-24).

FIGURE 14-20: OK, IDW AND NN SPECIFIC GRAVITY DISTRIBUTION

Tonnage factors were calculated from the SG values using the formula TF = 2,000 / (SG * 62.42797) .

The final tonnage factors are shown below in Table 14-25. The tonnage factors have been used directly as the dry bulk tonnage factors to report the tonnage estimates of the mineral resource.

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TABLE 14-25: TONNAGE FACTORS BY LITHOLOGY AND OXIDATION STATE

 

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14.9        Block Model Validation

The Rosemont block model was validated to ensure appropriate honoring of the input data by the following methods:

• Visual inspection of the OK block model grades in plan and section views in comparison to composites grade
• Metal removed via grade capping methodology
• Comparison between the interpolation methods of NN using 50-foot composites and IDW were created to confirm the absence of global bias in the OK grade model
• Swath plot comparisons of the estimation methods to investigate local bias
• Review of block model OK quality control parameters
• Comparison of the grade tonnage curves and statistics by estimation method

14.10        Visual Inspection

Visual inspection of block grade versus composited data was conducted in section and plan view. The visual inspection of block grade versus composited data showed a good reproduction of the data by the model. An east-west oriented cross-section is provided in Figure 14-21 to Figure 14-24.

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FIGURE 14-21: VERTICAL E-W SECTION 11,554,900 SHOWING OK MODEL AND COMPOSITES - COPPER GRADE

Note: The resource pit is not indicative of the mine plan

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FIGURE 14-22: VERTICAL E-W SECTION 11,554,900 SHOWING OK MODEL AND COMPOSITES – ACID SOLUBLE COPPER GRADE

Note: The resource pit is not indicative of the mine plan.

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FIGURE 14-23: VERTICAL E-W SECTION 11,554,900 SHOWING OK MODEL AND COMPOSITES - MOLYBDENUM GRADE

Note: The resource pit is not indicative of the mine plan.

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FIGURE 14-24: VERTICAL E-W SECTION 11,554,900 SHOWING OK MODEL AND COMPOSITES – SILVER GRADE

Note: The resource pit is not indicative of the mine plan.

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14.11        Metal Removed by Capping

The impact of capping was evaluated by estimating uncapped and capped grade models. Generally, the amounts of metal removed by capping in the models are consistent with the difference of the capped and uncapped assays. The percentages of metal removed by capping from the assays, NN, IDW and OK models in the blocks above $5.7/ton NSR contained within the resource pit are shown in Table 14-26 to Table 14-29. The amount of capping appears appropriate within the Horquilla, Earp (upper and lower) and Epitaph with a difference of approximately 0 to 3% for copper, 0 to 13% for ASCu, 4 to 22% for molybdenum and 1 to 8% for silver.

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TABLE 14-26: ASSAY, NN, IDW AND OK MODEL, COPPER REMOVED BY CAPPING IN BLOCKS WITHIN THE RESOURCE PIT AND ABOVE $5.7/TON NSR

TABLE 14-27: ASSAY, NN, IDW AND OK MODEL, ACID SOLUBLE COPPER REMOVED BY CAPPING IN BLOCKS WITHIN THE RESOURCE PIT AND ABOVE $5.7/TON NSR

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TABLE 14-28: ASSAY, NN, IDW AND OK MODEL, MOLYBDENUM REMOVED BY CAPPING IN BLOCKS WITHIN THE RESOURCE PIT AND ABOVE $5.7/TON NSR

TABLE 14-29: ASSAY, NN, IDW AND OK MODEL, SILVER REMOVED BY CAPPING IN BLOCKS WITHIN THE RESOURCE PIT AND ABOVE $5.7/TON NSR

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14.12        Global Bias Checks

A comparison between the interpolation methods estimates was completed on all the blocks within the resource pit shell that have NSR values greater than $5.7/ton for global bias in the grade estimates. Differences between the NN, IDW and OK grades are acceptable in Horquilla, Earp (lower and upper) and Epitaph, with differences within 0% to 3% for copper, 0% to 3% for ASCu, 0% to 5% for molybdenum and 0% to 6% for silver. The differences are summarized in Table 14-30 to Table 14-33.

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TABLE 14-30: NN, IDW AND OK MODEL STATISTICS MEAN BLOCK GRADE COMPARISONS FOR COPPER IN BLOCKS WITHIN THE RESOURCE PIT AND ABOVE $5.7/TON NSR

TABLE 14-31: NN, IDW AND OK MODEL STATISTICS MEAN BLOCK GRADE COMPARISONS FOR ACID SOLUBLE COPPER IN BLOCKS WITHIN THE RESOURCE PIT AND ABOVE $5.7/TON NSR

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TABLE 14-32: NN, IDW AND OK MODEL STATISTICS MEAN BLOCK GRADE COMPARISONS FOR MOLYBDENUM SILVER IN BLOCKS WITHIN THE RESOURCE PIT AND ABOVE $5.7/TON NSR

TABLE 14-33: NN, IDW AND OK MODEL STATISTICS MEAN BLOCK GRADE COMPARISONS FOR SILVER IN BLOCKS WITHIN THE RESOURCE PIT AND ABOVE $5.7/TON NSR

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14.13        Local Bias Checks

A local bias check was performed by plotting the average total copper, acid soluble copper, molybdenum and silver of the NN, IDW and OK models in swath plots oriented along the model northing, easting and elevation.

In reviewing the swath plots, only minor discrepancies were found between the different grade models. In areas where there is extrapolation beyond the drill holes, the swath plots indicate less agreement for all variables. The copper, acid soluble copper, molybdenum and silver swath plots for Measured and Indicated blocks are shown below in Figure 14-25 to Figure 14-36.

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FIGURE 14-25: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, COPPER SWATH PLOT BY EASTING

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-26: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, COPPER SWATH PLOT BY NORTHING

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-27: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, COPPER SWATH PLOT BY ELEVATION

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-28: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, ACID SOLUBLE COPPER SWATH PLOT BY EASTING

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-29: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, ACID SOLUBLE COPPER SWATH PLOT BY NORTHING

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-30: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, ACID SOLUBLE COPPER SWATH PLOT BY ELEVATION

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-31: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, MOLYBDENUM SWATH PLOT BY EASTING

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-32: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, MOLYBDENUM SWATH PLOT BY NORTHING

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-33: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, MOLYBDENUM SWATH PLOT BY ELEVATION

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-34: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, SILVER SWATH PLOT BY EASTING

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-35: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, SILVER SWATH PLOT BY NORTHING

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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FIGURE 14-36: MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL, SILVER SWATH PLOT BY ELEVATION

Note: Line charts show the grades and histogram shows the tons. Green line represents IDW model. Red line represents NN model. Blue line represents OK model.

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14.14        Block Model Quality Control

The closest distance of a composite (“CDIST”), the maximum distance of a composite (“MDIST”), the average distance of composites (“ADIST”), the number of composites (“NCOMP”), the number of holes (“NHOLE”), the number of quadrants (“QUAD”) used for the OK interpolation of copper were recorded in the block model.

The standard deviation of the kriging (“KSTD”), the regression slope (“RSLOP”), the local error (“LOCA”L), the relative variance (“RELVA”) and the relative variance index (“RVI”) were also recorded in the block model. Table 14-34 presents the quality control parameters recorded in the block model from the OK resource estimation.

TABLE 14-34: QUALITY CONTROL STATISTICS OF THE COPPER INTERPOLATION IN MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL

Overall, in the Horquilla, Earp (lower and upper) and Epitaph units of the sulfides, the average distance of the composites used to interpolate the grades in the measured and indicated blocks ranges from 298 to 319-foot, with an average of 10 composites used from more than 3 drill holes. The average standard deviation of the kriging for the bulk of the mineralization ranges from 0.67 to 0.69, indicating a high variability of the copper mineralization which is to be expected in a skarn deposit.

14.15        Grade-Tonnage Statistics

The 50 ft x 50 ft x 50 ft block size is considered suitable for a large scale open pit mining operation with production rates between 75,000 to 100,000 tons per day. Table 14-35 presents the grade-tonnage statistics of copper for each interpolation method at different cut-offs.

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TABLE 14-35: GRADE-TONNAGE STATISTICS, COPPER

The grade-tonnage curve for total copper is shown in Figure 14-37 as a way to present the overall assessment of the Measured and Indicated resources.

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FIGURE 14-37: NN, IDW AND OK COPPER GRADE -TONNAGE CURVES, ALL LITHOLOGIES IN MEASURED AND INDICATED BLOCKS ABOVE $5.7/TON NSR WITHIN THE RESOURCE PIT SHELL

Note: Solid lines represent tons, dashed lines represent grades, green represents IDW model, red represents NN model and blue represents OK model

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14.16        Classification of Mineral Resource

The resource category classification used for Rosemont relies on the relative difference between the kriged grade and the composites grades, and the Resource Classification Index (RCI) which uses the following formula³:

C is a calibration factor based on the distance of the composites, the number of composites, number of quadrants and number of drill holes using the following formula:

The RCI values corresponding to the 50^th (0.216) and 95^th (0.971) percentiles of the distribution of blocks with total copper grade above 0.1% contained within the resource pit were determined and used as thresholds for the Measured and Indicated resource categories, respectively.

Under this classification system, in order for a block to be considered as a Measured resource, the RCI value must be less than 0.216, the relative difference less than 0.15 and have a CDIST of less than 500 feet. In order for a block to be considered as an indicated resource, the RCI value must be less than 0.971, the relative difference less than 0.15 and have a CDIST of less than 500 feet. Blocks were classified as inferred resources when at least two drill holes were used to interpolate the grades within one of the three interpolation passes.

A smoothing algorithm was applied to remove isolated blocks of Measured within areas of mostly indicated category or isolated indicated blocks within areas of mostly Measured category blocks. Proportions of Measured and Indicated category blocks were not changed significantly by this process. Figure 14-38 presents the resource categories for a typical cross section.

^{______________________________________________________
3} Arik, A. 2002,” Resource Classification Index”, MineSight in the Foreground.

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FIGURE 14-38: VERTICAL E-W SECTION 11,554,600 SHOWING RESOURCE CLASSIFICATION AND DRILL HOLES

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14.17        Third Party Review

Hudbay requested Tim Maunula & Associates Consulting Inc. to perform an independent validation of the block model. The following minor issues were highlighted by the third party validation:

• Interpolation plan of the NN model was different than the IDW and OK models
• A small portion of blocks in the model estimated IDW values, but no estimated OK values and vice versa

These issues were corrected and the content of this Technical Report, including the tons and grades estimate, reflects these changes.

Based on the review, the third party has concluded that the mineral resources for the Project have been prepared using industry standard best practices and the methodology for the resource classification conforms to the requirements of the CIM Definition Standards.

14.18        Internal Peer Review

Hudbay Peru Technical Service group performed a full validation of the block model. The following minor issues were highlighted by the peer review:

• The Escabrosa formation should be interpolated with the lower plate units
• A low grade shell should be use to separate the high grade and low grade zones in Earp formation
• Resource classification assignation should include a distance component. Measured or Indicated blocks with CDIST > 500 feet should be reclassified as Inferred. Inferred blocks with ADIST < 300 feet should be reclassified as Indicated.

These recommendations were implemented and the content of this report, including the tons and grade estimate, reflects these changes.

14.19        Reasonable Prospects of Economic Extraction

The component of the mineralization within the block model that meets the requirements for reasonable prospects of economic extraction was based on the application of an LG cone pit algorithm. The mineral resources are therefore contained within a computer generated open pit geometry.

The following assumptions were applied to the determination of the mineral resources:

• Economic benefit was applied to Measured, Indicated and Inferred classified material within the resource cone.

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• No effort was made to establish a pit with maximum return on investment; consequently, the mineral resource cone was the direct result of the following metal prices: $3.15/lb copper, $11.00/lb molybdenum, $18.00/oz silver with a revenue ratio of 1.0, i.e. break- even logic.
• A constant 45-degree pit slope was used for the resource estimate.
• No haulage increment or bench discounting was applied to the resource estimate.

The LG cone input parameters are summarized in Table 14-36.

TABLE 14-36: LERCHS-GROSSMAN CONE INPUTS

			Unit	Value
Mining	Mineralized Material		$/ton mined	$1.00
	Waste Material		$/ton mined	$1.00
Processing	Oxide		$/ton processed	$4.00
	Mixed and Sulfide		$/ton processed	$5.00
G&A			$/ton processed	$0.70
Recovery	Copper	Oxide	%	65
		Mixed	%	40
		Sulfide	%	85
	Molybdenum	Oxide	%	0
		Mixed	%	30
		Sulfide	%	60
	Silver	Oxide	%	0
		Mixed	%	40
		Sulfide	%	75
Metal Price	Copper		$/lb	$3.15
	Molybdenum		$/lb	$11.00
	Silver		$/troy oz	$18.00
Slope Angle			Constant degrees	45

Note: The recoveries in oxide present the recoverable portion of the sulfides by the process plan flotation. The cost and price inputs are considered approximation and were used to test the economic viability of the resource. The cost and price inputs differ from the ones used for the reserve, which used more accurate numbers.

The reporting of the mineral resource by NSR within the LG pit shell, reflects the combined benefit of producing copper, molybdenum and silver as per the following equations based on mineralized type, in addition to mine operating and processing costs:

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The copper equivalency is calculated, using metal contributions and economic inputs as noted above, for each block using the following formula:

In oxide material, since molybdenum and silver are not considered, the copper equivalency value equals the copper value.

Table 14-37 presents the economic parameters used in addition to the LG cone inputs to calculate the NSR and CuEq formulas mentioned above.

TABLE 14-37: ECONOMIC PARAMETERS

Freight	Copper	$/ton of Concentrate	$130.00
Smelter Charges	Copper	$/ton of concentrate	$72.00
	Molybdenum	$/lb of molybdenum includes freight	$1.50
Refining Charges	Copper	$/lb of copper	$0.08
	Silver	$/troy oz of silver	$0.50
Smelting Terms	Copper	% concentrate	30
		% payable	96
	Molybdenum	% concentrate grade	50
		% payable	99
	Silver	% payable	90
Royalty		%	3

Note: The cost and price inputs are considered approximation and were used to test the economic viability of the resource. The cost and price inputs differ from the ones used for the reserve, which used more accurate numbers.

14.20        Mineral Resource Statement Inclusive of Mineral Reserve

Mineral resources for the Rosemont deposit were classified under the 2014 CIM Definition Standards for Mineral Resources and Mineral Reserves⁴ by application of a NSR that reflects the combined benefit of producing copper, molybdenum and silver in addition to mine operating, processing and off-site costs.

The mineral resources, classified as Measured, Indicated and Inferred and are inclusive of the mineral reserves. Table 14-38 summarizes the resource estimate.

Mineral resources that are not mineral reserves do not have demonstrated economic viability. Due to the uncertainty that may be associated with Inferred mineral resources it cannot be assumed that all or any part of Inferred resources will be upgraded to an Indicated or Measured resource.

^{_________________________________________________________________
4} Ontario Securities Commission web site (http://www.osc.gov.on.ca/en/15019.htm)

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TABLE 14-38: RESOURCE BY CATEGORY, MINERALIZED ZONE AND NSR CUT-OFF
^{(1)(2)(3)(4)(5)(6)(7)(8)(9)(10)}

Notes:

1.	The above mineral resources include mineral reserves.
2.	Domains were modelled in 3D to separate mineralized rock types from surrounding waste rock. The domains were based on core logging, structural and geochemical data.
3.	Raw drill hole assays were composited to 25-foot lengths broken at lithology boundaries.
4.	Capping of high grades was considered necessary and was completed for each domain on assays prior to compositing.
5.	Block grades for copper, molybdenum and silver were estimated from the composites using OK interpolation into 50 ft x 50 ft x 50 ft blocks coded by domain.
6.	Tonnage factors were interpolated by lithology and mineralized zone. Tonnage factors are based on 2,066 measurements collected by Hudbay and previous operators.
7.	Blocks were classified as Measured, Indicated or Inferred in accordance with CIM Definition Standards 2014.
8.	Mineral resources are constrained within a computer generated pit using the LG algorithm. Metal prices of $3.15/lb copper, $11.00/lb molybdenum and $18.00/troy oz silver. Metallurgical recoveries of 85% copper, 60% molybdenum and 75% silver were applied to sulfide material. Metallurgical recoveries of 40% copper, 30% molybdenum and 40% silver were applied to mixed material. A metallurgical recovery of 65% for copper was applied to oxide material. NSR was calculated for every model block and is an estimate of recovered economic value of copper, molybdenum, and silver combined. Cut-off grades were set in terms of NSR based on current estimates of process recoveries, total process and G&A operating costs of $5.70/ton for oxide, mixed and sulfide material.
9.	The oxide resource will be processed in the mill via flotation
10.	Totals may not add up correctly due to rounding.

14.21        Sensitivity of the Mineral Resource

The sensitivity of the mineral resource was assessed for changes in copper, molybdenum and silver by reporting the estimation at lower and several higher NSR cut-offs, as shown in Table 14-39, Table 14-40, and Table 14-41. The results show that the mineral resource is not highly sensitive to small increases in NSR cut-offs (a proxy for changes in metal prices), therefore concluding that the mineral resource is robust with respect to the inputs used to estimate.

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TABLE 14-39: MEASURED RESOURCE BY MINERALIZED ZONE AND MULTIPLE NSR CUTOFFS

Note: Using a $5.70 per ton baseline NSR cut-off for oxide, mixed and sulfide material.

TABLE 14-40: INDICATED RESOURCE BY MINERALIZED ZONE AND MULTIPLE NSR CUTOFFS

Note: Using a $5.70 per ton baseline NSR cut-off for oxide, mixed and sulfide material.

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TABLE 14-41: INFERRED RESOURCE BY MINERALIZED ZONE AND MULTIPLE NSR CUTOFFS

Note: Using a $5.70 per ton baseline NSR cut-off for oxide, mixed and sulfide material.

14.22        Comparison with the 2012 Resource Estimates

A review and comparison of 2017 Hudbay mineral resource and 2012 Augusta mineral resource was completed. The results (Table 14-42) of measured and indicated resources show that Hudbay reports a tonnage 29% higher, with copper grades 8% lower to those estimated in 2012. Molybdenum and silver grades are 17% and 4% lower than those reported in 2012. The 2017 oxide tonnage shows a difference of +137% with +106% copper grades.

TABLE 14-42: MEASURED AND INDICATED, COMPARISON TO 2012 AUGUSTA ESTIMATE

The difference in sulfide measured and indicated tonnage is partially a result of reinterpretation of the oxide blanket surface, as discussed in Section 7 of this report. Molybdenum grades are lower as a result of factoring of historical molybdenum assays, as discussed in Section 11 of this report. Silver grades are lower as a result of using regression against copper to assign values to samples with missing silver assays.

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The significantly higher oxide tonnage and grade compared to the 2012 estimate was a result of lowering of the oxide blanket surface. The reduction of tons in the Inferred category is a direct result of the infill drilling completed by Hudbay in 2014 and 2015. The comparison is shown in Table 14-43.

TABLE 14-43: INFERRED, COMPARISON TO 2012 AUGUSTA ESTIMATE

14.23        Factors That May Affect the Mineral Resource Estimate

Areas of uncertainty that may materially impact the mineral resource estimate includes:

• Long-term commodity price assumptions
• Operating cost assumptions
• Metal recovery assumptions used and changes to the metallurgical recovery assumptions as a result of new metallurgical testwork
• Changes to the tonnage and grade estimates may vary as a result of more drilling, new assay and tonnage factor information
• Assumptions as to the ability to maintain patented mining claims and surface rights, access to the site, obtain environmental and other regulatory permits and obtain social license to operate

14.24        Conclusions

The mineral resource estimation is well-constrained by three-dimensional wireframes representing geologically realistic volumes of mineralization. Exploratory data analysis conducted on assays and composites shows that the wireframes are suitable domains for mineral resource estimation. Grade estimation has been performed using an interpolation plan designed to minimize bias in the average grade.

As a result of validation steps conducted on the mineral resource block model the following was concluded:

• Visual inspection of block grade versus composited data shows a good reproduction of the data by the model.
• Checks for global bias in the grade estimates of the block model show differences within acceptable levels (typically less than 3%) between the NN, IDW and the OK models.
• Checks for local bias (swath plots) indicate good agreement between the NN, IDW and OK for all variables.

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• A review and comparison of the 2017 Hudbay estimate and 2012 Augusta estimate showed that the 2017 block model has greater tonnage and lower grades than previously reported. The oxide tonnage and grade are significantly higher in the 2017 model mainly due to lowering of the oxide blanket surface.

The impact of grade capping was evaluated by estimating uncapped and capped assay data. Generally, the amounts of metal removed by capping does not exceed 5%.

Mineral resources are constrained and reported using economic and technical criteria such that the mineral resource has reasonable prospects of economic extraction.

The estimated mineral resources for the Project conform to the requirements of 2014 CIM Definition Standards – for Mineral Resources and Mineral Reserves and requirements in Form 43-101F1 of NI 43-101, Standards of Disclosure for Mineral Projects.

14.25        Recommendations

The author recommends that Hudbay further investigate the cause(s) of the differences in average molybdenum grade of the historical assays. Hudbay should also evaluate the application of nonlinear interpolation or wireframing methods in the minor geological units.

It is also recommended that Hudbay further investigate change-of-support correction and alternative approach to resource classification taking into account the high production rate. This should be performed to ensure that the resource classification properly reflect the reduced risk when a large volume is mined and delivered to the mill on a quarterly and annual basis.

Finally, in order to better understand the distribution of gold with sufficient confidence, the following steps should be taken:

1.	Select drill hole intervals located in zones that will be mine as ore and sent to the mill and perform gold analysis on the pulps to test the robustness of the proxy model.
2.	Perform a variography analysis of the new dataset and re-interpolate the gold grade following the same method as described in this Technical Report.

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15        MINERAL RESERVES ESTIMATE

The Mineral Reserves estimate for the Project are based on a LOM which uses the block model described in Section 14, Mineral Resource Estimates, with economic value calculation per block (NSR in $/ton) and mining, processing, and engineering detail parameters. The mineral reserve economics are described in Section 22.

This Mineral Reserves estimate has been determined and reported in accordance with NI 43-101 and the classifications adopted by CIM Council in November 2014. NI 43-101 defines Mineral Reserves as “the economically mineable part of measured and indicated mineral resources.”

The Mineral Reserves estimate for the Project, which is presented in this report, was prepared by Hudbay (Javier Toro - Director, Technical Services) and under the supervision of Cashel Meagher.

This Technical Report includes refinements of certain aspects of the Project’s mine plan. While consistency with issued and pending environmental permits and analysis related thereto has always been a key requirement for this effort, updates to the original mine plan will be necessary. To the extent that any regulatory agency concludes that the current plan requires additional environmental analysis or modification of an existing permit, the intent will be to work with that agency to either complete the required process or to adjust the current mine plan as necessary.

15.1        Pit Optimization

Revenue created from the Project will be generated from the sale of copper and molybdenum concentrates to smelters and roasters who will further refine the product. In addition, the copper concentrate contains payable silver quantities.

Pit optimization of multi-element revenue generating deposits like Rosemont can either be performed on the grade equivalent of all the revenue generating elements expressed in terms of the predominant metal (copper in Rosemont), or on the NSR.

A copper grade equivalent optimization model is simpler to implement than a NSR model but is not able to adequately represent the many variables used in the calculation of revenues as a NSR model can. Hudbay has therefore decided to use a NSR optimization model despite its additional complexity.

LG analyses were conducted using the Rosemont deposit model (described in Section 14) to determine the ultimate pit limits and best extraction sequence for open pit mine design (six pit phases were selected). Only mineral resources classified as Measured or Indicated were considered as potential ore in the LG analyses; all inferred resources were treated as waste.

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15.1.1        Block model

The Block Model used for the Mineral Reserves estimation has the original Mineral Resources estimation described in Section 14 as a base, which has a Selective Mining Unit (“SMU”) of 50x50x50 feet.

The optimized models, which were created to simulate the actual mining practice by utilizing the SMU block sizes, were considered undiluted models.

An economic subroutine was developed to compute a NSR value for each block in the deposit model. This computer algorithm incorporates block grades, expected smelting/refining contracts (i.e., payables and deductions), metallurgical recoveries and projected market prices for each metal (Cu, Mo and Ag) to yield a net revenue value expressed in terms of US Dollars per ton. The subroutine also applies to mining, ore processing and general/administration costs to calculate a net dollar value per block, which includes adjustments for surface topography. Concurrently, a NSR value in $ per ton is computed and stored in the block model.

15.1.2        Metallurgical Recoveries

Metal recoveries were derived from metallurgical testwork conducted by XPS. These tests included: grinding and flotation testwork. The metallurgical testwork is fully described in Section 13.

Based on results from this testwork, Table 15-1 presents the metallurgical recoveries used in the LG evaluations and subsequent mineral reserve estimation. Only the three primary metals, copper, molybdenum and silver were modelled and used in the revenue calculations. No recovery of molybdenum or silver from oxide ore was projected.

TABLE 15-1: METALLURGICAL RECOVERIES USED IN LERCHS-GROSSMAN EVALUATIONS

Metal	Oxide Ore	Sulfide Ore	Mixed Ore
Copper1	90.0 %	90.0 %	90.0 %
Molybdenum	-	63.0 %	30.0 %
Silver	-	75.5 %	38.0 %

Note: 1. Expressed as recoveries of the quantity of copper contained in sulfides.

15.1.3        Economic Parameters

Table 15-2 summarizes the economic parameters and offsite costs used in the base-case LG evaluations of the Rosemont deposit.

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TABLE 15-2: BASE-CASE LERCHS-GROSSMAN ECONOMIC PARAMETERS

Parameter	Units	Value
Revenue
Metal Price
Copper	$/lb	3.15
Molybdenum	$/lb	11.00
Silver	$/oz	18.00
Payable Contained Metal
Copper	%	96.5%
Molybdenum	%	99.0%
Silver	%	90.5%
Concentrate grades
Copper	%	30%
Moly concentrate grade
Molybdenum	%	45%
Concentrate Moisture Content
Copper concentrate	%	8.0%
Moly concentrate	%	8.0%
Smelting Charges
Smelting charges - Cu conc (dry)	$/dst Cu conc	72.57
Roasting charges - Mo conc (dry)	$/dst Mo conc	1.50
Marketing Cost	$/dst Cu conc	5.08
Selling Cost (Freight)
Transport Cu conc	$/dst conc	137.55
Refining charges
Cu	$/lb Cu	0.08
Ag	$/oz Ag	0.50
S+T+R cost	$/lb Cu	0.4517
Royalties
Royalties	% of NSR	3.0%
Cost
Mining Cost
Ore	$/tmined	1.14
Waste	$/tmined	1.14
Incremental Cost by Bench
Up	$/tmined	-
Down	$/tmined	0.024
G&A Cost
Ore	$/milled	1.00
Process Cost
Sulfide	$/milled	5.00
Mixed	$/milled	5.00
Oxide	$/milled	5.00

The in-situ NSR value is first calculated and coded into each block in the model. This is to allow the pit optimization of the multi-element Rosemont deposit to be carried out on the in-situ NSR values. The following process is the procedure that was developed in order to achieve the NSR calculation:

• In-situ NSR is the net value of metals contained in a concentrate produced from an ore block after smelting and refining. Using the concentrator recovery of the metals with the concentrate and the grade of the concentrate produced, the mass pull of each block in the resource model expressed in terms of tons of concentrate per ton of ore processed is first estimated.
  
  

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• The value of the payable metals in the concentrate is then calculated based on agreed payable metal content in the concentrate subject to deductions with smelters, refineries and roasters. In the case of the copper concentrate, the payable precious metal silver is added to the value of the payable copper. For the molybdenum concentrate, only the molybdenum metal is payable.
  
  
• From the value of the payable metals, the selling costs, which include marketing costs, transportation costs, port charges, insurance costs, shipping costs, and smelting charges expressed in $/dmt concentrate and other deductions like the refining charges and price participation (if applicable) expressed in $/payable metal are taken out to obtain the gross concentrate NSR value (before royalties).
  
  
• The applicable royalties are then deducted from the gross concentrate NSR value to obtain the net concentrate NSR value (after royalty). The concentrate NSR value calculations described above are applied for both the copper and molybdenum concentrates.
  
  
• The concentrate NSR value after royalty for the copper and molybdenum concentrates are then multiplied by their respective mass pull expressed in tons of concentrate produced per tonne of ore processed to obtain the contribution of each metal in the concentrate to the in- situ NSR value.
  
  
• The in-situ NSR of each block in the normalized resource model is the sum of the in-situ NSR value from the copper concentrate and the molybdenum concentrate.

Only Measured and Indicated Resource model block categories with NSR values greater than their processing costs are considered potential ore while blocks which have NSR values less than their processing costs are considered waste.

Process plant recoveries, throughput, operating costs, and concentrate grades vary by ore type. Consistent with ore reserve reporting guidelines, only Measured and Indicated resources are coded to generate revenues in the NSR model. Inferred resources are coded and reported as waste.

Processing metal recoveries for copper and silver are fixed numbers depending on metallurgical domain while molybdenum is calculated by formula linked to molybdenum and copper feed grades, and copper recovery. Copper and silver grades in the copper concentrate are calculated by formula. The grade of molybdenum in the molybdenum concentrate is a fixed number.

15.1.4        NSR Input Parameters

The revenue, recovery and cost input parameters used for pit optimization are shown in Table 15-1 and Table 15-2.

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15.1.5        Pit Slope Guidance

Overall slope angles used on the LG evaluations were derived from the geotechnical recommendations made by CNI & Hudbay for pit slope designs. The overall slopes were adjusted to accommodate the recommended slope angles and the anticipated placement of internal haulage ramps along the pit walls in certain design sectors to be used as berms (step outs). Hudbay assigned slopes angles for each block of resources model, and a slope code was assigned to the block representing each of the pit slopes. The slope codes and pit slopes are then read as input to the LG analysis. The plan view and the design parameters by sector are shown in Figure 15-1 and Table 15-3 respectively.

FIGURE 15-1: PLAN VIEW CONTOURS OF SELECTED LERCHS-GROSSMAN PIT SHELL

TABLE 15-3: OVERALL SLOPE ANGLES USED IN LERCHS-GROSSMAN ANALYSIS

Geotechnical Sector	Bench Height, feet	Bench Face Angle°	Inter-Ramp Slope Angle°	Catch Bench, feet	Overall Slope Angle°
1	100	70	50	48	42
2	100	65	46	50	40
3	100	65	48	44	45
4	100	65	48	44	45
5	50	65	46	25	43
6	50	65	44	29	41
7	50	55	33	42	31
8	50	55	33	42	31

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15.1.6        Lerchs-Grossman Analyses

All LG analyses were restricted to prevent the pit shells from crossing the topographic ridge immediately west of the deposit. This was done due to permit constraints.

The base-case LG pit shell 40 is defined by the recoveries and economic parameters listed in Table 15-1 and Table 15-2, respectively. This pit shell contains about 710 million tons of Measured and Indicated mineral resource above an internal NSR cut-off of $6.00/ton. The resulting stripping ratio is about 2.24:1 (tons waste per ton of ore). However, this is not the pit shell selected for pit design Several economic analyses were developed for each nested pit. The purpose of this assessment was to evaluate free discounted cash flow, revenue, stripping ratio, development and sustaining capital. Figure 15-2 presents the results of the LG price and price sensitivity analyses, respectively.

FIGURE 15-2: ROSEMONT WHITTLE RESULTS, REVENUE FACTOR SENSITIVITY

Pit shell 30 was generated at a 0.80 revenue factor and contains approximately 622 M st of Ore and 1,269 M st of Waste. The pit shell captures about 99.3% of the Net Cash flow of the base revenue factor (RF) 1 pit shell 40. This pit generates a lower stripping ratio, better economics greater total revenue, and capital costs than other pits evaluated for the Project.

The selected LG pit shell 30 is shown in plan view in Figure 15-3 and in cross-section view in Figure 15-4.

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FIGURE 15-3: PLAN VIEW CONTOURS OF SELECTED LERCHS-GROSSMAN PIT SHELL

FIGURE 15-4: AA’ SECTION VIEW OF SELECTED LERCHS-GROSSMAN PIT SHELL

15.1.7        Pit Design Criteria

Design criteria for final pit takes into consideration the geotechnical recommendations summarized earlier in Table 15-3, pushback (mine phase) width, phase sequence, haul roads and access, berms, ditches as well as other engineering considerations. Mine phase, or pushback, widths are typically

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320 feet. The summary parameters used in the design of the ultimate pit are presented in Table 15-4.

TABLE 15-4: PIT DESIGN PARAMETERS

Parameter	Value
Bench height	50 – 100 feet
Bench face angle	55 – 70°
Catch bench interval	25 – 50 feet
Road width (including ditch & safety berm)	110 feet
Nominal road gradient	10 %
Minimum pushback width	320 feet

15.2        Mineral Reserves

The Rosemont Mineral Reserves estimation is based on Measured and Indicated resources. Therefore, the potential exists for Inferred Mineral Resources within the ultimate pit to be included and reported as waste, as they currently do not meet the economic and mining requirements to be categorized as Mineral Reserves. It cannot be assumed that all or any part of Inferred mineral resources will ever be upgraded to a higher category.

The mining phase and ultimate pit designs were applied to the 3D resource block model of the deposit described in Section 14 to estimate contained tonnages and grades.

15.2.1        Ore Definition Parameters

The base-case price and operating cost estimates presented in Table 15-2 are used as the economic envelope to define ore in the mineral reserve estimates.

Mineralized oxide and mixed materials that are indicated to be economic (above an internal NSR cut-off of $6.00/ton) in the optimized pit analysis are included in the pit ore reserves for this study.

15.2.2        Material Densities

Bulk material densities, which vary by rock type, were read from values stored in the resource block model. These assignments are described in more detail in Section 14. Generally, rock tonnage factors range between 11.7 ft³/ton and 12.4 ft³/ton, with an average of 12.10 ft³/ton for the rock contained within the ultimate pit.

15.2.3        Dilution

The Rosemont deposit is a well-disseminated polymetallic deposit that has large ore zones above the anticipated internal cut-off grade. With the planned bulk mining method, external ore dilution along the ore - waste contact edges is generally assessed to determine whether the feed grade from the run of mine production is adequately represented by those predicted from the resource block model.

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The resource block model dimensions are 50x50x50 feet. The interpolated metal grade is averaged for the entire block. When the Project commences operations, ore feed will be delineated by implementing a detailed blasthole sampling program. Drill blast patterns will be smaller, 30 feet to 30 feet, than the resource block dimensions, thereby providing a better definition than from the resource model. This new definition will be provided by a new block model built by assays from blastholes projects, dynamic or short range block model, which is a common practice in Hudbay operations.

The author has confirmed that enough geological dilution is already incorporated in the resource model due to the smoothing effect of kriging. Based on experience in similar types of skarn deposits and scale of operation, it is reasonable to use the resource tonnes and grade from the individual 50 ft x 50 ft x 50 ft blocks from the resource model without any additional adjustment for ore losses or mining dilution.

15.2.4        Mineral Resource and Mineral ReserveStatement

Proven and probable mineral reserves for the Rosemont deposit are summarized in Table 15-5. Proven and probable mineral reserves within the designed final pit total 592 million tons grading 0.45% Cu, 0.012% Mo and 0.13 oz Ag/ton. There are 1.25 billion tons of waste material, resulting in a stripping ratio of 2.1:1 (tons waste per ton of ore). Total material in the pit is 1.84 billion tons. Contained metal in proven and probable mineral reserves is estimated at 5.30 billion pounds of copper, 142 million pounds of molybdenum and 79 million ounces of silver.

Nearly 80% of the mineral reserves in the Rosemont ultimate pit are classified as proven with the remaining 20% identified as probable. The classifications are based on the exploration drilling in the Rosemont deposit. All of the mineral reserves estimate reported are contained in the mineral resource estimates presented in Section 14.

The Rosemont ultimate pit contains approximately 10 million tons of inferred mineral resources that are above the $6.00/ton NSR cut-off value for ore. Inferred mineral resources are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves.

The mineral reserves estimate presented in this report is dependent on market prices for the contained metals, metallurgical recoveries and ore processing, mining and general/administration cost estimates. Mineral reserve estimates in subsequent evaluations of the Rosemont deposit may vary according to changes in these factors. As of the effective date of this report, there are no other known mining, metallurgical, infrastructure or other relevant factors that may materially affect the mineral reserve estimates.

Proven and Probable mineral reserves for the Rosemont deposit are summarized in Table 15-5 and classified by ore type in Table 15-6. Illustrations of the ultimate pit in plan and section view against economic shell 30 are shown in Figure 15-5 and Figure 15-6.

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TABLE 15-5: PROVEN AND PROBABLE MINERAL RESERVES IN ROSEMONT FINAL PIT

	Short Tons	TCu1 %	SCu2 %	ASCu3 %	Mo %	Ag opt	NSR $/t	CuEq4 %
Proven	469,708,117	0.48	0.43	0.05	0.012	0.14	22.1	0.56
Probable	122,324,813	0.31	0.28	0.03	0.010	0.09	14.7	0.38
Total	592,032,930	0.45	0.40	0.05	0.012	0.13	20.57	0.53

Notes:

1.	TCu % corresponds to the total copper grade.
2.	SCu % grade corresponds to the sulfide copper in the Ore. As per formula SCU = TCU – ASCu
3.	ASCu % grade corresponds to the soluble copper.
4.	CuEq% is calculated based on metal prices of $3.15/lb Cu, $11.00/lb Mo and $18.00/oz Ag.

TABLE 15-6: PROVEN AND PROBABLE MINERAL RESERVES IN ROSEMONT FINAL PIT BY ORE TYPE

Ore Type	Short Tons	TCu1 %	SCu2 %	ASCu3 %	Mo %	Ag opt	NSR $/t	CuEq4 %
Sulfide	542,969,276	0.45	0.41	0.04	0.013	0.14	21.5	0.53
Proven	431,620,325	0.49	0.45	0.04	0.013	0.15	23.1	0.57
Probable	111,348,950	0.31	0.29	0.03	0.011	0.09	15.1	0.38
Mixed	26,477,889	0.33	0.25	0.08	0.007	0.07	11.8	0.37
Proven	18,743,016	0.34	0.25	0.09	0.008	0.08	12.0	0.39
Probable	7,734,873	0.30	0.24	0.06	0.007	0.05	11.3	0.34
Oxide5	22,585,766	0.50	0.24	0.26	-	-	9.8	0.50
Proven	19,344,776	0.52	0.24	0.28	-	-	9.9	0.52
Probable	3,240,990	0.38	0.22	0.17	-	-	8.9	0.38
Total	592,032,930	0.45	0.40	0.05	0.012	0.13	20.6	0.53

Notes:

1.	TCu % corresponds to the total copper grade.
2.	SCu % grade corresponds to the sulfide copper in the Ore. As per formula SCU = TCU – ASCu
3.	ASCu % grade corresponds to the soluble copper.
4.	CuEq% is calculated based on metal prices of $3.15/lb Cu, $11.00/lb Mo and $18.00/oz Ag.
5.	Oxide ore refers only to the sulfide copper species.

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FIGURE 15-5: PLAN VIEW OF ROSEMONT FINAL PIT AND ECONOMIC SHELL 30

FIGURE 15-6: SECTION VIEW BB’ OF ROSEMONT FINAL PIT AND ECONOMIC SHELL 30

Table 15-7 presents the mineral resource estimates exclusive of the mineral reserve estimates, i.e. the mineral resources located inside the resource pit shell and outside of the pit design. It represents

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the portion of the mineral resources estimate with potential for economic extraction after the current mineral reserves estimate has been mined and processed.

TABLE 15-7: ROSEMONT MINERAL EXCLUSIVE RESOURCE ESTIMATES

Measured	TONS	NSR Cut Off	CuEq (%)	Cu (%)	Mo (%)	Ag (opt)
Oxide	54,000,000	> = $5.70	0.41	0.41
Mix	5,000,000	> = $5.70	0.45	0.41	0.008	0.047
Hypogene	118,700,000	> = $5.70	0.44	0.36	0.014	0.117
Summary	177,700,000		0.43	0.38	0.009	0.079
Indicated	TONS	NSR Cut Off	CuEq (%)	Cu (%)	Mo (%)	Ag (opt)
Oxide	18,600,000	> = $5.70	0.27	0.27
Mix	2,600,000	> = $5.70	0.36	0.34	0.005	0.037
Hypogene	392,000,000	> = $5.70	0.31	0.25	0.012	0.080
Summary	413,200,000		0.31	0.25	0.011	0.076
Measured + Indicated	TONS	NSR Cut Off	CuEq (%)	Cu (%)	Mo (%)	Ag (opt)
Oxide	72,700,000	> = $5.70	0.38	0.38
Mix	7,600,000	> = $5.70	0.42	0.38	0.007	0.044
Hypogene	510,700,000	> = $5.70	0.34	0.27	0.012	0.088
Summary	591,000,000		0.35	0.29	0.011	0.077
Inferred	TONS	NSR Cut Off	CuEq (%)	Cu (%)	Mo (%)	Ag (opt)
Oxide	3,500,000	> = $5.70	0.33	0.33
Mix	1,300,000	> = $5.70	0.47	0.45	0.004	0.019
Hypogene	63,900,000	> = $5.70	0.35	0.29	0.011	0.049
Summary	68,700,000		0.35	0.30	0.010	0.046

Notes:

1.	Domains were modelled in 3D to separate mineralized rock types from surrounding waste rock. The domains were based on core logging, structural and geochemical data.
2.	Raw drill hole assays were composited to 25-foot lengths broken at lithology boundaries.
3.	Capping of high grades was considered necessary and was completed for each domain on assays prior to compositing.
4.	Block grades for copper, molybdenum and silver were estimated from the composites using OK interpolation into 50 ft x 50 ft x 50 ft blocks coded by domain.
5.	Tonnage factors were interpolated by lithology and mineralized zone. Tonnage factors are based on 2,066 measurements collected by Hudbay and previous operators.
6.	Blocks were classified as Measured, Indicated or Inferred in accordance with CIM Definition Standards 2014.
7.	Mineral resources are constrained within a computer generated pit using the LG algorithm. Metal prices of $3.15/lb copper, $11.00/lb molybdenum and $18.00/troy oz silver. Metallurgical recoveries of 85% copper, 60% molybdenum and 75% silver were applied to sulfide material. Metallurgical recoveries of 40% copper, 30% molybdenum and 40% silver were applied to mixed material. A metallurgical recovery of 65% for copper was applied to oxide material. NSR was calculated for every model block and is an estimate of recovered economic value of copper, molybdenum, and silver combined. Cut-off grades were set in terms of NSR based on current estimates of process recoveries, total process and G&A operating costs of $5.70/ton for oxide, mixed and sulfide material.
8.	The oxide resource will be processed in the mill via flotation.
9.	Totals may not add up correctly due to rounding.

15.2.5        Factors That May Affect the Mineral Reserves Estimate

Areas of uncertainty that may materially impact the mineral resource estimate includes:

• Long-term commodity price assumptions.
• Operating cost assumptions.
• Metal recovery assumptions used and changes to the metallurgical recovery assumptions as a result of new metallurgical testwork.
• Changes to the tonnage and grade estimates may vary as a result of more drilling, new assay and tonnage factor information.

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	Form 43-101F1 Technical Report

15.2.6         Comparison with the 2012 Mineral Reserves

A review and comparison of 2017 Hudbay mineral reserves and 2012 Augusta mineral reserves was completed. The results (Table 15-8) of proven and probable reserves show that Hudbay reports a tonnage 11% lower, with copper grades 2% higher. Molybdenum and silver grades are 17% lower and 11% higher, respectively, to those estimated in 2012.

TABLE 15-8: PROVEN AND PROBABLE, COMPARISON TO 2012 AUGUSTA RESERVE ESTIMATE

Category	Hudbay Reserves 2017				Augusta Reserves 2012 Model
	Tons	Cu (%)	Mo (%)	Ag (opt)	Tons	TCu (%)	Mo (%)	Ag (opt)
Proven	469,708,117	0.48	0.012	0.14	308,075,000	0.46	0.015	0.12
Probable	122,324,813	0.31	0.010	0.09	359,131,000	0.42	0.014	0.12
TOTAL	592,032,930	0.45	0.012	0.13	667,206,000	0.44	0.014	0.12

The changes between 2012 and 2017 Reserves estimate can be mostly attributed to a revision of the mining, processing and general & administration cost assumptions, resulting in a marginally higher cut-off in 2017.

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16        MINING METHODS

16.1        Mine Overview

The Rosemont deposit is a large tonnage, skarn-hosted, porphyry-intruded, copper-molybdenum deposit located in close proximity to the surface. The Project will be a traditional open pit shovel/truck operation. The Project consists of open pit mining and flotation of sulfide minerals to produce commercial grade concentrates of copper and molybdenum. Payable silver will report to the copper concentrate.

The proposed pit operations will be conducted from 50-foot-high benches using large-scale mine equipment, including: 10-5/8-inch-diameter rotary blast hole drills, 60 yd³ class electric mining shovels, 46 yd³ class hydraulic shovel, 25 yd³ front-end loader, and 260-ton capacity off-highway haul trucks.

The Rosemont final pit will measure approximately 6,000 feet east to west, 6,000 feet north to south, and have a total depth of approximately 2,900 feet down to 3,100 feet (AMSL). There is one primary WRSA, which is located 1,200 feet south east of the Rosemont final pit. The processing facility is located approximately 1,000 feet east of the final pit, while the dry stack tailings facility (“DSTF”) is located 1,500 feet southeast of the Rosemont pit. The final pit and facilities can be seen in Figure 16-1.

The mine production plan contains 592 million tons of ore and approximately 1.25 billion tons of waste, yielding a life of mine waste to ore stripping ratio of 2.1 to 1. The mine has a 19-year life (including pre-stripping period but excluding initial haul road development), with ore to be delivered to the processing plant at a throughput of 90,000 tpd. Mine operations are scheduled for 24 hours per day, 365 days per year. A mining rate of 132 million tons per year through year 11 will be required to provide the assumed nominal process feed rate of 32.9 million tons of ore per year. From year 12 through year 18, the annual mining rate decreases due to lower stripping ratios, starting with an average of 50 million tons per year and ending with approximately 33 million tons in production year 18. Ore shortfall will be made up from stockpiled ore.

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FIGURE 16-1: ROSEMONT MINE PLAN SITE LAYOUT

16.2        Mine Phases

16.2.1        Design Criteria

Mine phases and ultimate pit for the Project are designed for large-scale mining equipment (specifically, 60 yd³ class electric shovels and 260-ton haulage trucks) and are derived from the selected LG pit shells described in the previous section. The design process included smoothing pit walls, eliminating or rounding significant noses and notches that may affect slope stability, and providing access to working faces by developing internal ramps (dual ramp for final pit). The summarized parameters used in the design of mine pit phases are presented in Table 16-1.

TABLE 16-1: PIT DESIGN PARAMETERS

Parameter	Value
Bench height	50 – 100 feet
Bench face angle	55 – 70°
Catch bench interval	25 – 50 feet
Road width (including ditch & safety berm)	110 feet
Nominal road gradient	10 %
Minimum pushback width	320 feet

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16.2.2        Pit Slopes Angles

For the pit design, the targeted minimum mining width is 320 feet and employs the wall slope design provided by CNI and Hudbay. Table 16-2 lists the configuration of the recommended pit slope configuration for each sector, and Figure 16-2 shows the Ultimate Pit Slope Design with the corresponding Geotechnical Sectors.

TABLE 16-2: ROSEMONT SLOPE GUIDANCE

Geotechnical Sector	Bench Height, feet	Bench Face Angle°	Inter-Ramp Slope Angle°	Catch Bench, feet	Overall Slope Angle°
1	100	70	50	48	42
2	100	65	46	50	40
3	100	65	48	44	45
4	100	65	48	44	45
5	50	65	46	25	43
6	50	65	44	29	41
7	50	55	33	42	31
8	50	55	33	42	31

FIGURE 16-2: ROSEMONT GEOTECHNICAL SECTORS

16.2.3        Mine Phases and Ultimate Pit

Six mining phases define the extraction sequence for the Rosemont deposit. The phase development strategy consists of extracting the higher metal grades along with minimum strip ratios during the initial years to maximize the economic benefits of the ore-body, while enabling smooth transitions in waste stripping throughout the life of the mine to ensure enough ore exposure for mill feed.

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Mine Phase 1

The starter pit, Phase 1, is fit approximately to the LG pit shell defined by a $1.26/lb Cu price (equivalent to 40% of base metal price sensitivity case). This pit is located about 3,500 feet west of the primary crusher and ranges in elevation from 5,800 to 4,350 feet AMSL. The phase is approximately 3,000 feet wide east-west and 4,000 feet north-south. The upper benches will be dozed down until haul road access can be developed to the 5,700 feet elevation (AMSL). Phase 1 will develop approximately 85 million tons of ore at a stripping ratio of 2.2:1 (tons waste per ton of total ore). An illustration of the Phase 1 pit is shown in Figure 16-3.

Phase 1 material will be accessed via a haul road, 2C, which will be constructed from the pit exit eastward to the primary crusher. This road will also branch off towards the WRSA. These roads will be used for the life of the Project, and will also be extended to access the DSTF.

The pit entrance is at the 5,150 feet elevation (AMSL), and a ramp from that location enters the pit in a counter clockwise direction. The ramp switches back at the 4,950, and 4,650 feet elevations (AMSL) before reversing to a counter clockwise direction to the bottom of the pit. All benches are accessed by a double lane width haul road.

FIGURE 16-3: PLAN VIEW OF MINING PIT PHASE 1

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Mine Phase 2

Mining Phase 2 will expand the pit roughly 600 feet to the north, 400 feet to the east and 500 feet to the southeast. Bench toe elevations will range from 5,450 to 4,050 feet (AMSL). The phase is 2,500 feet wide east-west and 4,000 feet north-south. Phase 2 will supply over 88 million tons of ore. The average stripping ratio for this pushback is 1.3:1. An illustration of the Phase 2 pit is shown in Figure 16-4.

The pit entrance is at the 5,150 feet elevation (AMSL), and a ramp from that location enters the pit in a counter clockwise direction. The ramp switches back at the 4,950, and 4,650 feet elevations (AMSL).

FIGURE 16-4: PLAN VIEW OF MINING PIT PHASE 2

Mine Phase 3

The open pit is further expanded 500 to 600 feet to the east with the development of Phase 3. The easternmost limits of this pushback lie about 2,500 feet west of the primary crusher. Benches will range between 5,500 and 3,750 feet toe elevations (AMSL). The phase is 3,400 feet wide east-west and 5,000 feet long north-south. Over 75 million tons of ore will be generated by Phase 03 at an average stripping ratio of 2.4:1.

The pit entrance is at the 5,150 feet elevation (AMSL), and a ramp from that location enters the pit in a counter clockwise direction. The ramp switches back at the 4,950 feet elevation (AMSL) to the bottom of the pit. An illustration of the Phase 3 pit is shown in Figure 16-5.

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FIGURE 16-5: PLAN VIEW OF MINING PIT PHASE 3

Mine Phase 4

Phase 4 will expand the open pit about 600 feet to the east and 400 feet to the north. The easternmost limits of this pushback lie about 2,000 feet west of the primary crusher. Phase 4 benches range in elevation between 5,300 and 3,650 feet AMSL. The phase is 2,500 feet wide east-west and 5,500 feet north-south. Phase 4 will produce nearly 64 million tons of ore at a stripping ratio of 2.9:1. Phases 2, 3 and 4 fit approximately to the LG pit shell defined by a $1.30/lb Cu price (equivalent to 41% of base case metal price sensitivity). This expansion from the Phase 1 pit is split into 3 separate pushbacks, all in the same general direction. For each phase expansion, the ramp on the east side of the pit is re-developed. An illustration of the Phase 4 pit is shown in Figure 16-6.

The pit entrance is at the 5,100 feet elevation (AMSL), and a ramp from that location enters the pit in a counter clockwise direction. The ramp switches back at the 4,850, 4,450, and 4,200 feet elevations (AMSL) before reversing to a clockwise direction to the bottom of the pit.

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FIGURE 16-6: PLAN VIEW OF MINING PIT PHASE 4

Mine Phase 5

Phase 5 is fit approximately to the LG pit shell defined by a $1.50/lb Cu price (equivalent to 47% of base case metal price value sensitivity). Mining Phase 5 expands the pit approximately 300 feet to the north and 600 feet to the east. The easternmost limits of this pushback lie about 1,200 feet west of the primary crusher. Phase 5 bench elevations range between 5,300 and 3,450 feet (AMSL). The phase is 3,000 feet wide east-west and 5,000 feet north-south. Phase 5 will produce nearly 60 million tons of ore at a stripping ratio of 2.5:1. The ramp on the east side of the pit is developed for this phase. An illustration of the Phase 5 pit is shown in Figure 16-7.

The pit entrance is at the 5,050 feet elevation (AMSL), and a ramp from that location enters the pit in a counter clockwise direction. The ramp switches back at the 4,800, 4,450, 4,200 and 3,900-feet elevations (AMSL) before reversing to a counter clockwise direction to the bottom of the pit.

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FIGURE 16-7: PLAN VIEW OF MINING PIT PHASE 5

Mine Phase 6 and Ultimate Pit

The final pushback, Phase 6, extends the open pit from 300 to 600 feet along the east side to its ultimate limits and down to its maximum depth at the 3,100 feet elevation (AMSL). The ultimate pit will be about 6,000 feet wide east-west and 6,500 feet wide north-south. Phase 6 is fit approximately to the LG pit shell defined by a $2.52/lb Cu price (equivalent to 80% of base case metal price value sensitivity). Phase 6 will generate nearly 220 million tons of ore at a stripping ratio of 2.0:1. An illustration of the Phase 6 pit, or final pit, is shown in Figure 16-8.

Total ore reserves extracted from the six mining phases are estimated to be 592 million tons and will generate 1.25 billion tons of waste material. Approximately 55 million tons of medium and low grade oxide, mixed and sulfide ore will be stockpiled.

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FIGURE 16-8: PLAN VIEW OF MINING PIT PHASE 6 (ULTIMATE PIT)

Final configuration of mine phases is presented in plan view in Figure 16-9 and in cross section in Figure 16-10. Mineral reserves for the Rosemont deposit by mine phase are summarized in Table 16-3 and classified by ore type in Table 16-4.

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	Form 43-101F1 Technical Report

FIGURE 16-9: PLAN VIEW OF ROSEMONT MINE PHASES

FIGURE 16-10: AA’ SECTION VIEW OF ROSEMONT MINE PHASES

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	Form 43-101F1 Technical Report

TABLE 16-3: ROSEMONT MINE PHASES MINERAL RESERVES

	Ore M Tons	TCu %	SCu %	ASCu %	Mo %	Ag opt	NSR $/t	CuEq %	Waste M Tons	Total M Tons	S.R.
PH01	84.8	0.49	0.43	0.06	0.011	0.16	21.80	0.57	190.3	275.1	2.24
PH02	88.3	0.43	0.38	0.05	0.010	0.15	19.77	0.51	115.6	203.9	1.31
PH03	74.8	0.50	0.45	0.04	0.012	0.15	23.18	0.58	177.9	252.7	2.38
PH04	63.5	0.53	0.50	0.03	0.014	0.13	25.26	0.62	182.5	246.0	2.87
PH05	59.4	0.47	0.44	0.03	0.014	0.12	22.65	0.56	150.3	209.8	2.53
PH06	221.2	0.39	0.34	0.05	0.012	0.12	17.64	0.46	431.9	653.1	1.95
Total	592.0	0.45	0.40	0.05	0.012	0.13	20.57	0.53	1,248.6	1,840.6	2.11

Notes:

1.	TCu % corresponds to the total copper grade.
2.	SCu % grade corresponds to the sulfide copper in the Ore. As per formula SCU = TCU – ASCu
3.	ASCu % grade corresponds to the soluble copper.
4.	CuEq% is calculated based on metal prices of $3.15/lb Cu, $11.00/lb Mo and $18.00/oz Ag.

TABLE 16-4: ROSEMONT MINE PHASES, MINERAL RESERVES BY ORE TYPE

	Ore M Tons	TCu %	SCu %	ASCu %	Mo %	Ag opt	NSR $/t	CuEq %
PH01	84.8	0.49	0.43	0.06	0.011	0.16	21.80	0.57
Sulfide	74.8	0.50	0.45	0.05	0.011	0.16	23.27	0.58
Mixed	4.2	0.28	0.21	0.08	0.008	0.09	10.15	0.34
Oxide	5.8	0.56	0.28	0.28	0.006	0.13	11.33	0.56
PH02	88.3	0.43	0.38	0.05	0.010	0.15	19.77	0.51
Sulfide	78.8	0.44	0.40	0.04	0.010	0.16	20.84	0.52
Mixed	7.5	0.30	0.23	0.07	0.007	0.09	11.34	0.35
Oxide	2.0	0.43	0.22	0.20	0.004	0.13	9.23	0.43
PH03	74.8	0.50	0.45	0.04	0.012	0.15	23.18	0.58
Sulfide	70.7	0.50	0.47	0.04	0.012	0.15	23.86	0.59
Mixed	2.7	0.32	0.26	0.06	0.007	0.06	12.29	0.36
Oxide	1.4	0.41	0.25	0.16	0.005	0.08	10.18	0.41
PH04	63.5	0.53	0.50	0.03	0.014	0.13	25.26	0.62
Sulfide	62.3	0.54	0.50	0.03	0.015	0.13	25.55	0.62
Mixed	1.1	0.30	0.22	0.08	0.008	0.04	10.69	0.34
Oxide	0.1	0.36	0.21	0.15	0.003	0.10	8.62	0.36
PH05	59.4	0.47	0.44	0.03	0.014	0.12	22.65	0.56
Sulfide	58.1	0.48	0.45	0.03	0.015	0.12	22.90	0.56
Mixed	1.2	0.29	0.24	0.05	0.011	0.05	11.53	0.34
Oxide	0.1	0.37	0.27	0.10	0.002	0.09	11.02	0.37
PH06	221.2	0.39	0.34	0.05	0.012	0.12	17.64	0.46
Sulfide	198.3	0.38	0.35	0.03	0.013	0.12	18.44	0.46
Mixed	9.8	0.38	0.27	0.10	0.007	0.05	12.86	0.42
Oxide	13.2	0.49	0.22	0.27	0.004	0.09	9.10	0.49
Grand Total	592.0	0.45	0.40	0.05	0.012	0.13	20.57	0.53

Notes:

1.	TCu % corresponds to the total copper grade.
2.	SCu % grade corresponds to the sulfide copper in the Ore. As per formula SCU = TCU – ASCu
3.	ASCu % grade corresponds to the soluble copper.
4.	CuEq% is calculated based on metal prices of $3.15/lb Cu, $11.00/lb Mo and $18.00/oz Ag.
5.	Oxide ore refers only to the sulfide copper species.

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16.3        Mine Schedule and Production Plan

16.3.1        Production Scheduling Criteria

The operating and scheduling criteria used to develop the mining sequence plans are summarized in Table 16-5 below.

TABLE 16-5: MINE PRODUCTION SCHEDULE CRITERIA

Parameter	Value
Annual Ore Production Base Rate	32,850,000 tons
Daily Ore Production Base Rate	90,000 tons
Operating Hours per Shift	12
Operating Shifts per Day	2
Operating Days per Week	7
Scheduled Operating Days per Year	365
Number of Mine Crews	4

Pit and mine maintenance operations will be scheduled around-the-clock. Allowances for downtime and weather delays have been included in the mine equipment and manpower estimations.

A mill ramp up period for concentrator start-up has been considered. Provisions are included to reach full and steady production (throughput) by the end of the sixth month of year one of operation. The mill production targets schedule is presented in Table 16-6. The author believes this period is attainable, considering Hudbay’s recent experience in building a similar project in Peru (ramp-up to full production was also approximately 6 months).

TABLE 16-6: MILL RAMP-UP SCHEDULE

Month		Days	Efficiency	Design tph	Ramp Up Factor	Mill Ore 000 tpm	Mill Ore 000 tpd
1	Q1	31	92.1%	4,073	30%	837	27.0
2		28	92.1%	4,073	40%	1,008	36.0
3		31	92.1%	4,073	75%	2,093	67.5
4	Q2	30	92.1%	4,073	87%	2,349	78.3
5		31	92.1%	4,073	95%	2,651	85.5
6		30	92.1%	4,073	100%	2,700	90.0

16.3.2        Mill Feed and Cut-Off Grade Strategy

An elevated cut-off grade strategy has been implemented to bring forward the higher grade ore from the pit into the early part of the ore production schedule. Delivering higher grade ore to the mill in the early years will improve the net present value and internal rate of return of the Project.

NSR values are calculated for each block in the resource model to represent the net Cu, Mo, and Ag metal values. The pit reserves are estimated based on a cut-off with an NSR value of $6.00/ton. This

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is the minimum value of mineralized material that will cover the processing and G&A costs, and is therefore reserved for mill feed.

Priority plant feed will consist of high grade material (NSR above $12.00/ton) . The medium and low grade material (NSR between $6.00 and $12.00/ton) will be fed as needed and will otherwise be stockpiled.

Mill feed strategy considers that high grade ore stockpiled during the pre-stripping period will be processed during the first year of plant production, utilizing a dynamic stockpile located near the primary crushing facility.

16.3.3        Overburden Stripping Requirements

Mineral reserve tabulations by bench and by phase, and a mine production scheduling program (MSSO, a module from MineSight® software) were used to analyze long-term stripping requirements for the Project. Elevation and phase order dependencies and sinking rate controls were used in conjunction with mill ore production targets and an internal NSR cut-off of $6.00/ton to simulate open pit mining. The program, through successive iterations, allows the user to examine waste stripping rates over the life of the mine and their impact on ore exposure and mill head grades.

The stripping analysis determined that a minimum preproduction stripping of approximately 94 million tons of waste was required. Approximately 11 million tons of ore will also be mined and stockpiled during this period. The estimated Year 1 waste stripping total is 100 million tons, followed by 87 million tons for Year 2. The estimated waste stripping from Year 3 through Year 11 will average about 95 million tons per year to maintain a minimum of six months of ore exposure levels for uninterrupted ore deliveries to the mill. Waste stripping rates will decline to an annual average of 32 million tons for the next 3-year period, and then drop to an average of 5 million tons for the last 3 production years as the final mining phase approaches the pit bottom.

Preproduction stripping is planned to be conducted over a 12-month timeframe and will ramp up according to the delivery of mining equipment (particularly electric shovels) and the hiring and training of work crews. The long-term and peak mining rates suggest the use of at least two large (60 yd³ class) electric shovels, one large (25 yd³) front-end loader and a hydraulic shovel (46 yd³). Ramp-up for the mine is a part of the mine schedule target for total movement capacity by period.

16.3.4        Mine Plan

Mining sequence plans have been developed on a quarterly basis from preproduction through to the end of year 5, and on an annual basis through to year 19. The preproduction period consists of four quarters, or 12 months. 

A mine life of approximately 19 years of production is projected by this development plan. Peak mining rates of 367,000 tpd of total material are planned in year 1 until year 11. Average mining rates during years 12-14 are planned to be 180,000 tpd d of total material, and then reduce to an average of 105,000 tpd from years 15 – 17 as the strip ratio drops.

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During the pre-production period before the first ore is delivered to the mill, the pit will be pre-stripped of waste to expose ore and develop the upper benches for subsequent pushbacks. Specifically, pre-stripping will occur in pit Phase 1 for ore exposure and in Phase 2 to 6 for development. By the end of pre-production, the Phase 1 pit will be down to the 5,150-foot bench. At the end of this period, the ore will be exposed to deliver uninterrupted ore to the mill. Phase 2 will be stripped sufficiently ahead to ensure a supply of ore for mill feed by Year 2.

The mine schedule drawings for life of the mine are shown in Figure 16-11 to Figure 16-31. Mine schedule details regarding total annual movement, stripping ratios, mill feed by lithology and by ore type are presented in Figure 16-32 to Figure 16-34. The estimated mine production schedule is summarized in Table 16-7.

The development of the WRSA (Section 16.4.1) and dry stack tailing facility buttres ss (Section 16.4.2) was designed to allow concurrent reclamation of the facilities. Materials are placed at final slope angles required by the reclamation and closure plan conceptually approved by the USFS. Revegetation will start once the next lift is completed, this provides the opportunity for early bond release in some areas of the facility minimizing closure requirements at the end of the life of the facility. All elevations shown are in feet (AMSL).

FIGURE 16-11: MINE PLAN END OF PERIOD PRE-PRODUCTION

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FIGURE 16-12: MINE PLAN END OF PERIOD YEAR 1

FIGURE 16-13: MINE PLAN END OF PERIOD YEAR 2

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FIGURE 16-14: MINE PLAN END OF PERIOD YEAR 3

FIGURE 16-15: MINE PLAN END OF PERIOD YEAR 4

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FIGURE 16-16: MINE PLAN END OF PERIOD YEAR 5

FIGURE 16-17: MINE PLAN END OF PERIOD YEAR 6

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FIGURE 16-18: MINE PLAN END OF PERIOD YEAR 7

FIGURE 16-19: MINE PLAN END OF PERIOD YEAR 8

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FIGURE 16-20: MINE PLAN END OF PERIOD YEAR 9

FIGURE 16-21: MINE PLAN END OF PERIOD YEAR 10

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FIGURE 16-22: MINE PLAN END OF PERIOD YEAR 11

FIGURE 16-23: MINE PLAN END OF PERIOD YEAR 12

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FIGURE 16-24: MINE PLAN END OF PERIOD YEAR 13

FIGURE 16-25: MINE PLAN END OF PERIOD YEAR 14

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FIGURE 16-26: MINE PLAN END OF PERIOD YEAR 15

FIGURE 16-27: MINE PLAN END OF PERIOD YEAR 16

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FIGURE 16-28: MINE PLAN END OF PERIOD YEAR 17

FIGURE 16-29: MINE PLAN END OF PERIOD YEAR 18

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FIGURE 16-30: MINE PLAN END OF PERIOD YEAR 19

FIGURE 16-31: MINE PLAN, FINAL TOPOGRAPHY

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	Form 43-101F1 Technical Report

TABLE 16-7: MINE PRODUCTION SCHEDULE – LOM RP16AUG

		Yr -1	Yr1	Yr2	Yr3	Yr4	Yr5	Yr6	Yr7	Yr8	Yr9	Yr10	Yr11	Yr12	Yr13	Yr14	Yr15	Yr16	Yr17	Yr18	Yr19	Total
Mill	M Tons	-	28.1	32.9	32.9	32.9	32.9	32.9	32.9	32.9	32.9	32.8	32.9	32.9	32.8	32.9	32.9	32.9	32.9	32.9	5.5	592.0
	SCu %	-	0.43	0.50	0.50	0.44	0.51	0.55	0.48	0.55	0.39	0.46	0.32	0.30	0.34	0.37	0.39	0.36	0.24	0.14	0.12	0.40
	TCu %	-	0.51	0.55	0.54	0.49	0.55	0.59	0.51	0.61	0.44	0.49	0.37	0.36	0.40	0.40	0.42	0.39	0.28	0.19	0.18	0.45
	ASCu%	-	0.08	0.05	0.05	0.05	0.05	0.05	0.04	0.06	0.05	0.04	0.05	0.06	0.06	0.04	0.03	0.03	0.04	0.06	0.06	0.05
	Ox%	-	16%	9%	9%	12%	10%	9%	9%	11%	14%	10%	14%	15%	15%	11%	9%	9%	14%	24%	24%	12%
	Mo %	-	0.01	0.01	0.01	0.01	0.01	0.01	0.01	0.02	0.01	0.01	0.01	0.01	0.01	0.01	0.02	0.02	0.01	0.01	0.00	0.01
	Ag opt	-	0.16	0.18	0.16	0.17	0.17	0.18	0.13	0.14	0.10	0.13	0.10	0.10	0.12	0.12	0.14	0.14	0.11	0.07	0.07	0.13
	CuEq %	-	0.59	0.64	0.63	0.58	0.64	0.69	0.60	0.70	0.50	0.58	0.43	0.43	0.48	0.48	0.52	0.49	0.34	0.23	0.21	0.53
	NSR $/t	-	21.8	25.6	25.3	22.6	25.7	28.0	24.2	27.8	19.3	23.3	16.4	15.6	17.9	19.1	20.6	19.6	12.9	7.2	6.3	20.6
	SWCL %	-	6.0	7.1	8.3	8.4	8.2	8.9	8.9	8.7	7.9	7.0	7.1	7.6	8.5	7.7	6.5	6.3	6.9	8.5	7.9	7.7
	MGCL %	-	1.4	2.6	3.2	2.5	4.6	4.8	4.4	5.7	2.6	4.5	3.5	3.8	3.0	3.9	3.7	4.2	2.3	1.5	1.2	3.4
	BWI Kw-Hr/t	-	13.6	13.2	12.9	12.8	12.3	13.3	12.6	13.1	12.7	12.7	13.1	12.9	12.2	11.7	12.2	12.6	13.2	13.3	12.4	12.8
SPM-	M Tons	-	0.1	-	-	-	-	-	-	-	-	-	0.4	-	-	-	-	-	1.7	4.6	1.2	8.0
SPM+	M Tons	1.7	1.2	0.9	2.2	0.4	1.0	0.3	0.1	0.2	-	-	-	-	-	-	-	-	-	-	-	8.0
SPO-	M Tons	-	0.3	-	-	-	-	-	-	-	-	-	0.2	-	-	-	-	-	1.0	3.7	0.8	5.9
SPO+	M Tons	1.9	0.7	0.6	0.6	0.3	0.7	0.1	-	1.0	-	-	-	-	-	-	-	-	-	-	-	5.9
SPS-	M Tons	-	3.2	-	-	-	-	-	-	-	-	-	2.9	-	-	-	-	-	14.0	24.6	3.5	48.3
SPS+	M Tons	7.7	2.3	10.9	10.4	2.1	4.2	6.1	1.7	3.0	-	-	-	-	-	-	-	-	-	-	-	48.3
WRSA	M Tons	78.2	45.5	58.5	47.3	75.0	93.3	23.0	97.3	23.7	47.2	27.1	1.0	9.9	15.8	8.4	3.6	5.8	5.7	-	-	666.3
DSTF	M Tons	15.6	54.1	28.2	38.7	21.3	-	69.6	-	71.4	51.9	72.1	98.1	51.7	9.5	-	-	-	-	-	-	582.3
Total	M Tons	105.0	132.0	132.0	132.0	132.0	132.0	132.0	132.0	132.0	132.0	132.0	132.0	94.0	58.0	41.0	36.0	39.0	39.0	33.0	5.0	1,903.0

Notes: CuEq% is calculated based on metal prices of $3.15/lb Cu, $11.00/lb Mo and $18.00/oz Ag.

             Ox%: Oxide Ratio, between Soluble Copper and total copper, as per formula: ASCu% / TCu%.

             SWCL%: Swelling clays grade.

             MGCL%: Magnesium clays grade.

             BWI: Bond Work Index.

             SPM-: Mixed Ore Stockpile (Out).

             SPM+: Mixed Ore Stockpile (In).

             SPO-: Oxide Ore Stockpile (Out).

             SPO+: Oxide Ore Stockpile (In).

             SPS-: Sulfide Ore Stockpile (Out).

             SPS+: Sulfide Ore Stockpile (In).

             WRSA: Waste rock facility destination.

             DSTF: Dry stack tailings facility destination.

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	Form 43-101F1 Technical Report

FIGURE 16-32: ROSEMONT MINE SCHEDUULE, MATERIAL MOVEMENT

FIGURE 16-33: ROSEMONT MINE SCHEDULE, MILL FEED ORE BY LITHOLOGY

Note: Ore Mill in others lithology are Abrigo 9.6Mt@0.42%Cu, Andesite 16Mt@0.23% Cu, Arkose 13.5Mt@0.21%Cu, Bolsa 6.3Mt@0.3 8%Cu, Escabrosa 9.8Mt@0.54%Cu, Glance 4.7Mt@0.22%Cu, Granodiorite 2Mt@0.62%Cu, Martin 5.7Mt@0.31%Cu, QMP 17.7M t@0.40%Cu and Scherrer 7.5Mt @0.35%Cu, which represents 16% of the total Ore mill.

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FIGURE 16-34: ROSEMONT MINE SCHEDULE, MILL FEED ORE BY ORE TYPE

16.4        Mine Facilities

16.4.1        WRSA

Overburden and other waste rock encountered in the course of mining will be placed into the WRSA located to the south and southeast of the planned open pit and into landform area. The design criteria for the WRSA area and associated haul roads are summarized in Table 16-8 below. The general mine site layout is shown in Figure 16-1.

TABLE 16-8: WRSA DESIGN CRITERIA

Parameter	Value
Angle of Repose	37°
Average Tonnage Factor (with swell)	16.02 ft3/ton
Overall Slope Angle	3.5H:1V
Total Height, feet	600
Haul Road, feet	120
Max Elevation, feet (AMSL)	5700

One of the objectives in the early years of operation (specifically, Years 1 to 5) is to construct a series of buttresses and berms around the eastern and southern perimeters of the DSTF and WRSA respectively, for permit commitment. These buttresses and berms will also allow re-grading and re-vegetation of the facilities side slopes at much earlier time periods than with traditional mine waste rock closure plans.

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The WRSA berms and internal loading plan are designed to facilitate subsequent re-grading and concurrent reclamation. Side slopes in the WRSA will be re-graded to a maximum of 3:1 (horizontal: vertical) slopes. The WRSA loading plan will consist of haul trucks end-dumping waste rock in 100-foot lifts at the angle of repose (approximately 37°). The WRSA crests will be set back to allow simple dozing of the crests down to meet the target re-graded slope angles to support concurrent reclamation.

16.4.2        DSTF Buttress

Dry stack tailing resulting from processing mine ore will be placed behind the buttresses constructed from mine waste rock. Any acid generating waste will be disposed in the DSTF buttress. The DSTF is north of the WRSA and east-northeast of the pit. The design criteria for the DSTF and associated haul roads are summarized in Table 16-9 below. The general mine site layout is shown in Figure 16-1.

TABLE 16-9: DSTF BUTTRESS ROCK STORAGE DESIGN CRITERIA

Description	Unit
Angle of Repose	37°
Average Tonnage Factor (with swell)	16.02 ft3/ton
Overall Slope Angle	3.5H:1V
Total Height, feet	700
Haul Road, feet	120
Max Elevation, feet (AMSL)	5,490

The DSTF and WRSA are described in more detail in Section 18.4 of this report. As the mine matures, waste rock generation declines which forces maximum utilization during many of the early years to construct the buttress in the DSTF and berms in the WRSA. The DSTF buttress construction is planned to be finished by the end of year 13 with sufficient capacity to store the remainder of the tailing material generated during the 19-year mine life. Table 16-10 summarizes LOM Waste distribution for WRSA and DSTF. A cross section N-S view of DSTF buttress by year is shown in Figure 16-35.

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TABLE 16-10: LOM WASSTE ROCK DISTRIBUTION AND LANDFORMI NG STORAGE PLAN

FIGURE 16-35: DSTF NS SECTION VIEW, LOM BUTTRESS BY YEAR

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	Form 43-101F1 Technical Report

16.5        Mine Equipment

16.5.1        Equipment Operating Parameter

Mine equipment was selected based on the production requirements shown in Table 16-7. During the first quarter of preproduction, a 46 yd³ hydraulic excavator and a 25 yd³ loader will be matched with 260-ton-class haul trucks; supported with dozers, graders and water trucks to develop the initial mine area. At the end of the second quarter of preproduction, the first 60 yd³ class electric shovel will come on line followed by one more in preproduction third quarter.

The mine will operate two 12-hour shifts per day, for 365 days a year. No significant weather delays are expected and the mine will not be shut down for holidays. The craft work schedule will consist of a standard four crew rotation.

Material characteristics used to determine productivity calculations are listed in Table 16-11. Although there are several different rock types at Rosemont, the weighted average of all rock types was used for production estimation. Major loading and haulage equipment will be equipped with electronic load monitors, which will ensure optimum loading. All equipment production is reported in dry short tons, which is consistent with the reserve model. Moisture content is expected to range between 3 and 4 percent; for haulage calculations 3.5 percent was used.

TABLE 16-11: MATERIAL CHARACTERISTICS

Parameter	Value
In Situ Bulk Density	11.85 cubic feet per ton
Material Swell	40 Percent
Loose Density	16.02 cubic feet per ton
Moisture Content	3.5 Percent

16.5.2        Mine Equipment Calculation

Mine equipment requirements were developed based on the annual tonnage movement projected by the mine production schedule in Table 16-7, bench heights of 50 feet, two twelve hour shifts per day, 365 days per year operation, with manufacturer machine specifications and material characteristics specific to the deposit.

Specific manufacturer’s models used in this study are only intended to represent the size and class of equipment selected. The final equipment manufacturer selection will be done as required to meet delivery dates and current need of the operation.

A summary of fleet requirements by time period for major mine equipment is shown in Table 16-12. Furthermore, Table 16-13 lists equipment KPI’s, Availability and Utilization, and equipment productivity used to dimension equipment fleets for the different mine operations.

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This represents equipment necessary to perform the following mine tasks:

• Mine site clearing and topsoil salvage and stockpiling
• Construction of the main haul roads
• Production and pre-split drilling
• Loading and hauling of sulfide ore to the primary crusher (located on the east side of the pit), and waste rock to WRSA and DSTF areas
• Maintaining mine haulage and access roads
• Maintaining WRSA, DSTF, berms, and re-grading of slopes and final surfaces

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TABLE 16-12: MAJOR FLEET REQUIREMENTS FOR LOM

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16-13: MAJOR EQUIPMENT KPI AND PRODUCTIVITY

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16.6        Mine Operations

16.6.1        Drilling and Blasting

Production drilling will be done using 10-5/8 inch holes on a 30-foot by 30-foot pattern for ore and 33-foot by 33-foot pattern for waste. Blast hole depth will be 50 feet with 5 feet of sub-drilling. Subgrade drilling in limestones and skarns may be increased if hard toe conditions are encountered.

Drilling speed rates will vary between 129-153 feet per hour depending on the rock type and mineralization. The penetration rates are consistent with rates being used by other mines like Constancia which is currently in operation and has characteristics similar to Rosemont.

Powder factors varied between 0.29 – 0.43 pounds per ton depending on rock type and mineralization. Ammonium nitrate and fuel oil (“ANFO”) blasting agents will be loaded in dry holes, while wet holes will be pumped dry and sleeved before loading with ANFO. If this cannot be accomplished, emulsion will be used as a wet hole explosive.

Drills will be outfitted with GPS and electronic sensing units to allow recording of penetration rates in drill holes to assist in determining decking requirements for individual holes. Drill productivities are expected to range between 8,500 and 10,500 tons per hour (“tph”), depending on rock type.

16.6.2        Loading

Major loading equipment consists of two 60 yd³ class electric shovels, one 46 yd³ hydraulic excavator and a 25 yd³ front-end loader. On average, 71% of total material movement will be handled by the electric shovels, 22% by the hydraulic shovel and 7% by the front end loader.

The equipment was selected to work a 50-foot bench height and load 260 ton-class trucks. For this study, the 260-ton-class trucks were chosen based on economics, but the loading fleet is sized for the larger trucks to give the operator flexibility in fleet selection at a later date.

Loading 260-ton trucks with a 60 yd³ class shovel requires three passes at 35 seconds per cycle, 30 second spot and queuing for a total load time of 2.30 minutes per truck. Loading the 260-ton trucks with 46 yd³ hydraulic excavator requires four passes at 35 seconds per pass, a 30-second spot time and queuing time, for total load time of 2.8 minutes. Finally, 260-ton trucks loaded with a 25 yd³ FEL require seven passes at 40 seconds per pass, a 30-second spot time and queuing time, for total load time of 5.2 minutes.

Loading equipment production rates vary during equipment start up, and according to operator training and experience. After reaching a steady state, the 60 yd³ class shovel productivity will be 6,800 tph, hydraulic shovel will be 5,400 tph and the loader productivity will be 2,800 tph. Loading productivity is directly related to how well the shovel-loader/trucks match, the material being loaded and the haulage profile.

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16.6.3        Hauling

The 260-ton class truck was chosen based on an economic evaluation and as a result of the support in the region. Main factors influencing the study were fuel burn, tire costs and repair costs. Truck fleet requirements vary from 23 units at the start of pre-production to 38 by year 6. The fleet remains constant from year 6 until year 8, when the waste volumes start to decrease and only 18 units are required. In year 18, the truck requirements decrease to 6 units. An average load factor of 260 tons was used for production calculations for haulage trucks.

16.6.4        Support Equipment

Major support equipment includes mine equipment that is not directly responsible for production, but which is scheduled on a regular basis to maintain in-pit and ex-pit haul roads, pit benches, WRSA and DSTF and to perform miscellaneous construction work as needed. Equipment operating requirements were estimated for this equipment based on the major mine equipment support requirements and WRSA slope re-grading schedules. Equipment in the mine support fleet includes:

● Crawler dozers, D10T2 class

● Rubber-tired dozers, 834K class

● Motor graders, 14H class

● Water trucks, 777G class

In general, the rubber-tired 834K-class dozers will be used in the pit to clean up around the primary loading units, with the track dozers used for haul road construction, pit development, WRSA and DSTF management, and final re-grading requirements. The graders and water trucks will be used to maintain roads and control dust.

16.7        Mine Engineering

16.7.1        Geotechnical and Mine Planning

CNI was contracted by Hudbay to provide an update of their geotechnical recommendations for slope angles for the open pit development of the Rosemont deposit. The current and previous work included geologic and geotechnical mapping, drilling, rock strength testing and slope stability analysis to determine pit slope design criteria that is consistent with industry norms for safety and cost effectiveness. CNI provided a report in May 2016 - Feasibility-Level Geotechnical Study for The Rosemont Deposit.

Based on the CNI report, Hudbay worked to find the best strategy to combine geotechnical engineering, pit design, mine planning and operational point of view. With regards to geotechnical engineering and pit design, the following considerations have been made:

● Concave pit design which is more stable that a convex pit

● Height of ultimate pit is more than 2,900 feet deep at an elevation of 3,100 feet (AMSL)

● A two ramp ingress/egress system

 

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●  Drainage system following the main haulage ramps

●  Mining sequence, by phases and periods:

o Final pit wall will be established during year 08 (Figure 16-19)

o On-going evaluation of new data resulting from actual pit development

With respect to the general mine development sequence, Hudbay has developed the following strategy:

● As part of the pit dewatering plan, three pumping wells will be installed close to the Phase I development area. As currently planned, these holes will be core drilled using PQ diameter to obtain additional geological, geotechnical, and hydrogeological information. During year 5, one additional pumping well will be developed with the same strategy (multifunction hole).

●  Pre-stripping will expose several geological faults identified during the geotechnical study, allowing for better definition, exact location, geotechnical properties and behavior.

●  The strategy will remain the same as the mine progresses and other faults are encountered. Mine development will include specific design parameters to minimize the unintended structural issues, specifically:

o Inter ramp angle controls and review for optimization (wall phases)

o Bench face angles

o Control wall damage with blasting analyses

o Blasting control (VPP)

o Ground control (survey, water level)

o Slope monitoring system

RQD (%) and the hardness block model have been developed to support a geotechnical strategy for mine design to be implemented for mine planning and operations. Figure 16-36 to Figure 16-39 show plan and section views of the block model and final Rosemont pit.

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FIGURE 16-36: ROSEMONT GEOTEECHNICAL SECTORS

FIGURE 16-37: SECTION AA’ SHOWING RQD VAALUES IN FINAL ROSEMONT PIT

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FIGURE 16-38: ROSEMONT FINAL PIT, LITHOLOGY IN FINAL WALL

FIGURE 16-39: SECTION BB’ SHOWING HARD VALUES IN FINAL ROSEMONT PIT

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16.7.2        Hydrogeology and Mine Planning

Neirbo Hydrogeology was contracted by Hudbay to provide a hydrogeological study. Based on a refined and localized version of the 2010 Regional Groundwater Model prepared for the Environmental Impact Statement (EIS) process, Neirbo provided a report in May 2016 –Hydrogeological Study for The Rosemont deposit. Based on the Neirbo report, Hudbay worked to find the best strategy to combine: pit dewatering, pit design, mine planning and operational objectives.

The following general strategy has been considered:

● Starting the drilling and pumping before pre-stripping and continuing during the pre- stripping

● Dynamic updating of the hydrogeological parameters and model for each well

● Monitoring wells focused on dewatering

● Active and passive depressurization verification according to mining advance

● Updating the areas indicating high and low conductivity

● Establishing an operational correlation between the geological, geotechnical and hydrogeological parameters

The pit dewatering plan consists of:

● 10 wells during pre-production (14,850 feet)

● 10 additional wells between year 1 to year 5 (10,800 feet)

● 10 additional wells after year 5

● Annual horizontal drain sustaining capital cost ($ 3M) was considered for the arkose material as it will be mined every year in the mine’s life

● Limited by pre-production pumping described in the EIS as 18,500 acre-feet of water

The pit opens up in the central-western zone; away from the final walls which is expected to provide an opportunity to:

● Pre-mining

o To control the inter ramp angle (“IRA”) and the bench face angle (“BFA”)

o To manage the water with wells and superficial water management

o To install and monitor the impact of 14 pumping wells

● Year 1

o As the mine expands through Phase 1, monitoring of the effects of the pumping wells will continue. This will include water captured via the in-pit ponds

o Phase 2 will begin in year one and monitoring of the pumping wells and surface ponds will continue.

● Year 2

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o Phases 1 and 2 will remain the active mining areas supported by the original pumping wells and ponds.

● Year 3

o Beginning in Year 3 and continuing for the remainder of the mine’s life, additional wells will be drilled to achieve the drawdown and depressurization requirements necessary to safely advance the ore extraction sequence.

Figure 16-40 presents LOM well holes in the final Rosemont pit, and Table 16-13 summarizes LOM well holes for the pre-production and operating stages.

FIGURE 16-40: SECTION AA’ SHOWING HARD VALUES IN FINAL ROSEMONT PIT

16.8        Manpower Requirements

16.8.1        Mine Operations Manpower

Mine supervision, technical staff, mine maintenance, wor rkshop personnel and equipment operator requirements over the life of the mine are based on the mine plan. During the Pre-Production period, direct (workshop and operators) and indirect (Staff, supervision and technicians) requirements total 337, building up to 424 in the steady production (132 M ton per year) and is shown in Table 16-14.

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Hourly mine operation personnel requirements are calculated based on equipment operating hour requirements, fleet estimation, shown in Table 16-12. Maintenance personnel are calculated based on estimated maintenance repair time for mining equipment. The percent of maintenance to total hourly personnel averages 32% throughout the mine life.

The schedule assumed that the hourly personnel would be hired two months prior to the date they were actually required on-site to facilitate training requirements for MSHA, Safety, Environmental and other required training which is captured in the mining costs. Operation training is captured in Operational Readiness estimates.

Mine staff manpower employees and salaries were developed for Mine Administration, Mine Geology, Mine Operations, and Mine Maintenance. Salaries were a composite of information provided by Hudbay which was calibrated against local mine salaries. Salary information includes wages, burden and bonus for staff employees.

TABLE 16-14: LABOR ESTIMATION FOR ROSEMONT MINE OPERATIONS

Item	Category	Mine Operations Labor Requirements (Steady Production)
Mine Operations	Direct Labor	227 (Operators for shovels, trucks, drills and auxiliary equipment)
Workshop Personnel	Direct Labor	90 (Mechanics, Welder, Electrician and Helpers)
General Mine Staff	Indirect Labor	34 (Management, Supervision & Laborer)
Technical Services	Indirect Labor	26 (Mine Engineer, Geologist, Geotechnical and Surveyor)
Mine Maintenance	Indirect Labor	47 (Management, Supervision and Technicians)
TOTAL		424

16.8.2        Process Operations Manpower

Hourly process plant operation personnel requirements are estimated based on hourly equipment operating criterion, and fleet estimation as shown in Table 16-15. Maintenance personnel are determined based on estimated maintenance repair time for equipment. The percent of maintenance to total hourly personnel averages 32% throughout for the mine life.

Operating labor costs are based on staffing levels developed by the Project for a copper molybdenum concentrator with a DSTF.

Staff positions were grouped into four broad categories:

1.	Mill Management
2.	Mill Operations
3.	Mill Maintenance
4.	Mill Technical Services

Staff numbers are based on 24-hour operating coverage with personnel working 12-hour shifts on a four crew roster system. Labor positions categorized are summarized in Table 16-15.

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No labor has been included for an analytical laboratory because there will be no on-site facility for this purpose. Mine, plant, and concentrate quality samples will transported off site to a contract laboratory.

Besides staff labor, allowances have been made for contract labor for major periodic tasks. This is to free up staff maintenance personnel and allow them to perform required/preventative maintenance tasks while the affected circuits are down.

TABLE 16-15: LABOR ESTIMATION FOR ROSEMONT PROCESS OPERATIONS

Item	Mine Operations Labor Requirements (Steady Production)
Mill Management	22 (Management and Supervision)
Mill Operations	76 (Operations Technicians)
Mill Maintenance	38 (Trade Workers and Technicians)
Mill Technical Services	15 (Engineers and Technicians)
TOTAL	151

16.8.3        General and Administration (G&A)

General and Administration has been derived from each area within the G&A group. The estimate for each area was built up using project manpower inputs and industry standards values.

The G&A group consist of the following departments and are showed in Table 16-16:

1.	General Administration - includes general office, legal and corporate affairs areas.
2.	Procurement - included in this area are administration and logistic.
3.	Safety
4.	IT- Information and Technology and support
5.	Accounting – includes controller, payroll, taxes and finance staff.
6.	Environmental – included in this area are permitting, monitoring, and mitigation personnel.
7.	Human Resources

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TABLE 16-16: LABOR ESTIMATION FOR ROSEMONT G&A OPERATIONS

Item	G&A Labor Requirements (Steady Production)
General Administration	16 (Management, Legal, and Corporate Affairs)
Procurement	17 (Administration and Logistic)
Safety	8
IT	8
Accounting	12 (Controller, Payroll, Taxes, Finance)
Environmental	22 (Monitors, Technicians)
Human Resources	5
TOTAL	88

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17        RECOVERY METHODS

17.1        Introduction

The Project process plant is a conventional copper-molybdenum concentrator and process design is typical of concentrators treating low sulfur copper porphyry-skarn style ores. The process involves crushing, grinding, flotation, molybdenum separation, concentrate dewatering, and tailings dewatering.

With minor modifications, the plant is designed to process on average 90,000 ton/d (32.8 million ton/y) of ore. The Project details included in this section were specifically designed and evaluated to fall within the permitted facility constraints included in the EIS and State of Arizona permits while optimizing production, minimizing costs and ramping up production as quickly as technically and environmentally as possible. State of Arizona Department of Environmental Quality (ADEQ) permits that were issued based on early designs will be amended to include designs included in the EIS and this section; these amendments are customary in the state.

This Technical Report includes refinements of certain aspects of the Project’s mine plan. While consistency with issued and pending environmental permits and analysis related thereto has always been a key requirement for this effort, updates to the original mine plan will be necessary. To the extent that any regulatory agency concludes that the current plan requires additional environmental analysis or modification of an existing permit, the intent will be to work with that agency to either complete the required process or to adjust the current mine plan as necessary.

17.1.1        Facility Layout and Location

The plant will be located east of the open pit and has been arranged in a north-south orientation. Ore flow is from south to north with the ROM stockpile and dump pad located to the south and concentrate filtration and load-out located at the north end of the facility.

ROM ore will be transported from the mine to the primary crusher by off-highway haulage trucks. After crushing the ore, it will be conveyed to the Coarse Ore Stockpile to then be conveyed to the concentrator facility.

Copper concentrate produced at the concentrator facility will be loaded onto highway haul trucks for transportation to smelting and refining facilities. Molybdenum concentrate will be bagged and loaded onto trucks for shipment to market.

An overall view of the processing facilities is given in Figure 17-1.

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FIGURE 17-1: OVERALL VIEW OF PROCESS PLANT LOOKING NORTH

17.1.2        Facility Description

The process plant is modelled after the Constancia Copper Project design with changes made in certain areas. In some cases, ore characteristics, local conditions and permitting constraints have dictated that changes be made such as the filter plant for dry stack tailings.

With minor modifications, the plant is designed to process 90,000 tpd (32,850,000 tpy). Production during the first year of operation is expected to fall short of full capacity (28,114,000 tons) to account for a staggered commissioning schedule and brief ramp-up and optimization period. Annual concentrate production is expected to reach an average of 344,000 tons.

The coarse ore stockpile, grinding areas, pebble crushers, conveyors, molybdenum plant, and filter plant process areas will be covered and painted to match the natural landscape. Design reviews were held to ensure that safety, environmental, permitting, technical, quality, constructability, maintainability and operability considerations had been correctly addressed.

Key facility design criteria are summarized in Table 17-1.

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TABLE 17-1: KEY FACILITY DESIGN CRITERIA

Parameter	Units	Value
Plant capacity	Mton/yr	32.9
Copper feed grade	Avg %	0.48
Copper feed grade	Max %	0.66
Molybdenum feed grade	Avg %	0.014
Molybdenum feed grade	Max %	0.023
Copper concentrate grade	%	32
Molybdenum concentrate grade	%	45

The design grades were selected on the basis of the prevailing mine plan in order to properly size flotation cells, pumps and pump boxes, pipelines, tanks, and concentrate de-watering and handling equipment to ensure adequate unit operation capacities under the vast majority of circumstances. While it is possible for feed grade to the plant to fall outside of range of these parameters, these excursions are expected to be both rare and brief, and can be managed with relative ease. Average and monthly maximum copper head grades in the current mine plan are 0.447% and 0.670%, respectively.

17.2        Buildings

The facility will include buildings for the following:

• Administration Facilities comprised of the Administration Building, Medical Clinic/Emergency Response Building, vehicular staging pads and dedicated vehicular parking areas.
• Security Gatehouse and Weighbridge (truck scale) are located at the entrance to the plant.
• Plant Maintenance Building where all routine plant maintenance activities are conducted. Additional ancillary space includes a secure tool crib, parts store and electrical and instrument shop.
• Metallurgical Lab which includes space for sample preparation and testing.
• Plant Change House that provides space for plant workers to shower, change and use the restroom.
• Plant Production & Maintenance Office for plant and maintenance staff.
• Plant Control Room, a modular building housing the plant control operators.
• Plant Warehouse Building with Plant Warehouse Office and external warehouse yard.

17.3        Processing Plant

The process plant design is based on a combination of metallurgical testwork, Project production plan and in-house information. Benchmarking has been used to define and support the design parameters. This includes the copper-molybdenum separation circuit where testwork has been limited to a few tests. This is due to the relatively large sample mass required for a more detailed molybdenum testwork program and analysis.

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The molybdenum plant design is based primarily on projected mass flows, grades and densities as well as the recent Constancia plant design.

The flowsheet has been developed from previous feasibility study work, value engineering studies and recent testwork. The Rosemont process plant includes the following unit processes and facilities:

• Primary crushing.
• Crushed ore stockpile and reclaim.
• Parallel SABC grinding lines with pebble crushing.
• Copper flotation comprising rougher flotation, concentrate regrind, and two stage cleaning.
• Cu-Mo concentrate thickening.
• Molybdenum flotation including roughing, concentrate regrind and five stages of cleaning.
• Molybdenum concentrate thickening, filtration and drying.
• Copper concentrate thickening and filtration.
• Copper concentrate load out and storage.
• Tailings thickening, filtration and dry stacking.
• Reagents storage and distribution (including lime slaking, flotation reagents, water treatment and flocculant).
• Grinding media storage and addition.
• Water services (including fresh water, fire water, gland water, cooling water and process water).
• Potable water treatment and distribution.
• Air services (including high pressure air and low pressure process air).
• Plant control rooms.

A generalized process flow diagram (PFD) is provided in Figure 17-2.

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FIGURE 17-2: PROCESS PLANT PROCESS FLOW DIAGRAM

The flowsheet consists of primary crushing, followed by two parallel SAG, ball milling and pebble crushing (SABC) circuits, copper flotation with regrinding ahead of cleaning, a moly separation circuit, concentrate thickening and filtering and tailings thickening, filtering and dry stacking.

With minor modifications, the process plant is designed to treat on average 90,000 tons/d (or 32.8 million ton/y).

Key design criteria used in the plant design are summarized in Table 17-2.

TABLE 17-2: KEY DESIGN CRITERIA

Parameter	Units	Value
Plant capacity	tons/day	90,000
Flotation feed size, P80	µm	140
Flotation feed density, nominal	% solids (w/w)	34
Flotation feed density, minimum1	% solids (w/w)	28
Tailings thickener underflow density	% solids (w/w)	65
Tailings filter cake moisture	%	15

Notes: 1. Minimum density at which design rougher minimum residence time can be achieved.

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The process plant capacity was determined by a combination of ore characterization testwork, environmental factors and regulatory constraints, mine planning, engineering estimation, and financial analysis to define best economic return for the Project. Flotation feed size was selected on the basis of the best balance of moderating energy input in the grinding circuit and achievement of recovery targets. Flotation feed density design values are based on the results of numerous flotation tests and typical industry practice. Tailings dewatering targets are selected on the basis of thickening and filtration testwork, capacity requirements, and environmental compliance limitations. Equipment and unit operations throughout the plant have been designed to meet these requirements on a routine basis.

The overall annual plant operating schedule is 8,059 hours (92% of available hours). Operating availability is summarized in Table 17-3.

TABLE 17-3: PLANT UTILIZATION SUMMARY

Description	Units	Value
Crusher Utilization	%	75
Grinding and Flotation Availability	%	93.0
Concentrate Filter Utilization	%	84
Tailings Filtration and Dry Stack Utilization1	%	98.6
Overall Concentrator Asset Efficiency2	%	92.1

¹factored on a mill runtime basis ²assumes production ceases when ore flow to the SAG mill is interrupted

Availability estimates are based on typical industry experience for plants of similar size and configuration and utilizing typical maintenance and operating practices. A 99% utilization factor is applied to the availability to derive the asset efficiency factor, which accounts for non-productive time unrelated to mechanical maintenance or failure such as shut-down (grind-out) and start-up, lack of ore, or other upstream/downstream constraints.

17.4        Crushing

17.4.1        Primary Crushing

The Primary Crusher is a 60 x 113 size Gyratory Crusher fitted with manganese steel concave and mantle liners. The primary crusher was selected based on the required mill feed rate, expected ROM feed size distribution, ore bulk density, crushing work index, stockpile capacity, and SAG feed size. The design crusher feed rate is 6,000 tons/h with a capacity of 90,000 tons/d, based on a crusher availability of 75% and a 20% catch-up capacity factor. Design parameters including the expected range of feed size distributions and crushing work indices were provided to vendors to confirm throughput performance and motor power requirements. The primary crusher dump pocket is designed to allow two trucks to dump simultaneously, one from each side. It has the capacity to hold two truckloads, approximately 520 tons in total.

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FIGURE 17-3: PRIMARY CRUSHER

A modular Crusher Control Room is located above the ROM wall to provide a direct line of sight to the dump pocket as well as the Stockpile Feed Conveyor. The Control Room includes space for crushing plant operators and two mine fleet controllers who manage the truck fleet. It is fitted with washroom and break facilities. A water spray system is fitted at each corner of the dump pocket for dust control when trucks are dumping. Dust generated in the transfer point between the Feeder and the Stockpile Feed Conveyor is captured by a dedicated Cartridge Dust Collector. Dust generated in the crusher vault is vented to the dump hopper and controlled by the water spray system.

17.4.2        Stockpile Feed Conveyor

Crushed ore is transferred from the Primary Crusher to the Crushed Ore Stockpile by a Stockpile Feed Conveyor. The conveyor belt is 72 inches wide and 1,014 feet in length with a 211-foot lift and has capacity to convey 6,600 tons/h of crushed feed to the stockpile (i.e. crusher capacity +10%). The conveyor is covered outside the Stockpile Dome to minimize dust emissions.

The conveyor is driven by two 1,200 HP drives (one mounted on each side of the drive pulley) complete with high speed disc brake and variable speed drive for controlled start-up. The drives are located adjacent to the gravity take-up.

17.4.3        Coarse Ore Stockpile and Reclaim

Crushed ore for both grinding lines is stored in a single Conical Ore Stockpile. The stockpile is enclosed by a dome structure with an impervious colored fabric cover (approx. 380 feet diameter and 164 feet high) as shown in Figure 17-4.

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The Stockpile Cover consists of a structural steel multi-arch frame pinned to a circular ring beam at the center of the dome and pinned at ground level to a concrete ring beam. A 20-foot-wide path around the stockpile perimeter provides access for dozer and equipment travel inside the cover. Two sets of access doors are included to allow machinery access to the stockpile area.

Live capacity of the stockpile will range between 23,000 to 51,000 tons (6 to 12 hours at full production rate and dependent on prevailing ore characteristics), representing 14% to 28% of the total stockpile capacity of 198,000 tons. Stockpile ore in dead storage can be reclaimed by heavy equipment (dozer and/or excavator) to allow for up to two additional days’ interruption of feed from the primary crusher.

FIGURE 17-4: STOCKPILE FABRIC COVER

Coarse ore is reclaimed from the stockpile by four 72-inch wide Apron Feeders, two for each grinding line. Each Apron Feeder is fitted with a variable-speed drive and has the capacity to provide 100% of the full tonnage rate to its respective SAG mill.

17.5        Grinding

17.5.1        SAG and Ball Mill Grinding

The selected grinding circuit consists of two parallel SABC grinding lines, each comprising one SAG mill in closed circuit with a sizing screen and pebble crusher followed by a ball mill in closed circuit with hydrocyclones. The selected grinding mills are summarized as follows:

• SAG mills – 36 feet diameter (inside shell) x 24 feet effective grinding length (EGL) with 18 MW twin pinion drives (9 MW per pinion). Motors are controlled via an SER / hyper synchronous variable speed drive.

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• Ball mills – 26 feet diameter (inside shell) x 40.5 feet effective grinding length (EGL) with 16.4 MW GMD.

Mill size and power requirements were determined via Ausgrind, Ausenco’s proprietary power-based comminution calculation program using the 75^th percentile values of ore parameters (ore competency and hardness). RQD was also used to adjust SAG feed size based on a correlation identified between RQD and sample depth. A +10% design factor was also added to the SAG motor specification to ensure mill power limitations would not be a significant factor for achieving throughput targets.

The mills are positioned at right angles to the Feed Conveyor as shown Figure 17-5 to minimize footprint. A full-width platform is located in between the Grinding Mills to provide sufficient space for Mill relining activities.

FIGURE 17-5: GRINDING BUILDING FROM STOCKPILE LOOKING NORTH (ROOF AND WALLS REMOVED)

17.5.2        Pebble Crushing and Conveyor Systems

Measured ore competency shows that an SABC grinding circuit with a pebble crusher is required. Pebble crushers were selected based on crusher feed rate, ore characteristics and competency factors, pebble top size, and crusher product size. Pebble rate is expected to vary from 510 tph (15% to 25% of new SAG mill feed nominally), up to 700 tph (30% with worn SAG mill grates).

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Cone crushers (one per grinding line) were selected for pebble crushing duty and will have sufficient capacity of 495 tph to over 1000 tph depending on liner profile and gap setting.

FIGURE 17-6: PEBBLE CRUSHING LOOKING SOUTH FROM THE GRINDING AREA

Pebbles from each SAG Mill are transferred to the Pebble Crusher Bin via the 54-inch Pebble Conveyors. Tramp metal is captured by two cross-belt self-cleaning magnets arranged in series for each crusher.

The Pebble Conveyor discharges onto a diverter gate which directs the material to the Pebble Crushing Bin or allows bypass to the Pebble Conveyor Bypass Bunker or the Pebble Crusher Product Conveyor.

17.6        Copper Flotation

The Copper Flotation Area is positioned perpendicular to the Grinding Area taking advantage of the ground sloping west to east as shown in Figure 17-7.

The circuit consists of two parallel trains of Rougher Flotation Cells. A third train of Cells includes the first and second stage Cleaner Cells as well as the Cleaner Scavenger Cells and is arranged parallel to the Rougher Cells.

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FIGURE 17-7: COPPER FLOTATION

The copper flotation circuit consists of two parallel trains of four rougher flotation cells followed by rougher concentrate regrind and two stages of cleaner flottation. Flotation feed, from primary cyclone overflow, reports to two rougher flotation trains each consisting of four forced air mechanical flotation tank cells. The Copper Rougher Flotation Cells are conventional 630 m³ (22,000 ft³) tank cells with 650 HP direct drive arrangement and are fed low pressure air by blowers. The two lines provide a total of 34 minutes’ residence time at the nominal feed density (34% solids) and 28 minutes’ residence time at 28% solids feed density at a potential 90,000 tpd.

The rougher flotation cells produce a low grade copper-molybdenum (Cu-Mo) concentrate that requires further liberation and upgrading. Copper rougher concentrate is combined in the copper regrind feed hopper and pumped to the regrind cyclones for classification.

The copper rougher tailings stream from each train are combined and gravitate to the flotation tailings thickener feed distributor via a cross-cut sampler r. The cleaner scavenger concentrate can also be directed to the regrind feed hopper if required. The underflow from the copper regrind cyclone cluster reports to the regrind mill, which overflows back into the copper regrind feed hopper.

The overf flow from the regrind mill cyclone cluster is pumped directly to the copper cleaner flotation circuit. The copper cleaner circuit consists of two stages of cleaning and one bank of cleaner scavenger cells.

The first cleaner consists of four forced-air mechanical staged flotation reactor (“SFR”) flotation cells complete with froth wash water system to minimize non-sulfide gangue entrainment. The second cleaners also comprise four forced-air mechanical SFR cellls with froth wash water systems.

The Copper Cleaner 1, Copper Cleaner 2 and Cleaner Scavenger flotation cells will be SFRs. The Copper Cleaner 1 and Cleaner Scavenger cells will have approximately 3,900 feet³ of volumetric capacity, while the Copper Cleaner 2 will have 1600 feet³ of volumetric capacity.

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SFR flotation cells were selected for the cleaner duty due to their ability to achieve high upgrade ratios with a relatively small footprint and reduced air and power consumption.

The online stream analyzer (“OSA”) is located in a dedicated area on the west side of the facility. Major concentrate and tailings streams are pumped to the OSA to allow optimization of reagent additions and flotation performance. Samples are collected using Gravity Samplers or Pressure Samplers at the pump discharges and transferred to the OSA using peristaltic pumps. The samples are sorted by a multiplexer and are returned to the process by horizontal centrifugal pumps after analysis.

17.6.1        Copper Regrind

The Copper Regrind circuit consists of a Copper Regrind Mill Feed Hopper and a Copper Regrind Cyclone Cluster in a closed circuit with the Copper Regrind Mill, as shown inFigure 17-8.

FIGURE 17-8: COPPER REGRIND AREA LOOKING WEST

The concentrate regrind mill was selected based on the expected range of concentrate feed rate (rougher mass pull), estimated feed size and product size (40 µm). Regrind product size was selected based on mineralogy studies and locked cycle flotation tests.

A typical regrind cyclone partition curve was used to estimate cyclone underflow size distribution and cyclone mass split. The estimated nominal recirculating load based on simulated size distributions was 170% with regrind mill feed size F₈₀ (cyclone underflow) approximating 70-75 µm.

Regrind mill power requirements were estimated based on the overall flotation circuit mass balance, grind-size target as determine by testwork programs (40 µm), and Ausenco’s in-house regrind specific energy data set calculations. A VertiMill with 3000 kWh motor was selected for regrinding duty.

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17.6.2        Copper-Molybdenum Concentrate Thickening

The bulk copper-molybdenum (Cu-Mo) concentrate is pumped from the copper flotation circuit to the Cu-Mo concentrate thickener via a trash screen. The trash screen removes coarse oversize that can damage or block downstream equipment, e.g. copper pressure filter ports and metallurgical samplers. Trash reports to a collection box via a chute. Trash screen undersize gravitates to the copper-molybdenum concentrate thickener via a thickener feed box.

Cu-Mo concentrate is thickened to reduce the volume of residual copper flotation reagents in the molybdenum flotation feed.

Thickener overflow from the Cu-Mo concentrate thickener is pumped to the tailings thickener where it is recovered as process water. Cu-Mo concentrate thickener underflow is pumped to the molybdenum flotation circuit by centrifugal pumps.

The Cu-Mo concentrate thickener and molybdenum flotation circuit can be bypassed if required by diverting the bulk Cu-Mo concentrate (from copper flotation) direct to the copper concentrate thickener.

The area is bunded and can contain the entire volume of a Filter Feed Tank combined with the Copper Area bund. Any further spillage overflows into the site drainage system and reports to the primary settling basin.

17.7        Copper-Molybdenum Separation

The molybdenum separation circuit consists of rougher flotation cells followed by a regrind circuit, five stages of cleaner flotation and one stage of cleaner scavenger flotation. All flotation cells, with the exception of the fifth cleaner cell (Jameson cell), are self-aspirated mechanical flotation cells fitted with covers. All self-aspirated cells are driven by a v-belt drive.

The Molybdenum Plant is shown in Figure 17-9.

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FIGURE 17-9: MOLYBDENUM PLANT LOOKING WEST (ROOF AND WALLS REMOVED)

Flotation cells selected for the molybdenum separation circuit are summarized as follows:

• Rougher – six 988 ft³ (28 m³) self-aspirated inert gas cells in a 2-2-2 configuration.
• Cleaner 1 – five 495 ft³ (14 m³) self-aspirated inert gas cells in a 1-2-2 configuration.
• Cleaner scavenger – four 495 ft³ (14 m³) self-aspirated inert-gas cells in a 2-2 configuration.
• Cleaner 2 – four 495 ft³ (14 m³) self-aspirated inert gas-cells in a 2-2 configuration.
• Cleaner 3 – five 300 ft³ (8.5 m³) self-aspirated inert-gas cells in a 2-3 configuration.
• Cleaner 4 – four 152 ft³ (4.3 m³) self-aspirated inert-gas cells in a 2-2 configuration.
• Cleaner 5 – single Jameson flotation cell.

The layout takes advantage of gravity used from Molybdenum Cleaner 5 through to Molybdenum Cleaner 2 with only the concentrate flows being pumped. Gravity transfer is used from Molybdenum Cleaner 1 to the Molybdenum Cleaner Scavenger and the rougher cells tails.

The molybdenum flotation circuit separates molybdenum and copper minerals as separate concentrates from a bulk Cu-Mo concentrate.

An OSA is used to monitor metal contents and solids concentrations in the feed, final concentrate, cleaner scavenger tailings and rougher tailings streams and allow operators to optimize reagent additions and flotation performance.

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17.7.1        Molybdenum Scrubbing System

The scrubbing system comprises of a Molybdenum Scrubber and uses caustic soda. The scrubbing system is mounted between the NaHS and plant diesel area and includes redundancy of all components for safety reasons. The scrubber removes gas from all the flotation cells, hopper and tanks via an elevated pipe network, converts gases to NaHS by using caustic soda addition (through intermediate bulk containers “IBCs”) and discharges clean air through the main stack.

17.8        Molybdenum Concentrate Thickening, Filtration and Drying

The molybdenum concentrate handling circuit consists of a small concentrate thickener, pressure filter, dryer and concentrate bag loading system. Molybdenum concentrate gravitates from the molybdenum plant OSA to the molybdenum concentrate thickener via a thickener feed box.

The concentrate thickener overflow reports to the tailings thickener. Molybdenum concentrate solids settle for collection at the underflow cone at a density of 60% w/w solids. The thickener underflow stream is pumped to an agitated filter feed tank by peristaltic pumps. A trash screen is located prior to the filter feed tank. The trash screen removes coarse oversize that may damage or block the filter.

High pressure air for the concentrate filter is supplied by a dedicated air compressor. High pressure air for drying is stored in a dedicated air receiver. Filter membrane pressing is supplied by the filter pressing water pump. Molybdenum filter cake is discharged from the filter and directed to a concentrate dryer via a screw feeder. The molybdenum concentrate dryer is a Holoflite dryer with a thermal oil heater and off-gas scrubber. The Holoflite drier reduces concentrate moisture to approximately 5% w/w.

Dried concentrate is stored in a bin ready to be bagged in the bagging station. The storage bin is sized to allow bagging of concentrate on dayshift only. Bagged concentrate is weighed and labelled prior to being loaded onto trucks by fork lift.

17.9        Copper Concentrate Dewatering and Storage

17.9.1        Copper Concentrate Dewatering

Concentrate thickeners were sized using benchmarked typical unit settling rates for copper concentrates with comparable size distribution and mineral composition. Thickener feed rates were based on maximum design copper head grades at 90,000 tpd. Copper-molybdenum and final copper concentrate thickeners with diameters of 79 feet (24 meters) were selected based on a unit settling rate of 2 ft²/tpd (0.2 tonne/m²/h).

Two pressure filters were selected for the copper concentrate filter duty. The nominated filters are expected to operate with sufficient design margin; however, if additional filter capacity is required, each filter can be expanded from 1,162 ft² (108 m²) to 1,550 ft² (144 m²) with the installation of additional plates.

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17.9.2        Copper Concentrate Storage and Load Out

Copper concentrate filter cake is discharged by gravity to a covered stockpile. A front-end loader (FEL) is used to maximize concentrate storage within the covered building. The covered building provides storage capacity for up to 5,000 tons at average production rates.

Copper concentrate is loaded into containers which are then covered and sealed. These containers sit on top of trucks licensed for use on the highway for transport from the mine site to a storage terminal.

An automated truck wash washes concentrate from the road trucks as they leave the concentrate storage building. Wash water and solids are recovered in a sump and pumped to the copper concentrate thickener.

Road trucks are weighed on a weighbridge located at the main security gate prior to leaving the mine site.

The Copper Concentrate area consists of the Concentrate Filter building and storage shed. Compressed air services are located to the west and the Electrical Rooms are located on the east side.

The Concentrate Filter Building, shown in Figure 17-10, houses filters and ancillary equipment such as control panels and mufflers.

FIGURE 17-10: COPPER CONCENTRATE FILTRATION, STORAGE AND LOAD OUT LOOKING SOUTH

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17.10        Tailings Thickening

The copper rougher tailings streams are combined with the copper cleaner scavenger tailings and the concentrate thickener overflow streams and gravitate to the tailings thickener feed distributor via a metallurgical cross-cut sampler. Tailings slurry is split into two uniform feed streams and directed to each thickener. Flocculant is added to the thickener feed streams to enhance settling.

The tailings thickening circuit consists of a tailings feed distribution box and two 213 feet (65 meters) diameter high compression thickeners to thicken flotation tailings to 65% w/w solids and recover process water to the process water tank.

FIGURE 17-11: TAILING THICKENERS LOOKING EAST

Tailings thickeners were selected as part of an overall tailings dewatering strategy that involved optimising underflow density to maximize downstream filtration rates, and were sized based on settling rate data derived from testwork conducted by FLSmidth, Outotec, Bilfinger, and Pocock Industrial. Optimum underflow density targets a narrow range below the point where the slurry yield stress inhibits handling by centrifugal pumps (without shear thinning systems).

17.11        Tailings Filtration Plant

The tailings filter area as shown in Figure 17-12 is located on the east side of the plant. The design consists of two identical parallel trains of filter feed tanks, and utilities such as compressed air and water.

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FIGURE 17-12: TAILINGS FILTER PLANT (ROOF AND WALLS REMOVED)

Thickened underflow is pumped from the tailings thickeners to two lines of agitated filter feed tanks. Filter feed tank design has a 4.5 -hour residence time.

The design allows for a total of 20 filters to be installed in order to process ore requirements over the life of mine. Nevertheless, as a risk mitigation strategy, additional space has been allowed for the installation of 4 additional filters, in the event that they may be required. Filter cake from the filters discharges to a single belt feeder via a set of ‘bomb-bay’ doors. The belt feeder operates continuously but at a low rate to deliver filter cake to the downstream conveyor continuously over the full cycle time of the filter.

The overland conveyor to the discharge point for the tailings stacking system is 2,171 feet long with 211-foot lift. The drive is installed at the head end, as is the gravity take-up system. The conveyor is driven by two 1200 HP drives mounted one on each side of the drive pulley.

The control room for this area is mounted on the upper deck central to the building on the west end. This bay includes a drive-through for trucks and other mobile equipment. The east end includes a drop down zone for filter components loading onto a truck.

17.11.1        Tailings Stacking

Dried tailings from each tailing transfer conveyor is discharged to a single overland tailings conveyor. This conveyor delivers tailings via a bifurcated chute to either the primary or secondary stacking system, which comprise the material handling system for the DSTF.

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FIGURE 17-13: TAILINGS SHIFTABLE CONVEYOR/MOBILE TRIPPER

The detailed design and selection philosophy for the stacking equipment considers the following drivers:

• Minimize stacker downtime for changeovers between lifts and sweeps.
• Maintain stacker progression rates within practical limitations.
• Maximize evaporative drying of the stacked tailings.
• Minimize stacker and conveyor costs.
• Equipment should be mechanically robust to minimize unplanned downtime.
• Stack material in a way which addresses geotechnical considerations, including equipment setback distances
• Permit constraints.

Dry Tailings Stacking Mobile Conveyors operate in series and transport tailings from the Mobile Tripper to the Extendable Mobile Dry Tailings Stacker. During operation, the number of these conveyors required is dependent on the final tailings deposition location relative to the position of the Shiftable Conveyor and Mobile Tripper.

17.12        Reagents and Consumables

The main reagent area is located at the west end of the copper flotation area, with the flocculant area being located next to the tailings thickeners. All of the reagents required for the molybdenum plant are handled inside the molybdenum building.

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FIGURE 17-14: REAGENTS AREA

Major process reagents and consumables are received and stored on site as either dry product or bulk liquids. Where required, dedicated mixing, storage and dosing facilities are provided for each reagent.

• Lime is used to increase slurry pH and subsequently depress minor pyrite in copper flotation.
• Collector (SIBX) - Sodium isobutyl xanthate (“SIBX”) is used as the collector in the bulk copper-molybdenum flotation circuit.
• Promoter-Cytec AP3894 promoter is added as a secondary collector in the bulk copper- molybdenum flotation circuit.
• Frother-Methyl isobutyl carbinol (“MIBC”) is used to provide a stable froth in the copper flotation cells.
• Diesel fuel oil is used as a collector/promoter to assist flotation of molybdenum minerals in the copper and molybdenum flotation circuits.
• Sodium hydrosulfide (“NaHS”) is used in the molybdenum flotation circuit as a depressant for non-molybdenum minerals.
• Sodium silicate can be used in the molybdenum flotation circuit as a dispersant for fine non-sulfide gangue and clay-bearing minerals.
• Flocculant is used as a settling aid in the concentrate and flotation tailings thickeners.
• Carbon dioxide is used to control pH to 9.0-10.0 in the molybdenum rougher and cleaner conditioning tanks.

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• Nitrogen is used in the molybdenum flotation cells to minimize oxygen entrainment and maximize depression of copper. A small pressure swing adsorption (“PSA”) plant supplies nitrogen to the molybdenum flotation circuit.
• Additional miscellaneous reagents are required in the plant; however, these are expected to be used in relatively small quantities.

17.13        Plant Services

17.13.1        Process Water Services

Fresh water is sourced from wells located on the western side of the Santa Rita Mountains and is pumped through a series of booster tanks and pumps to the fresh water tank located above the plant site.

• Fire water.
• Potable water including a water treatment plant.
• Gland seal water.
• Cooling water including chillers and closed loop cooling system.

Process water for ‘general use’ is sourced from the tailings thickener overflow (including concentrate thickener overflow and tailings filter filtrate water) and the fresh water tank as required. Supplementary water sources used for process water make-up include:

• Plant site run-off collected in the primary settling basin.
• Crusher stormwater pond.
• Mine pit sump and pit storm water pond.
• Mine ground water dewatering wells.

Process water is stored in the process water pond. Process water pond pumps transfer water from the storage pond to the process water tank. Excess water in the process water tank overflows back to the process water pond.

The tailings thickener overflow streams report directly to the process water tank for immediate distribution and use. Process water pumps distribute process water to the grinding mills, copper flotation, regrind circuits and lime slaking plant. Cloth wash water pumps distribute process water to the tailings filters for automated cloth washing and manifold flushing.

17.13.2        Air Services

Three separate plant air compressors provide air service throughout the plant. Due to its remote location, the primary crusher is serviced by a dedicated air compressor with an air dryer and filter system.

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17.14        Process Control Strategy

17.14.1        Process Plant PCS

The process control system (“PCS”) is an integrated plant-wide design, enabling the start-up, monitoring and control and shutdown of equipment from the plant control rooms.

The process plant is monitored and controlled from three separate control rooms:

• Crusher control room.
• Main control room (located in the flotation area).
• Tailings control room (located in the tailings filtration building).

Operators can control the plant via PC-based human machine interface (“HMI”) stations. Each HMI station provides dynamic graphical representation of the plant operation; equipment control functions; alarm displays; event logging; trending; data collection and reporting to assist in analysis of plant operations.

Where specific equipment forms part of an approved vendor package and drives are controlled from a vendor control panel, a communications interface is used to enable remote control and monitoring from the PCS. This includes digital and analogue signals for alarms; faults; instrumentation and monitoring; motor and valve control; process variables and interlock controls.

The crusher control room contains a single HMI station with two monitors. The HMI station provides dedicated control of the crushing plant area.

The main control room contains three HMI stations, each with two monitors. The main control room provides dedicated control of the main plant areas. Control of the crusher and tailings areas is also possible from the main control room. An engineering development workstation is also located in the main control room building.

The tailings control room contains one HMI station with two monitors. The HMI station provides dedicated control of the tailings filtration and dry stack areas.

17.14.2        On-Stream Analysers

Plant instrumentation includes OSA that are used to continuously monitor copper, molybdenum, iron and density in key process streams and assist with optimizing concentrate grade and recovery.

Dedicated OSA systems are provided in the copper and molybdenum flotation circuits. Each OSA unit is centrally located in the respective plant and elevated to allow gravity discharge of samples to sample return pumps. Each analyzer has two 6-channel multiplexers. Sub-samples for shift composites are collected automatically.

A single PSA is installed to continuously measure the particle size of the copper regrind cyclone overflow.

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17.14.3        Closed Circuit Television (CCTV) Systems

A closed-circuit television (“CCTV”) system is used to assist control room operators in monitoring the operation of plant and equipment.

The CCTV system provides real-time monitoring with archived recording for a nominal period. Camera types include fixed cameras and cameras with remote pan-tilt and/or zoom functions accessible by the control room operators.

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18        PROJECT INFRASTRUCTURE

This section addresses the infrastructure facilities that will support the Rosemont mine and processing facilities. The infrastructure facilities include the access roads into the plant site, source of electrical power and power distribution, source of fresh water and water distribution, DSTF, WRSA, transportation and shipping, communications, and mobile equipment.

This Technical Report includes refinements of certain aspects of the Project’s mine plan. While consistency with issued and pending environmental permits and analysis related thereto has always been a key requirement for this effort, updates to the original mine plan will be necessary. To the extent that any regulatory agency concludes that the current plan requires additional environmental analysis or modification of an existing permit, the intent will be to work with that agency to either complete the required process or to adjust the current mine plan as necessary.

18.1        Access Roads, Plant Roads and Haul Roads

Access and plant roads consist of an access road into the plant from State Highway 83, in-plant roads, haul roads and a security patrol road around the toe of the WRSA and DSTF. The plant and access roads are shown in Figure 18-1.

FIGURE 18-1: PLANT AND ACCESS ROAD

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The access road to the property starts at State Highway 83 at a point between mile markers 46 and 47 and ends at the main guard house at the entrance to the plant. The intersection of the access road with State Highway 83 will be modified to provide safe ingress and egress from the access road in compliance with ADOT and AASHTO standards. Modifications will include a northbound acceleration lane, northbound left turn lane and a southbound right turn lane.

In-plant roads extend from the plant entrance both through and around the perimeter of the process facilities. Secondary roads, such as the utility maintenance road, leave this perimeter road to serve the main substation, water storage tank, and access the utility corridor. As per the State of Arizona Air Quality Control Permit and the EIS analysis, specific in-plant roads will be paved to reduce dust emissions.

Haul roads used for access to and construction of perimeter waste rock buttresses shall be a minimum of 150 feet wide to allow trucks to turn around with the roadway surface.

A security patrol road will be provided around the toe of the WRSA and the DSTF along the security fence line for security to monitor the plant boundaries and provide maintenance access to the WRSA and DSTF.

18.2        Power Supply and Distribution

Pursuant to the Certificate of Environmental Compatibility (“CEC”) issued by the Arizona Corporation Commission (ACC) on June 12, 2012, TEP will provide the electrical power supply for the Rosemont mine and process facilities. The total connected load for the Rosemont mine and process facilities is estimated to be approximately 183 MVA and will require a transmission voltage of 138 kV.

The proposed Toro Switchyard, located approximately 3 miles south of Sahuarita Road and 3.5 miles east of I-19 near the Country Club Road and Corto Road alignments will tap into the existing 138kV transmission line that extends from the South Substation to the Green Valley Substation. The transmission line follows a 13.2 -mile-long route originating at the Toro Switchyard and terminating on Rosemont private property at the Rosemont Switchyard, Figure 18-2.

Distribution power tapping from the Rosemont Switchyard to the substation will provide power to the process plant and the mine.

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FIGURE 18-2: CEC APPROVED UTILITY CORRIDOR FOR 138KV TRANSMISSION LINE

18.3        Water Supply and Distribution

The fresh water design requirement for the Rosemont facilities is 3,500 gallons per minute (“gpm”) and peak of 5,000 gpm. The delivery requirements are based on the draft overall site water balance developed by Ausenco which takes into consideration dust control, process make-up water, process fresh water requirements, and potable water. The source of water supply identified for the Project is groundwater in the basin-fill deposits of the upper Santa Cruz basin, which lies west of the Project and the Santa Rita Mountains. The Project has a permit to withd draw groundwater for Mineral Extraction and Metallurgical Processing in the amount of 6,000 acre-feet per year for 20 years.

Rosemont Copper has acquired a 53-acre land parcel near Santa Rita and Davis Roads (Sanrita West), and a 20-acre parcel near Santa Rita Road and Country Club Drive (Sanrita South, or Station No. 1 site), for the purpose of constructing and operating a production well field for the Rosemont water supply.

The wells will deliver water to pump station no. 1 located at Sanrita South. There are three (3) other pump stations located strategically along the alignment of the water piipeline to pump the necessary water to the storage tank located at the mine site, Figure 18-3.

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FIGURE 18-3: UTILITY CORRIDOR FOR WATER LINE

The pipeline will discharge to the Rosemont fresh/fire water tank, which serves to provide storage and reserve for the operations. The lower portion of the tank, with an approximate capacity of 300,000 gallons, will be reserved for the fire water system. Flow of fire water and fresh water is provided by gravity.

Water will be provided to a potable water system, fresh water system, process water system, and fire water system.

The potable water system consists of a potable water treatment package, potable water tank and a distribution network delivering potable water by gravity to all ancillary buildings, process facilities, restrooms, and safety showers. The fresh water system consists of the gravity distribution network from the fresh water storage tank to the process facilities requiring fresh water. The fresh water usage is for gland water pump seals, fresh water make-up to the mills, flotation plant make-up, and reagent make-up. The process water system consists of a process water pond that collects process water from the concentrate and tailings de-watering equipment for recycling back into the circuit. The fire water system consists of a gravity distribution network from the fresh water / fire water storage tank to a system of hydrants around the ancillary buildings and process facilities.

Rosemont has voluntarily committed to recharging 105% of the groundwater used during operations. Thus far, 45,000 acre-feet of water resulting in nearly 42,600 acre-feet of storage credits have been recharged back into the Tucson Active Management Area, the area of planned withdrawal. Additionally, in an effort to reduce water usage, Rosemont is committed to use dry stack tailings instead of conventional tailings. The tailings dewatering system is expected to recycle about 15,000,000 gallons of water per day.

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18.4        Tailings Management

The Rosemont DSTF has been designed to receive dewatered tailings from the processing plant at a nominal rate of 90,000 dry tpd. This material will be stacked behind large containment buttresses constructed from pit run waste rock.

The deposition of dewatered tailings, waste rock and overburden will be initiated with a series of perimeter buttresses and berms. The staging of these buttresses will also allow reclamation to begin early in the operation. Soil will be salvaged from pit and WRSA and DSTF for use as a vegetation growth medium. The dewatered tailings deposition will incorporate staged waste rock buttresses for visual screening and to improve mechanical and erosional stability of the tailings.

18.4.1        DSTF Location and Design

Design criteria and objectives for the original dry stacked tailings facility included:

• Provision of secure long-term storage of a minimum 500 million tons (Mt), which is sufficient for the ore to be mined and processed during the Project life;
• Location within the immediate general area of the mine (approximately five-mile radius from the proposed mine pit);
• Prevention of airborne release of tailings solids to the environment by provision of dust suppression measures;
• Compliance with all applicable regulations including Arizona Best Available Demonstrated Control Technology (“BADCT”) standards;
• Creation of a site-specific design that accounts for local factors including climate, geology, hydrogeology, seismicity and vegetation; and
• Establishment of an effective and efficient reclamation program, with a focus on concurrent reclamation.

Advantages of the dry stacked tailings over a conventional tailings impoundment is that it eliminates the need for an engineered embankment and seepage containment system, maximizes water conservation and minimizes water makeup requirements, results in a very compact site limiting disturbance to a single drainage, and allows opportunities for concurrent reclamation and provisions for dust control.

The selected site is located just east of the proposed mill site in Barrel Canyon. The DSTF site is characterized by terrain sloping generally east from the plant area to the Barrel Canyon, which generally runs north south in this area.

The design was developed based on hydrological and geotechnical studies that included review of regional climate data, drilling and testing programs, and laboratory characterization of subsurface and tailings samples.

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18.4.2        DSTF Stability

The DSTF is designed as a low hazard facility with waste rock placed as buttressing material. The filtered tailings lifts will densify under successive controlled conveyor lift placement and will result in an increase in the lower lift fill strength over time. The DSTF stability analyses considered the maximum ultimate height at the maximum section through the facility for downstream and upstream stability.

The tailings will be placed in a dewatered state for acceptable handleability during conveyance and trafficability of the tailing surface, which will limit susceptibility to liquefaction under dynamic loading. However, limited higher moisture zones within the tailings mass created by meteoric water may occur. This condition was considered in the stability modelling by applying reduced shear strength to thin layers within the tailings mass at various levels to simulate these higher moisture zones and to evaluate the subsequent earthquake resistance of the facility.

Thus, adequate factors of safety for static and pseudo-static were obtained from the stability analyses based on the selected parameters and proposed facility. The use of dry stack tailings as oppose to conventional tailing impoundments eliminates the danger of dam failure typically seen with tailings ponds.

The slope stability analyses performed on the outer slope indicate the dry stack tailings operations can be constructed with stable 3H:1V inter-bench slopes and an overall stable slope of approximately 3.5H:1V.

18.4.3        Hydrologic Modeling

Modelling calculations performed for the EIS indicate that infiltration of rain into the DSTF did not develop. The resistance to infiltration is a function of the fine-grained nature of the crushed and ground tailings material and the compaction that will occur with placement and facility construction.

Much of the DSTF will ultimately lie above the ultimate groundwater capture zone predicted by the groundwater models. Within this zone, any seepage that may occur would ultimately flow via groundwater to the open pit. The portions of the facility not included in this capture zone will generate seepage and is permitted under the Aquifer Protection Permit program in the state of Arizona. Water entrained in the DSTF that comprises drain-down has been chemically analyzed and modeled and is not expected to exceed AWQS at compliance locations, located around the perimeter of the facilities. The ADEQ evaluated the potential for seepage and issued permits based on their analysis. Modelling and analytical results indicate that seepage constituent concentrations will be below the AWQS for regulated constituents. This modeling is supported by 174 geochemical samples of waste rock evaluated during the EIS process as well as ten tailings samples specifically reviewed by the Arizona Department of Environmental Quality during the Aquifer Protection Permit process. Samples were subject to combinations of testing to determine their acid generating potential, whole rock analysis, synthetic precipitation leach procedure testing, meteoric water mobility testing, and humidity cell testwork. The tests performed on the waste rock showed a net neutralization potential of 225. This is considered highly buffering (zero or less being acid generating) ensuring that the buttress materials will not generate acid rock drainage (ARD) The tailings testing resulted in analysis of seepage that supported the ADEQ determination.

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Later work continues to support this finding and is consistent with prior testwork. Mineralogical analysis of 107 drill core composite samples (Section 7.6) indicating that Rosemont ore will contain less than 3% by weight sulfide (potentially acid generating) minerals, most of which will be recovered as valuable concentrates in the process. The analysis also indicates that the ore will contain approximately 20% carbonates (calcite, dolomite) which are alkaline and will serve to neutralize any acidic species that could be generated by decomposition residual sulfide minerals in the process plant tailings. Pit-run waste rock will consist largely of limestone and skarn rock types, with some andesite, quartz monzonite porphyry, and arkose. The presence of substantial quantities of limestone and skarn (97%) along with low-sulfide content, supports the analysis in the EIS that determined there is a large buffering capacity within the buttress materials which will minimize the potential generation of ARD.

18.4.4        Surface Water Control

Once the perimeter buttresses/berms are placed across the drainages and washes, stormwater run-on will be limited by ponding stormwater upstream of the dry stack areas. Stormwater runoff sediments from the waste rock buttresses will be captured in sediment basins located downstream of the tailings facility. During operations, the tailings surface will be sloped away from the waste rock buttresses to limit potential water impoundment against the buttresses. Perimeter ditches will be constructed at the upgradient outer edges of the tailings surface to retain and evaporate water.

18.4.5        DSTF Operations

Dewatered tailings will be delivered by conveyor and placed with a radial stacker. A dozer might be used to spread the dry tailings to provide a suitable surface for the conveyor and stacker as needed.

An initial starter buttress will be constructed with waste rock. Concurrent tailings placement and buttress construction using waste rock placement will occur throughout the life of the tailings facility. Waste rock will be advanced ahead of the tailings level in successive lifts using the upstream construction method. The waste rock buttresses will accommodate haul traffic and outer slopes generally of 3H:1V with benches to achieve an overall sloped facility of 3.5H:1V.

18.5        Communications

There are requirements for accounting, purchasing, maintenance, and general office business as well as specialized requirements for control systems.

The two most common options are to design separate data networking and telecommunication systems or to integrate the two into a common infrastructure. For this Project, the proposed approach is integration.

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A voice over I/P (VoIP) phone system will be a part of the office network and VoIP handsets will be used for voice communication.

The office ethernet network will support accounting, payroll, maintenance, and other servers as well as individual user computers. High bandwidth routers and switches will be used to logically segment the system and to provide the ability to monitor and control traffic over the network.

The control system ethernet network will support the screen, historian, and alarm servers and connect to the Control Room computers as well as the Programmable Logic Controllers and other control systems provided with ethernet communication capabilities. This system will incorporate redundancy and will be designed to minimize traffic and latencies. No phone or user computer will be connected to this system.

A security system will also be incorporated into the plant network. Using a dedicated video server and monitors, I/P cameras utilizing power over ethernet connections will be plugged into dedicated switches. Security cameras are typically located in storerooms, parking lots, visitor lobbies, warehouses, and areas where sensitive materials are kept.

Mobile radios will also be used by the mine and plant operation personnel for daily control and communications while outside the offices.

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19        MARKET STUDIES AND CONTRACTS

Hudbay has a marketing division that is responsible for establishing and maintaining all marketing and sales administrations of concentrates and metals. The Project’s copper concentrates are expected to be a clean, high grade concentrate containing small gold and silver by-product credits which will be suitable as a feedstock for smelters globally. Approximately 50% of the copper concentrate production has been contracted under long term sales contracts.

Table 19-1 below summarizes the key assumptions for the sale of Rosemont’s copper concentrate.

TABLE 19-1: COPPER CONCENTRATE

	Units	LOM Total / Average
Copper Concentrate Base Treatment Charge	$ / dry short ton con	$73
Copper Refining Charge	$ / lb Cu	$0.08
Silver Refining Charge	$ / oz Ag	$0.50
Copper Concentrate Transport & Freight	$ / wet short ton con	$127
LOM Copper Grade in Copper Concentrate	% Total Cu	34.3%
Moisture Content of Copper Concentrate	% H2O	8.0%

No deleterious elements are expected to be produced in quantities which would result in material selling penalties.

A precious metals stream agreement with Silver Wheaton Corporation for 100% of payable gold and silver from the Project was entered into on February 11, 2010. Under the agreement, Hudbay will receive payments equal to the lesser of the market price and $450 per ounce for gold and $3.90 per ounce for silver, subject to 1% annual escalation after three years.

Rosemont is expected to produce a marketable 45% molybdenum concentrate. Table 19-2 summarizes the key assumptions for the sale of Rosemont’s molybdenum concentrate.

TABLE 19-2: MOLYBDENUM CONCENTRATE

	Units	LOM Total / Average
Molybdenum Concentrate Base Treatment Charge	$ / lb Mo	$1.50
Molybdenum Concentrate Transport & Freight	$ / wet short ton con	$124
LOM Molybdenum Grade in Molybdenum Concentrate	% Mo	45.0%
Moisture Content of Molybdenum Concentrate	% H2O	8.0%

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20        ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL 
            OR COMMUNITY IMPACT

The Project permitting status is well advanced and continues to progress since July 2007. The final approvals required include the Final Record of Decision (“ROD”) from the U.S. Forest Service (“USFS”) and the 404 Permit from the U.S. Army Corps of Engineers (“USACE”). There have been over 450 days of public comment associated with this Project that have culminated in over 43,500 comments. All comments have been reviewed, categorized, and either incorporated or answered by the various State and Federal agencies.

Since issuing the Final Environmental Impact Statement (“FEIS”), the USFS has issued two Supplemental Information Reports (“SIR”) and a Supplemental Biological Assessment (“SBA”), both of which were required after the Draft ROD was issued in December 2013. The SIRs considered whether new information or changed circumstances remained within the scope of the effects disclosed in the FEIS. The second SIR was issued to summarize on-going analysis and reporting completed since the first SIR was produced in 2015. Nothing disclosed to date would indicate that the information in either SIR falls outside the effects considered in the FEIS, which would require the FEIS to be supplemented.

The SBA evaluated the sighting of a jaguar and ocelot and resulted in the reinitiation of consultation with the U.S. Fish and Wildlife Service (“USFWS”). This consultation process included species such as the ocelot, the yellow-billed cuckoo, the Mexican garter snake, and other aquatic species. The reinitiated consultation was completed in April 2016 and culminated in an Amended Final Reinitiated Biological and Conference Opinion (“BO”) similar to the one produced in 2013. Both BO’s stated that the Project would not jeopardize the existence of any endangered species. The next step will be to issue a Final ROD. The Final ROD will trigger a requirement for an updated operating plan ( MPO) that will include measures to be taken so the Project will meet the requirements of the ROD, including measures to mitigate adverse environmental impacts.

The USACE Division Offices are evaluating the 404 permit application, the record, and the mitigation package that will go into making their permit decision. The mitigation package is designed to mitigate impacts to a total of 68.8 acres, which consists of 40.4 acres of ephemeral channels on the Project site plus the 28.4 acres of off-site indirect impacts. This mitigation incorporates the restoration of a floodplain that was impacted by agriculture; mitigation for two sites impacts by grazing, poor roadway maintenance, and other activities; as well as preservation of sites near to the Project site. Once the USACE evaluation is complete, and a positive permit decision is made, the terms and conditions of the permit and appropriate financial assurance will be negotiated.

At this time, State of Arizona environmental permits and approvals have been issued for the Project, and these permits remain in force and are current. Two of the permits (air and groundwater protection) will need to be amended to match the applicable Federal permits. The Project continues to comply with these current State permit terms and conditions.

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As stated above, the Project details included in this document were specifically designed and evaluated to fall within the permit constraints included in the EIS and State of Arizona permits with amendments. This Technical Report includes refinements of certain aspects of the Project’s mine plan. While consistency with issued and pending environmental permits and analysis related thereto has always been a key requirement for this effort, updates to the original mine plan will be necessary. To the extent that any regulatory agency concludes that the current plan requires additional environmental analysis or modification of an existing permit, the intent will be to work with that agency to either complete the required process or to adjust the current mine plan as necessary.

Certain permits issued for the site have specific design and monitoring requirements built into the permits. In particular, the Project meets (and in some design elements exceeds) the Arizona Department of Environmental Quality (ADEQ) Best Available Demonstrate Control Technology (BADCT) requirements. BADCT covers specific requirements for any “discharging facility” and includes items such as specific liner requirements for ponds, design requirements for WRSA and DSTF including seismic design requirements as well as geochemical characterization requirements for possible discharges. In addition, the aquifer protection permit (APP) issued by ADEQ has specific groundwater monitoring requirements that requires quarterly monitoring for various parameters in specific wells.

The USFS has incorporated the permit requirements required by ADEQ, as well as other agencies, into their Mitigation Measures listed in the FEIS (Appendix B of the FEIS). FEIS mitigation measures also cover mitigation measures requirements to cover areas of interest/concern to the USFS. These requirements will be incorporated into the Final ROD and into the updated MPO. Mitigation Measures in the FEIS include categories such as:

• Geology, Minerals, and Paleontology
• Soils and Revegetation
• Groundwater Quality and Quantity
• Surface Water Quantity and Quality
• Seeps, Springs, and Riparian
• Geology, Minerals, and Paleontology
• Biological Resources
• Landownership and Boundary Management
• Dark Skies
• Visual Resources
• Recreation and Wilderness
• Hazardous Materials
• Transportation/Access
• Noise
• Public Health and Safety
• Cultural Resources
• Air Quality and Climate Change
• Fire and Fuels

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• Power Use
• Community Programs

Certain permits issued require financial assurance to ensure the success of mitigation, while others are solely to ensure that adequate funds are available at closure. The requisite bonds for the Project are expected to be obtained from the surety market with an estimated annual bond fee of 2% of the bond’s notional value.

Currently the Project has bonds in place to cover $40,000 in surface reclamation costs for the Arizona State Mine Inspector (AMSI), and $4,300,000 for the ADEQ APP permit specifically for closure of discharging facilities. It is expected that the bond for ADEQ will be reduced once the permit amendment is completed. Additionally, the reclamation bond with ASMI is expected to be incorporated into the USFS bonding.

In addition, a USFS bond will be required to cover reclamation and closure costs and will be updated in 3-year increments throughout the life of the Project. The required bonding for the initial 3-year period will be negotiated with the USFS during their review of the MPO. Hudbay has estimated and included fees for a $65 million USFS bond through construction and operations with some curtailment in the final four years of operations as certain reclamation activities are completed. Although the final amount of the USFS bond remains to be negotiated, the notional bond value is not expected to differ materially from this estimate.

The USACE bond will be a performance bond with a long-term management component to ensure the mitigation proposed is successful. The overall USACE bonding has been estimated at approximately $50 million; however, this bond package will need to be negotiated and the estimate includes items that may be eliminated in the final negotiation. Initial bond amounts have been estimated at $35 million with $15 million long-term bonding added during the first ten years of operation. The USACE bond is not expected to be required after the tenth year of operations.

It is also expected that the bond for ADEQ will be reduced once the permit amendment is completed. Additionally, the reclamation bond with ASMI, currently at $40,000, is expected to increase to $4 million once mine construction commences. However, Hudbay expects the requirements related to this bond to be covered by the USFS bonding pending future negotiation with ASMI and USFS. At this time, Hudbay has assumed that a $4M ASMI bond will stay in place during the construction and operation of the mine.

With regard to community outreach and other social commitments, the following provides a summary of costs associated with those items:

• $650,000 for the relocation of the Arizona Trail;
• $6,500,000 to $7,500,000 to repair the pavement, install bus pull outs, and replace guard rails on a 12-mile stretch of State Route 83, the main roadway connecting I-10 to the entrance road into the Project;

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• $25 million endowment fund for conservation purposes. The development of this endowment trust will fund priority community projects including recreation, cultural, and environmental conservation projects. Currently proposals under consideration by the Forest Service set annual funding of up to $1.5 million depending on annual production and metal prices and is subject to a cap of $25 million. By agreement, 10% of this endowment is available to the Mescalero Apache tribe for grants for specific cultural and educational projects.
• $500,000 in annual community donations.

In addition, there are specific mitigation and data recovery obligations related to archaeological (cultural) sites associated with the Project. These specific requirements are a culmination of negotiations between the USFS and the Tribes with input from various state agencies, other cooperators, and Hudbay.

Details on permit status and authorizations for current project activities are included in Appendix A3-1.

20.1        Reclamation and Closure Plan

The Reclamation and Closure Plan is based on several key components, referred herein as initiatives. These initiatives provide the physical and philosophical foundation that will remain constant throughout the operation of the facility. As related to this Plan, some of these initiatives include:

• Beginning with the end in mind. The placement of materials in the various storage areas is based on the final closure configuration. For example, the overall out-slopes of the Landform will generally be 3.5H:1V (H: Horizontal, V: Vertical) with inter-bench slopes at 3.0H:1V. The placement of waste rock that comprises the outer shell of the Landform, i.e., the outer slopes of the waste rock buttresses/berms, will incorporate setbacks to facilitate efficient regrading to achieve the final design slopes.
• Constructing outer facility berms/buttresses. During the initial years of mine development, waste rock from stripping operations will be preferentially placed along portions of the outer footprint of the WRSA (berms) and the DSTF (buttresses). In addition to defining the outer footprint, placement of these outer berms/buttresses will be used to help screen active mining operations from vantage points in the area, such as from SR 83, and will serve to isolate mining activities and protect down-gradient surface water quality. Regarding the DSTF area, an outer waste rock buttress will serve to stabilize the dry tailings stacks and prevent the erosion of tailings into the down-gradient drainages.

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• Concurrently reclaiming the outer surfaces of the WRSA and DSTF. Reclamation of the Landform will not be deferred to the end of the Project. Concurrent reclamation, as practicable, is planned for the outer shell encompassing the WRSA and DSTF. Growth media (soil) and woody debris will be salvaged from the facility footprints and placed directly on reclaimed areas or stockpiled for future use. Reclamation of the outer shell of the Landform also includes the construction of stormwater management features, such as channels and drop structures, in order to divert as much stormwater down-gradient as practicable. The final reclaimed surfaces will have a soil (growth media) cover or a combination of rock/soil cover and will be revegetated using a seed mixture appropriate for the Project area.
• Using modern technology to minimize the generation of impacted water. The Rosemont operation will include milling operations for sulfide ores. Conventional slurry line and settling pond technology for tailings disposal will not be used at Rosemont. Tailings will instead be mechanically filtered to 18% (or less) moisture content by weight and stacked behind a waste rock buttress. The formation and migration of seepage from this tailings disposal system is negligible. The dry stack tailings disposal method also facilitates the ability to concurrently reclaim the facility as described above.

To the maximum extent practicable, the final landform is graded to route as much stormwater runoff off the reclaimed surface and into the down-gradient flow system. Bench channels and drop chutes are constructed on the surface to direct stormwater down-gradient toward lower Barrel Canyon drainage. Building facilities within the Plant Site are removed and the area regraded. The Plant Site area will be also regraded with the intent to route as much stormwater down-gradient as practicable – in this case to the McCleary Canyon drainage which feeds the lower Barrel Canyon drainage. Reclaimed areas are covered with growth media (soil salvaged from the facility footprints) and revegetated.

The reclamation and closure of the Utility Corridor includes the removal of facilities (such as the water and power lines and pump stations) and the regrading and revegetation of disturbed areas.

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21        CAPITAL AND OPERATING COSTS

21.1        Introduction

Capital costs are estimated in constant 2016 US dollars.

21.2        Capital Costs

The total initial capital required to construct the processing plant, purchase mining equipment and pre-strip the pit is estimated to be $1,921 million including 15% contingency on all items as shown in Table 21-1.

TABLE 21-1: INITIAL CAPITAL COST SUMMARY

Initial Capital	000 US $
Site Wide	42,433
Mining	474,070
Process Plant	670,525
Site Services and Utilities	21,802
Internal Infrastructure	127,300
External Infrastructure	113,954
Common Construction Facilities	50,914
EPCM Services	107,009
Owner’s Cost	312,895
Total Initial Capital	1,920,903

The initial capital investment represents the total project cost; including facility costs, infrastructure costs and Owner’s Costs. The estimate is based on inputs from various organizations as follows:

• Ausenco: Process Plant
• Ausenco and M3: Plant Infrastructure
• Ausenco and M3: DSTF
• Knight Piésold: Engineering quantities relating to Heavy Civil Works (“HCW”): geotechnical, haul roads, buttresses, WRSA and diversion channels
• Stantec: Water Infrastructure
• Tucson Electric Power: Power Infrastructure
• Hudbay: Owner’s Costs

The estimate was produced using Prism Project Estimator based on a bottom-up approach. The overall capital cost estimate meets the Association for Advancement of Cost Engineers (“AACE”) Class 3 requirement of an accuracy range between -10% and +20% of the final project cost (excluding contingency). It has a base date of end of March 2016 with no allowance for escalation. This assessment is based on:

• Project maturity 10 – 40%
• Engineering progress on the processing plant was approximately 38% complete

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     •    Engineering progress on the off-site infrastructure was approximately 50% complete

Process plant costs were estimated by Ausenco with input from various consulting firms including Knight Piésold. Construction labor rates and productivities were developed on a discipline by discipline basis with input from major industrial contractors in the Southwest U.S. Labor rates, supply rates and productivities were benchmarked against projects of similar size and scope.

External infrastructure for the water supply system was designed and estimated by Stantec. External infrastructure for the power supply was designed and estimated by costs were provided by Tucson Electric Power.

Direct cost estimate quantities were derived from engineering lists, material take-offs, consultant databases (previous projects) and vendor input as shown in the table below:

Description	WBS	Earthworks	Concrete	Structural	Platework	Piping	E & I
Geology and Mine Design /Mining (infrastructure only)	1000+2000	Preliminary Design	Preliminary Design	Preliminary Design	Preliminary Design	Factor	Factor
Process Plant	3000	Preliminary Design	Preliminary Design	Preliminary Design	Preliminary Design	Preliminary Design and Previous Project	Preliminary Design
Primary Crusher	3110
Stockpile Feed Conveyor	3120			Vendor	Vendor
Stockpile and Reclaim	3210			Preliminary Design	Preliminary Design
Grinding	3220
Pebble Crushing	3230			Previous Project
Pebble Crushing Conveying	3240			Vendor	Vendor
Copper Flotation	3250			Preliminary Design	Preliminary Design
Copper Regrind	3260
Copper Concentrate Thickening	3270
Copper Concentrate Filtration and Loadout	3280
Tailings Thickening	3290
Molybdenum Plant	3300		Previous Project	Preliminary Design and Previous Project	Previous Project
			Preliminary Design	Preliminary Design	Preliminary Design
Reagents	3400
Plant Services	3500		Previous Project
			Preliminary Design
Tailings Filter Plant	3600
Site Services and Utilities	4000		Preliminary Design and Previous Project		Previous Project
Site Stormwater Ponds	4100		Preliminary		N/A

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Description	WBS	Earthworks	Concrete	Structural	Platework	Piping	E & I
Plant Fuel Storage and Distribution	4200		Design	N/A
Sewerage and Waste Management	4300		Factor
Communications and IT	4400		N/A			N/A	N/A
Site Infrastructure	5000		Preliminary Design	Preliminary Design			Preliminary Design
Tailings Management	5500		N/A	Vendor	Vendor	Vendor	Vendor
External Infrastructure	6000		Preliminary Design	Preliminary Design	N/A	Preliminary Design	Preliminary Design

Project indirect costs were developed based on construction sequencing plans and quotes for EPCM service providers. Owner’s Costs were developed by Rosemont Copper based on quoted equipment costs, third-party technical experts and in-house estimates.

A contingency component of $251M or 15% is included in the initial capital cost. No contingency is included in sustaining capital. All sunk costs are excluded from the capital cost estimates.

The estimate was reviewed against Constancia’s actual costs by the Hudbay project team. An independent review of the capital cost estimate was performed by M3 Engineering and Technology; who has experience with mining projects in the US and benchmarked the estimate against other local projects of similar size.

The LOM sustaining capital costs are estimated to be $387 million excluding capitalized stripping and $1,168 million including capitalized stripping. Sustaining capital costs associated with mining include new mine equipment purchases and major rebuilds, ongoing haul road construction and expansions to the truck shop and heavy truck fuel facility. Sustaining capital for the processing plant includes three tailings stacking expansions, upgrades to the regrind mill, and the installation of a SHMP mixing facility. Sustaining capital also includes the installation of buttress drop structures.

TABLE 21-2: SUSTAINING CAPITAL

Sustaining Capital	000 $
Mining – Equipment and Rebuilds	138,898
Mining – Equipment Major Repair	170,123
Mining – Heavy Civil Works	19,540
Mining – Facilities	6,445
Plant – Tailing Stacking Expansions	23,855
Plant – Upgrades	3,417
Plant – Ramp-up Support	4,970
Buttress Drop Structures	12,515
Light Vehicles & Misc	7,100
Total Sustaining Capital
(Excluding Capitalized Stripping)	386,865
Capitalized Stripping	780,897
Total Sustaining Capital Expenditures (Including Capitalized Stripping)	1,167,762

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21.3        Operating Costs

Operating costs were developed by Hudbay based on a bottom-up approach and utilizing budget quotes from local suppliers, Arizona operations experience, and labor costs within the region. Site visits were conducted to other facilities currently utilizing dry stack technology to better understand the operations and maintenance requirements. Mining operating costs were validated against actual costs at Constancia.

The total LOM operating costs, including off-site costs (transport, TCRCs, etc.) are estimated at $12.34/ton milled (before deducting capitalized stripping) and $11.02/ton milled (after deducting capitalized stripping) as shown in Table 21-3. The mining costs below do not include pre-stripping costs as these are included in the development capital cost.

TABLE 21-3: OPERATING COST SUMMARY

	Before Deducting Capitalized Stripping		After Deducting Capitalized Stripping
	Unit Cost ($/ton milled)	Annual Average 19 Year LOM (000 $)	Unit Cost ($/ton milled)	Annual Average 19 Year LOM (000 $)
Mining ($/ton moved)	$1.08	$101,997	$0.64	$60,897
Mining	$3.27	$101,997	$1.95	$60,897
Processing	$4.71	$146,723	$4.71	$146,723
G&A, Other	$1.26	$39,245	$1.26	$39,245
Total Mine / Mill / G&A	$9.24	$287,965	$7.92	$246,865
Off-Site Costs	$3.10	$96,993	$3.10	$96,993
Total	$12.34	$384,958	$11.02	$343,858

Reclamation costs of approximately $25M (which are incurred over the 19 year LOM) and closure costs of approximately $9M are included in operating G&A costs.

The total C1 cash costs and sustaining cash costs (net of by-product credits at stream prices) over the LOM and over the first 10 years are shown in Table 21-4. C1 cash costs include mining, milling, G&A and offsite costs. Sustaining cash costs include C1 costs plus royalties and sustaining capital.

TABLE 21-4: CASH COSTS (NET OF BY-PRODUCT CREDITS AT STREAM PRICES)

Cash Costs (Net of By-Product Credits at Stream Prices)	Units	Before Deducting Capitalized Stripping	After Deducting Capitalized Stripping
Years 1-10 Average C1 Cash Costs	$ / lb Cu in con	$1.40	$1.14
Years 1-10 Average C1 Cash Costs + Royalties + Sustaining Capex	$ / lb Cu in con	$1.59	$1.59
LOM C1 Cash Costs	$ / lb Cu in con	$1.47	$1.29
LOM C1 Cash Costs + Royalties + Sustaining Capex	$ / lb Cu in con	$1.65	$1.65

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21.4        Working Capital Costs

Working capital for accounts receivable and accounts payable will vary over the mine life based on revenue, operating costs and capital costs. The working capital estimate is based on 33% of revenue as accounts receivable and 33% of cost of goods sold and capital costs as accounts payable in a given quarter. This is equivalent to an approximate 30-day delay in converting revenue to cash and an approximate 30-day delay in paying cash for capital and operating costs. All of the working capital is assumed to be recaptured by the end of the mine life and the closing value of the account is zero.

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22      ECONOMIC ANALYSIS

22.1        Key Model Assumptions

All figures in the economic analysis are shown on an unlevered 100% Project basis (except where indicated in Section 22.5) and include the impact of the existing precious metals stream with Silver Wheaton.

22.1.1       Metal Prices

The economic viability of the Project has been evaluated using the metal prices outlined in Table 22-1. The metal prices used in the economic analysis are based on a blend of consensus metal price forecasts from over 30 well known financial institutions and Wood Mackenzie.

TABLE 22-1: METAL PRICE ASSUMPTIONS

Metal	Unit	Price
Spot Copper	$/lb	$3.00
Spot Molybdenum	$/lb	$11.00
Spot Silver	$/oz	$18.00
Streamed Silver1	$/oz	$3.90

1. Subject to a 1% annual inflation adjustment

The terms of the existing precious metals streaming agreement with Silver Wheaton were included in the analysis. Silver Wheaton will make upfront cash payments totaling $230 million to fund initial development capital in exchange for 100% of the silver and gold production from Rosemont. Silver Wheaton will make ongoing payments of $3.90 per ounce of silver and $450 per ounce of gold subject to a 1% inflation adjustment starting on the third anniversary of production.

Although gold is not part of the current reserve estimate, metallurgical testing has demonstrated economic concentrations of gold in copper concentrate as outlined in Section 13. Over the LOM, approximately 309 thousand ounces of gold are expected to be recovered in copper concentrate (although the financial impact has not been included).

At the effective realized prices including the impact of the stream, the revenue breakdown at Rosemont is approximately 92% copper, 6% molybdenum, and 2% silver.

22.1.2        Life of Mine Model Summary

The key mine and mill operating assumptions used in the cash flow model are outlined in Table 22-2.                                                                                                                                                     

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TABLE 22-2: MINE AND MILL OPERATING ASSUMPTIONS USED IN THE FINANCIAL MODEL (100% PROJECT BASIS)

	Units	LOM Total / Average
Mine Plan
Ore Mined Waste Mined (Excluding Pre-Strip)1	M short ton M short ton	592 1,155
Strip Ratio (Excluding Pre-Strip)1 Total Copper Grade Milled	Waste/Ore % TCu	2.0 0.45%
Sulfide Copper Grade Milled	% Sulfide Cu	0.40%
Molybdenum Grade Milled	% Mo	0.012%
Silver Grade Milled	oz/short ton Ag	0.133
LOM Metallurgical Recoveries
Sulfide Copper Recovery	% Sulfide Cu	90.0%
Effective Total Copper Recovery Molybdenum Recovery	% TCu % Mo	80.4% 53.4%
Silver Recovery	% Ag	74.4%
LOM Concentrate Specifications
Copper Grade in Copper Concentrate Molybdenum Grade in Molybdenum Concentrate	% TCu % Mo	34.3% 45.0%
Moisture Content of Copper & Molybdenum Concentrates	% H2O	8.0%
Production & Mill Throughput
Design Mill Throughput	k short ton / day	90
Mine Life (Including Processed Stockpiles) Years 1-10 Average Annual Copper Production	Years k short ton Cu in con	19 140
LOM Average Annual Copper Production LOM Average Annual Molybdenum Production	k short ton Cu in con k short ton Mo in con	112 2.0
LOM Average Annual Silver Production	k oz Ag in con	3,095

1. Pre-stripping costs included as part of the development capital costs

Key capital and operating cost assumptions used in the cash flow model are outlined in Table 22-3.

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TABLE 22-3: CAPITAL AND OPERATING COST ASSUMPTIONS USED IN THE FINANCIAL MODEL (100% PROJECT BASIS)

	Units	LOM Total / Average
Capital Costs
Development Capital (Including Pre-Strip) 1	$M	$1,921
Development Capital Less Upfront Stream Proceeds	$M	$1,691
LOM Sustaining Capital (Excluding Capitalized Stripping)	$M	$387
Capitalized Stripping	$M	$781
LOM Sustaining Capital (Including Capitalized Stripping)	$M	$1,168
Onsite Operating Costs (Before Deducting Capitalized Stripping)
Mining (Before Deducting of Capitalized Stripping)	$ / short ton moved	$1.08
Mining (Before Deducting of Capitalized Stripping)	$ / short ton milled	$3.27
Milling	$ / short ton milled	$4.71
On-Site G&A2	$ / short ton milled	$1.26
Total Onsite Operating Costs	$ / short ton milled	$9.24
Onsite Operating Costs (After Deducting Capitalized Stripping)
Mining (After Deducting Capitalized Stripping)	$ / short ton moved	$0.64
Mining (After Deducting Capitalized Stripping)	$ / short ton milled	$1.95
Milling	$ / short ton milled	$4.71
On-Site G&A2	$ / short ton milled	$1.26
Total Onsite Operating Costs	$ / short ton milled	$7.92
Offsite Operating Costs
Copper Concentrate Base Treatment Charge	$ / dry short ton con	$73
Copper Refining Charge	$ / lb Cu	$0.08
Silver Refining Charge	$ / oz Ag	$0.50
Copper Concentrate Transport & Freight	$ / wet short ton con	$127
Molybdenum Concentrate Base Treatment Charge	$ / lb Mo	$1.50
Molybdenum Concentrate Transport & Freight	$ / wet short ton con	$124
Cash Costs (Net of By-Product Credits at Stream Prices) Before Deducting Capitalized Stripping
Years 1-10 Average C1 Cash Costs	$ / lb Cu in con	$1.40
Years 1-10 Average C1 Cash Costs + Royalties + Sustaining Capex	$ / lb Cu in con	$1.59
LOM C1 Cash Costs	$ / lb Cu in con	$1.47
LOM C1 Cash Costs + Royalties + Sustaining Capex	$ / lb Cu in con	$1.65
Cash Costs (Net of By-Product Credits at Stream Prices) After Deducting Capitalized Stripping
Years 1-10 Average C1 Cash Costs	$ / lb Cu in con	$1.14
Years 1-10 Average C1 Cash Costs + Royalties + Sustaining Capex	$ / lb Cu in con	$1.59
LOM C1 Cash Costs	$ / lb Cu in con	$1.29
LOM C1 Cash Costs + Royalties + Sustaining Capex	$ / lb Cu in con	$1.65

1.	15% contingency is included.
2.	G&A also includes property tax, reclamation and closure costs.

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Rosemont’s annual copper production (contained copper in concentrate) and C1 cash costs (net of by-products at stream prices after deducting capitalized stripping) are shown below in Figure 22-1. Over the first 10 years, annual production is expected to average 140 thousand tons of copper at an average C1 cash cost of $1.14/lb. Over the 19 year LOM, annual production is expected to average 112 thousand tons of copper at an average C1 cash cost of $1.29/lb.

FIGURE 22-1: ROSEMONT ANNUAL COPPER PRODUCTION AND C1 CASH COSTS

22.1.3        Taxes and Royalties

22.1.3.1    Applicable Tax Rates

The Project will be subject to a federal income tax rate of 35% (effectively reduced to 32% by the Section 199 deduction, a manufacturing and production credit available in the U.S.) and an alternative federal minimum tax rate of 20% (effectively reduced to 17% by the Section 199 deduction). A state income tax rate of 4.9% is also applicable to the Project. The amount of state tax payable in a given period reduces the amount of taxable income that is subject to federal income tax.

A depletion allowance of 15% has been utilized to reduce taxable income. It is determined as a percentage of gross income from the property, not to exceed 50% of taxable income before the depletion deduction. The gross income from the property is defined as metal revenue less downstream costs (smelting, refining and transportation). Taxable income is defined as gross income minus operating expenses, overhead expenses, depreciation and state taxes.

A 2.5% severance tax is imposed in Arizona in lieu of sales tax on mining minerals. The net severance base is 50% of the difference between gross value of production and the production cost. The amount of tax is calculated by multiplying the net severance base by 2.5% . Severance tax is considered an income tax and is included in cash income taxes.

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Rosemont has accrued approximately $123 million in net operating losses and approximately $333 million in depreciable capital that can be used to offset future income taxes.

The combined effective tax rate (excluding property tax which is included as part of G&A costs) varies in any given year but amounts to approximately 37% of taxable income over the life of mine.

22.1.4        Depreciation

Tax depreciation was applied to the development capital costs depending on the classification of capital as detailed in Table 22-4. 70% of the development capital associated with the pre-strip is expensed in the year it occurs and the remaining 30% is depreciated on a 5-year straight line basis.

TABLE 22-4: TAX DEPRECIATION FOR DEVELOPMENT CAPITAL

	Development Capital (Excluding Infrastructure & Pre-Strip)	Pre-Strip Development Capital	Infrastructure Development Capital
Year 1	10.71%	20.00%	5.00%
Year 2	19.13%	20.00%	9.50%
Year 3	15.03%	20.00%	8.55%
Year 4	12.25%	20.00%	7.70%
Year 5	12.25%	20.00%	6.93%
Year 6	12.25%	-	6.23%
Year 7	12.25%	-	5.90%
Year 8	6.13%	-	5.90%
Year 9	-	-	5.91%
Year 10	-	-	5.90%
Year 11	-	-	5.91%
Year 12	-	-	5.91%
Year 13	-	-	5.90%
Year 14	-	-	5.91%
Year 15	-	-	5.90%
Year 16	-	-	2.95%
Total	100%	100%	100%

Sustaining capital was depreciated on 7-year straight-line basis for tax purposes (14.29% annually).

22.1.4.1    Royalties

A 3% NSR royalty exists on the Project and is included in the economic analysis.

22.2        Annual Cash Flow Model

A summary of the annual cash flow model is outlined in Table 22-5.

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TABLE 22-5: ANNUAL CASH FLOW MODEL

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22.3        Financial Analysis (100% Project Basis)

Rosemont (on a 100% basis) has an unlevered after-tax NPV8% of $769 million and a 15.5% after-tax IRR at a copper price of $3.00/lb as summarized in Table 22-6. The Project NPV and IRR are calculated using end of period quarterly discounting in the quarter before initial development capital is spent.

TABLE 22-6: LIFE OF MINE FINANCIAL METRICS (100% PROJECT BASIS)

	Units	LOM Total
Gross Revenue (Stream Prices)	$M	$13,377
TCRCs	$M	($1,837)
On-Site Operating Costs (after deducting capitalized stripping)	$M	($4,691)
Royalties	$M	($368)
Operating Margin	$M	$6,480
Development Capital	$M	($1,921)
Stream Upfront Payment	$M	$230
Sustaining Capital (excludes capitalized stripping)	$M	($387)
Capitalized Stripping	$M	($781)
Pre-Tax Cash Flow	$M	$3,622
Cash Income Taxes	$M	($718)
After-Tax Free Cash Flow	$M	$2,903
After-Tax NPV8%	$M	$769
After-Tax NPV10%	$M	$496
After-Tax IRR	%	15.5%
After-Tax Payback Period	Years	5.2

22.4        Sensitivity Analysis (100% Project Basis)

The NPV8% (100% Project basis) was sensitized based on percentage changes in various input assumptions above or below the base case. Each input assumption change was assumed to occur independently from changes in other inputs. The sensitivity analysis is summarized in Figure 22-2. The Project is most sensitive to the copper price followed by initial capital costs, on-site operating costs, and the molybdenum price.

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FIGURE 22-2: NPV8% SENSITIVITY (100% BASIS)

Table 22-7 below reports the after-tax NPV8%, NPV10%, IRR and payback of the Project (on a 100% basis) at various flat copper prices assuming all other inputs remain constant.

TABLE 22-7: AFTER-TAX NPV8%, NPV10% AND IRR SENSITIVITY AT VARIOUS FLAT COPPER PRICES (100% BASIS)

	Flat Copper Price ($/lb)
	$2.50	$2.75	$3.00	$3.25	$3.50
After-Tax NPV8% ($M)	$45	$412	$769	$1,115	$1,448
After-Tax NPV10% ($M)	($122)	$192	$496	$792	$1,076
After-Tax IRR (%)	8.5%	12.2%	15.5%	18.5%	21.2%
After-Tax Payback (years)	6.9	5.9	5.2	4.4	4.3

22.5        Project Ownership Impact on Valuation

The existing Joint Venture Agreement requires cash payments from UCM totaling $106 million to the JV in order for UCM to complete its earn-in for 20% ownership of the Project. The payments will be made on an installment basis to fund the initial development capital, and payments will commence once certain milestones are achieved. The NPV attributable to Hudbay is improved beyond 80% of the standalone Project NPV due to the JV payments, and the IRR attributable to Hudbay is improved beyond the standalone Project IRR as a result of the reduced time period between development capital spending and positive Project cash flow. Table 22-8 shows the adjusted key financial metrics attributable to Hudbay.

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TABLE 22-8: KEY FINANCIAL METRICS ATTRIBUTABLE TO HUDBAY

	Units	LOM Total
Development Capital (100% Basis)	$M	$1,921
Stream Upfront Payment	$M	($230)
Joint Venture Earn-in Payment	$M	($106)
JV Share of Remaining Capital (20%)	$M	($317)
JV Loan Repayment to Hudbay1	$M	($20)
Hudbay's Share of Development Capital	$M	$1,248
After-Tax NPV8% to Hudbay	$M	$719
After-Tax NPV10% to Hudbay	$M	$499
After-Tax IRR to Hudbay	%	17.7%
After-Tax Payback Period to Hudbay	Years	4.9

1. Hudbay is funding the JV’s share of project expenditures until receipt of material permits and approximately $20M in principal and accrued interest is due to Hudbay.

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23      ADJACENT PROPERTIES

The author is not aware of any relevant work on properties immediately adjacent to the Project.

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24      OTHER RELEVANT DATA AND INFORMATION

A unpublished draft feasibility study was completed for the Project which included information on the basis of design, infrastructure, design strategies, Project Execution Strategy, risks assessments and recommendations. The EPCM team has also completed a draft construction execution plan.

24.1        Project Implementation

A draft project execution plan was delivered to Hudbay as part of the draft DFS. The Project will be executed following a classical “EPCM” (Engineering, Procurement and Construction Management) model, which is an industry standard for projects of the magnitude and complexity of the Project. Figure 24-1 describes the Project delivery strategy. The project execution strategy will remain preliminary until the Project is funded and approved to begin after which Hudbay’s Project Team will be developed, roles defined, positions filled and strategies confirmed or modified.

Hudbay’s Project Technical Services is self-delivering the mine development scope for the Project and Hudbay’s Project Team is responsible for the offsite facilities and for the management of the specialist engineering and consulting services providers. The EPCM services provider is responsible for the process plant and related facilities, as well as for the overall project construction management and Health Safety Environmental Community (“HSEC”).

FIGURE 24-1: OVERALL CONSTRUCTION MANAGEMENT & HSEC (EPCM)

The general intent of Hudbay’s Project Team is to provide managerial and technical resources to safely and responsibly manage and control scope, cost, schedule and the quality of the work in the field. The Hudbay Project Team has been assigned responsibility for the management of all aspects of engineering, procurement and construction, with oversight in compliance from the Arizona Business Unit leadership. All communications, coordination and control of services providers and contractors are directed through Hudbay’s Project Team.

The roles and responsibilities for the Hudbay’s Project Management group are detailed below.

24.1.1        Project Director

The Project Director has the overall responsibility for meeting the Owner’s requirements and completing the Project within budget and on schedule.

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24.1.2        Area Managers

There are four Area Managers reporting to the Project Director who are accountable for the Project Management of specific scope including:

• Process Plant and Infrastructure

• Mining

• Heavy Civil Works

• Offsite Infrastructure

24.1.3 F       unctional Managers

The Project Director and Area Managers are supported by functional managers and teams in the areas of:

• Project Services including cost control, administration and document control

• Project Planning and Scheduling

• Commercial including Risk, Procurement and Contracts

• Engineering

• Systems Integration

• Operational Readiness

24.1.4        Hudbay’s Rosemont Project Internal Stakeholders

Hudbay’s Project internal stakeholders include the Project Services group, Business Development, Social Responsibility, Environmental, Finance, Legal, Internal Audit and Technical Services Teams.

24.1.5        Project Sponsors Team

The purpose of the Sponsors team is to remove barriers to success, provide guidance and direction and facilitate the alignment of the integrated team, providing a culture driven to meet the project requirement and to celebrate the project successes.

24.1.6        EPCM Services Provider

EPCM scope is principally the detail engineering of the Process Plant and related facilities, from the dump pocket of the primary crusher to the discharge of tailing conveying systems at the tailing management facility. The scope also includes all water piping and pumping systems, as well as infrastructure such as distribution power and communication. On the concentrate side, the scope extends to the discharge of the concentrate filters, from which point the concentrate will be loaded onto trucks using front-end loaders for shipment.

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Procurement scope includes all of the major equipment packages plus supply of major bulk commodities such as structural steel, pipe electrical cable and cable ladder. These will be free-issued to the Contractors for installation. Contract packages prepared by EPCM service provider will detail the scope of construction work by package, as well as applicable interfaces and battery limits including responsibility for traffic and logistics, receipt, storage and issue of provided materials and equipment.

The EPCM service provider has responsibility for construction management for the overall Project. Other service providers will provide services in general support of the CM effort. An example is a specialty civil engineering firm who will provide field quality assurance services for Heavy Civil Works (“HCW”) projects.

Through the process of advancing the Project to its current state, the EPCM scope also includes bringing in management systems and oversight to ensure delivery of a project that meets Project business objectives and commitments. During the course of the work, Hudbay’s Project Team will monitor and manage EPCM’s performance in this respect.

24.1.7        Project Execution Plan

Various plans have been preliminarily developed to be included in the overall Project Execution Plan such as:

• Project Delivery Strategy

• Health, Safety, Environmental and Community Plans – to include implementation of systems, orientations, security, badging, transportation, emergency response, solid waste removal, compliance inspections, site specific programs, drug/alcohol testing,

• Construction Management Environmental Plan – The plan is based on compliance with the Arizona Business Unit’s Environmental Plan and details implementation of management systems.

• Engineering Execution Plan – Details strategy, processes and standards for delivery of the Engineering deliverables to ensure HSEC, schedule, cost and quality objectives.

• Procurement Management Plan – The main objective is to purchase the equipment and bulk materials required for the Project and manage risks. It will detail out sole source tender procedures, bidding, supplier relations and supply chain management, packaging plan, bulks, and procurement management software. Included in this plan is a sub-plan for materials and equipment management.

• Contracting Management Plan – The Contracting Strategy has been developed considering the strengths of local Contractors and their ability to provide skilled labor, engineered and bulk materials such as earthworks and concrete and in some instances equipment such as concrete batch plants, as an example. In general, horizontal packages favoring Contractor’s strengths have been preferred over vertical packages where a single Contractor coordinates the work of numerous subcontractors or self performs work that is not a proven core capability.
  

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• Construction Execution Plan – The Construction Execution Plan is developed to safely deliver a quality finished product in a cost effective and timely manner. The Project will be carried out in accordance with EPCM and Hudbay’s Rosemont Copper Construction Management and HSEC Management Policies to maximise construction efficiency, do no harm to the health and safety of the Project team and minimize impact on the environment and surrounding community.

24.2        Risk Assessments

The Project has undergone various risk assessments and workshops during the years. These risk assessments were mainly isolated to project specific scopes, therefore, facilitated and maintained by Hudbay’s engineering consultant firms.

TABLE 24-1: RISK ASSESSMENTS

Document	Date	Authoring Entity
Risk Model Workshop Results	25 June 2015	M3/Amec
Project Risk Assessment from Geological Point of View, Rev 1	16 July 2015	Knight Piésold
Hazard Identification Review Results	3-4 Nov. 2015	Ausenco
Project Risk Register	11 March 2016	Ausenco

During June 2016, the Project team took ownership of the risk assessment and conducted an internal facilitated workshop. The facilitator took into consideration the previous assessments, interviews with Hudbay personnel, and a review of documentation (draft DFS, schedule, etc.) to compile a list of underlying assumptions. This list served as the basis for potential risk which were discussed, quantified/measured, and to some extent mitigated.

Following up quarterly with the risk assessments, a second session occurred in November 2016.

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25      INTERPRETATION AND CONCLUSIONS

The purpose of this Technical Report is to present Hudbay’s estimate of the mineral reserves and mineral resources for the Project based on the current mine plan, the current state of metallurgical testing, operating cost and capital cost estimates. This Technical Report includes refinements of certain aspects of the Project’s mine plan. While consistency with issued and pending environmental permits and analysis related thereto has always been a key requirement for this effort, updates to the original mine plan will be necessary. To the extent that any regulatory agency concludes that the current plan requires additional environmental analysis or modification of an existing permit, the intent will be to work with that agency to either complete the required process or to adjust the current mine plan as necessary.

The results of “feasibility study” level work conducted partly by external contractors and partly internally by Hudbay, completion of drill program and bench marking including Hudbay’s Constancia mine has resulted in the following fundamental conclusions:

• The Rosemont deposit consists of copper-molybdenum-silver-gold mineralization primarily hosted in skarn formed on a chemical/siliciclastic sedimentary sequence after the intrusion of Laramide quartz monzonite porphyry intrusions.
• A new geological model was built based on chemostratigraphy and lithogeochemistry using ICP multi-element assays (4 acid digestion) from 33,000 samples covering the full footage of the 2014 and 2015 Hudbay drilling programs. Geochemical based geological model reduces uncertainties in the formational, lithological and alteration logging. The updated geological model incorporated a revised structural framework based on a surface and downhole structural review.
• The resource economic parameters utilized are slightly different than those supporting the mineral reserve statement. The cost and price inputs for the mineral resource economic parameters are considered to represent reasonable estimates and were used to test the economic viability of the resource. Although these cost and prices differ from the ones used for the mineral reserves, it is the opinion of the QP that changing the resource parameters would not materially change the output of the reserve.
• A proven and probable reserve of 592 million tons has been identified and its economic viability demonstrated. An additional 591 million tons of measured and indicated resources and 69 million tons of inferred resources have been identified and outlined as having further potential for economic extraction once the mineral reserves have been extracted.
• Mining equipment performance and maintenance requirements and costs are benchmarked from Constancia’s actual operating information and are robust.
• Metallurgical testwork has confirmed that the Rosemont ores respond well to proven and widely used sulfide mineral processing techniques.
• The tailing properties have been sufficiently characterized as well as the dewatering performance of vendor equipment over the life of the operation to satisfy the estimated number, type and size of tailing filters for this Project. To be conservative, expansion space has been allocated for additional filtering equipment, to the extent that it may be necessary.

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• The Project process plant design has been modelled after the operating Constancia processing plant’s flow sheet and is sufficiently robust to routinely achieve key production targets such as throughput, recovery, and concentrate grade as stated in the production schedule, based on the metallurgical testing conducted.
• Flexibility exists within the mine plan to optimize plant recovery and performance through the management of feed types including clays, oxides and hardness.
• The Project is one of many large projects scheduled to be constructed. The author believes that schedule slip will be the principal pressure on cost should the Project experience construction delays. In the opinion of the author, the contingency allotted to construction capital cost should mitigate most cost risk. At the time of publication of this Technical Report, committed and spent dollars would raise this contingency to approximately 15% of the required project capital as estimated in the dDFS.
• Related to the capital costs, M3 Engineering and Amec Foster Wheeler (acting under a Joint Venture agreement) and Ausenco each completed a value engineering phase and independently produced capital estimates for the project which were within 5% of each other. Since then, Ausenco has also completed a feasibility quality capital estimate, benchmarked their findings with Constancia (and other similar projects), and engaged third party construction contractors and various consulting firms to provide additional input into the estimate. Hudbay has also completed an independent third party review of the feasibility study estimate.
• The net present value of the project is most sensitive to the price of copper. The resulting project NPV8% ($769 million) and IRR (15.5%) utilizing the current Hudbay long term view on metal prices, TCRC’s and other economic assumptions, in the opinion of the author, support the declaration of mineral reserves as outlined in the CIM guidelines.
• The project execution plan is modelled after the Constancia delivery method. Many key personnel who developed the Constancia mine are members of the Project who will be involved in the development of the Project and therefore considered a robust project execution plan.

The Project is uniquely located in a copper mining jurisdiction that has sustained economic copper production for close to 140 years. Since it is located approximately 30 miles from Tucson, it is expected to have a significant impact on employment and economic gain for the region. The proposed mining, processing, and logistics plan provides a step forward in innovation and sustainability. The dry stack tailings deposition proposed would be among the largest in size and address industry and stakeholder concerns regarding the use of water and the stability of tailings impoundment facilities. The proposed design and operating practice that will be applied in respect of the Project is expected to set a new standard by which other large mining projects are judged with respect to their impact on stakeholders, the ecology and the environment.

In recognition of the scarcity of world economic copper reserves in an environment of ever increasing consumption of the metal, Hudbay has carefully considered the ecological, environmental, and ethical extraction methods to be applied to the Project in an effort to set it apart from others in the world. This Project located in a first world leading nation, where extraction and production is governed by laws with due process and human rights fundamental to the consumer.

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This Technical Report also concludes that the estimated mineral reserves and mineral resources for the Project conform to the requirements of 2014 CIM Definition Standards – for Mineral Resources and Mineral Reserves and requirements in Form 43-101F1 of NI 43-101, Standards of Disclosure.

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26      RECOMMENDATIONS

The Author recommends the following:

• The author recommends that Hudbay further investigate the cause(s) of the differences in average molybdenum grade from the historical assays. Hudbay should also evaluate the application of non-linear interpolation or wireframing methods in the minor geological units.
• Consideration of a program of drill hole twinning of the pre-Augusta drilling to bolster the confidence of the re-assaying campaign as conducted by Augusta.
• Further investigate change-of-support correction and alternative approaches to resource classification taking into account the high production rate. This should be performed to ensure that the resource classification properly reflects the reduced risk when a large volume is mined and delivered to the mill on a quarterly and annual basis.
• It is recommended by the author that 5% of the samples should be sent for check assay in future drill hole campaigns.
• Future drilling campaigns should consider the by-product credit contribution of gold noted contained in copper concentrate produced through testing and increase the confidence in geological continuity such that it can be included in the reserve statement.
• Metallurgical testwork has confirmed that marketable copper concentrates can be produced. Mine and mill production sequencing and planning will require care to manage clay content that can adversely affect flotation and tailings filtration.
• A geometallurgical program is recommended as a component of the operating plan to further refine, monitor and optimize the mine to mill performance. It is concluded that fluorine can be readily rejected from copper concentrate, and further study is recommended to develop understanding of ore conditions and indicators that trigger elevated fluorine content in concentrate.
• Ongoing optimization of pit slope designs should be conducted during operation accounting for more detailed mapping of local alterations, jointing, and faulting. These observations and compilation can then provide the basis for revised slope geometry and pushback (mining phase) configurations that have the potential to increase mineral reserves and reduce overall stripping requirements.

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27      REFERENCES

Anzalone, S.A., 1995, The Helvetia Area Porphyry Systems, Pima County, Arizona; in Pierce, F.W., and Bolm, J.G. eds., Porphyry Copper Deposits of the American Cordillera: Tucson, Arizona Geological Society Digest 20, p. 436-441.

AACE International, 2012. AACE International Recommended Practice No.47R-11. Cost Estimate Classification System – As applied in the Mining and Mineral Processing Industries. Rev. July 6, 2012.

Barra, F., Ruiz, J., Valencia, V. A., Ochoa-Landín, L., Chesley, J. T., & Zurcher, L. (2005). Laramide porphyry Cu-Mo mineralization in northern Mexico: Age constraints from Re-Os geochronology in molybdenite. Economic Geology,100(8), 1605-1616.

Base Met Laboratories (BML), 2016. Verification of Rosemont Process Parameters. BL0065. January 8, 2016.

Briggs D. F., 2014. History of Helvetia-Rosemont Mining District, Pima County, Arizona. June 2014.

Call & Nicholas, Inc. (CNI), 2016. Feasibility-Level Geotechnical Study for Rosemont Deposit. May 2016.

Copper State Engineering, Inc. 2014. Geotechnical Investigation Report for West Side Portion of Proposed Water Transmission Pipeline, Rosemont Copper, Pima County, Arizona. June 19, 2014.

Drewes, Harald, Fields, R.A., Hirschberg, D.M., and Bolm, K.S., 2002, Spatial digital database for the tectonic map of southeast Arizona: U.S. Geological Survey Miscellaneous Investigations Series Map I-1109, Digital database, ver. 2.0, published scale 1:125.000, 2 sheets accessed September 9, 2012 at http://pubs.usgs.gov/imap/i1109/i1109_e.pdf [sheet 1 of 2] and http://pubs.usgs.gov/imap/i1109/i1109_w.pdf [sheet 2 of 2].

Kermack, K.A., Haldane, J.B.S., 1950. Organic Correlation and Allometry. Biometrika 37 (1/2), 30-41.

Knight Piésold Consulting, 2016. Rosemont Project, Dry Stack Tailing Facility Preliminary Geotechnical Slope Stability Analysis, Rev. 1. January 28, 2016.

M3 Engineering & Technology Corporation, 2009. Rosemont Copper Project Updated Feasibility Study.

M3 Engineering and Technology Corporation, 2012. Rosemont Copper Project: NI 43-101 Technical Report, Updated Feasibility Study, Pima County, Arizona, USA. Revision 0. Project No. M3-PN 08036. Prepared for Augusta Resource Corporation. Tucson, AZ: M3 Engineering and Technology Corporation. August 28, 2012.

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Maher, D. J. Reconstruction of middle Tertiary extension and Laramide porphyry copper systems, east-central Arizona: Unpublished Ph. D. Diss. thesis, University of Arizona, 2008.

Neirbo Hydrogeology, 2016. Prefeasibility Dewatering Plan and Analyses. Rosemont Project. Prepared for Hudbay. May 2016.

Ramussen, J., Hoag, C. and Horstman, K., 2012. Geology of the Northern Santa Rita Mountains, Arizona. Arizona Geological Society Fall Field Trip September 15, 2012.

SGS Canada Inc., 2015. An investigation into The Grindability Characteristics of Three Samples from the Rosemont Deposit. Prepared for XPS Consulting and Testwork Services. June 17, 2015.

Simón, A., 2004. Evaluation of Twin and Duplicate Samples: The Hyperbolic Method. AMEC Peru Ltda, Internal Document, 2pp.

Terracon Consultants, Inc. 2015. Preliminary Geotechnical Engineering Report. Rosemont Well System, South Old Nogales Highway, Sahuarita, Arizona. December 7, 2015.

Tetra Tech, 2009. Plant Site Geotechnical Report, Rosemont Copper Project. Prepared for Rosemont Copper. August 2009.

Wardrop, 2005. Technical Report on the Rosemont Property, Pima County, Arizona, June 3, 2005, 145 pp.

XPS Consulting & Testwork Services, 2015a. BWi and SAG Variability Analysis of Rosemont Composites. MET 1A, MET 1B, MET 2. Prepared for Hudbay Mining. July 16, 2015.

XPS Consulting & Testwork Services, 2015b. Rosemont Project Phase II Geometallurgical Program Technical Review. September 10, 2015.

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28      SIGNATURE PAGE

This NI 43-101 technical report titled “Mineral Resource Estimate, Rosemont Project, Pima County, Arizona, USA”, dated and effective as of March 30, 2017 was prepared and signed by the following author:

Dated this 30th day of March, 2017.

(signed) Cashel Meagher

_________________________Signature of Qualified Person

Cashel Meagher, P.Geo. Senior Vice President & Chief Operating Officer, Hudbay

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29      CERTIFICATES OF QUALIFIED PERSONS

CASHEL MEAGHER

CERTIFICATE OF QUALIFICATION

Re: Rosemont Project Technical Report, March 30, 2017

I, Cashel Meagher, B. Sc., P. Geo, of Toronto, Canada, do hereby certify that:

1.	I am currently employed as Senior Vice President and Chief Operating Officer, with Hudbay Minerals Inc., 25 York St, Suite 800, Toronto Ontario.
2.	I graduated from Saint Francis Xavier University with a Joint Advanced major in Geology and Chemistry in 1994.
3.	I am a member in good standing with the Association of Professional Geoscientists of Ontario, member #1056.
4.	I have practiced my profession continuously over 20 years and have been involved in mineral exploration, project evaluation, resource and reserve evaluation, and mine operations in underground and open pit mines for base metal and precious metal deposits in North and South America.
5.	I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education and affiliation with a professional association and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purpose of NI 43-101.
6.	I have reviewed and approved the Summary of the Technical Report and I am responsible for the preparation of this Technical Report titled “NI 43-101 Technical Report, Rosemont Project, Pima County, Arizona, USA” (the “Technical Report”), dated March 30, 2017.
7.	I last visited the property on April 21, 2016. I also visited it several times prior to that date.
8.	As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
9.	I am not independent of the Issuer. Since I am an employee of the Issuer, a producing issuer, I fall under subsection 5.3(3) of NI 43-101 where “a technical report required to be filed by a producing issuer is not required to be prepared by or under supervision of an independent qualified person”.
10.	I have been directly involved with the Rosemont Project property, which is the subject of the Technical Report, continuously since January, 2016.
11.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with the instrument and form.
12.	I consent to the public filing of the Technical Report with any stock exchange, securities commission or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

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Dated as of March 30, 2017.

Original signed by:

Cashel Meagher, P.Geo.

Senior Vice President & Chief Operating Officer, Hudbay

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30      A1-1 LAND TENURE

The following tables identify the patented claims, unpatented claims, and fee owned associated lands.

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A1-2 Rosemont Project Patented Claims

COUNT	PARCEL NO.	PROPERTY NAME	SECTION-TOWNSHIP-RANGE	ASSESSED ACRES	2015 FEES
1	305540020	BLACK BESS	13-18-15	13.540	$12.21
2	305540030	FLYING DUTCHMAN	13-18-15	20.380	$12.21
3	305540040	WISCONSIN	13-18-15	20.660	$12.21
4	305540050	EXCHANGE	13-18-15	20.660	$12.21
5	305540060	EXCHANGE NO. 2	13-18-15	6.590	$12.21
6	305540070	COPPER WORLD	13-18-15	20.660	$12.21
7	305540080	OWOSKO	13-18-15	20.660	$12.21
8	305540090	BLACK HORSE	13-18-15	13.810	$12.21
9	305540100	BRUNSWICK	13-18-15	18.660	$12.21
10	305540110	ANTELOPE	13-18-15	17.360	$12.21
11	305550010	NEWMAN	14-18-15	16.500	$12.21
12	305550040	CHANCE	14-18-15	20.160	$12.21
13	305550050	BLACK HAWK	14-18-15	11.360	$12.21
14	305550060	TELEMETER	14-18-15	8.150	$12.21
15	305550070	WEST END	14-18-15	19.530	$12.21
16	305550080	HATTIE	14-18-15	12.190	$12.21
17	305550090	SILVER SPUR	14-18-15	8.610	$12.21
18	305550100	SLIDE	14-18-15	12.880	$12.21
19	305550110	BACK BONE	14-18-15	19.070	$12.21
20	305550130	BUZZARD	14-18-15	20.660	$12.21
21	305550140	HEAVY WEIGHT	14-18-15	20.660	$12.21
22	305550150	LIGHT WEIGHT	14-18-15	20.660	$12.21
23	305560040	PEACH	15-18-15	18.070	$12.21
24	305560050	SOUTH END	15-18-15	17.810	$12.21
25	305560060	MONITOR	15-18-15	13.320	$12.21
26	305560070	GAP	15-18-15	16.250	$12.21
27	305580080	WATER WISH	23-18-15	20.660	$12.21
28	305580090	NEW MEXICO	23-18-15	15.130	$12.21
29	305580100	GRIZZLY	23-18-15	20.660	$12.21
30	305580110	OLD DICK	23-18-15	20.130	$12.21
31	305580120	AMERICAN	23-18-15	20.100	$12.21
32	305580130	RECORDER	23-18-15	6.700	$12.21
33	305580140	MOHAWK	23-18-15	13.550	$12.21
34	305580150	WEDGE	23-18-15	19.310	$12.21
35	305580160	DAN	23-18-15	2.480	$12.21
36	305580170	GENERAL	23-18-15	9.170	$12.21
37	305580180	ELGIN	23-18-15	14.000	$12.21
38	305580190	SUNSETE	23-18-15	0.667	$12.21
39	305580200	TELEPHONE	23-18-15	18.660	$12.21
40	305580220	ELGIN MILLSITE	23-18-15	4.994	$12.21
41	305580250	DAN MILLSITE	23-18-15	2.856	$12.21
42	305580260	WEDGE MILLSITE	23-18-15	4.987	$12.21
43	305580270	OLD DICK MILLSITE	23-18-15	2.196	$12.21
44	305590060	ARCOLA	24-18-15	20.660	$12.21
45	305590070	BONNIE BLUE	24-18-15	20.660	$12.21
46	305590080	KING	24-18-15	20.660	$12.21
47	305590090	EXILE	24-18-15	16.020	$12.21
48	305590100	VULTURE	24-18-15	15.730	$12.21
49	305590110	ISLE ROYAL	24-18-15	20.660	$12.21
50	305590120	INDIAN CLUB	24-18-15	19.200	$12.21
51	305590130	A.O.T.	24-18-15	14.200	$12.21
52	305590140	BALTIMORE	24-18-15	9.620	$12.21
53	305590150	PILOT	24-18-15	14.700	$12.21
54	305590160	LITTLE DAVE	24-18-15	20.660	$12.21
55	305590170	COPPER FEND	24-18-15	20.660	$12.21

Page 30-2

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	PARCEL NO.	PROPERTY NAME	SECTION-TOWNSHIP-RANGE	ASSESSED ACRES	2015 FEES
56	305590180	TALLY HO	24-18-15	20.380	$12.21
57	305590190	LEADER	24-18-15	20.660	$12.21
58	305590200	OMEGA	24-18-15	20.660	$12.21
59	305590220	ECLIPSE COPPER	24-18-15	20.660	$12.21
60	305590230	SCHWAB	24-18-15	9.261	$12.21
61	305590240	NARRAGANSETT BAY	24-18-15	12.428	$12.21
62	305590250	LANDOR	24-18-15	15.870	$12.21
63	305590260	WARD	24-18-15	17.693	$12.21
64	305590270	ALTA COPPER	24-18-15	18.180	$12.21
65	305590280	BROAD TOP	24-18-15	17.150	$12.21
66	305590290	MALACHITE	24-18-15	20.660	$12.21
67	305600040	YORK	25-18-15	13.380	$12.21
68	305600050	OLCOTT	25-18-15	5.485	$12.21
69	305600060	HILO CONSOLIDATED	25-18-15	12.190	$12.21
70	305600070	ELDON	25-18-15	18.984	$12.21
71	305600080	RAINBOW	25-18-15	7.765	$12.21
72	305600090	AJAX CONSOLIDATED	25-18-15	13.980	$12.21
73	305600100	CUBA	25-18-15	12.030	$12.21
74	305600110	FALLS	25-18-15	16.340	$12.21
75	305600130	OLD PUT CON	25-18-15	20.650	$12.21
76	305600140	FRANKLIN	25-18-15	20.540	$12.21
77	305600150	CUSHING	25-18-15	15.040	$12.21
78	305600160	CENTRAL	25-18-15	17.860	$12.21
79	305600170	POTOMAC	25-18-15	20.620	$12.21
80	305610010	MARION	36-18-15	20.660	$12.21
81	305610030	EXCELSIOR	36-18-15	20.575	$12.21
82	305610040	EMPIRE	36-18-15	10.210	$12.21
83	305610050	ALTAMONT	36-18-15	20.610	$12.21
84	305610060	ERIE	36-18-15	19.610	$12.21
85	305610080	CHICAGO	36-18-15	16.660	$12.21
86	305610090	COCONINO	36-18-15	14.100	$12.21
87	305630020	OLUSTEE	19-18-16	20.520	$12.00
88	305630040	AMOLE	19-18-16	17.921	$12.00
89	305640020	CHICAGO MILLSITE	29-18-16	5.000	$13.70
90	305640030	COCONINO MILLSITE	29-18-16	5.000	$13.70
91	305640040	OLD PUT MILLSITE	29-18-16	5.000	$13.70
92	305640050	OREGON MILLSITE	29-18-16	5.000	$13.70
93	305640060	OLD PAP MILLSITE	29-18-16	5.000	$13.70
94	305640070	AJAX CONSOLIDATED MILLSITE	29-18-16	5.000	$13.70
95	305650020	R. G. INGERSOLL	30-18-16	20.620	$12.00
96	305650040	PATRICK HENRY	30-18-16	19.050	$12.00
97	305660050	MOHAWK SILVER	01-19-15	19.760	$7.89
98	305660060	TREMONT	01-19-15	12.860	$7.89
99	30554012A	BLUE POINT	13-18-15	19.288	$12.21
100	30555012A	HEAVY WEIGHT MILLSITE	14-18-15	5.000	$12.21
101A	30558021A	TELEPHONE MILLSITE	23-18-15	4.610	$12.21
102	30558023A	RECORDER MILLSITE	23-18-15	2.640	$12.21
101B	30558023B	TELEPHONE, RECORDER & AMERICAN MILLSITE	23-18-15	3.830	$12.21
103	30558024A	AMERICAN MILLSITE	23-18-15	4.540	$12.21
104	30559021A	OMEGA FIRST EXTENSION SOUTH	24-18-15	20.660	$12.21
105A	30560003A	DAYLIGHT	25-18-15	13.210	$12.21
105B	30560003B	DAYLIGHT	30-18-16	5.960	$12.21
106	30560012A	OLD PAP COPPER	25-18-15	20.650	$12.21
107	30560012D	FALLS NO. 2	25-18-15	7.320	$12.21
108	30560012F	WEDGE NO. 2	25-18-15	1.280	$12.21

Page 30-3

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	PARCEL NO.	ROPERTY NAME	SECTION-TOWNSHIP-RANGE	ASSESSED ACRES	2015 FEES
109	30560012G	WEDGE	25-18-15	6.600	$12.21
110	30560012H	SANTA RITA FRACTION	25-18-15	0.980	$12.21
111A	30560012J	SANTA RITA #13	25-18-15	10.520	$12.21
112	30561007A	OREGON COPPER	36-18-15	16.080	$12.21
113A	30561007D	SANTA RITA #15	36-18-15	13.590	$12.21
114	30561007E	SANTA RITA #14	36-18-15	19.160	$12.21
115	30561007F	SANTA RITA #12	36-18-15	19.620	$12.21
116	30561007G	LAST CHANCE NO. 1	36-18-15	15.600	$12.21
117	30561007H	LAST CHANCE NO. 2	36-18-15	18.270	$12.21
118	30561007J	SANTA RITA #26	36-18-15	20.030	$12.21
119	30561007K	SANTA RITA #27	36-18-15	18.760	$12.21
120A	30561007L	SANTA RITA #28	36-18-15	18.570	$12.21
121	30562034C	SANTA RITA #16	31-18-16	18.920	$12.00
113B	30562034D	SANTA RITA #15	31-18-16	6.440	$12.00
120B	30562034E	SANTA RITA #28	31-18-16	2.010	$12.00
111B	30562034F	SANTA RITA #13	31-18-16	7.510	$12.00
122	30563003A	CUPRITE	19-18-16	20.660	$12.00
123	30564008A	FRANKLIN MILLSITE	29-18-16	5.000	$12.00
124	30565003A	LA FAYETTE	30-18-16	13.950	$12.00
125	30565003D	SANTA RITA #4	30-18-16	19.000	$12.00
126	30565003E	SANTA RITA #5	30-18-16	19.020	$12.00
127	30565003F	SANTA RITA #6	30-18-16	18.990	$12.00
128A	30565003G	SANTA RITA #8A	25-18-15	3.660	$12.00
129A	30565003H	SANTA RITA #9	30-18-16, 31-18-16	19.580	$12.00
130	30565003J	SANTA RITA #10	30-18-16, 31-18-16	20.560	$12.00
131	30565003K	SANTA RITA #11	31-18-16	20.560	$12.00
128B	30565003L	SANTA RITA #8A	25-18-15,(S/B 30-18-16)	10.750	$12.00
129B	30565003M	SANTA RITA #9	25-18-15	1.020	$12.00
132A	30565005A	DAN WEBSTER	30-18-16	15.190	$12.00
132B	30565005B	DAN WEBSTER	25-18-15	3.770	$12.00
PATENTED CLAIM TOTALS				2003.520	$1,705.08

Page 30-4

	Rosemont Project
	Form 43-101F1 Technical Report

A1-3 Rosemont Project Unpatented Claims

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
1	York Fraction	AMC2198	$155.00
2	Travis #1	AMC2199	$155.00
3	Jim	AMC2200	$155.00
4	Isle Royal Fraction	AMC2201	$155.00
5	Indian Club Fraction	AMC2202	$155.00
6	Pilot Fraction	AMC2203	$155.00
7	A.O.T. Fraction	AMC2204	$155.00
8	Malachite Fraction	AMC2211	$155.00
9	MAX 121 B/Relocation	AMC13284	$155.00
10	MAX 123 B/Relocation	AMC13286	$155.00
11	MAX 125 B/Relocation	AMC13288	$155.00
12	MAX 126 B/Relocation	AMC13289	$155.00
13	MAX 127 B/Relocation	AMC13290	$155.00
14	MAX 128 B/Relocation	AMC13291	$155.00
15	MAX 129 B/Relocation	AMC13292	$155.00
16	MAX 130 B/Relocation	AMC13293	$155.00
17	MAX 131 B/Relocation	AMC13294	$155.00
18	MAX 132 B/Relocation	AMC13295	$155.00
19	MAX 133 B/Relocation	AMC13296	$155.00
20	MAX 134 B/Relocation	AMC13297	$155.00
21	MAX 135 B/Relocation	AMC13298	$155.00
22	MAX 136 B/Relocation	AMC13299	$155.00
23	MAX 137 B/Relocation	AMC13300	$155.00
24	MAX 138 B/Relocation	AMC13301	$155.00
25	MAX 139 B/Relocation	AMC13302	$155.00
26	MAX 140 B/Relocation	AMC13303	$155.00
27	MAX 141 B/Relocation	AMC13304	$155.00
28	MAX 142 B/Relocation	AMC13305	$155.00
29	MAX 143 B/Relocation	AMC13306	$155.00
30	MAX 144 B/Relocation	AMC13307	$155.00
31	MAX 145 B/Relocation	AMC13308	$155.00
32	MAX 146 B/Relocation	AMC13309	$155.00
33	MAX 147 B/Relocation	AMC13310	$155.00
34	MAX 148 B/Relocation	AMC13311	$155.00
35	MAX 149 B/Relocation	AMC13312	$155.00
36	MAX 150 B/Relocation	AMC13313	$155.00
37	MAX 151 B/Relocation	AMC13314	$155.00
38	MAX 152 B/Relocation	AMC13315	$155.00
39	MAX 153 B/Relocation	AMC13316	$155.00
40	MAX 154 B/Relocation	AMC13317	$155.00
41	MAX 155 B/Relocation	AMC13318	$155.00
42	MAX 156 B/Relocation	AMC13319	$155.00
43	Rosaland	AMC14972	$155.00
44	Michael M	AMC14973	$155.00
45	Lydia J	AMC14974	$155.00
46	Ida D	AMC14975	$155.00
47	D & D #1	AMC14976	$155.00
48	D & D II	AMC14977	$155.00
49	Frijole	AMC14978	$155.00
50	Frijole II	AMC14979	$155.00

Page 30-5

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
51	Frijole III	AMC14980	$155.00
52	Frijole IV	AMC14981	$155.00
53	Frijole V	AMC14982	$155.00
54	Frijole VII	AMC14984	$155.00
55	Frijole VIII	AMC14985	$155.00
56	Frijole IX	AMC14986	$155.00
57	Frijole X	AMC14987	$155.00
58	Frijole XI	AMC14988	$155.00
59	Frijole XI Extension	AMC14989	$155.00
60	Deering Springs No. 2 A/Relocation	AMC15002	$155.00
61	Deering Springs No. 4 A/Relocation	AMC15003	$155.00
62	Deering Springs No. 6 A/Relocation	AMC15004	$155.00
63	Deering Springs No. 8 A/Relocation	AMC15005	$155.00
64	Deering Springs No. 10 A/Relocation	AMC15006	$155.00
65	Deering Springs No. 12 A/Relocation	AMC15007	$155.00
66	Deering Springs No. 14 A/Relocation	AMC15008	$155.00
67	Deering Springs No. 15 A/Relocation	AMC15009	$155.00
68	Deering Springs No. 16 A/Relocation	AMC15010	$155.00
69	Deering Springs No. 17 A/Relocation	AMC15011	$155.00
70	Deering Springs No. 21 A/Relocation	AMC15012	$155.00
71	Deering Springs No. 22 A/Relocation	AMC15013	$155.00
72	Deering Springs No. 23 A/Relocation	AMC15014	$155.00
73	Deering Springs No. 24 A/Relocation	AMC15015	$155.00
74	Deering Springs No. 25 A/Relocation	AMC15016	$155.00
75	Deering Springs No. 26 A/Relocation	AMC15017	$155.00
76	Deering Springs No. 27 A/Relocation	AMC15018	$155.00
77	Deering Springs No. 28 A/Relocation	AMC15019	$155.00
78	Deering Springs No. 29 A/Relocation	AMC15020	$155.00
79	Deering Springs No. 30 A/Relocation	AMC15021	$155.00
80	Deering Springs No. 31 A/Relocation	AMC15022	$155.00
81	Deering Springs No. 32 A/Relocation	AMC15023	$155.00
82	Deering Springs No. 33 A/Relocation	AMC15024	$155.00
83	Deering Springs No. 34 A/Relocation	AMC15025	$155.00
84	Deering Springs No. 35 A/Relocation	AMC15026	$155.00
85	Deering Springs No. 36 A/Relocation	AMC15027	$155.00
86	Deering Springs No. 37 A/Relocation	AMC15028	$155.00
87	Deering Springs No. 38 A/Relocation	AMC15029	$155.00
88	Deering Springs No. 39 A/Relocation	AMC15030	$155.00
89	Deering Springs No. 42 A/Relocation	AMC15031	$155.00
90	Deering Springs No. 51 A/Relocation	AMC15032	$155.00
91	Deering Springs No. 52 A/Relocation	AMC15033	$155.00
92	Kid 1	AMC25210	$155.00
93	Kid 2	AMC25211	$155.00
94	Kid 3	AMC25212	$155.00
95	Kid 4	AMC25213	$155.00
96	Kid 5	AMC25214	$155.00
97	Kid 6	AMC25215	$155.00
98	Kid 7	AMC25216	$155.00
99	Kid 8	AMC25217	$155.00
100	Kid 9	AMC25218	$155.00

Page 30-6

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
101	Kid 10	AMC25219	$155.00
102	Kid 11	AMC25220	$155.00
103	Kid 12	AMC25221	$155.00
104	Kid 13	AMC25222	$155.00
105	Kid 14	AMC25223	$155.00
106	Kid 15	AMC25224	$155.00
107	Kid 16	AMC25225	$155.00
108	Kid 17	AMC25226	$155.00
109	Kid 18	AMC25227	$155.00
110	Kid 19	AMC25228	$155.00
111	Kid 20	AMC25229	$155.00
112	Kid 21	AMC25230	$155.00
113	Kid 22	AMC25231	$155.00
114	Kid 23	AMC25232	$155.00
115	Kid 24	AMC25233	$155.00
116	Kid 25	AMC25234	$155.00
117	Kid 26	AMC25235	$155.00
118	Kid 27	AMC25236	$155.00
119	Kid 28	AMC25237	$155.00
120	Kid 29	AMC25238	$155.00
121	Kid 34	AMC25243	$155.00
122	Kid 35	AMC25244	$155.00
123	Kid 36	AMC25245	$155.00
124	Kid 37	AMC25246	$155.00
125	Kid 38	AMC25247	$155.00
126	Kid 39	AMC25248	$155.00
127	Kid 40	AMC25249	$155.00
128	Kid 41	AMC25250	$155.00
129	Kid 42	AMC25251	$155.00
130	Kid 43	AMC25252	$155.00
131	Kid 44	AMC25253	$155.00
132	Kid 45	AMC25254	$155.00
133	Kid 46	AMC25255	$155.00
134	Kid 47	AMC25256	$155.00
135	Wasp 52	AMC25257	$155.00
136	Wasp 53	AMC25258	$155.00
137	Wasp 54	AMC25259	$155.00
138	Wasp 55	AMC25260	$155.00
139	Wasp 56	AMC25261	$155.00
140	Wasp 57	AMC25262	$155.00
141	Wasp 58	AMC25263	$155.00
142	Wasp 60	AMC25264	$155.00
143	Wasp 61	AMC25265	$155.00
144	Wasp 101	AMC25268	$155.00
145	Wasp 102	AMC25269	$155.00
146	Wasp 103	AMC25270	$155.00
147	Wasp 104	AMC25271	$155.00
148	Wasp 105	AMC25272	$155.00
149	Wasp 106	AMC25273	$155.00
150	Wasp 107	AMC25274	$155.00

Page 30-7

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
151	Wasp 111	AMC25275	$155.00
152	Wasp 112	AMC25276	$155.00
153	Wasp 113	AMC25277	$155.00
154	Wasp 114	AMC25278	$155.00
155	Wasp 115	AMC25279	$155.00
156	Wasp 116	AMC25280	$155.00
157	Wasp 117	AMC25281	$155.00
158	Wasp 118	AMC25282	$155.00
159	Wasp 119	AMC25283	$155.00
160	Wasp 120	AMC25284	$155.00
161	Wasp 121	AMC25285	$155.00
162	Wasp 122	AMC25286	$155.00
163	Wasp 123	AMC25287	$155.00
164	Wasp 124	AMC25288	$155.00
165	Wasp 125	AMC25289	$155.00
166	Wasp 126	AMC25290	$155.00
167	Wasp 127	AMC25291	$155.00
168	Wasp 128	AMC25292	$155.00
169	Wasp 129	AMC25293	$155.00
170	Wasp 130	AMC25294	$155.00
171	Wasp 201	AMC25295	$155.00
172	Wasp 202	AMC25296	$155.00
173	Wasp 203	AMC25297	$155.00
174	Wasp 204	AMC25298	$155.00
175	Wasp 205	AMC25299	$155.00
176	Wasp 206	AMC25300	$155.00
177	Wasp 207	AMC25301	$155.00
178	Wasp 208	AMC25302	$155.00
179	Wasp 209	AMC25303	$155.00
180	Wasp 210	AMC25304	$155.00
181	Wasp 211	AMC25305	$155.00
182	Wasp 212	AMC25306	$155.00
183	Wasp 213	AMC25307	$155.00
184	Wasp 214	AMC25308	$155.00
185	Wasp 215	AMC25309	$155.00
186	Wasp 216	AMC25310	$155.00
187	Wasp 217	AMC25311	$155.00
188	Wasp 218	AMC25312	$155.00
189	Wasp 313	AMC25349	$155.00
190	Wasp 315	AMC25351	$155.00
191	Wasp 317	AMC25353	$155.00
192	Wasp 319	AMC25355	$155.00
193	Wasp 321	AMC25357	$155.00
194	Wasp 323	AMC25359	$155.00
195	Wasp 325	AMC25361	$155.00
196	Wasp 327	AMC25363	$155.00
197	Wasp 329	AMC25365	$155.00
198	Wasp 331	AMC25367	$155.00
199	Wasp 333	AMC25369	$155.00
200	Wasp 335	AMC25371	$155.00

Page 30-8

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
201	Wasp 337	AMC25373	$155.00
202	Wasp 339	AMC25375	$155.00
203	Wasp 341	AMC25377	$155.00
204	Wasp 343	AMC25379	$155.00
205	Wasp 344	AMC25380	$155.00
206	Wasp 345	AMC25381	$155.00
207	Wasp 346	AMC25382	$155.00
208	Wasp 347	AMC25383	$155.00
209	Wasp 348	AMC25384	$155.00
210	Wasp 349	AMC25385	$155.00
211	Wasp 350	AMC25386	$155.00
212	Wasp 351	AMC25387	$155.00
213	Wasp 352	AMC25388	$155.00
214	Wasp 353	AMC25389	$155.00
215	Wasp 354	AMC25390	$155.00
216	Max 41	AMC25662	$155.00
217	Max 43	AMC25664	$155.00
218	Max 45	AMC25666	$155.00
219	Max 47	AMC25668	$155.00
220	Max 49	AMC25670	$155.00
221	Max 71	AMC25692	$155.00
222	Max 72	AMC25693	$155.00
223	Max 73	AMC25694	$155.00
224	Max 74	AMC25695	$155.00
225	Max 75	AMC25696	$155.00
226	Max 76	AMC25697	$155.00
227	Max 77	AMC25698	$155.00
228	Max 78	AMC25699	$155.00
229	Max 79	AMC25700	$155.00
230	Max 80	AMC25701	$155.00
231	Max 81	AMC25702	$155.00
232	Max 82	AMC25703	$155.00
233	Max 83	AMC25704	$155.00
234	Max 84	AMC25705	$155.00
235	Max 85	AMC25706	$155.00
236	Max 86	AMC25707	$155.00
237	Max 87	AMC25708	$155.00
238	Max 88	AMC25709	$155.00
239	Max 89	AMC25710	$155.00
240	Max 90	AMC25711	$155.00
241	Max 91	AMC25712	$155.00
242	Max 93	AMC25714	$155.00
243	Max 95	AMC25716	$155.00
244	Max 97	AMC25718	$155.00
245	Max 99	AMC25720	$155.00
246	Max 101	AMC25722	$155.00
247	Max 102	AMC25723	$155.00
248	Max 103	AMC25724	$155.00
249	Max 104	AMC25725	$155.00
250	Max 105	AMC25726	$155.00

Page 30-9

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
251	Max 106	AMC25727	$155.00
252	Max 107	AMC25728	$155.00
253	Max 108	AMC25729	$155.00
254	Max 109	AMC25730	$155.00
255	Max 110	AMC25731	$155.00
256	Max 111	AMC25732	$155.00
257	Max 112	AMC25733	$155.00
258	Max 113	AMC25734	$155.00
259	Max 114	AMC25735	$155.00
260	Max 115	AMC25736	$155.00
261	Max 116	AMC25737	$155.00
262	Max 117	AMC25738	$155.00
263	Max 118	AMC25739	$155.00
264	Max 119	AMC25740	$155.00
265	Max 120	AMC25741	$155.00
266	Elk 1	AMC27423	$155.00
267	Elk 2	AMC27424	$155.00
268	Elk 3	AMC27425	$155.00
269	Elk 4	AMC27426	$155.00
270	Elk 5	AMC27427	$155.00
271	Elk 6	AMC27428	$155.00
272	Elk 35	AMC27451	$155.00
273	Elk 36	AMC27452	$155.00
274	Elk 37	AMC27453	$155.00
275	Elk 39	AMC27455	$155.00
276	Elk 41	AMC27457	$155.00
277	Elk 43	AMC27459	$155.00
278	Elk 45	AMC27461	$155.00
279	Elk 70	AMC27465	$155.00
280	Elk 71	AMC27466	$155.00
281	Elk 72	AMC27467	$155.00
282	Elk 73	AMC27468	$155.00
283	Elk 74	AMC27469	$155.00
284	Elk 75	AMC27470	$155.00
285	Elk 76	AMC27471	$155.00
286	Elk 77	AMC27472	$155.00
287	Elk 78	AMC27473	$155.00
288	Elk 79	AMC27474	$155.00
289	Elk 80	AMC27475	$155.00
290	Elk 81	AMC27476	$155.00
291	Elk 83	AMC27478	$155.00
292	Elk 85	AMC27480	$155.00
293	Elk 87	AMC27482	$155.00
294	Alpine #5	AMC27513	$155.00
295	Alpine #6	AMC27514	$155.00
296	Alpine #7	AMC27515	$155.00
297	Alpine #8	AMC27516	$155.00
298	Alpine #9	AMC27517	$155.00
299	Alpine #10	AMC27518	$155.00
300	Alpine #11	AMC27519	$155.00

Page 30-10

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
301	Alpine #12	AMC27520	$155.00
302	Alpine #13	AMC27521	$155.00
303	Alpine #14	AMC27522	$155.00
304	Alpine #15	AMC27523	$155.00
305	Alpine #16	AMC27524	$155.00
306	Alpine #17	AMC27525	$155.00
307	Alpine #18	AMC27526	$155.00
308	Alpine #19	AMC27527	$155.00
309	Alpine #20	AMC27528	$155.00
310	Alpine #21	AMC27529	$155.00
311	Alpine #22	AMC27530	$155.00
312	Alpine #23	AMC27531	$155.00
313	Alpine #24	AMC27532	$155.00
314	Santa Rita Wedge	AMC28871	$155.00
315	Buzzard No. 5	AMC36021	$155.00
316	Shadow #4	AMC36025	$155.00
317	John 1	AMC36026	$155.00
318	John 2	AMC36027	$155.00
319	Flying Dutchman No. 2	AMC36028	$155.00
320	Flying Dutchman No. 3	AMC36029	$155.00
321	Flying Dutchman No. 4	AMC36030	$155.00
322	Flying Dutchman No. 5	AMC36031	$155.00
323	Flying Dutchman No. 6	AMC36032	$155.00
324	Black Bess No. 2	AMC36034	$155.00
325	K.W.L.	AMC36036	$155.00
326	G.E.J.	AMC36037	$155.00
327	R.F.E.	AMC36038	$155.00
328	R.C.M.	AMC36039	$155.00
329	Sycamore #1	AMC36040	$155.00
330	Sycamore #2	AMC36041	$155.00
331	Sycamore #3	AMC36042	$155.00
332	Sycamore #4	AMC36043	$155.00
333	Sycamore #5	AMC36044	$155.00
334	Sycamore #6	AMC36045	$155.00
335	Sycamore #7	AMC36046	$155.00
336	Sycamore #8	AMC36047	$155.00
337	Sycamore #9	AMC36048	$155.00
338	Sycamore #10	AMC36049	$155.00
339	Sycamore #11	AMC36050	$155.00
340	Sycamore #12	AMC36051	$155.00
341	Naragansett Extension #1	AMC36052	$155.00
342	Naragansett Ext. #2	AMC36053	$155.00
343	Naragansett Extension #3	AMC36054	$155.00
344	Naragansett Extension #4	AMC36055	$155.00
345	Naragansett Extension #5	AMC36056	$155.00
346	Naragansett Extension #6	AMC36057	$155.00
347	Naragansett Extension #7	AMC36058	$155.00
348	Naragansett Extension #8	AMC36059	$155.00
349	Narragansett Ext. No. 9	AMC36060	$155.00
350	Schwab Extension #1 North West	AMC36061	$155.00

Page 30-11

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
351	Rocky 1	AMC36062	$155.00
352	Amole No. 2	AMC36063	$155.00
353	Falls No. 3	AMC36065	$155.00
354	Falls No. 4	AMC36066	$155.00
355	Perry No. 1	AMC36067	$155.00
356	Perry #2	AMC36068	$155.00
357	Perry #3	AMC36069	$155.00
358	Perry #4	AMC36070	$155.00
359	Perry #7	AMC36073	$155.00
360	Perry #8	AMC36074	$155.00
361	Perry #9	AMC36075	$155.00
362	Perry #10	AMC36076	$155.00
363	Perry #11	AMC36077	$155.00
364	Perry #12	AMC36078	$155.00
365	Perry #15	AMC36081	$155.00
366	Perry #16	AMC36082	$155.00
367	Perry #17	AMC36083	$155.00
368	Perry #18	AMC36084	$155.00
369	Gunsite 1-A	AMC36086	$155.00
370	Gunsite No. 2	AMC36087	$155.00
371	Gunsite No. 3	AMC36088	$155.00
372	Gunsite No. 4	AMC36089	$155.00
373	Gunsite 5A	AMC36090	$155.00
374	Gunsite 6-B	AMC36091	$155.00
375	Gunsite No. 7	AMC36092	$155.00
376	Gunsite 7A	AMC36093	$155.00
377	Gunsite No. 8	AMC36094	$155.00
378	Gunsite No. 9	AMC36095	$155.00
379	Gunsite No. 10	AMC36096	$155.00
380	Gunsite No. 11	AMC36097	$155.00
381	Gunsite No. 12	AMC36098	$155.00
382	Gunsite No. 13	AMC36099	$155.00
383	Gunsite No. 14	AMC36100	$155.00
384	Gunsite No. 15	AMC36101	$155.00
385	Gunsite No. 16	AMC36102	$155.00
386	Gunsite No. 17	AMC36103	$155.00
387	Gunsite No. 18	AMC36104	$155.00
388	Gunsite No. 19	AMC36105	$155.00
389	Gunsite No. 20	AMC36106	$155.00
390	Gunsite No. 21	AMC36107	$155.00
391	Gunsite No. 22	AMC36108	$155.00
392	Gunsight No. 23	AMC36109	$155.00
393	Gunsite No. 24	AMC36110	$155.00
394	Gunsite No. 25	AMC36111	$155.00
395	Gunsite No. 26	AMC36112	$155.00
396	Gunsite No. 27	AMC36113	$155.00
397	Gunsight No. 28	AMC36114	$155.00
398	Gunsight No. 29	AMC36115	$155.00
399	Gunsight No. 30	AMC36116	$155.00
400	Gunsight No. 31	AMC36117	$155.00

Page 30-12

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
401	Gunsight No. 32	AMC36118	$155.00
402	Gunsight No. 33	AMC36119	$155.00
403	Gunsight No. 35	AMC36121	$155.00
404	Gunsight No. 36	AMC36122	$155.00
405	Gunsight No. 37	AMC36123	$155.00
406	Gunsight No. 38	AMC36124	$155.00
407	Gunsight No. 39	AMC36125	$155.00
408	Gunsight No. 40	AMC36126	$155.00
409	Gunsight No. 41	AMC36127	$155.00
410	Gunsight No. 42	AMC36128	$155.00
411	Gunsight No. 43	AMC36129	$155.00
412	Gunsight 44	AMC36130	$155.00
413	Gunsight #45	AMC36131	$155.00
414	Gunsight #46	AMC36132	$155.00
415	Gunsight #47	AMC36133	$155.00
416	Gunsight #48	AMC36134	$155.00
417	Gunsight #49	AMC36135	$155.00
418	Gunsight #50	AMC36136	$155.00
419	Williams Folly	AMC36137	$155.00
420	Williams Folly #2	AMC36138	$155.00
421	Santa Rita #1	AMC46740	$155.00
422	Santa Rita #2	AMC46741	$155.00
423	Santa Rita #3	AMC46742	$155.00
424	Santa Rita #7	AMC46746	$155.00
425	Santa Rita #17	AMC46756	$155.00
426	Santa Rita #18	AMC46757	$155.00
427	Santa Rita #19	AMC46758	$155.00
428	Santa Rita #20	AMC46759	$155.00
429	Santa Rita #21	AMC46760	$155.00
430	Santa Rita #22	AMC46761	$155.00
431	Santa Rita #23	AMC46762	$155.00
432	Santa Rita #24	AMC46763	$155.00
433	Santa Rita #25	AMC46764	$155.00
434	Santa Rita #29	AMC46768	$155.00
435	Santa Rita #30	AMC46769	$155.00
436	Santa Rita #31	AMC46770	$155.00
437	Catalina #1	AMC46771	$155.00
438	Catalina #2	AMC46772	$155.00
439	Catalina #3	AMC46773	$155.00
440	Catalina #4	AMC46774	$155.00
441	Catalina #5A	AMC46775	$155.00
442	Catalina #6A	AMC46776	$155.00
443	Catalina #7	AMC46777	$155.00
444	Catalina #8	AMC46778	$155.00
445	Fred Bennett	AMC46779	$155.00
446	Fred Bennett	AMC46780	$155.00
447	Rosemont #9	AMC46781	$155.00
448	Rosemont #11	AMC46782	$155.00
449	Rosemont 11-A	AMC46783	$155.00
450	Rosemont #12	AMC46784	$155.00

Page 30-13

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
451	Rosemont #13	AMC46785	$155.00
452	Rosemont #15	AMC46786	$155.00
453	Rosemont #16	AMC46787	$155.00
454	Rosemont #17	AMC46788	$155.00
455	Rosemont #18	AMC46789	$155.00
456	Rosemont 21	AMC46790	$155.00
457	Fred Bennett Fraction	AMC46791	$155.00
458	Last Chance No. 3/Relocation	AMC46794	$155.00
459	Cave	AMC46796	$155.00
460	Strip	AMC46800	$155.00
461	Cuba Fraction	AMC46801	$155.00
462	Patrick Henry Fraction/Relocation	AMC46802	$155.00
463	R. G. Ingersoll Fraction	AMC46803	$155.00
464	Daylight Fraction	AMC46804	$155.00
465	Travis #2	AMC46805	$155.00
466	Travis #3	AMC46806	$155.00
467	Travis #4	AMC46807	$155.00
468	Travis #5	AMC46808	$155.00
469	Travis #6	AMC46809	$155.00
470	Art	AMC46810	$155.00
471	Al	AMC46811	$155.00
472	Sam	AMC46812	$155.00
473	Fred	AMC46813	$155.00
474	Bert	AMC46814	$155.00
475	Bob	AMC46815	$155.00
476	Canyon No. 34	AMC47482	$155.00
477	Canyon No. 35	AMC47483	$155.00
478	Canyon No. 36	AMC47484	$155.00
479	Canyon No. 37	AMC47485	$155.00
480	Canyon No. 38	AMC47486	$155.00
481	Canyon No. 39	AMC47487	$155.00
482	Canyon No. 40	AMC47488	$155.00
483	Canyon No. 41	AMC47489	$155.00
484	Canyon No. 42	AMC47490	$155.00
485	Canyon No. 43	AMC47491	$155.00
486	Canyon No. 64	AMC47512	$155.00
487	Canyon No. 65	AMC47513	$155.00
488	Canyon No. 66	AMC47514	$155.00
489	Canyon No. 67	AMC47515	$155.00
490	Canyon No. 68	AMC47516	$155.00
491	Canyon No. 69	AMC47517	$155.00
492	Canyon No. 70	AMC47518	$155.00
493	Canyon No. 71	AMC47519	$155.00
494	Canyon No. 72	AMC47520	$155.00
495	Canyon No. 73	AMC47521	$155.00
496	Canyon No. 74	AMC47522	$155.00
497	Canyon No. 75	AMC47523	$155.00
498	Canyon No. 76	AMC47524	$155.00
499	Canyon No. 77	AMC47525	$155.00
500	Canyon No. 78	AMC47526	$155.00

Page 30-14

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
501	Canyon No. 79	AMC47527	$155.00
502	Telemeter Fraction	AMC62785	$155.00
503	West End Fraction	AMC62786	$155.00
504	Hattie Fraction	AMC62787	$155.00
505	Cactus	AMC64123	$155.00
506	Travis #7	AMC64124	$155.00
507	Fox #1	AMC64125	$155.00
508	Fox #2	AMC64126	$155.00
509	Fox #7	AMC64131	$155.00
510	Fox #13	AMC64133	$155.00
511	Cloud Rest	AMC64134	$155.00
512	Big Windy	AMC64135	$155.00
513	Big Windy Fraction	AMC64136	$155.00
514	Blue Wing	AMC64137	$155.00
515	Cloud Rest No. 1	AMC64138	$155.00
516	Kent #1 Long John	AMC66835	$155.00
517	Kent #2 Patricia C.	AMC66836	$155.00
518	Kent #3 Little Joe	AMC66837	$155.00
519	Belle of Rosemont	AMC66838	$155.00
520	John	AMC74390	$155.00
521	Joe	AMC74391	$155.00
522	Ben	AMC74392	$155.00
523	Pete	AMC74393	$155.00
524	Adolph Lewisohn	AMC74394	$155.00
525	Adolph Lewisohn	AMC74395	$155.00
526	Rosemont	AMC74396	$155.00
527	Rosemont	AMC74397	$155.00
528	Albert Steinfeld	AMC74398	$155.00
529	Albert Steinfeld	AMC74399	$155.00
530	Hugh Young	AMC74400	$155.00
531	Hugh Young	AMC74401	$155.00
532	Ethel	AMC74402	$155.00
533	Albert	AMC74403	$155.00
534	Rosemont #1	AMC74404	$155.00
535	Rosemont #2	AMC74405	$155.00
536	Rosemont #3	AMC74406	$155.00
537	Rosemont #4	AMC74407	$155.00
538	Rosemont #7	AMC74408	$155.00
539	Rosemont #8	AMC74409	$155.00
540	Rosemont #14	AMC74410	$155.00
541	Rosemont #19	AMC74411	$155.00
542	Rosemont #20	AMC74412	$155.00
543	Rosemont #20	AMC74413	$155.00
544	Rosemont #22	AMC74414	$155.00
545	Rosemont #23	AMC74415	$155.00
546	Rosemont #24	AMC74416	$155.00
547	Rosemont #25	AMC74417	$155.00
548	RX	AMC74418	$155.00
549	Flying Dutchman #7A	AMC75181	$155.00
550	Blue Point No. 2A	AMC75182	$155.00

Page 30-15

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
551	Alpine #1A	AMC75183	$155.00
552	Alpine #2A	AMC75184	$155.00
553	Alpine #3A	AMC75185	$155.00
554	Alpine #4A	AMC75186	$155.00
555	Frijole VI A	AMC95315	$155.00
556	Falcon 1A	AMC99789	$155.00
557	Falcon 2A	AMC99790	$155.00
558	Falcon 3A	AMC99791	$155.00
559	Falcon 4A	AMC99792	$155.00
560	Falcon 5A	AMC99793	$155.00
561	Falcon 6A	AMC99794	$155.00
562	Falcon 7A	AMC99795	$155.00
563	Falcon 8A	AMC99796	$155.00
564	Falcon 9A	AMC99797	$155.00
565	Falcon 10A	AMC99798	$155.00
566	Falcon 11A	AMC99799	$155.00
567	Falcon 12A	AMC99800	$155.00
568	Falcon 13A	AMC99801	$155.00
569	Falcon 14A	AMC99802	$155.00
570	Falcon 15A	AMC99803	$155.00
571	Falcon 16A	AMC99804	$155.00
572	Falcon 17A	AMC99805	$155.00
573	Falcon 18A	AMC99806	$155.00
574	Falcon 19A	AMC99807	$155.00
575	Falcon 20A	AMC99808	$155.00
576	Falcon 21A	AMC99809	$155.00
577	Falcon 22A	AMC99810	$155.00
578	Falcon 27A	AMC99811	$155.00
579	Falcon 28A	AMC99812	$155.00
580	Falcon 29A	AMC99813	$155.00
581	Falcon 30A	AMC99814	$155.00
582	Falcon 31A	AMC99815	$155.00
583	Falcon 32A	AMC99816	$155.00
584	Wasp 62A	AMC99817	$155.00
585	Wasp 63A	AMC99818	$155.00
586	Wasp 219A	AMC99819	$155.00
587	Wasp 220A	AMC99820	$155.00
588	Wasp 221A	AMC99821	$155.00
589	Wasp 222A	AMC99822	$155.00
590	Tecky	AMC99823	$155.00
591	MIA 1A	AMC117293	$155.00
592	MIA 2A	AMC117294	$155.00
593	MIA 3A	AMC117295	$155.00
594	MIA 4A	AMC117296	$155.00
595	MIA 5A	AMC117297	$155.00
596	MIA 6A	AMC117298	$155.00
597	MIA 7A	AMC117299	$155.00
598	MIA 8A	AMC117300	$155.00
599	MIA 9A	AMC117301	$155.00
600	MIA 12A	AMC117304	$155.00

Page 30-16

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
601	MIA 13A	AMC117305	$155.00
602	MIA 14A	AMC117306	$155.00
603	BILLY C.	AMC129394	$155.00
604	Hope-1	AMC303950	$155.00
605	Hope 2	AMC303951	$155.00
606	Hope-3	AMC303952	$155.00
607	Hope-4	AMC303953	$155.00
608	Hope-5	AMC303954	$155.00
609	Hope-6	AMC303955	$155.00
610	Hope-7	AMC303956	$155.00
611	Hope 8	AMC303957	$155.00
612	Hope-9	AMC303958	$155.00
613	Hope 10	AMC303959	$155.00
614	Hope-10A	AMC303960	$155.00
615	Hope-11	AMC303961	$155.00
616	Hope-12	AMC303962	$155.00
617	Hope-13	AMC303963	$155.00
618	Hope-14	AMC303964	$155.00
619	Hope-15	AMC303965	$155.00
620	Hope-16	AMC303966	$155.00
621	Hope-17	AMC303967	$155.00
622	Hope-18	AMC303968	$155.00
623	Hope-19	AMC303969	$155.00
624	Hope-20	AMC303970	$155.00
625	Hope-21	AMC303971	$155.00
626	Hope-22	AMC303972	$155.00
627	Hope 23	AMC303973	$155.00
628	Hope-24	AMC303974	$155.00
629	Hope-25	AMC303975	$155.00
630	Hope-26	AMC303976	$155.00
631	Hope-27	AMC303977	$155.00
632	Hope-28	AMC303978	$155.00
633	H-29	AMC303979	$155.00
634	Hope-30	AMC303980	$155.00
635	Hope-31	AMC303981	$155.00
636	Hope 32	AMC303982	$155.00
637	Hope-33	AMC303983	$155.00
638	Hope-34	AMC303984	$155.00
639	Hope-35	AMC303985	$155.00
640	Hope-36	AMC303986	$155.00
641	Hope-37	AMC303987	$155.00
642	H-38A	AMC313532	$155.00
643	H-39A	AMC313533	$155.00
644	H-40A	AMC313534	$155.00
645	H-41A	AMC313535	$155.00
646	H-42A	AMC313536	$155.00
647	H-43A	AMC313537	$155.00
648	H-44A	AMC313538	$155.00
649	H-45A	AMC313539	$155.00
650	H-46A	AMC313540	$155.00

Page 30-17

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
651	H-47A	AMC313541	$155.00
652	H-48A	AMC313542	$155.00
653	H-49A	AMC313543	$155.00
654	H-50A	AMC313544	$155.00
655	H-51A	AMC313545	$155.00
656	H-52A	AMC313546	$155.00
657	H-53A	AMC313547	$155.00
658	H-54A	AMC313548	$155.00
659	H-55A	AMC313549	$155.00
660	H-56A	AMC313550	$155.00
661	H-57A	AMC313551	$155.00
662	H-58A	AMC313552	$155.00
663	H-59A	AMC313553	$155.00
664	H-60A	AMC313554	$155.00
665	H-61A	AMC313555	$155.00
666	H-62A	AMC313556	$155.00
667	H-63A	AMC313557	$155.00
668	H-64A	AMC313558	$155.00
669	H-65A	AMC313559	$155.00
670	H-66A	AMC313560	$155.00
671	H-67A	AMC313561	$155.00
672	H-68A	AMC313562	$155.00
673	H-69A	AMC313563	$155.00
674	H-70A	AMC313564	$155.00
675	H-71A	AMC313565	$155.00
676	H-72A	AMC313566	$155.00
677	H-73A	AMC313567	$155.00
678	H-74A	AMC313568	$155.00
679	H-75A	AMC313569	$155.00
680	H-76A	AMC313570	$155.00
681	H-77A	AMC313571	$155.00
682	H-78A	AMC313572	$155.00
683	H-79A	AMC313573	$155.00
684	H-80A	AMC313574	$155.00
685	H-81A	AMC313575	$155.00
686	H-82A	AMC313576	$155.00
687	H-83A	AMC313577	$155.00
688	H-84A	AMC313578	$155.00
689	H-85A	AMC313579	$155.00
690	H-86A	AMC313580	$155.00
691	H-87A	AMC313581	$155.00
692	H-88A	AMC313582	$155.00
693	H-89A	AMC313583	$155.00
694	H-90A	AMC313584	$155.00
695	H-91A	AMC313585	$155.00
696	H-92A	AMC313586	$155.00
697	H-93A	AMC313587	$155.00
698	H-94A	AMC313588	$155.00
699	H-95A	AMC313589	$155.00
700	H-96A	AMC313590	$155.00

Page 30-18

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
701	H-97A	AMC313591	$155.00
702	H-98A	AMC313592	$155.00
703	H-99A	AMC313593	$155.00
704	H-100A	AMC313594	$155.00
705	H-101A	AMC313595	$155.00
706	H-102A	AMC313596	$155.00
707	H-103A	AMC313597	$155.00
708	H-104A	AMC313598	$155.00
709	H-105A	AMC313599	$155.00
710	H-106A	AMC313600	$155.00
711	H-107A	AMC313601	$155.00
712	H-108A	AMC313602	$155.00
713	H-109A	AMC313603	$155.00
714	H-110A	AMC313604	$155.00
715	H-111A	AMC313605	$155.00
716	H-112A	AMC313606	$155.00
717	H-113A	AMC313607	$155.00
718	H-114A	AMC313608	$155.00
719	H-115A	AMC313609	$155.00
720	H-116A	AMC313610	$155.00
721	H-117A	AMC313611	$155.00
722	H-118A	AMC313612	$155.00
723	H-119A	AMC313613	$155.00
724	H-120A	AMC313614	$155.00
725	H-121A	AMC313615	$155.00
726	H-122A	AMC313616	$155.00
727	H-123A	AMC313617	$155.00
728	H-124A	AMC313618	$155.00
729	H-125A	AMC313619	$155.00
730	H-126A	AMC313620	$155.00
731	H-127A	AMC313621	$155.00
732	H-128A	AMC313622	$155.00
733	H-129A	AMC313623	$155.00
734	H-130A	AMC313624	$155.00
735	H-131A	AMC313625	$155.00
736	H-132A	AMC313626	$155.00
737	H-133A	AMC313627	$155.00
738	H-134A	AMC313628	$155.00
739	H-135A	AMC313629	$155.00
740	H-136A	AMC313630	$155.00
741	H-137A	AMC313631	$155.00
742	H-138A	AMC313632	$155.00
743	H-139A	AMC313633	$155.00
744	H-140A	AMC313634	$155.00
745	H-141A	AMC313635	$155.00
746	H-142A	AMC313636	$155.00
747	H-143A	AMC313637	$155.00
748	H-144A	AMC313638	$155.00
749	H-145A	AMC313639	$155.00
750	H-146A	AMC313640	$155.00

Page 30-19

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
751	H-147A	AMC313641	$155.00
752	H-148A	AMC313642	$155.00
753	H-149A	AMC313643	$155.00
754	H-150A	AMC313644	$155.00
755	H-151A	AMC313645	$155.00
756	H-152A	AMC313646	$155.00
757	H-153A	AMC313647	$155.00
758	H-154A	AMC313648	$155.00
759	H-155A	AMC313649	$155.00
760	H-156A	AMC313650	$155.00
761	H-157A	AMC313651	$155.00
762	H-158A	AMC313652	$155.00
763	H-159A	AMC313653	$155.00
764	H-160A	AMC313654	$155.00
765	H-161A	AMC313655	$155.00
766	H-162A	AMC313656	$155.00
767	H-163A	AMC313657	$155.00
768	H-164A	AMC313658	$155.00
769	H-165A	AMC313659	$155.00
770	H-166A	AMC313660	$155.00
771	H-167A	AMC313661	$155.00
772	H-168A	AMC313662	$155.00
773	H-169A	AMC313663	$155.00
774	H-170A	AMC313664	$155.00
775	H-171A	AMC313665	$155.00
776	H-177A	AMC313671	$155.00
777	H-178A	AMC313672	$155.00
778	H-179A	AMC313673	$155.00
779	H-180A	AMC313674	$155.00
780	H-181A	AMC313675	$155.00
781	H-182A	AMC313676	$155.00
782	H-183A	AMC313677	$155.00
783	H-187A	AMC313678	$155.00
784	H-188A	AMC313679	$155.00
785	H-189A	AMC313680	$155.00
786	H-190A	AMC313681	$155.00
787	H-191A	AMC313682	$155.00
788	H-192A	AMC313683	$155.00
789	H-194A	AMC313684	$155.00
790	H-195A	AMC313685	$155.00
791	H-196A	AMC313686	$155.00
792	H-197A	AMC313687	$155.00
793	H-198A	AMC313688	$155.00
794	H-199A	AMC313689	$155.00
795	Hope No. 201	AMC330891	$155.00
796	Hope 201A	AMC330892	$155.00
797	Hope No. 202	AMC330893	$155.00
798	Hope No. 203	AMC330894	$155.00
799	Hope No. 204	AMC330895	$155.00
800	Hope No. 205	AMC330896	$155.00

Page 30-20

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
801	Hope No. 206	AMC330897	$155.00
802	Hope No. 207	AMC330898	$155.00
803	Hope No. 208	AMC330899	$155.00
804	Hope No. 209	AMC330900	$155.00
805	Hope No. 210	AMC330901	$155.00
806	Hope No. 211	AMC330902	$155.00
807	Hope No. 212	AMC330903	$155.00
808	Hope No. 213	AMC330904	$155.00
809	Hope No. 214	AMC330905	$155.00
810	Hope No. 215	AMC330906	$155.00
811	Hope No. 216	AMC330907	$155.00
812	Hope No. 222	AMC330910	$155.00
813	Hope No. 223	AMC330911	$155.00
814	Hope No. 224	AMC330912	$155.00
815	Hope No. 225	AMC330913	$155.00
816	Hope 226A	AMC330914	$155.00
817	Hope 227A	AMC330915	$155.00
818	Hope 228A	AMC330916	$155.00
819	Hope 229A	AMC330917	$155.00
820	Hope No. 230	AMC330918	$155.00
821	Hope No. 231	AMC330919	$155.00
822	Hope No. 232	AMC330920	$155.00
823	Hope No. 233	AMC330921	$155.00
824	Hope No. 234	AMC330922	$155.00
825	Hope No. 235	AMC330923	$155.00
826	Hope No. 236	AMC330924	$155.00
827	Hope No. 237	AMC330925	$155.00
828	Hope No. 238	AMC330926	$155.00
829	Hope No. 239	AMC330927	$155.00
830	Hope No. 240	AMC330928	$155.00
831	Hope No. 241	AMC330929	$155.00
832	Hope No. 242	AMC330930	$155.00
833	Hope No. 243	AMC330931	$155.00
834	Hope No. 244	AMC330932	$155.00
835	Hope No. 245	AMC330933	$155.00
836	Hope No. 246	AMC330934	$155.00
837	Hope No. 250	AMC330935	$155.00
838	Hope No. 251	AMC330936	$155.00
839	Hope No. 252	AMC330937	$155.00
840	Hope No. 253	AMC330938	$155.00
841	Hope No. 254	AMC330939	$155.00
842	Hope No. 255	AMC330940	$155.00
843	Hope No. 256	AMC330941	$155.00
844	Hope No. 257	AMC330942	$155.00
845	Elk 47/Relocation	AMC330943	$155.00
846	H-172 B/Relocation	AMC331308	$155.00
847	H-173 B/Relocation	AMC331309	$155.00
848	H-174 B/Relocation	AMC331310	$155.00
849	H-175 B/Relocation	AMC331311	$155.00
850	H-176 B/Relocation	AMC331312	$155.00

Page 30-21

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
851	MMRE	AMC367652	$155.00
852	HV 1	AMC380250	$155.00
853	HV 2	AMC380251	$155.00
854	HV 3	AMC380252	$155.00
855	HV 4	AMC380253	$155.00
856	ROSE 1	AMC385174	$155.00
857	ROSE 2	AMC385175	$155.00
858	ROSE 3	AMC385176	$155.00
859	ROSE 4	AMC385177	$155.00
860	ROSE 5	AMC385178	$155.00
861	ROSE 6	AMC385179	$155.00
862	ROSE 7	AMC385180	$155.00
863	ROSE 8	AMC385181	$155.00
864	ROSE 9	AMC385182	$155.00
865	HV 6	AMC387231	$155.00
866	HV 7	AMC387232	$155.00
867	HV 8	AMC387233	$155.00
868	HV 9	AMC387234	$155.00
869	HV 10	AMC387235	$155.00
870	HV 11	AMC387236	$155.00
871	HV 12	AMC387237	$155.00
872	HV 13	AMC387238	$155.00
873	HV 23	AMC387241	$155.00
874	HV 24	AMC387242	$155.00
875	HV 25	AMC387243	$155.00
876	HV 16	AMC390077	$155.00
877	HV 17	AMC390078	$155.00
878	HV 18	AMC390079	$155.00
879	HV 19	AMC390080	$155.00
880	HV 20	AMC390081	$155.00
881	HV 21	AMC390082	$155.00
882	HV 22	AMC390083	$155.00
883	WAIT-1	AMC390084	$155.00
884	WAIT-2	AMC390085	$155.00
885	WAIT-3	AMC390086	$155.00
886	WAIT-4	AMC390087	$155.00
887	WAIT-5	AMC390088	$155.00
888	WAIT-6	AMC390089	$155.00
889	WAIT-7	AMC390090	$155.00
890	WAIT-8	AMC390091	$155.00
891	WAIT-9	AMC390092	$155.00
892	WAIT-10	AMC390093	$155.00
893	WAIT-11	AMC390094	$155.00
894	WAIT-12	AMC390095	$155.00
895	WAIT-13	AMC390096	$155.00
896	WAIT-14	AMC390097	$155.00
897	WAIT-15	AMC390098	$155.00
898	WAIT-16	AMC390099	$155.00
899	WAIT-17	AMC390100	$155.00
900	WAIT-18	AMC390101	$155.00

Page 30-22

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
901	WAIT-19	AMC390102	$155.00
902	WAIT-20	AMC390103	$155.00
903	WAIT-21	AMC390104	$155.00
904	WAIT-22	AMC390105	$155.00
905	WAIT-23	AMC390106	$155.00
906	WAIT-24	AMC390107	$155.00
907	WAIT-25	AMC390108	$155.00
908	WAIT-26	AMC390109	$155.00
909	WAIT-27	AMC390110	$155.00
910	WAIT-28	AMC390111	$155.00
911	WAIT-29	AMC390112	$155.00
912	WAIT-30	AMC390113	$155.00
913	WAIT-31	AMC390114	$155.00
914	WAIT-32	AMC390115	$155.00
915	FALLS FRACTION	AMC391154	$155.00
916	H-69B	AMC391155	$155.00
917	NO CHANCE No. 3	AMC391156	$155.00
918	SCHWAB FRACTION	AMC391157	$155.00
919	H FRAC. 1	AMC392445	$155.00
920	H FRAC. 2	AMC392446	$155.00
921	H FRAC. 3	AMC392447	$155.00
922	H FRAC. 4	AMC392448	$155.00
923	H FRAC. 5	AMC392449	$155.00
924	H FRAC. 6	AMC392450	$155.00
925	H FRAC. 7	AMC392451	$155.00
926	H FRAC. 8	AMC392452	$155.00
927	BILLY FRAC.	AMC393532	$155.00
928	DSM 1	AMC393533	$155.00
929	DSM 2	AMC393534	$155.00
930	DSM 3	AMC393535	$155.00
931	DSM 4	AMC393536	$155.00
932	DSM 5	AMC393537	$155.00
933	DSM 6	AMC393538	$155.00
934	DSM 7	AMC393539	$155.00
935	DSM 8	AMC393540	$155.00
936	DSM 9	AMC393541	$155.00
937	DSM 10	AMC393542	$155.00
938	HV5 A	AMC393543	$155.00
939	MIA FRAC 1	AMC393544	$155.00
940	MIA FRAC 2	AMC393545	$155.00
941	SON OF GUN 34	AMC394006	$155.00
942	RMT FRAC 1	AMC394561	$155.00
943	RMT FRAC 2	AMC394562	$155.00
944	RMT FRAC 3	AMC394563	$155.00
945	RMT FRAC 4	AMC394564	$155.00
946	NC-CF	AMC396422	$155.00
947	Thankful	AMC404128	$155.00
948	RCC-1	AMC411964	$155.00
949	RCC-2	AMC411965	$155.00
950	RCC-3	AMC411966	$155.00

Page 30-23

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
951	RCC-4	AMC411967	$155.00
952	RCC-5	AMC411968	$155.00
953	RCC-6	AMC411969	$155.00
954	RCC-7	AMC411970	$155.00
955	RCC-8	AMC411971	$155.00
956	RCC-9	AMC411972	$155.00
957	RCC-10	AMC411973	$155.00
958	RCC-11	AMC411974	$155.00
959	RCC-12	AMC411975	$155.00
960	RCC-13	AMC411976	$155.00
961	RCC-14	AMC411977	$155.00
962	RCC-15	AMC411978	$155.00
963	RCC-16	AMC411979	$155.00
964	RCC-17	AMC411980	$155.00
965	RCC-18	AMC411981	$155.00
966	RCC-19	AMC411982	$155.00
967	RCC-20	AMC411983	$155.00
968	RCC-21	AMC411984	$155.00
969	RCC-22	AMC411985	$155.00
970	RCC-23	AMC411986	$155.00
971	RCC-24	AMC411987	$155.00
972	RCC-25	AMC411988	$155.00
973	RCC-26	AMC411989	$155.00
974	RCC-27	AMC411990	$155.00
975	RCC-28	AMC411991	$155.00
976	RCC-29	AMC411992	$155.00
977	RCC-30	AMC411993	$155.00
978	RCC-31	AMC411994	$155.00
979	RCC-32	AMC411995	$155.00
980	RCC-33	AMC411996	$155.00
981	RCC-34	AMC411997	$155.00
982	RCC-35	AMC411998	$155.00
983	RCC-36	AMC411999	$155.00
984	RCC-37	AMC412000	$155.00
985	RCC-38	AMC412001	$155.00
986	RCC-39	AMC412002	$155.00
987	RCC-40	AMC412003	$155.00
988	RCC-41	AMC412004	$155.00
989	RCC-42	AMC412005	$155.00
990	RCC-43	AMC412006	$155.00
991	RCC-44	AMC412007	$155.00
992	RCC-45	AMC412008	$155.00
993	RCC-46	AMC412009	$155.00
994	RCC-47	AMC412010	$155.00
995	RCC-48	AMC412011	$155.00
996	RCC-49	AMC412012	$155.00
997	RCC-50	AMC412013	$155.00
998	RCC-51	AMC412014	$155.00
999	RCC-52	AMC412015	$155.00
1000	RCC-53	AMC412016	$155.00

Page 30-24

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
1001	RCC-54	AMC412017	$155.00
1002	RCC-55	AMC412018	$155.00
1003	RCC-56	AMC412019	$155.00
1004	RCC-57	AMC412020	$155.00
1005	RCC-58	AMC412021	$155.00
1006	RCC-59	AMC412022	$155.00
1007	RCC-60	AMC412023	$155.00
1008	RCC-61	AMC412024	$155.00
1009	RCC-62	AMC412025	$155.00
1010	RCC-63	AMC412026	$155.00
1011	RCC-64	AMC412027	$155.00
1012	RCC-65	AMC412028	$155.00
1013	RCC-66	AMC412029	$155.00
1014	RCC-67	AMC412030	$155.00
1015	RCC-68	AMC412031	$155.00
1016	RCC-69	AMC412032	$155.00
1017	RCC-70	AMC412033	$155.00
1018	RCC-71	AMC412034	$155.00
1019	RCC-72	AMC412035	$155.00
1020	RCC-73	AMC412036	$155.00
1021	RCC-74	AMC412037	$155.00
1022	RCC-75	AMC412038	$155.00
1023	RCC-76	AMC412039	$155.00
1024	RCC-77	AMC412040	$155.00
1025	RCC-78	AMC412041	$155.00
1026	RCC-79	AMC412042	$155.00
1027	RCC-80	AMC412043	$155.00
1028	RCC-81	AMC412044	$155.00
1029	RCC-82	AMC412045	$155.00
1030	RCC-83	AMC412046	$155.00
1031	RCC-84	AMC412047	$155.00
1032	RCC-85	AMC412048	$155.00
1033	RCC-86	AMC412049	$155.00
1034	RCC-87	AMC412050	$155.00
1035	RCC-88	AMC412051	$155.00
1036	RCC-89	AMC412052	$155.00
1037	RCC-90	AMC412053	$155.00
1038	RCC-91	AMC412054	$155.00
1039	RCC-92	AMC412055	$155.00
1040	RCC-93	AMC412056	$155.00
1041	RCC-94	AMC412057	$155.00
1042	RCC-95	AMC412058	$155.00
1043	RCC-96	AMC412059	$155.00
1044	RCC-97	AMC412060	$155.00
1045	RCC-98	AMC412061	$155.00
1046	RCC-99	AMC412062	$155.00
1047	RCC-100	AMC412063	$155.00
1048	AGAVE-1	AMC412064	$155.00
1049	AGAVE-2	AMC412065	$155.00
1050	AGAVE-3	AMC412066	$155.00

Page 30-25

	Rosemont Project
	Form 43-101F1 Technical Report

COUNT	UNPATENTED CLAIM NAME	BLM SERIAL NUMBER	ANNUAL FEE
1051	AGAVE-4	AMC412067	$155.00
1052	AGAVE-5	AMC412068	$155.00
1053	AGAVE-6	AMC412069	$155.00
1054	CONTINENTAL-1	AMC412070	$155.00
1055	CONTINENTAL-2	AMC412071	$155.00
1056	CONTINENTAL-3	AMC412072	$155.00
1057	CONTINENTAL-4	AMC412073	$155.00
1058	CONTINENTAL-5	AMC412074	$155.00
1059	CONTINENTAL-6	AMC412075	$155.00
1060	TAILOR	AMC423213	$155.00
1061	AGAVE-7	AMC429429	$155.00
1062	AGAVE-8	AMC429430	$155.00
1063	AGAVE-9	AMC429431	$155.00
1064	RECORDER FRACTION	AMC429432	$155.00
UNPATENTED CLAIM TOTALS			$164,920.00

Page 30-26

	Rosemont Project
	Form 43-101F1 Technical Report

A1-4 Rosemont Project Fee Owned (Associated) Lands

COUNT	PARCEL NO.	PROPERTY NAME	SECTION-TOWNSHIP-RANGE	ASSESSED ACRES	2015 FEES
1	305580280	HELVETIA RANCH (KILGORE/ANDERSEN)	23-18-15	10.080	$460.00
2	305580330	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT	23-18-15	40.000	$20.53
3	305580350	HELVETIA RANCH ANNEX (DE LA OSSA)	23-18-15	10.000	$76.72
4	305580360	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT)	23-18-15	10.000	$12.21
5	305580370	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT)	23-18-15	20.000	$12.21
6	305580420	HELVETIA RANCH ANNEX (MAATR)	23-18-15	40.000	$20.53
7	30553002D	HELVETIA RANCH ANNEX NORTH (TERRA BELLA)	10-18-15	20.000	$12.21
8	30553002F	HELVETIA RANCH ANNEX NORTH (TERRA BELLA)	10-18-15	120.000	$61.69
9	30553002G	HELVETIA RANCH ANNEX NORTH (TERRA BELLA)	10-18-15	310.000	$159.51
10	30553002H	HELVETIA RANCH ANNEX NORTH (AVRA VALLEY/LEBRECHT)	10-18-15	108.420	$55.73
11	30553004D	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT)	27-18-15	40.000	$20.53
12	30553004H	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT)	27-18-15	40.000	$20.53
13	30556001B	HELVETIA RANCH ANNEX NORTH (AVRA VALLEY/LEBRECHT	15-18-15	313.110	$161.04
14	30556001C	HELVETIA RANCH ANNEX NORTH (AVRA VALLEY/LEBRECHT)	15-18-15	67.800	$869.00
15	30557004B	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT)	22-18-15	5.000	$12.21
16	30557004C	HELVETIA RANCH ANNEX (ADC/CALICA)	22-18-15	52.480	$12.81
17	30557004D	HELVETIA RANCH (KILGORE/ANDERSEN)	22-18-15	10.000	$87.44
18	30557005B	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT)	22-18-15	20.000	$12.21
19	30557013B	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT)	22-18-15	35.000	$18.09
20	30557013C	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT)	22-18-15	40.000	$20.53
21	30557013D	HELVETIA RANCH ANNEX (SUTTLES)	22-18-15	20.000	$341.75
22	30557013E	HELVETIA RANCH ANNEX (AVRA VALLEY/LEBRECHT)	22-18-15	40.000	$20.53
23	30557022C	HELVETIA RANCH ANNEX (MAATR)	22-18-15	40.000	$20.53
24	30558034C	HELVETIA RANCH ANNEX (PIPELINE TRIANGLE)	23-18-15	2.190	$12.21
25	30562006B	ROSEMONT RANCH	14-18-16	34.120	$17.31
26	30562007D	ROSEMONT RANCH	15-18-16	40.000	$20.17
27	30562007F	ROSEMONT RANCH	15-18-16	40.000	$20.17
28	30562007G	ROSEMONT RANCH	15-18-16	70.590	$35.64
29	30562007H	ROSEMONT RANCH	15-18-16	160.000	$80.97
30	30562008C	ROSEMONT RANCH (HIDDEN VALLEY)	21-18-16	60.150	$23.35
31	30562008F	ROSEMONT RANCH (HIDDEN VALLEY)	21-18-16	35.060	$380.24
32	30562008G	ROSEMONT RANCH (HIDDEN VALLEY)	21-18-16	5.010	$472.95
33	30562008H	ROSEMONT RANCH (HIDDEN VALLEY	21-18-16	24.880	$12.99
34	30562008J	ROSEMONT RANCH (HIDDEN VALLEY)	21-18-16	35.270	$13.70
35	30562009A	ROSEMONT RANCH	23-18-16	160.000	$92.38
36	30562011A	ROSEMONT RANCH	27-18-16	40.000	$23.01
37	30562012A	ROSEMONT RANCH	32-18-16	20.000	$13.70
38	30562012C	ROSEMONT RANCH	32-18-16	180.000	$3,602.68
FEE OWNED (ASSOCIATED) TOTALS				2319.160	$7,330.01

Page 30-27

	Rosemont Project
	Form 43-101F1 Technical Report

A2-1 Permits and Authorizations

Details on permit status and authorizations for current project activities.

PERMIT AND AUTHORIZATIONS TABLE
Permit	Lead Agency and Description	Submittal Date	Status	Issue Date	Term
Federal Permits and Authorizations Issued
MSHA ID Number	Mine Safety and Health Administration	Paperwork filed for the Rosemont Copper Project	Issued 02-03256	July 21, 2010
Hazardous Waste Identification Number	Environmental Protection Agency (EPA) – Issued for hazardous waste can be generated and transported off site in quantities in excess of 100 pounds.	Submitted Requires a contingency plan.	Received RCRA EPA ID Number: AZR000509976	Sept 14, 2011	Life of the facility
State Permits and Authorizations Issued
Groundwater withdrawal permits	ADWR – Groundwater withdrawal rights	Mineral Extraction Right	Issued Permit No. 59- 215979.0000	Jan 18, 2008	20 years
Dam Safety Permit	ADWR – Regulates the construction and operation of large containment structures	Storage Analysis submitted Feb 17, 2012 for review by State Engineer	No permit necessary as designed
Well Drilling Permit	Arizona Department of Water Resources (ADWR) – Issued anytime drilling may intercept water table	On-going submittals for mineral exploration, geotechnical, and hydrologic investigation activities	Issued for current, still needed for future activities Issued for current wells 55-225120 55-225121 55-225122 55-225123 55-225124 55-225125 55-225126 55-225127		Until well or borehole closed
Water Right for Hydrologic testing	ADWR	Issued for production water well testing	Expired – June 25, 2016 55-225120 55-225121	June 16, 2015	One year

Page 30-28

	Rosemont Project
	Form 43-101F1 Technical Report

PERMIT AND AUTHORIZATIONS TABLE
Permit	Lead Agency and Description	Submittal Date	Status	Issue Date	Term
Class II Air Permit	ADEQ	Application submittal on Nov 23, 2011, Activities must follow dust control plan	Issued Permit no. 55223	Jan 31, 2013	Five years
Aquifer Protection Permit	Arizona Department of Environmental Quality (ADEQ) – Groundwater discharge permit (includes Landfill)	Application submitted to ADEQ February 27, 2009, Administratively complete on May 21, 2009	Issued Permit No. P- 106100 Place ID 135845 LTF 49639	April 3, 2012	Life of facility once issued
Type 2.02 General Permit	ADEQ – Intermediate Stockpile Permit	Submitted for three stockpiles	Issued Inventory 106100 LTF(s) 54136, 54138 and 59138	Feb 8, 2012 and  Jun 12, 2013 All revised May 26, 2016	7 years May 2, 2018 and May 14, 2020
Type 3.03 General Permit	ADEQ – Vehicle/Equipment Wash	Submitted for Southwest Energy nd Vehicle Wash 2 Quarter 2011	Issued Inventory 106100 USAS No 509976-03 LTF 64358 Place ID 19883	Sep 12, 2011 Renewal form Jul 11, 2016	5 years Sept 12, 2021
Stormwater permit	ADEQ – Regulate stormwater discharge quality	AZPDES MSGP for all onsite activities	Covered AZMSG 2010- 003 AZMSG- 74939	Feb 7, 2013	Five years (2015 or until new general permit issued)
Construction Stormwater General Permit	Arizona Department of Environmental Quality (ADEQ) – Issued for construction activities	The Notice of Intent (NOI) was received by ADEQ on July 10, 2015 – inactive and unstaffed status	Covered AZCON 86646	Jul 10, 2015	Dec 31, 2018
401 Certification	ADEQ	Application submitted Jan 12, 2012 Surface Water Mitigation Plan accepted Dec 2014	Issued Application NO SPL-2008- 00816-MB ADEQ LTF 55425	Feb 3, 2015	No expiration

Page 30-29

	Rosemont Project
	Form 43-101F1 Technical Report

PERMIT AND AUTHORIZATIONS TABLE
Permit	Lead Agency and Description	Submittal Date	Status	Issue Date	Term
Arizona Mined Land Reclamation Permit	Arizona State Mine Inspector – Permit for reclamation activities at a site.	Submitted planned May 2008	Complete	July 10, 2009	Life of facility – annual updates
Start-up Notice for Mine Operations	Arizona State Mine Inspector	Registers mine with Arizona State Mine Inspector	Filed	Sept 9, 2009
Agricultural Land Clearing Permit	Arizona Department of Agriculture – Permit to clear land	Submittal prior to construction of facilities or disturbance of state- protected native plants	As needed
Certificate of Environmental Compatibility	Arizona Corporation Commission and the Line Siting Committee	Application for power line route submitted by Tucson Electric Power on November 2	Issued (to TEP) Docket No. L-00000C-11- 0400-00164 Case No. 164	CEC signed Dec 19, 2011 Approved - Mar 21, 2012 Amended - Jun 12, 2012	Life of facility (Certificate expires in 7 years unless line carries power.)
County Permits and Authorizations Issued
Pima County Flood Control District Permit	Pima County Flood Control	Submitted planned Nov 2013	Issued and renewed FPUP 13-640 (original) FPUP 15- 170RP (renewal)	May 2, 2016	One Year from date of issue
Town Permits and Authorizations Issued
Right of Way Encroachment	Town of Sahuarita	This is a 35’ encroachment is for a waterline along Santa Rita Road	Issued License Contract No. CO13-0029	June 24, 2013	none

Page 30-30

	Rosemont Project
	Form 43-101F1 Technical Report

PERMIT AND AUTHORIZATIONS TABLE
Permit	Lead Agency and Description	Submittal Date	Status	Issue Date	Term
Federal Permits and Authorizations – NOT Issued
Mine Plan of Operation	Forest Service – This permit is needed for mining operations on public lands. Because all actions are connected to EIS produced for this the MPO is needed to start ground disturbing activities	Plan was submitted in sections: •    Mine Plan of Operations       and supporting       documents        – July 11, 2007 •    Reclamation Plan –       August 7, 2007 •    Infrastructure Plan –       July 25, 2007	Sufficient for NEPA Analysis, NOI issued in FR Mar 13, 2008 Final EIS issue Nov 2013 Draft ROD issued Dec 2013 Supplemental BO issued Apr 22, 2016		Life of mine
404 permit	Army Corps of Engineers – Allow operations in Waters of the U.S. There is no formal Corps delineation to date Needed for powerline, waterline, access road and discharge activities	Preliminary Jurisdictional Delineation submitted in May 2009 Application for 404 permit re-submitted October 11, 2011 Habitat Mitigation and Management Plans submitted – April to Aug 2014	Pubic Notice published Dec 5, 2011
Hazardous Materials Transportation Permit	Department of Transportation (DOT) – Permit needed to transport or received “hazardous materials” under the DOT definitions.	Registration and plan only - will be dependent upon construction schedule but number should be in place prior to construction start.			1-3 years dependent upon permit
Radio Licenses	Federal Communications Commission (FCC)	License already exists for current on- site use, additional uses will require additional licenses that will be managed by the contractor (Empire)	Issued for current use – need larger project use license
Blasting License	Bureau of Alcohol Tobacco and Firearms	Submittal will be dependent upon development of on- site facilities and blasting personnel

Page 30-31

	Rosemont Project
	Form 43-101F1 Technical Report

PERMIT AND AUTHORIZATIONS TABLE
Permit	Lead Agency and Description	Submittal Date	Status	Issue Date	Term
State Permits and Authorizations – NOT Issued
State Land Right of Way	Arizona State Land Department	Applications Submitted in Nov 8, 2010 All studies, surveys, and appraisals complete	30-day notice letters issued in August 2012 awaiting auction
Encroachment Permit	Arizona Department of Transportation	Traffic Impact Assessment was submitted in July 2010 and the Encroachment Permit application submitted Feb 28, 2011	Permit in draft, issuance keyed on ROD
AZPDES DeMinimis Storm water Permit	ADEQ	Submitted as needed for project – well development, drilling, etc.	Issued as necessary and closed when not in use
Septic System APP	ADEQ – Onsite Wastewater Permit	Awaiting design and percolation testing			Life of facility (will be combined with facility- wide APP)
Hazardous Waste Identification Number	ADEQ – Issued by EPA/ADEQ so hazardous waste can be generated and transported off site in quantities in excess of 100 pounds.	RCRA EPA ID Number: AZR000509976 Requires a contingency plan.	Issued – need to register number with State when activities start		Life of the facility
County Permits and Authorizations – NOT Issued
Hazardous Waste Management	PCDEQ – registration with PCDEQ for all EPA ID Nos.	RCRA EPA ID Number: AZR000509976	Issued – need to register number with County when activities start		Life of facility
Drinking Water System Registration	PCDEQ – system plans need to be approved prior to installation, registration for all non-community non-transient drinking water systems	Submittal prior to construction of system. Sampling and emergency plans required.

Page 30-32

EX-99.6

CASHEL MEAGHER

CERTIFICATE OF QUALIFICATION

Re: Rosemont Project Technical Report, March 30, 2017

I, Cashel Meagher, B. Sc., P. Geo, of Toronto, Canada, do hereby certify that:

1.	I am currently employed as Senior Vice President and Chief Operating Officer, with Hudbay Minerals Inc., 25 York St, Suite 800, Toronto Ontario.
2.	I graduated from Saint Francis Xavier University with a Joint Advanced major in Geology and Chemistry in 1994.
3.	I am a member in good standing with the Association of Professional Geoscientists of Ontario, member #1056.
4.	I have practiced my profession continuously over 20 years and have been involved in mineral exploration, project evaluation, resource and reserve evaluation, and mine operations in underground and open pit mines for base metal and precious metal deposits in North and South America.
5.	I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education and affiliation with a professional association and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purpose of NI 43-101.
6.	I have reviewed and approved the Summary of the Technical Report and I am responsible for the preparation of this Technical Report titled “NI 43-101 Technical Report, Rosemont Project, Pima County, Arizona, USA” (the “Technical Report”), dated and effective as of March 30, 2017.
7.	I last visited the property on April 21, 2016. I also visited it several times prior to that date.
8.	As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
9.	I am not independent of the Issuer. Since I am an employee of the Issuer, a producing issuer, I fall under subsection 5.3(3) of NI 43-101 where “a technical report required to be filed by a producing issuer is not required to be prepared by or under supervision of an independent qualified person”.
10.	I have been directly involved with the Rosemont Project property, which is the subject of the Technical Report, continuously since January, 2016.
11.	I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with the instrument and form.
12.	I consent to the public filing of the Technical Report with any stock exchange, securities commission or other regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated as of March 30, 2017.

Original signed by:
Cashel Meagher
Cashel Meagher, P.Geo., Senior Vice President & Chief Operating Officer, Hudbay

EX-99.7

CONSENT OF QUALIFIED PERSON

March 30, 2017

British Columbia Securities Commission
Alberta Securities Commission
Financial and Consumer Affairs Authority of Saskatchewan
The Manitoba Securities Commission
Ontario Securities Commission
Autorité des marchés financiers
New Brunswick Financial and Consumer Services Commission
Nova Scotia Securities Commission
Office of the Superintendent of Securities, Prince Edward Island
Office of the Superintendent of Securities, Service Newfoundland and Labrador
Office of the Superintendent of Securities, Northwest Territories
Office of the Yukon Superintendent of Securities
Nunavut Securities Office

Dear Sir/Madam:

Re: Technical Report entitled “NI 43-101 Technical Report, Feasibility Study, Updated Mineral Resource, Mineral Reserve and Financial Estimates, Rosemont Project, Pima County, Arizona, USA” dated and effective as of March 30, 2017

I, Cashel Meagher, consent to the public filing of the technical report titled “NI 43-101 Technical Report, Feasibility Study, Updated Mineral Resource, Mineral Reserve and Financial Estimates, Rosemont Project, Pima County, Arizona, USA” dated and effective as of March 30, 2017 (the “Technical Report”) by Hudbay Minerals Inc. (“Hudbay”).

I also consent to the use of any extracts from, or a summary of, the Technical Report in the annual information form for the year ended December 31, 2016, dated March 30, 2017 (the “AIF”) of Hudbay.

I certify that I have read the AIF and that it fairly and accurately represents the information in the Technical Report for which I am responsible.

(signed) Cashel Meagher

Cashel Meagher P. Geo.

Senior Vice-President and Chief Operating Officer, Hudbay