BOLEO PROJECT
	MINERA Y METALURGICA DEL BOLEO, SA DE CV
	FEASIBILITY STUDY SUMMARY REPORT UPDATE

1 SUMMARY

1.1 Geology & Mineral Resource Modelling

The Boleo Cu-Co-Zn-Mn deposit is located near the town of Santa Rosalía, Baja California Sur, Mexico.

The deposit, which occurs in the El Boleo Formation, comprises seven mineralized units called “Mantos” (manto is a Spanish term used in mining parlance for a generally mineralized layer or stratum).  The mineralized mantos dip gently to the east but faulting, which is common throughout the project area, produces a step-like pattern in the present position of the mantos.

The deposit was continuously mined predominantly by underground methods from 1886 to 1972, during which time an estimated 18 Mt of ore were treated. After 1972 both underground and open pit mining was carried out sporadically until the copper smelter at Santa Rosalía closed in 1985.  Since cessation of mining operations International Curator Ltd (Curator) carried out exploration between the years 1993 to 1998 and Minera y Metalúrgica del Boleo S.A. de C.V. (MMB) has been carrying out exploration and project development activities from 2001 to the present. When MMB re-acquired the property in 2001 it also acquired the Curator drill hole assay database which has been used in the current resource study.

Independent geologic consultants, Wardrop Engineering Inc.(Wardrop), were engaged in 2009 to update the geologic model and resource estimates of the Boleo deposit in accordance with NI 43-101 standards. Updated resource estimates include all new drilling data.  This was an update from previous work performed by Hellman & Schofield (H&S) of Sydney, Australia.

The deposit will be mined by a combination of underground and open pit mining methods and as planning and operating requirements are quite different, separate models were developed for each mining method.

A gridded block model for each Manto was developed for open pit mine design in preference to a gridded seam model because it allows greater versatility to evaluate mining methods at various cut-off grades.  Block dimensions used were 50 m EW x 50 m NS and 0.9 m vertical.

Gridded pseudoseam models incorporating minimum and maximum mining height considerations have been created from the block model data for Mantos 1, 2, 3, and 4 for underground mine planning. Block dimensions used were 50 m EW x 50 m NS with a variable vertical height based on cut-off grade and mining constraints.

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Resource classification was determined by a combination of the number of composite data sets available for grade estimation and distance to data source employing search ellipses of increasing dimensions.

Table 1-1: Block Model Mineral Resource Estimate at 0.5% CuEq Cut-off

	Cut-off CuEq (%)	Tonnes (106)	Cu (%)	Co (%)	Zn (%)	Mn (%)	CuEq (%)
Measured	0.50	85.82	0.82	0.074	0.50	3.04	1.59
	0.75	73.31	0.93	0.081	0.52	3.20	1.75
	1.00	59.83	1.07	0.088	0.53	3.29	1.95
Indicated	0.50	178.86	0.74	0.053	0.71	3.32	1.46
	0.75	143.88	0.87	0.059	0.77	3.55	1.66
	1.00	112.55	1.03	0.065	0.83	3.65	1.88
M+I	0.50	264.67	0.76	0.060	0.64	3.23	1.50
	0.75	217.20	0.89	0.067	0.69	3.43	1.69
	1.00	172.38	1.04	0.073	0.72	3.53	1.91
Inferred	0.50	159.85	0.47	0.045	0.70	2.93	1.15
	0.75	109.48	0.59	0.052	0.87	3.42	1.39
	1.00	77.52	0.70	0.058	0.99	3.66	1.61

Table 1-2: Pseudo-seam Model Resource Estimate at 0.5% CuEq Cut-off

	Cut-off CuEq (%)	Tonnes (106)	Cu (%)	Co (%)	Zn (%)	Mn (%)	CuEq (%)
Measured	0.50	60.67	1.07	0.083	0.52	3.19	1.89
	0.75	57.44	1.12	0.086	0.52	3.22	1.96
	1.00	52.30	1.20	0.090	0.51	3.21	2.07
Indicated	0.50	129.37	0.89	0.056	0.73	3.38	1.63
	0.75	112.77	0.99	0.061	0.76	3.51	1.78
	1.00	95.66	1.11	0.065	0.79	3.56	1.94
M+I	0.50	190.04	0.94	0.065	0.66	3.32	1.71
	0.75	170.21	1.03	0.069	0.68	3.41	1.84
	1.00	147.97	1.14	0.074	0.69	3.44	1.99
Inferred	0.50	95.35	0.60	0.050	0.76	3.13	1.33
	0.75	77.85	0.68	0.055	0.83	3.38	1.50
	1.00	60.88	0.78	0.061	0.91	3.55	1.67

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1.2 Metallurgy & Process Design

A hydrometallurgical process has been developed that is suitable for the recovery of Cu, Co, Zn and possibly Mn from mined and blended Boleo ores.  The metal recovery circuits and equipment envisaged for the Boleo process plant are typical of those deployed in numerous operations worldwide.

The proposed plant for the process consists of crushing and milling the ore in sea water followed by atmospheric leaching of the whole ore stream at elevated temperatures.  The leach circuit consists of both an oxidative and a reductive leach to ensure the recovery of both oxide and sulphide forms of Boleo’s pay metals.

The leached metals are separated from the leach slurry in a counter current decantation (CCD) washing circuit.  The dissolved metals are recovered at high efficiency from the wash solution and concentrated in four separate solvent extraction circuits, two electrowinning circuits and a fluid bed drying operation to produce high-quality copper and cobalt metal cathode and zinc sulphate monohydrate crystals respectively.  A key feature of the process is the use of direct solvent extraction technology (DSX®) in the second solvent extraction circuit to separate Mn from Co and Zn in the process solutions.

The process plant is being designed to treat 3.1 Mdt/a of mined ore and produce product at the following rates of production:

· up to 62,000 t/a of copper cathode

· up to 2,400 t/a of cobalt cathode

· up to 37,000 t/a of zinc sulphate monohydrate salt.

In addition, there is the potential to produce up to 220,000 t/a of a manganese carbonate powder (up to 100,000 t/a of contained Mn) via the relatively simple addition of a small number of unit operations to the proposed plant.

The process and individual unit operations have been tested successfully in extensive testwork campaigns over the past six years and the development of the process plant has progressed to a detailed design stage.  There is therefore a high level of confidence that the metallurgical process development initiatives undertaken to date have been successful and are appropriate for the design and eventual operation of the plant.

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1.3 Mining 

Based on resource models by Wardrop, mine engineering consultant, Agapito Associates Inc (AAI) prepared a 23 year life of mine plan and schedule using mainly underground mining methods supplemented on an ongoing basis with ore mined from surface operations.  A number of different mining methods were evaluated before options of using predominantly underground mining operations supported by a number of small, open cut mines was selected as the technique which would best deliver the targeted plant feed grade at the required rate. Approximately 95% of all ore processed over the current life of mine is from underground production and the remaining 5% is from surface mining.

In addition to progressing the optimization of mine planning and scheduling AAI also prepared mine designs, evaluated equipment requirements, estimated production rates and prepared the initial mine capital cost and provided the basis for the mine operating and sustaining capital cost estimates which were subsequently adjusted by Baja Mining in certain scheduling cost categories for this update to the Feasibility Study.

The seam formation and low material strength of the mantos suggested conventional “soft rock” underground mining methods such as used in coal, potash, or salt mining would be appropriate for Boleo underground conditions and accordingly a number of methods were examined for their suitability.  Longwall mining was discounted due to the faulted and dipping manto structure and high initial capital cost.  Shortwall mining was also discounted when efficient mine layouts could not be readily designed and accommodated within the complex, fault disturbed structure of mantos.  After extensive study, room-and-pillar mining using continuous miners was selected because the method requires relatively low capital cost and can accommodate variations and undulations in the footwall as well as variations in mining height. Moreover, room and pillar mining can achieve relatively high productivity and panels can be laid out in such a way that the mineral recovery from relatively small, fault bounded mining blocks can be maximized.

The resource pseudoseam model was used to define areas where room and pillar mining could be carried out. The criteria used for this determination were the following;

· Minimum mining height of 1.8 m to allow working room for the machines.  If the economic thickness of the manto was less than this it was diluted by the lower grade blocks above up to 1.8 m height.

· Maximum mining height of 4.2 m, matching the proposed reach of the continuous miner.  Economic blocks above this height were ignored.

· The composite copper equivalent grade of the manto over the mining height must exceed a Cut-off grade of 0.5% CuEq.  (For mining purposes, the copper equivalent Cut-off grade was calculated using metal prices of $1.50/lb Cu; $15.00/lb Co; $1.20/lb Zn and plant recoveries for Cu, Co and Zn of 91.2%, 78.5 and 65.5% respectively.)

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	MINERA Y METALURGICA DEL BOLEO, SA DE CV
	FEASIBILITY STUDY SUMMARY REPORT UPDATE

· An allowance for voids in old works and recovery of “retaque”, or previously mined material, was made in terms of both recovery and density.

Underground mining trials to test equipment, working methods and geotechnical ground responses to the chosen method were undertaken in two stages in the years 2005 to 2006 under the supervision of Australian Mine Design and Development (AMDAD) and AAI.  These tests confirmed the suitability of the proposed continuous mining method and provided valuable field data and information concerning ground behaviour during mining and methods of ground support.

1.4           Economic Assessment

A financial model was created utilizing the Wardrop resource model and the AAI mining schedule, capital and operating costs as set out herein and US Securities and Exchange Commission (SEC) guidelines for metal prices.  Financial modelling, using the Case 1 prices and 23 years for the project life, shows that the project could generate a net after- tax Internal Rate of Return (IRR) of 25.6% and an after tax NPV (using 8% discount rate) of $1,306 million.

In addition, sensitivity analysis was also conducted at various metal prices.  The project is sensitive to four key variables; copper price; cobalt price; capital costs and operating costs.  The sensitivity of the After-Tax IRR and NPV (at 8% discount rate) relative to the Case 1 is shown in Table 1-3 to indicate the effect of a plus or minus 10% change in the key variables.

Table 1-3: Project Economic Sensitivities Summary

	After Tax IRR			NPV @ 8% ($Millions)
	-10%	Case 1	+10%	-10%	Case 1	+10%
Copper price	23.2%	25.6%	27.9%	$1,111	$1,306	$1,505
Cobalt price	25.0%	25.6%	26.1%	$1,246	$1,306	$1,367
Capital cost	27.7%	25.6%	23.7%	$1,361	$1,306	$1,250
Operating cost	26.3%	25.6%	24.8%	$1,379	$1,306	$1,234

This modelling, based on the current mine schedule, indicates that the economic sensitivities confirm the financial attractiveness of the project.

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	MINERA Y METALURGICA DEL BOLEO, SA DE CV
	FEASIBILITY STUDY SUMMARY REPORT UPDATE

2 INTRODUCTION

This document is being issued as an update to the NI 43-101 compliant technical report entitled “El Boleo Project Feasibility Study Summary Report” dated July 11, 2007 that was issued (and filed on SEDAR) by the Company in support of Boleo Project Feasibility Study.  The study, as issued, was current at the time and, as stated in it, some of the estimates of Capital Costs were based on quotes received in mid-2006.  Since issuing the Feasibility Study, Baja Mining Corp. (the “Company”) has advanced its efforts to secure financing for construction of the project. Recognizing potential changes in construction costs since completion of the feasibility study and in contemplation of entering into a contract for the engineering, procurement and construction management (“EPCM) of the project facilities, the Company elected to bring all of its capital and operating cost estimates up to current dates.  This update incorporated revised capital and operating costs, as of the fourth quarter of 2009, as well as an update to the construction schedule, prices of consumables and products available for sale.

Most sections of this report remain unchanged from the Feasibility Study.  Specifically Sections of the original Feasibility Study, with reference to Location, History, Geology, Environment and Community Development, required no change from the original Feasibility Study.  Updates made to all the other sections included all available new information.  The economic assessment of the project has been updated to include changes made to all factors of future construction and operation activities expected for the Boleo Project.  There include but are not limited to capital costs, construction schedule, and prices of consumables and metal products.

This report incorporates capital and operating costs prepared by the Company.  A revised geological model, in accordance with NI 43-101, has been prepared by Wardrop Engineering (Wardrop) and used by Agapito Associates, Inc. (AAI) to prepare the current mine plan.  The initial capital cost for the mine was prepared by AAI, however mine sustaining capital and operating costs were prepared by the Company.

All qualified persons noted have inspected the property as per NI 43-101 requirements.

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3 RELIANCE ON OTHER EXPERTS

A number of specialist organisations, experts and consultants have contributed to defining and assessing the feasibility and economics of the Boleo project.

The capital cost for development of the Boleo Project has been developed by a number of specialist organizations.  These organizations are listed below in a table that summarises areas of significant capital cost and the organizations responsible for development of capital costs for these respective areas.

Major Cost Area	Consultant	Location
Open Cut Mining	Agapito Associates Inc	Grand Junction, Colorado
Underground Mining	Agapito Associates, Inc	Golden, Colorado
Mining Surface Infrastructure	Wardrop Engineering, AMEC Earth & Environmental	Vancouver, Canada Denver, Colorado
Tailings Dam	Arcadis Geotecnica, Klohn Crippen Berger Inc Amec Earth & Environmental.	Santiago, Chile Vancouver, Canada Denver, Colorado
Co-Generation Plant	SNC Lavalin	Toronto, Canada
Acid Plant	Fenco Pty Ltd	Toronto, Canada
SO2 Gas Production Facility	Noram Engineering & Constructors Ltd	Vancouver, Canada
Barging Facility / Operations Terminal	Westmar	Vancouver, Canada
Sulphur Infrastructure	ICEC Canada Ltd	Calgary, Canada

Operating costs were developed by the Company working in conjunction with suppliers of services, consumables, materials and equipment.  The following table lists a number of the organisations that assisted the Company in preparing estimates of operating costs.

Major Cost Area	Consultant	Location
Mining Operating Costs	Agapito Associates Inc	Grand Junction, Colorado
Process Consumable Cost	Emmett Consulting FLSmidth Bateman Litwin Bateman Engineering	Tuscon, Arizona Bethlehem, Pennsylvania Amsterdam, Netherlands Perth, Australia
Laboratory Operating Cost	GMAR Development	Nevada, USA

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	MINERA Y METALURGICA DEL BOLEO, SA DE CV
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With the exception of Agapito Associates the consultants listed above cannot be classified as qualified persons for the purposes of this report.  The Company has relied on the consultants listed above for the generation of capital and operating estimates in their particular areas of expertise.  The initial capital costs of the mine were prepared by Agapito Associates, however mine sustaining capital and operating costs were prepared by the Company.

The portions of the report to which the above disclaimer applies are Section 19.1.3, Capital Cost Estimate and Section 19.1.4, Operating Cost Estimate.

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4 PROPERTY DESCRIPTION & LOCATION

4.1 Location

The Boleo Project is located along the east coast of the Baja peninsula centred on the port town of Santa Rosalía in Baja California Sur, Mexico (Figure 4-1).  The town is approximately 850 km south of San Diego, California, USA.  Coordinates for the project are Latitude 27°14' to 27°25' N, Longitude 112°14' to 112°22' W.

Figure 4-1: El Boleo Location Map

1 Information based on Mehner 2003, reviewed and updated by T. Albinson (MMB) Nov 2004.  Maps supplied by MMB.

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4.2 Description

The Boleo property consists of 25 total mineral concessions covering 20,490.9 ha, of which 24 concessions are contiguous.

The San Bruno concession is not contiguous and is located 30 km south of St. Rosalía in the San Bruno basin area.  The titled concessions are listed in Table 4-1 and shown in Figure 4-2.

Table 4-1: El Boleo Property –Concessions as of September 2009

Claim	Title No.	Surface Area (ha)	Date Initiated	Expiry Date	Annual Taxes (MX$)
El Boleo	218082	4,975.6132	29-Sep-2000	28-Sep-2050	285,004
El Boleo I	218092	72.4463	31-Aug-2000	30-Aug-2050	4,150
El Boleo19, 256.1872 II fracc. Uno	218179	1,296.6156	29-Sep-2000	28-Sep-2050	74,270
El Boleo II Fracc. Uno A	218180	507.2841	29-Sep-2000	28-Sep-2050	29,058
Boleo III	212148	224.6410	31-Aug-2000	30-Aug-2050	12,868
Nuevo San Luciano	214189	150.0000	10-Aug-2001	09-Aug-2051	4,272
Boleo II Fracc. IV	218975	267.1579	28-Jan-2003	27-Jan-2053	7,608
Boleo X Fracc. 5	211055	1.3829	24-Mar-2000	23-Mar-2050	80
Boleo X Fracc. 8	211058	3.9486	24-Mar-2000	23-Mar-2050	226
Boleo X Fracc. 9	211059	9.9612	24-Mar-2000	23-Mar-2050	570
Boleo X Fracc 12	211062	3.1241	24-Mar-2000	23-Mar-2050	178
Boleo X Fracc 16	211066	0.0068	24-Mar-2000	23-Mar-2050	2
Biarritz B	219819	0.0055	16-Apr-2003	15-Apr-2053	2
San Luciano 2	220740	670.0000	30-Sep-2003	29-Sep-2053	9,220
San Luciano 3	221073	1,899.0000	19-Nov-2003	18-Nov-2053	26,130
San Bruno	222772	8,783.0000	27-Aug-2004	26-Aug-2054	120,854
San Luciano 4	223358	392.0000	03-Dec-2004	02-Dec-2054	5,394
San Luciano 5	229314	262.9861	04-Apr-2007	03-Apr-2057	2,420
Boleo II Fraccion IV A	218967	148.0000	28-Jan-2003	27-Jan-2053	9,360
Nuevo Biarritz	225299	576.2215	12-Aug-2005	11-Aug-2055	8,758
El Boleo X Fracc. 1	211051	114.6770	24-Mar-2000	23-Mar-2050	14,500
El Boleo X Fracc. 7	211057	19.5073	24-Mar-2000	23-Mar-2050	2,466
El Boleo X Fracc. 14	211064	40.0431	24-Mar-2000	23-Mar-2050	5,064
El Boleo X Fracc. 15	211065	73.2996	24-Mar-2000	23-Mar-2050	9,268
Biarritz A	219820	0.0096	16-Apr-2003	15-Apr-2053	2
Total (MX$)		20,490.9314			631724
Total (US$)					58,430

Note: Exchange rate used above is MX$13.3721 = US$1

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The project includes three surface lots currently totalling 6,692.6 ha as shown in Table 4-2 and Figure 7-1.  Baja is currently consolidating the various surface lots into a single master deed and as part of this process, is streamlining its final boundaries with other property owners, and expects the final surface property size to change slightly from the above-mentioned amount.

Table 4-2:            El Boleo Project, Surface Property, and Annual Taxes

Surface Lot	Size (ha)	Annual Taxes
El Boleo	6,553.55	21,604
Soledad Property	99.91	23,786
San Luciano Property	39.12	5,301
Total (MX$)	6,692.58	50,691
Total (US$)		4,688

The surface property also contains one rented tract totalling 583.7 ha.  This tract is Ejido land located on the western edge of the Project Area and covers a portion of the anticipated maximum extent of the tailings disposal impoundment.  The rented period is 30 years and commenced on 18 April 2007.  The total amount paid for the entire lease life was US$35,000.  There is both a renewal option and the obligation of both parties to continue to work towards an outright purchase.

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Figure 4-2: El Boleo Property Map

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4.3 Ownership

The mineral concessions covering the Boleo copper-cobalt-zinc-manganese deposit are 100% owned by MMB, a Mexican company involved in mineral exploration and development.  MMB is owned as to 70% by Baja, a Toronto Stock Exchange (TSX) listed company, symbol BAJ, and 30% by a consortium of Korean companies consisting of Korea Resources Corporation, LS-Nikko Copper Inc., Hyundai Hysco Co., Ltd., SK Networks Co., Ltd., and Iljin Copper Foil Co., Ltd. (together referred to as the “Korean Consortium”).

4.4 Taxes and Assessment Work Requirements

4.4.1 Taxes

Total annual fees payable in September 2009 as of this report are MX$631,724 for the mineral concessions and MX$50,691 for the surface leases, or using the exchange rate as of 30 September 2009 (13.3721 MX$/US$), US$58,548 for the mineral concessions and US$4,688 for the surface lots (Table 4-1 and Table 4-2).  The calculated annual fees are based on the latest published government tax guides.  It is expected that during operation there will be ongoing taxes such as change in land use and federal zone taxes.  These have not been estimated at this stage.

4.4.2 Work Requirements

Work obligations on the property (known in Mexico as “Informes de Comprobaciones de Obras”) are in good standing.  Based on past work expenditures of approximately US$22 million (M), enough credits have been accrued to keep the property in good standing until 2050.

4.4.3 Option Payments

There are no royalties payable on the properties and there are no other agreements or encumbrances.

4.5 Permits and Liabilities

4.5.1 Permits

During 2006, MMB successfully completed a full Environmental Impact Assessment that covers the construction, operation, and closure phases of the Boleo Project.  Given the complexities of the project itself and the environmental sensitivity surrounding the project location, the Mexican Federal environmental agency, Secretaria de Medio Ambiente y Recursos Naturales, (SEMARNAT) requested the submittal of an Environmental Impact Manifest with a regional scope.

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Following SEMARNAT’s request additional fieldwork was undertaken to fully characterize the regional area of influence of the project.  The evaluation process also included a request from SEMARNAT to submit additional information relating to the project to better clarify the identified environmental impacts.  This request was given to MMB on July 5th, 2006.  The information was formally filed on October 2nd, 2006.

Finally, after incorporating the observations and recommendations from the National Commission for Natural Protected Areas: the Secretariat for Urban Planning, Infrastructure, and Ecology of the State Government of Baja California Sur and the Municipal Presidency of Mulegé at Baja California Sur, the environmental impact resolution was issued on November 27, 2006 and delivered to MMB on December 7, 2006.  This resolution authorizes the construction, operation, and closure of the Boleo Mining Project.  The official document number containing this resolution is S.G.P.A.-DGIRA.-DDT.-2395.06 and is signed by the General Director for Environmental Impact and Risk (DGIRA).

This authorization allows MMB to initiate the procedures to obtain permits that are more specific.  Work is proceeding in securing these additional permits and in managing the terms and conditions that were established in the environmental impact authorization.

4.5.2 Liabilities

There are no outstanding liabilities associated with the property.  The drilling and metallurgical sampling programs of International Curator in 1997 to 1998 and MMB in 2004 to 2007 caused the most recent disturbances, including the development and operation of the underground test mine.  All of the disturbed areas from 1997 to 1998 were remediated without any assessed environmental liabilities from that period.  The 2006 / 2007 MMB drilling program work has been remediated and no liabilities have been recorded at this time and no liabilities are expected to be incurred from this work going forward.  The test mine has been incorporated into the EIM permit as an identified disturbed area allowed under the permit.

The project is located within the “buffer zone” of the Vizcaino Biosphere Reserve, which is centered on the Desierto de Vizcaino on the west central coast of the Baja.  The Biosphere extends south to encompass the historic El Boleo Mining District and serves to protect certain environmental and cultural features in the town of Santa Rosalía.  It is believed the Biosphere is intended to protect the historic buildings dating from the late 1800’s associated with the early mining of the Boleo district.  It should be noted these buildings are outside the Boleo Project and study area boundary and will not be directly affected by project development.

Since the Boleo district has been mined for copper and cobalt since 1872, and with two large gypsum quarries currently operating in the region, the authorities have designated the local land suitable for mining and have established land management directives within the Biosphere for development.  The area within the reserve is therefore managed relative to a specific land usage description.

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There are no tailings ponds on the lands owned by MMB or on the referenced concessions.  There are 88 small mine waste dumps located at the portals of historic mine workings.

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5 ACCESSIBILITY, CLIMIATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Access

The Boleo property is located on tidewater on the east coast of the State of Baja California Sur, Mexico, county of Mulegé, adjacent to the town of Santa Rosalía.  Access for construction equipment would principally be by the Trans-peninsular highway extending some 850 km south from the US border.  This highway passes through Santa Rosalía and carries heavy traffic volumes year round.  Heavy construction equipment and project supplies would also be brought in by barge to the port of Santa Rosalía and the Pacific Coast marine facilities at Guerro Negro.  There are regular scheduled air services from the United States and mainland Mexico to both Loreto, which is a two hour drive to the south and La Paz which is a six hour drive to the south.  The closest private airstrip is at Palo Verde a half hour drive away.  Port facilities, which serviced the copper mine until 1985, are still being used twice a week by a ferry service to the mainland at Guaymas.

5.2 Climate

The project area is immediately adjacent to the Gulf of California, with a climate typical of the Sonora desert region with warm to hot temperatures and minimal seasonal precipitation.  Rainfall is confined mainly to heavy cloud bursts at intervals of several years during tropical cyclones.  Mining operations can be scheduled 365 days per year (d/a) save for rare heavy rainfall events, which occur every 10 years on average.

5.3 Local Resources

Santa Rosalía has a population of 10,000 people and services a large fishing fleet, fish processing facilities and two open pit gypsum mines.  A socio-economic study prepared in 1997 showed that almost all of the non-staff positions could be filled from the local population.

According to the most recent Mexican census, nearly 40,000 people live in Mulegé County, which is over 10 percent of the Baja California Sur population.  Population concentrations within the county include Santa Rosalía, Santa Agueda, Mulegé (town), Guerrero Negro, San Ignacio, Bahía Tortugas, San Marcos Island, Gustavo Diaz Ordaz and Bahía Asunción.  Santa Rosalía is the largest town in the county.

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Hotel accommodation, gasoline, groceries and various hardware goods can all be purchased in Santa Rosalía.  Other items including machinery and trained personnel are readily available from mainland Mexico via the ferry to Guaymas or airplane through Loreto or La Paz.  Services and supplies are also obtainable from California, USA.

Fresh water for domestic and drilling purposes is scarce and mostly obtained from wells in Palo Verde, 30 km away.  The planned process plant and mines are expected to use sea water for 100% of its requirements.  A desalination plant with a capacity of 200 m3/h will be required to supply process and potable water for domestic use at the mine as well as provide sufficient fresh water for the final stages of metal production.

5.4 Infrastructure

Aside from the many kilometres of drill road built along each arroyo to access the property, the other infrastructure on the property is a partially completed camp facility to house approximately 1,000 workers, a small de salination plant ready to be commissioned, a warehouse, two environmental shade houses, several box trailers and containers, some mining and construction equipment, purchased steel and forming materials inside a fenced yard, a guardhouse for the main entrance and several environmental monitoring stations.  These were either constructed or purchased since 2002 by MMB as site improvements and will be used for ongoing development.

5.5 Physiography

Property topography is best described as mesa-arroyo with relatively flat plateau cut by deeply incised arroyo valleys resulting in rugged, steep sided valleys with arroyos that drain into the Gulf of California.  Project site elevations vary from 0 to 350 metres above sea level.

The project site is very arid with vegetation consisting of a wide variety of cactus.  Over most of the project area, vegetation is quite sparse; it does occur in significant amounts only locally along the mesa tops, a few kilometres in from the coast.

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6 HISTORY

The Boleo Property was discovered in the 1860’s by local ranchers.  Limited production ensued until acquisition by the Companie du Boleo, a French company that operated the mine from 1885 to 1938.  There were sporadic operations thereafter by parastatal corporations in Mexico until the 1980’s when it was placed the Mexican Strategic National Mining Reserve.

After being released from the National Mining Reserve on December 31, 1991, it was acquired, by staking, in July 1992 by Minera Terra Gaia S.A. de C.V. (“MTG”), a Mexican company owned by Terratech Environmental Corporation (“Terratech”), a Barbados company.  In 1993, Terratech granted International Curator Resources Ltd. (“Curator”), a TSE listed company, the option to acquire an interest in the Boleo Property.  Curator’s interest in the property was held via a Barbados company, Mintec International Limited (“Mintec), which company held all the shares Curator’s Mexican subsidiary, Minera Curator S.A de C.V. (“Minera Curator”).  In 2001, ownership of the property was reacquired by Terratech acquiring all the issued shares of Mintec (and accordingly ownership of Minera Curator) from Curator.  Subsequently MTG transferred a residual interest held by it in the property to Minera Curator; such that Minera Curator held a 100% interest in the Boleo Property.  Minera Curator subsequently changed its name to Minera y Metalurgica del Boleo S.A. de C.V. (“MMB”).  Terratech dividended the Mintec shares to its shareholders in 2003.  The shares of Mintec, and accordingly a 100% interest in the Boleo Property, were acquired by the Company in a reverse takeover by the shareholders of Mintec in 2004. For historic reference, any work carried out under the control of direction of Curator will be referred to as work by Curator.  Any work subsequent to March 2001 by Minera Curator will be referred to as under the direction of MMB to acknowledge the change of ownership of Minera Curator.

The deposit was continuously mined predominantly by underground methods from 1886 to 1972, during which time an estimated 18 Mt of ore were treated. After 1972 both underground and open pit mining was carried out sporadically until the copper smelter at Santa Rosalía closed in 1985.  Since cessation of mining operations Curator carried out exploration between the years 1993 to 1998 and Minera y Metalúrgica del Boleo S.A. de C.V. (MMB) has been carrying out project development activities from 2001 to the present.  When MMB re-acquired the property in 2001 it also acquired the Curator drill hole assay database which has been used in the current resource study.

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Table 6-1: Historical Mining Activities at El Boleo

Period of Activity	Mined (t)	Average Cu Grade (%)	Copper Produced (t)
To 1884	~54,400	24.0%	~10,400
1888 – 1947	13,622,327	4.81%	540,342
1948 – 1952	817,300	3.95%	~27,000
1953 – 1972	1,118,200	3.95%	~36,500
1973 – 1985	720,900	3.02%	~18,000
1964 – 1972	2,500,000 *	1.40%	n/a

* Estimate of tonnage mined by poquiteros. (Source unknown).

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7 GEOLOGICAL SETTING

7.1 Regional Geology

The Boleo deposits occur within the Boleo sub-basin of the Santa Rosalía basin.  This basin formed because of Miocene rifting in the Gulf of California extensional province (Figure 7-1).  The northward extension of this province is the Basin and Range province of the southwest United States.

Figure 7-1: El Boleo Geological Setting

The timing of initial rifting varies from 13 Ma to 8 Ma.  In El Boleo District, which is located near the western edge of the Gulf of California extensional province, rifting is believed to have started sometime after 10 Ma (Sawlan and Smith, 1984).

The early rifting direction was east-northeast and produced north-northwest oriented basins and ranges in basement Miocene volcanic rocks flanking the rift axis.  

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The latest movement has been right-lateral oblique movement (Stock and Hodges, 1989).  This has moved Baja California approximately 350 km northwest relative to mainland Mexico and has created a number of deep pull-apart basins along the axis of the Gulf of California (Bailes et al., 2001).

Stratigraphically El Boleo copper-cobalt-zinc-manganese deposits occur within the late Miocene age succession of fine to coarse clastic sedimentary rocks of El Boleo Formation, lying unconformably on andesitic rocks of early to middle Miocene age called Comondú Volcanics.  El Boleo Formation is characterized by a number of coarsening upward cycles of sediments that are believed to represent deltaic deposition in a shallow, near-shore marine basin.  The upper part of the formation has been locally eroded and unconformably overlain by similar but barren and fossil-rich sedimentary successions of Pliocene and Pleistocene age delta and beach deposits, known as the Gloria, Infierno, and Santa Rosalía Formations.

El Boleo and overlying formations collectively make up the so-called El Boleo Basin.  Locally, the entire succession is capped by Pleistocene to recent flows and pyroclastic rocks of the Tres Vírgenes Volcanics.  The geology of the district has been described in detail by Wilson and Veytia (1949) and by Wilson and Rocha (1955), in privately prepared reports for International Curator Resources Ltd. (Curator) by Peatfield (1995) and Christoffersen (1997) and in numerous other published and unpublished papers and reports referred to in the above-mentioned documentation.

7.2           Property Geology

The oldest rocks outcropping on the property are andesitic volcanics of the Comondú Formation.  They include sub-aerially erupted flows and coarse explosive breccias, which grade into coeval epiclastic sediments to the west.  The volcanics have been dated from 24 Ma to 11 Ma and are underlain by cretaceous granodiorite (Schmidt, 1975).

The overlying El Boleo Formation consists of five coarsening upward cycles of sedimentation numbered “4” at the base and “0” at the top (Figure 7-2).  This interpretation is based on work by Curator from 1993 to 1997 and is different from that published by Wilson and Rocha (1955) who interpreted conglomerates, the coarsest units in the stratigraphy, to be the basal unit in each cycle.

The basal unit of El Boleo Formation is a 1 m to 5 m thick limestone unit.  It contains cherty lenses and non-diagnostic fossil fragments.  Its occurrence atop very steep paleo-surfaces, combined with banding parallel to its base, and also the cherty horizons, suggests it is, at least in part, a chemical sediment.

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Figure 7-2: El Boleo Formation Stratigraphic Column

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Overlying the limestone or laying directly on Comondú Formation over parts of the district, particularly over much of the coastal area, is an extensive gypsum deposit up to 80 m thick.  Although a few dome or mound structures have been noted, the gypsum unit is characteristically flat to shallow dipping exhibiting laminated to massive and even brecciated textures.  Intraformational carbonate beds are rare.

On top of the gypsum/limestone beds is the cyclic succession of clastic beds that average 150 m and range to 270 m thick.  Individual cycles range from 20 m to 140 m thick and consist of a basal mud and fine volcanic ash horizon (now altered to montmorillonite clay) that hosts the copper-cobalt-zinc-manganese mineralization (the manto).

The mantos are overlain by progressively coarser material of maroon coloured, tuffaceous claystone, siltstone, feldspathic sandstone, pebbly sandstone and eventually cobble to boulder orthoconglomerates.

Typically, the earliest cycles (enclosing manto 4, then 3) are thickest with each successive cycle being thinner.  The last cycle manto 0 is thin and believed to be incomplete.  All cycles thin over basement highs and wedge out toward the basin margins.  The copper-cobalt-zinc-manganese stratiform deposits only occur within El Boleo Formation rocks.

Unconformably overlying El Boleo clastics are fossiliferous marine sandstones and conglomerates of the lower Pliocene (about 5.3 Ma.), Gloria formation (Bailes et al., 2001).  These in turn are overlain above a slight unconformity by a sequence of fossiliferous marine sandstones and conglomerates of the Infierno Formation.  Unconformably overlying these are fossiliferous sandstone and conglomerate of the Pleistocene, Santa Rosalía Formation (Wilson and Rocha, 1955).

7.3 Structural Geology

El Boleo Formation rests on an irregular volcanic basement, with several distinct basement highs and intervening troughs.  In places, these basement highs were so pronounced that they influenced the deposition of the lower mantos, such that these pinch out against the volcanics and only the upper mantos are present.  There is also a tendency for the sediments of each cycle to thin towards the high ground, giving a stratigraphic compression and thus less vertical separation of mantos.

Faulting is common throughout the district.  The dominant faults are northwest to north-northwest striking and steeply dipping with normal movements.  These faults have downthrows to both east and west, with more of the major faults down dropping to the west.  This, coupled with the generally easterly dip of the mantos, yields a stepwise pattern of the present position of

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on the major faults, with much lesser movements, 1 m to 5 m, toward the ends of these faults and on lesser structures throughout the district.  Fault displacements will obviously be important in detailed mine planning; fortunately, in much of the district, the faults and their displacements are well documented in old mining records.

Major faults at El Boleo are, in most cases, laterally separated by several hundred metres to, in some cases, over 1,000 m.  Lesser faults are common and more closely spaced.  Faults displace mantos and because of their dip, may form “fault windows” in which the mantos are not present.  However, an order of magnitude calculation suggests that the windows may represent less than 2% of the total area.

Many of the major faults have zones a few metres to tens of metres wide in which the rocks, including the mineralized mantos, are highly disrupted.

Figure 7-3: Geological Cross-Sections

There has also been some oblique strike-slip movement on many of the faults.  The sense of this movement appears to be predominantly right-lateral.

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8 DEPOSIT TYPE

The Boleo District hosts a number of mineral deposit types that have the potential to be of sufficient size and grade to be economically mined and processed.  The most important of these and the subject of this report are the manto hosted copper-cobalt-zinc-manganese deposits, which occur in El Boleo Formation clastic sediments.

Other possible targets are the extensive gypsum beds, which occur over portions of the property, particularly north and east of Arroyo Saturno.  It is believed they occur along the same stratigraphic position as those currently being mined immediately north of El Boleo and 20 km southeast on San Marcos Island.

9 MINERALIZATION

Deposits of copper-cobalt-zinc-manganese mineralization in the Boleo District occur within widespread, stratiform clay-rich horizons or beds known as “mantos” (manto is a Spanish term used in mining parlance for a generally mineralized layer or stratum).  Within the El Boleo Formation stratigraphy there are up to seven mantos, including two of very limited extent, which occur as relatively flat to generally shallow dipping, stratabound and stratiform beds.  These include, with increasing depth, mantos 0, 1, 2, 3AA, 3A, 3, and 4.  Historically the major producing manto has been number 3, which yielded approximately 83% of all production between 1886 and 1985 when the plant shut down.  Most of the remaining production has come from manto 1 in the southeast portion of the Boleo area where manto 3 is absent.  A small amount of production has come from the widespread but generally thin manto 2 while an even smaller level of production has come from the relatively restricted manto 3A.  Based on previous studies and exploration work, mantos 1, 2, 3 and 4 offer the most potential for hosting significant mineral reserves.

The mantos themselves tend to be clay rich (ash altered to montmorillonite) with laminated basal zones generally less than 1 m thick overlain by intra-basin slump breccias up to 20 m thick.  Underlying lithologies vary from predominantly ortho-conglomerates in the heart of the Boleo basin to coarse sandstones typically containing pebbles of Comondú Formation volcanics.  The contact between the mantos and footwall rocks is sharp.

Overlying lithologies vary from fine to medium-grained sandstones.  The contact between them and the clay rich slump breccias is gradational.

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In a general sense, each manto has distinctive characteristics, especially with regard to copper-cobalt ratios and relative concentrations of zinc, manganese and carbonates.

Metals of interest in the mantos include copper, cobalt, zinc, and manganese.  Ore minerals include a fine-grained, complex assemblage of primary Cu-Fe-Zn sulphides including pyrite, chalcocite, chalcopyrite, bornite, covellite, carrolite, sphalerite and secondary Cu-Zn-Mn minerals including chrysocolla, malachite, azurite, smithsonite, cryptomelane, pyrolusite and the rare cobalt minerals of boleite, pseudoboleite, and cumengite.  Mineralization is generally finely disseminated over intervals up to 20 m thick in the slump breccias.  The richest material typically occurs in the laminated basal section of the manto, which was historically mined from 1872 to 1985 to an average thickness of about 80 cm and graded 4% to 5% Cu.

Based on chemical analyses and core descriptions, it is estimated that copper in the Boleo deposit is 60% to 70% oxides and 30% to 40% sulphides.  Azurite and chrysocolla are the principle copper oxides and chalcocite is the principle sulphide copper bearing mineral.  Chalcopyrite is rare and is primarily associated with the deeper manto 1 located in the southeast portion of the property.

Zoning of the principal economic metals occurs both vertically and laterally.  Within individual mantos, copper is enriched at the base, zinc towards the top and cobalt is more or less evenly distributed.  Stratigraphically, vertical zoning shows a trend of zinc enrichment from the lowest manto (4) to the uppermost mantos.  Lateral variations indicate the central core of the Boleo sub-basin is copper rich flanked by a zinc rich marginal zone.  Cobalt is variable and shows no clear correlation with copper or zinc.

Individual mantos and their enclosing strata are “time transgressive,” in that they are progressively younger toward the present Gulf of California.  One very distinctive unit, the “Cinta Colorada” or “red ribbon” is a layer of reddish andesitic-basaltic tuff up to 2 m thick.  This is interpreted to be the product of a single explosive volcanic event, which probably blanketed the entire region.  The Cinta Colorada represents a true “time horizon,” and can be seen to transgress stratigraphy.  In some places it lies within the unit 2 conglomerate (below manto 2) and elsewhere within the underlying unit 3 clastic succession.  Thus, it demonstrates the time transgressive nature of the enclosing stratigraphic units.

Individual mantos have great lateral continuity and relatively consistent thicknesses.  In the principal areas of interest, the lowest manto (4) lies at the base of El Boleo Formation, directly on the Comondú Formation.  Manto 3 is widespread and thick, and accounting for the largest proportion of the mineral resource.  Mantos 3A and 3AA are less continuous and thinner, lying higher in the succession.  In some places, especially in the Saturno-Jalisco area, 3A merges with 3.  Manto 2 is stratigraphically widely distributed; however, because of its higher stratigraphic position it is more commonly eroded.

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Manto 1 makes up the bulk of the mineral resource in the southeast portion of the Boleo property, where the lower mantos (3 and 4) were for the most part not deposited.  The beds dip to the southeast and as a consequence, manto 1 lies deeply buried in this area.  To the northwest, manto 1 overlies the well mineralized portion of manto 3 (and in many places, 3A, 3AA and 2).

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10 EXPLORATION

Since acquiring control of the Boleo Project in 2001, and prior to December 2004, MMB concentrated on carrying out a complete geological, mining, and processing review of the Boleo property in place of only conducting field exploration work.  This includes an independent assessment by Hellman and Schofield Pty Ltd. (H&S) of the copper-cobalt-zinc-manganese resources.  Bateman Engineering Pty Ltd. (BEPL) of Perth, Western Australia, completed an in-depth review of alternate flow sheets for the processing of ore from the Boleo property and Agapito Associates Ltd. (AAI) and Australian Mine Design and Development Ltd. (AMDD) carried out investigations into alternative mining methods, in particular the potential to use underground mining techniques.

Those reviews resulted in a new Pre-feasibility Study being issued by BEPL in February 2002, which focused principally on a new metallurgical flow sheet for the processing of ore from the Boleo property.

Drilling activities commenced on the property in December 2004 (discussed in Section 11).

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11 DRILLING

11.1 General

The current resource study uses all diamond drill holes (DDHs) completed by previous owners of the Boleo property, plus the additional holes of the infill drill programs carried out by MMB between December 2004 and September 2008 (DDHs 04-928 to 08-1321).

11.2 Historical

The oldest recorded drilling program was carried out between 1927 and 1940 when 10,237 m were drilled in 46 vertical churn holes (Wilson and Rocha, 1955).  Most of these holes were drilled in the southeast portion of the property to explore for manto 1 in the Ranchería, San Luciano, and Montado areas.  Further diamond drilling was carried during the latter years of mining operations when Fomento Minero was looking for high-grade reserves to exploit.  Records of this drilling were never kept or have been lost or destroyed.

The most extensive drilling program ever conducted in El Boleo District was that of International Curator between 1993 and 1998.  All exploration diamond drilling was completed using skid mounted Longyear 38 drill rigs moved with logging skidders.  Core size was HQ (63.5 mm diameter), reduced to NQ (47.6 mm diameter) when necessary because of drilling problems.  In areas where considerable thickness of overlying barren stratigraphy (Gloria, Infierno, and Santa Rosalía Formations) was expected, the upper portion of the hole was triconed before coring commenced near the target horizon.  A total of 963 holes for 76,169 m were completed by International Curator.

11.3 2004 – 2008

The objective of the previous property operator was to define an open cut resource, therefore the majority of their drilling was confined to or near the arroyos targeting near surface mineralization, amenable to low stripping ratio, open pit mining.  This left areas of mineralized mantos, with thick overlying cover rocks, with little drilling.

MMB plans to mine the mantos by a combination of underground and open cut methods. MMB’s primary objective for their drill programs was to fill-in poorly drilled areas to <200 m drill hole spacing thereby increasing the Measured and Indicated resource classes in these areas.  The location of all the drill holes is shown in Figure 11-1.

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The International Curator and MMB drilling and trenching programs are summarized in Table 11-1.  During the period 1993 to 1998, International Curator drilled 963 diamond drill holes totalling 76,169m.  Since a relatively small proportion of each hole tested manto material and some holes did not intersect a manto, only 8,940 m were sampled in 925 holes.  Since 2004, MMB has drilled 401 diamond drill holes and cut 37 vertical trenches totalling 51,710.26 m and 106.4 m respectively.  As with the Curator holes, not all of the holes and trenches tested manto material and a number of holes were drilled to collect geotechnical information.  Therefore, the MMB exploration produced 2,943 m of sample collected from 323 drill holes and 36 trenches.

Table 11-1: Drill and Trench Program Summary

Year	Company	Drilled		Sampled		Type
		Holes	Metres	Holes	Metres
1993	Curator	22	1,508.26	22	153.61	DDH
1994	Curator	70	6,197.51	70	577.54	DDH
1995	Curator	375	30,754.51	361	3,134.97	DDH
1996	Curator	307	27,227.64	288	3,169.14	DDH
1997	Curator	154	9,127.11	149	1,569.42	DDH
1998	Curator	35	1,353.93	35	335.42	DDH
2004	MMB	2	442.80	2	9.90	DDH
2005	MMB	31	5,297.05	31	400.04	DDH
2006	MMB	246	37,846.96	243	2,008.02	DDH
2006T	MMB	37	106.40	36	89.47	Trench
2007	MMB	74	5,287.00	29	249.00	DDH
2008	MMB	48	2,836.45	18	186.62	DDH
Total		1,401	127,985.77	1,284	11,882.94

Note: Pre-2004 drill holes are shown as open symbols.  MMB drill holes are shown as black solid symbols.

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Figure 11-1: Drill Hole Locations – Historical and Infill

Note:  Pre-2004 drill holes are shown as open symbols.  MMB drill holes are shown as black solid symbols.

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12 SAMPLING METHODS AND APPROACH

12.1 Sampling Methods and Approach

The samples used in the resource estimation are from diamond drill holes and vertical trenches.  Down hole sample intervals vary in length as intervals were selected on the basis of geology.  A total of 12,533 samples totalling 11,883.9 m were collected from the mineralized manto units and the mean sample interval was 0.95 m.  The 36 trenches produced 122 samples totalling 89.5 m and 1,284 drill holes produced 12,411 samples totalling 11,793.4 m.

Historical drilling activities prior to 2004 have not been observed.  Sampling procedures (adopted by Curator) were as follows:

· Core was transported from the drill site by either helpers or geologists to company warehouses in Santa Rosalía, where the boxes were labelled and core recoveries calculated.

· A company geologist marked out all sample intervals, then logged the core.

· A trained local helper split core with a mechanical splitter (or a knife in poorly consolidated material).

Core logging was based on geological intervals with detailed written descriptions for each interval.  Mineralogical, structural, and textural information was not recorded in dedicated fields, making it difficult to extract anything other than summary data from the logs.  Logged geology intervals did not always correspond to assay intervals and in many cases, overlaps exist between geology and assay intervals.

Drilling rigs were inspected during the recent drilling program in May 2006.

Procedures for drill site supervision and handling of core for this program were as follows:

· Drilling Supervision:At the time of manto intersection, geologists were on-site to take notes on core recovery and the conditions of the core.

· Transport of the Core: On completion of a hole the core boxes with lids are loaded into a pick up and secured with a rope to avoid core losses during transport.

· Layout of Core for Logging:  In the core shack, the core is placed in sequential order on benches.  The benches are designed for ease of use and are well lit.

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· Calculation of Core Recovery and RQD:  Measurement of the length of recovered core between each drilling interval is used to calculate the percent recovery of the core.  The total length of core that occurs in individual core pieces greater than 10 cm in length is measured in order to calculate the RQD.

· Logging Procedures:  The core is logged based on geological criteria such as lithology, formation number and members, paying special attention to subdivisions in the mineralized zones.  Geologists complete a fully descriptive written log of each interval.  This descriptive log is then computer coded for entry into a geological database.  Each sample interval corresponds exactly to an assay interval.

· Sample Selection:  Sample selection is carried out in the mineralized manto intervals with a black marker pen indicating where each sample initiates and ends and showing arrows that indicate the interval to be sampled.  The length of the sample varies depending on the type of material being sampled.  The maximum length for an individual sample does not exceed 1.5 m.

· Numbering of Samples:  The numbering sequence used for sampling is continuous and ascending with respect to depth in a specific drill hole.  Numbering control is recorded in a sampling book, sample numbers are also recorded on the full log and the computer coded log sheets.  Assay quality control items such as standards, blanks, and pulp duplicates are inserted at this stage.

· Sampling of the Core:  Splitting of the core is completed using a mechanical splitter.  One-half of each sample interval is sent to the laboratory for assaying the other half is retained in the original core box Figure 12-1.

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Figure 12-1: Mechanical Splitter (H&S Modified Design)

12.2 Core Recovery

Core recovery from the historical drill holes was measured for each retrieved core run.  Measured recoveries range from 82% to 90%, with a mean of 83% (Table 12-1).  Where recovery was less than 100% no attempt was made to specifically identify the location of lost core within the recovered interval.

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Table 12-1: Diamond Core Recovery – Curator

Manto	Intervals	Recovery (%)
0	49	90
1	128	87
2	381	83
3	772	82
3a	564	84
3aa	50	88
4	141	84
Total	2,085	83

Table 12-2 shows the core recovery for the current drill programs (DDH928-1233) and demonstrates better overall values with a mean recovery of 94.4%.

Table 12-2: Diamond Core Recovery – Current Programs (DDH928-1233)

Manto	Intervals	Recovery (%)
0	208	95.4
1	400	95.0
2	680	95.6
3	780	92.9
3a	277	94.9
3aa	22	89.5
4	396	94.3
Total	2,767	94.4

For the historical holes, the core run intervals over which recoveries were determined do not correspond to specific assay intervals so it is not possible to evaluate whether a relationship, of any sort, exists between recovery and copper or cobalt grade.

For the current drill holes, recoveries have been determined for each sample interval.  Figure 12-2 shows core recovery against copper grade.  There is no correlation between recovery and grade, although lower recoveries tend to be lower grade suggesting possible loss of friable mineralization.

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Figure 12-2:                      Core Recovery vs. Cu Grade – Current Drill Programs

12.3 Density Determinations

Data and methodology used to determine density were taken from an internal company report (Félix, 1996).  Measurements of specific gravity were taken during the exploration drilling campaigns of 1995 and 1996.  A total of 2,110 measurements were obtained of which 418 volumes were determined by a dimensional method (discussed below) and 1,694 by water immersion.

The dimensional method was used to obtain volume determinations in 1995 and the first weeks of 1996, after which only the water immersion method was used.  Approximately a hundred samples were measured for specific gravity using both methods.  The difference between the results averaged 0.4%.  The samples were taken from both mantos (993 determinations) and non-mineralized rocks (1,117 determinations).

Densities were determined on un-dried samples.  Dry density for mineralized manto samples was calculated from the wet density by factoring the density value by the proportion of water content lost during sample drying.  To determine water content, samples were dried at 110°C for a minimum of 6 hours.  SGS-XRAL Laboratories in Hermosillo calculated the water content, as the weight loss on drying and as a percentage of the original sample weight.

12.3.1 Dimensional (or Calliper) Method

The core sample was cut with a saw perpendicular to its axis; samples were nominally 30 cm in length.  A tape was used to measure the diameter of the two ends, with two measurements taken at each end with the average of the four taken to establish the diameter of the sample.  In

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the same way, the length of the core was measured with the same tape in two directions and the average taken as the length.  The volume is calculated using the formula for the volume of a cylinder and bulk density determined dividing the weight by the volume after adjusting for core recovery.

12.3.2 Immersion Method

The immersion is based on the Archimedes Principle, which states that the density of an object is equal to the weight divided by the difference between the weight in air and the weight in water.

In the laboratory, the samples were sprayed with lacquer to form a thin and uniform coat.  The lacquer is used to prevent water being absorbed by the sample.  The sprayed sample was allowed to dry a few minutes and was weighed in normal (in air) conditions and then weighed submerged in water.

12.3.3 Results

For the purposes of resource calculations, a global in situ dry bulk density of 1.41 t/m3 has been used for all mantos, based on an average calculated from 993 wet density measurements and the corresponding analysed water contents of the samples (Table 12-3).  A density of 1.24 t/m3 was applied to previously mined areas of mantos 1 and 3 (see Section 17.4.1).

Table 12-3: Bulk Density Summary by Manto

Manto	Wet (Natural)			Dry
	Bulk Density	Data	H2O (%)	Bulk Density	Data
Tbcu0	1.84	53	26.75	1.36	53
Tbcu1	1.88	86	26.15	1.40	86
Tbcu2	1.87	136	26.93	1.38	136
Tbcu2A	1.86	11	26.86	1.36	11
Tbcu3	1.89	543	25.71	1.41	489
Tbcu3A	1.90	133	26.15	1.41	128
Tbcu3AA	1.88	9	26.89	1.38	9
Tbcu4	1.91	78	24.19	1.46	78
Tbcu4A	1.93	3	25.48	1.44	3
All Mantos	1.89	1,052	25.93	1.41	993

Although there is some indication of increasing bulk density with sample depth, a graph of bulk density vs. depth (Figure 12-3) shows only a weak correlation.

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Figure 12-3:                      Bulk Density vs. Sample Depth

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13        SAMPLE PREPARATION, ANALYSIS & SECURITY

13.1 Historical Drill Holes – Pre-2004

Prior to 1997, assay samples were sent to the SGS-XRAL laboratory in Hermosillo where they were dried, crushed, and pulverized.  Analysis for copper, cobalt, zinc, iron, and manganese was carried out at the same laboratory, with a perchloric acid digest and AAS finish.  Samples that reported grades greater than 1% of Cu, Co, or Zn were re-assayed using a method more appropriate for higher-grade samples.

From 1997 onwards, sample preparation was carried out at Chemex facilities in Hermosillo with pulps being forwarded to the Chemex Laboratory (Chemex) in Vancouver, Canada for analysis.  A four acid digest with AAS finish was used for Cu, Co, Zn, Fe, and Mn.

The reject and remaining pulp material were returned to Santa Rosalía where they are securely and systematically stored in warehouses.

The reason for the change in assaying laboratory and technique was due to the identification of a systematic under-reporting of primarily cobalt by SGS-XRAL.  The extent of the analytical problems and remedial measures taken are discussed in more detail in Section 14.1.

All samples were stored, prior to shipping, in one of three locked warehouses in Santa Rosalía.  When shipped, the samples were taken to the ferry by company personnel where they were put on the ferry and shipped across the Gulf of California to Guaymas.  The representatives of the laboratory picked the samples up and delivered them directly to the laboratory in Hermosillo.

13.2 Post 2004 Drill Holes

The samples for assaying are placed into 30 kg sacks and kept in locked premises until they are transported to the laboratory.  A company vehicle is used to transport the samples from Santa Rosalía to Guaymas via a ferry and then driven directly to Hermosillo, Sonora.  The samples are driven and accompanied by a company employee.

Remaining core is stored in boxes in an underground storage area that is secured by a locked metal door.

The sample shipment is delivered to ALS Chemex, located at:

Ignacio Salazar 688 Local 5, Fracc. Los Viñedos,C.P. 83147, Hermosillo, Sonora, México.

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The sample preparation protocol used is as follows:

· samples are oven dried at 110° until dry

· samples are fine crushed to > 70% < 2 mm (TM Rhino or Terminator crusher)

· crushed sample is split using a riffle (Jones) splitter to 250 g of crushed material is pulverized (Labtech LM2)> 85% < 75 µm

After sample preparation, pulps are transported for assaying from Hermosillo to ALS Chemex in North Vancouver.

The assay method currently used for copper, cobalt, zinc, and manganese is ME-ICP61a (four acid “near-total” digestion).

· Sample Decomposition: HNO3-HClO4-HF-HCl digestion (ASY-4A02)

· Analytical Method: Inductively Coupled Plasma – Atomic Emission Spectroscopy (ICP – AES).

The sample is digested in a mixture of nitric, hydrochloric and hydrofluoric acids.  Perchloric acid is added to assist oxidation of the sample and to reduce the possibility of mechanical loss of sample as the solution is evaporated to moist salts.  Concentrations of elements are determined by inductively coupled plasma – atomic emission spectroscopy (ICP-AES).

Samples reporting high grades (> 100,000 ppm) for copper, cobalt, zinc, or manganese are re-assayed for those elements by method AA62.

· Sample Decomposition: HNO3-HClO4-HF-HCl digestion (ASY-4ACID)

· Analytical Method: Atomic Absorption Spectroscopy (AAS).

A prepared sample (0.4 g) is digested with nitric, perchloric, and hydrofluoric acids, and then evaporated to dryness.  Hydrochloric acid is added for further digestion, and the sample is again taken to dryness.  The residue is dissolved in nitric and hydrochloric acids and transferred to a volumetric flask (100 mL or 250 mL).  The resulting solution is diluted to volume with de-mineralized water, mixed, and then analyzed by atomic absorption spectrometry against matrix-matched standards.

Total sulphur has been assayed for using ME-ICP61a with sulphide sulphur determined by S-IR07 (Sodium carbonate dissolution of sulfates, Leco furnace and infrared spectroscopy.)  Moisture content was determined by OA-GRA05 (Gravimetric comparison of sample weight before and after drying.).

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14        DATA VERIFICATION

Data verification can be considered as having four separate components:

· Are the assay results reported for each sample accurate?

· Are the samples, used in assaying process, representative of the sample interval?

· Are the reported assay values, which are identified by a unique sample number, assigned to the correct down hole sample intervals?

· Are the drill hole locations known accurately, and locations correctly entered in the database?

To evaluate accuracy and quality of the analytical data, a number of quality control samples must be inserted into the sample stream.  These samples include standard reference materials (SRMs or “standards”) of known grade, blank samples with no grade, and possibly field duplicate samples.  Field duplicates of diamond core samples should be the un-sampled half of core retained after the original round of sampling.  To be most effective, the quality control samples should be blind to the assaying laboratory.

14.1 Historical Data – Pre-2004

14.1.1 Problems with the SGS-XRAL Pre-2004 Assay Results

During the original round of assaying by SGS-XRAL, no SRMs or field duplicate samples were used.  Some internal repeat assays were reported by the laboratory and on quick visual examination, these appeared to be adequate, but no statistical analysis was completed to support this.

At a late stage in the process, a group of samples were retrieved and sent to an independent laboratory, Bondar-Clegg, Vancouver, for check analysis.  The results of this check program revealed some considerable departures from the original values, though no standards were included with these samples, so the same samples were forwarded to Chemex laboratories, also in Vancouver.  The results of these second repeat check assays were again considerably different to both the original SGS-XRAL and Bondar-Clegg check assays, such that sufficient doubt was generated as to the validity of the entire Pre-1997 assay database at El Boleo.

Consequently, Curator made the decision to review the entire assaying process used for the Boleo samples.

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14.1.2             Summary of 1997 Process Review and Re-Assaying Program

The review was carried out on behalf of Curator by consultants G. Peatfield and B. Smee.  The process and subsequent re-assay program is reported in detail (Peatfield and Smee, 1997; Peatfield, 1997; Peatfield, 1998).  These reports are available and a limited synopsis is included below.

Several problems were identified with respect to the original SGS database:

· Despite repeated assurances from SGS that the perchloric acid digest would give a total metal extraction, data from check assaying indicated otherwise.

· Above 4% copper SGS assay values showed very poor correspondence with check assays from other laboratories.  It was thought that this was to do with a dilution step in the process; however, SGS failed to provide a rational explanation.

· SGS reporting left much to be desired, with issues such as changes in reporting format and numerous data errors.

Check assaying strongly suggested that copper values from SGS were reported systematically low, possibly by as much as 15%.

Preparation of Assay Standards

The use of standards to monitor assay accuracy was decided on immediately.  It was also deemed essential that these standards should be prepared from El Boleo material so that the matrix of the standards matched that of the samples.

Initially two standards (Interim Standards) were prepared from El Boleo material (El Boleo I, II).  These were prepared at the SGS laboratory.  Sub-samples were dispatched to seven laboratories for round-robin analysis.  The material was subjected to a four-acid digest (nitric, hydrofluoric, perchloric, and hydrochloric), with an AAS finish.  From the data received from each laboratory, the mean and relative standard deviations were determined.  The interim standards were used only as a stop-gap measure whilst the full review of the assay process and the preparation of formal standards were completed.

The formal standards (El Boleo III, IV, V) were prepared at the Colorado Minerals Research Institute and CDN Resource Laboratories, Burnaby, BC, Canada.  Composites of varying grade were combined to form three single 25 kg samples.  Samples were again dispatched to several laboratories in the USA, Canada, and to SGS in Mexico for round-robin analysis.  The same four-acid digest method was used.  The recommended grade and limits for each standard (Peatfield, 1997) are shown in Table 14-1.

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Table 14-1: El Boleo Assay Standards

Standard	Cu			Co			Zn
	Mean	+2sd	-2sd	Mean	+2sd	-2sd	Mean	+2sd	-2sd
Interim Standards
El Boleo I	1.546	1.682	1.410	0.0595	0.0640	0.0550	0.34	0.329	0.278
El Boleo II	3.813	4.159	3.466	0.0527	0.5677	0.4377	0.355	0.389	0.321
Formal Standards
El Boleo III	0.514	0.555	0.472	0.0520	0.0630	0.0490	0.322	0.338	0.306
El Boleo IV	1.124	1.202	1.047	0.0934	0.1082	0.0786	0.387	0.415	0.358
El Boleo V	7.405	7.900	6.910	0.0986	0.1146	0.0827	0.459	0.459	0.434

Analytical Method and Assaying Laboratory

To determine the most appropriate analytical method to use, 12 samples were selected, prepared in replicate pulps, and dispatched to several laboratories where they were assayed using four different techniques:

· aqua-regia digest

· sodium peroxide fusion

· perchloric acid/aqua-regia digest

· four-acid digest (nitric, hydrochloric, perchloric, hydrofluoric).

The sodium fusion method tended to report substantially lower grades indicating less complete dissolution.  The four-acid method gave the most consistent results although slightly lower than the aqua-regia or perchloric/aqua-regia digests.

The four-acid digest was chosen as the analytical technique and Chemex was chosen as the assaying laboratory.  A second laboratory, Mineral Environments Laboratories, was chosen for check assaying.

Re-Assay Program

All pulp and reject material held by SGS in Hermosillo were returned to Santa Rosalía.  Samples were selected for re-assay on the premise that any sample obviously or likely to be used in resource estimations would be re-assayed.  When in doubt samples were included rather than omitted.

Thirty-gram sub-samples of existing pulps were prepared and a completely new sample number sequence applied.  Standards, blanks and duplicate pulps were inserted into the sample stream in sequence so that every 40 samples sent to Chemex would include, two standards, two blanks and two duplicate pulps.  In addition, Chemex inserted two of their own standards and a sample blank.  In all about 6,800 samples were re-assayed.  To monitor quality 1,200 internal quality

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control samples (standards, blanks, and duplicates) were inserted by Curator and Chemex inserted an additional 800 laboratory samples.

The results of the internal standards and blanks were monitored as data was received by means of control charts, which show the acceptable value and upper/lower limits (±2sd).  No concerning results or trends were identified.  Blanks sample results were monitored by eye and no obvious contamination was seen.  Chemex standards were monitored in a similar way.

Sample duplicate data was analysed using a mathematical technique (Thompson and Howarth, 1978) to determine analytical precision.  The results indicated that at grades likely to be mined, levels of precision were very good (±<5%) for both copper and cobalt.

Second Laboratory Check Assaying

A total of 441 duplicate pulps were sent to a second laboratory, Mineral Environments Laboratories, (Min-En) for check assays.  The results were typically similar; however, there was a clear laboratory bias, with Min-En’s results reporting lower than Chemex’s results.

Similar results were seen in the data reported for the standards returned from Min-En.  It was concluded that the check laboratory did not match their performance in the original round-robin tests.

Core Duplicate Assaying

As part of the re-assaying program, pulp duplicates were routinely inserted into the sample stream.  However, because they are from the same original sample, they do not provide a means for determining overall precision resulting from all steps in the process, from sampling, sample preparation and assaying.  To do this, the remaining half-core has to be sampled and subjected to the same sample preparation and assaying regime.

One hundred samples were selected for core duplicate assaying.  For each of these samples the remaining half-core was collected from the core tray and sent for assay.  Half core was crushed and split into two equal samples.  From each of these, a 250-g sub-sample was taken and fine pulverized and a final 30-g sub-sample taken from each and sent to Chemex for assaying.  As a result, it was possible to compare the original assay against two identical assays of the remaining core, referred to as ‘A’ and ‘B’ core duplicates.

In addition, duplicate samples were also taken from the coarse residue of the original assay samples retained by the laboratory.

Standards were inserted into the sample stream at a rate of 1 in 20, as done previously.

The results showed, as expected, lower precision between the core duplicates and original samples than for the coarse residue duplicates and the original samples.  Surprisingly though, there were also noticeable differences between the calculated precision from the ‘A’ and ‘B’

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core duplicates, which lead to the decision to include core duplicate sampling programs as part of all future drilling and assaying programs.

Average grades from both the ‘A’ and ‘B’ core duplicates were lower than the original assays by ~10% for copper, 5% for cobalt and 2% for zinc.  No explanation was given for this difference.

Table 14-2shows a summary of the results of the ‘A’ and ‘B’ core duplicate sampling program.

Subsequently a further 55 core duplicate samples were assayed (Table 14-3).  Levels of precision clearly improve from core duplicates to coarse residue duplicates to pulp duplicates and, as in the first program, the core duplicate assays show lower average grades than the original assays, although in the second program this difference is less.

Table 14-2: El Boleo Core Duplicate Results – 1st Program

Core Duplicates	Range (%)	Average Precision	Original Assay Average	Dupl. Core Assay Average	% Diff.
Copper
‘A’	0.10 – 4.78	25.9	1.001	0.895.	-10.6
‘B’	0.10 – 4.78	26.8	1.001	0.903	-9.8
Cobalt
‘A’	0.010 – 0.0525	20.2	0.063	0.060	-4.8
‘B’	0.010 – 0.0525	20.1	0.063	0.060	-4.8
Zinc
‘A’	0.2 – 6.69	23.9	0.751	0.736	-2.0
‘B’	0.2 – 6.69	14.2	0.751	0.736	-1.5

Table 14-3: El Boleo Core Duplicate Results – 2nd Program

Duplicate Type	Range (%)	Average Precision			Original Assay (Average)			Duplicate Assay (Average)	Diff. (%)
Copper
Core	0.10 – 4.38	14.4			1.020			0.957	-6.2
Coarse Residue	0.10 – 3.34	5.7			0.816			0.811	-0.6
Pulp	0.10 – 3.72		1.7	0.595		0.587	-1.3
Cobalt
Core	0.010 – 0.206		22.3	0.045		0.044	-2.2
Coarse Residue	0.010 – 0.335		8.8	0.051		0.051	0.0
Pulp	0.010 – 0.187		3.7	0.053		0.052	-1.9
Zinc
Core	0.20 – 1.30		12.9	0.398		0.386	-3.0
Coarse Residue	0.20 – 0.65		1.5	0.326		0.325	-0.3
Pulp	0.20 – 1.48		1.8	0.441		0.440	-0.2

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One reason for the improved result in the second program may be that the duplicates and original samples were assayed in the same batch, while in the first program duplicates were submitted in two batches, one at a later date than the other.

Therefore, any problem unique to a batch or prevalent at a laboratory for a limited time can affect the entire program.  The average grades of both the coarse residue and pulp duplicates are almost identical to the original samples.

Conclusions Derived from the Pre-2004 Re-Assay Program

Peatfield and Smee arrived to the following conclusions:

· The initial assay method used by SGS was unreliable and uncontrolled, and therefore unacceptable.

· Tests carried out showed that a multi acid digestion provided the most consistent assay results.

· Round-robin assaying showed the original laboratory had difficulty in generating statistically acceptable results.

· The re-assay program was completed under controlled QC conditions, and provided revised database assay information with acceptable accuracy.

· Subsequent core duplicate sampling produced relatively high but acceptable levels of precision.

· Precision based on reject duplicates showed little difference to pulp duplicate precision, indicating that the error attributable to sample preparation was minimal.

· The systematic decrease in metal in the core duplicates is not readily explicable; the improvement between the 1st and 2nd core duplicate sampling program may be due to the fact that, in the 2nd program, duplicates were assayed in the same batch as the original samples.

· Routine monitoring of internal and Chemex standards indicates that there are no recognizable problems with the accuracy of copper, cobalt, zinc, and manganese assays in the Boleo database.

14.1.3 Comments on Assay Quality Control Results

Because of the problems identified in the original El Boleo assay database, the results from the re-assay program have been subjected to a high level of scrutiny.

The preparation of matrix-matched standards is commendable and is the best means available to monitor and ensure assay accuracy.

The control charts for each standard (Peatfield and Smee, 1997) show that although the copper results do fall within the acceptable limits, they are mostly higher than the accepted value,

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rather than being scattered above and below.  Conversely, the standards reported by the second laboratory were also within the acceptable limits but systematically low.  These patterns are consistent, in all cases, with the Chemex value for each standard reported in the round-robin assaying relative to the accepted value (the mean of seven participating laboratories).

However, H&S considered that a more rigorous evaluation of the round-robin results used in determining the standard values standards should have been applied.  Of the seven participating laboratories, results from six were used, in most cases, to determine the true values (Peatfield and Smee, 1997).  H&S believed that using results from fewer laboratories (3 or 4) of superior quality and consistency is best in determining the recommended values.

Figure 14-1shows graphically the copper data reported by each laboratory used for El Boleo V, while Figure 14-2 shows the data used by H&S.  A similar exercise was completed for copper and cobalt data for El Boleo III and IV for the results are compared in Table 14-4.  Copper values determined by H&S are slightly higher than Peatfield and Smee, while cobalt values are slightly lower.  If revised copper standard values were used, the potential assay bias of 2 to 4% would be reduced to less than 2%.

The core duplicate sampling programs show a consistent difference in the means of the original assays and the duplicate samples, by as much as 10% for copper.  This may be in part attributable to batch issues but this has not been confirmed.

Another aspect that can cause differences between reported grades is the method and completeness of the collection of material from the core tray.  It is unlikely that the fines in the bottom of a core tray are identical in make up as the larger segments of core.  If for example, fine-grained chalcocite is dislodged through handing of core and collects in the tray, it needs to be carefully collected; failure to do this may result in understating copper grade.

Figure 14-1:              El Boleo V Cu, Round-Robin Cu Assays used by Peatfield and Smee

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Figure 14-2: El Boleo V Cu, Round-Robin Cu Assays used by H&S

Table 14-4: El Boleo Formal Standards – Revised Values

Standard	Copper			Cobalt
	Mean	+2sd	-2sd	Mean	+2sd	-2sd
Formal Standards – Peatfield and Smee (1997)
El Boleo III	0.514	0.555	0.472	0.0520	0.0630	0.0490
El Boleo IV	1.124	1.202	1.047	0.0934	0.1082	0.0786
El Boleo V	7.405	7.900	6.910	0.0986	0.1146	0.0827
Revised Formal Standards – H&S
El Boleo III	0.525	0.530	0.519	0.0511	0.0567	0.0455
El Boleo IV	1.135	1.159	1.111	0.0929	0.0988	0.0871
El Boleo V	7.579	7.709	7.450	0.0986	0.1045	0.0927

The use of a soft brush is often needed to ensure that all the fines are collected from the bottom of the tray.  Unfortunately, sample collection was not observed.  The 55 duplicate core samples from the second program were dispatched to the laboratory at the same time as the original samples, suggesting that they were bagged immediately on splitting and not returned to the core tray to be re-sampled at a later date.  Yet results for these duplicate samples were still lower than for the original samples.

14.1.4 Database Verification

To ensure that the database is accurate (i.e. to ensure that assays are assigned to the correct sampled interval) an audit of original assay certificates against the database files was carried out.

The assay data is located in 10 Microsoft Excel spreadsheets (a separate file for each hundred holes).  Two holes were checked from each file (Table 14-5), with copper, cobalt, and zinc

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assays checked.  Geological logs and summary logs contained in the same Excel files were not audited.

Table 14-5: Drill Holes Audited for Data Entry Errors

Hole ID	Assay Job No.	Hole ID	Assay Job No.
022	A 9631180	491	A 9627709
57	A 9629734	526	A 9627711
115	A 9634051	592	A 9628562
173	A 9626536	656	A 9626533
226	A 9626538	695	A 9629737
287	A 9626541	746	A 9717119
324	A 9626542	765	A 9718436
361	A 9626544	818	A 9720883
410C	A 9626545	888	A 9737165

No data entry errors were detected.

14.1.5 Drill Hole Surveying

Pre-2004 drill hole collars at El Boleo have not been surveyed.  In place of conventional surveying, high resolution orthophotos have been acquired from which drill hole coordinates have been calculated.  To assist in identifying drill holes in the photos each collar was marked with a large white cross, with the hole at the centre.  Fifty pre-2004 drill holes were checked using high precision GPS to verify the collar coordinates.  Average results of this survey show easting, northing, and elevation to be within 0.4, 1.1, and 1.2 m respectively of the orthophoto coordinate.

The drill hole spacing, at its closest, is in the order of 150 m x 150 m so ~1 m accuracy for the easting and northing coordinates is adequate.  Mantos are generally only a few metres thick so accuracy of the elevation coordinate is more critical.  A tolerance of ±1.2 m that was obtained for this dimension is not considered serious.

In an open-pit environment, mining is carried out with strict grade control practices that are used to identify ore and waste prior to mining.  In an underground situation geological control, particularly the different lithologies that define the base of each manto should be adequate to avoid losses due to uncertainty in the collar surveys.

All the holes at El Boleo are vertical, which is appropriate.  No down hole surveys were conducted.

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14.2           Current Program

Assay standards, duplicate sampling, blank samples, and check assaying have been used during the 2004 to 2007 drilling programs to validate the assay results reported by ALS.

14.2.1 Assay Standards

The three El Boleo standards (El Boleo III, IV, V) discussed above have been used again during the current assaying program. @@  Figure 14-4 and Figure 14-5 show these results.  The reported results are combined and presented as a percentage difference from the revised accepted values of each standard (see Section 14.1.3).

There is a persistent under-stating of standard grades by between 5% and 11% for the 2006/2007 drill programs.  The reason for this is not at present well understood but the issue is currently under investigation and review by MMB staff.  Initial thoughts were that the sample preparation laboratory failed to adequately dry the standard samples thereby overstating the true weight of the sample aliquot taken for analysis.  A small batch of 18 standards was sent to the laboratory for check assay, with specific instructions to completely dry the pulps.  The samples returned similarly low values ranging from -6% to -9%.  A larger number of standard pulps (90) were sent to ALS and two other laboratories, ACME and Global Discovery, in an attempt to establish whether or not the recommended values are still applicable or whether the material has degraded over time.

The results are very similar to the routine standard results reported by ALS.  All laboratories consistently reported lower than expected copper grades (Figure 14-6 through to Figure 14-8).  It therefore appears that El Boleo assay standards have degraded (oxidized) since being constructed in 1998. New standards (El Boleo VI, VII, and VIII) are in now use.

Figure 14-3: El Boleo Assay Standards – Cu

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Figure 14-4: El Boleo Assay Standards – Co

Figure 14-5: El Boleo Assay Standards – Zn

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Figure 14-6: Check Assay El Boleo III – Cu

Figure 14-7: Check Assay El Boleo IV – Cu

Figure 14-8: Check Assay El Boleo V – Cu

14.2.2 Check Assays

Check assays from ACTLabs in Tucson however show very good correlation to the ALS-Chemex results (Figure 14-9 to Figure 14-11).  These results can also be used to make some comments regarding the problem with the assay standards reported above.  It would appear unlikely that both laboratories produced similarly poor results, to the extent that the inter laboratory correlation was as good as reported.

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Figure 14-9: El Boleo Check Assay Results – Cu

Figure 14-10: El Boleo Check Assay Results – Co

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Figure 14-11: El Boleo Check Assay Results – Zn

14.2.3 Duplicate Assays

Duplicate assays have been determined from crusher split reject material.  Although the total number of duplicates is low (< 50) the comparison is very good (Figure 14-12 to Figure 14-17)

Figure 14-12: El Boleo Duplicate Assay Results – Cu

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Figure 14-13: El Boleo Duplicate Assay Results – Co

Figure 14-14: El Boleo Duplicate Assay Results – Zn

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14.2.4 Blank Sample Assays

Blank samples, derived from the unmineralized Gloria Formation, were inserted into the sample stream at regular intervals.  These samples are designed to test for cross-contamination in the laboratory.

Figures 14.15 to 14.17 show the results showing consistent low levels of copper, cobalt, and zinc, with occasional elevated values.  The increased spread of copper values (on the right hand side of the graph) may be indicative of less thorough laboratory hygiene between samples at a time when reduced sample turnaround time was required.  There is however, no corresponding elevation in either cobalt or zinc.

Figure 14-15: El Boleo Blank Assay Results – Cu

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Figure 14-16: El Boleo Blank Assay Results – Co

Figure 14-17: El Boleo Blank Assay Results – Zn

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14.3 Project Database

14.3.1 Data

A new Microsoft Access database was customized for the Boleo Project.  This database was set up to accommodate the historical data and the new data from the current programs.

Historical data exists in 10 Excel spreadsheets, a single spreadsheet for each sequential hundred holes.  Each file contained several sheets including a drill hole coordinate sheet, assay sheet and summary geological logging sheet.  The data is stored in a number of non-active tables, i.e., no new data is to be added to the tables.  The data was taken as supplied and no modifications were made.

Data from the current drill programs is imported into a number of active tables to which the data is progressively added as it becomes available.  These tables are:

· Assay Header Table: Dispatch number, sample number sequence, elements assayed, method of assay, detection limits.

· Assay Table: Sample number, reported results for each sample.

· Header Table: Drill hole coordinates, hole depth, hole azimuth, dip, drilling dates, drilling type, bit diameter, etc.

· Geology Table: Logging and sampling intervals, sample number, dispatch number, geology logging data.

Assay files are loaded directly from digital files supplied by the laboratory without any modification or manipulation, therefore avoiding traditional errors from manual data entry.

Original geology logs are hand written in full descriptive text.  The relevant geologist then translates these logs into computer coding on paper data entry forms.  The data on these forms is then manually entered into Excel spreadsheets with identical column names and formats.  Finally, these Excel forms are automatically uploaded into the database.

Data is verified by means of a number of database queries run that detect such errors as mismatched ‘From’ and ‘To’ intervals, drill hole depth greater than last ‘To’ depth, missing sample numbers, duplicated sample numbers, duplicated sample intervals, missing key geology fields such as formation and lithology.  A list of acceptable codes for each geological field is used to prevent incorrect and inappropriate codes being used.

On completion of all data entry, the assay data from the assay table is merged with the geology data from the geology table.  Sample number and dispatch number are used to ensure correct data merging.  Historical data and new data are also combined together at this time to produce two tables that contain data for all holes, which are:

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· Collar and Survey Table: Easting, Northing, Elevation, Total depth, Azimuth, and Dip.

· Geology and Assay Table: Hole ID, Interval ‘From and To’, sample number, preferred assays fields, geology descriptors for each sample interval.

14.3.2 Drill Hole Surveying

Drill hole locations were surveyed using high precision GPS.  Twenty-seven drillholes were re-surveyed by high precision GPS to verify the original collar coordinates.  Easting and northing coordinates were within an average of ±10 mm of the original survey.  Elevation variation fell within two groups, with half-returning values averaging ±0.3 m of the original and the other half-returning values averaging ±2.7 m of the original.  The latter group was surveyed using a different survey base station which resulted in the larger variation measured.  A re-survey of these points will be carried out.

Based on the above results (and ignoring the elevation results from the questionable survey point) the tolerances obtained for all dimensions are not considered serious.

No down hole surveys were carried out.  However, Wardrop feels that the database can be used for resource estimation due to the shortness of the drillholes (average 94 m), the relatively wide drill spacing, and the fact that the vast majority of the holes were drilled vertical.

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15 ADJACENT PROPERTIES

No adjacent properties of similar mineralogy are known in the area.

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16 MINERAL PROCESSING & METALLURGICAL TESTING

16.1 Metallurgy

16.1.1 Background

Treatment strategies for the Boleo polymetallic mixed oxide/sulphide ore were studied in the mid 1990s during the Curator PFS development that resulted in a complex, high capacity flow sheet matching the requirements of a low-grade, ‘super-pit’ design.  The flow sheet featured a combination of roasting, leaching, precipitation, and metal refining.

After acquiring the Boleo property, MMB engaged Bateman of Western Australia (Bateman) to develop a simpler more cost effective method for recovering paymetals from mined ore.  In parallel with the adoption of selective mining of higher grade ore (and significantly reduced waste mining), Bateman proposed a simplified flow sheet via a more direct approach incorporating leaching, solid-liquid separation, solvent extraction and electrowinning.

Key to the revised process was the successful demonstration of the solid-liquid separation characteristics of the leached clay ore followed by an effective process for dealing with the manganese in the pregnant leach solution.  These changes in processing strategy resulted in a more robust, operable flow sheet with reduced operating and capital cost.

Significantly, the flowsheet has been successfully tested in two separate pilot plant campaigns held at the SGS Lakefield facility in Ontario, Canada.

The metal recovery circuits are typical of those deployed in numerous operations worldwide.  In developing the Boleo flow sheet, Bateman was able to incorporate the results of earlier testwork initiatives – some of which dated back by 15 years; supplementing these with more bench scale and pilot plant testwork results, as well as information from Bateman projects featuring similar processes and unit operations.  Further development of the Boleo process plant was progressed during detailed design of the project in 2008 and 2009.

The following summarises the major metallurgical testwork history since 2004 and goes on to describe the proposed process plant.

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16.1.2             Proof of Concept Pilot Campaign

Solid Liquid Separation Testwork

As a precursor to conducting a ‘Proof of Concept’ pilot campaign a bench scale testwork campaign was conducted at SGS Lakefield's facility in Ontario.  Batch leach testing was carried out on six different ore samples from the Boleo Reserve.  The samples varied widely in their copper and cobalt grades and were sourced from widely geographically dispersed outcrops of Manto 3.

The principal objective of the work was to generate leached pulp samples for subsequent settling testwork by specialist vendors (Outokumpu) and consultant representatives (Pocock Industrial).  The samples were tested for their amenability to high rate thickening in a series of tests, the success of which was key to the flowsheet development.

In essence the solid-liquid separation tests conducted at SGS Lakefield Research have demonstrated that Boleo ores can be settled and washed in a conventional CCD circuit utilizing high rate thickeners.  Washing of settled solids to recover metal values from the PLS solution is fundamental to the economic success of the project.

The samples were found to settle best when diluted to between 2.0% and 3.5% solids.  The optimum flocculant dosage for the leach varied from 3 ppm to 6 ppm.  Underflow densities of 20% to 22% were achieved in testwork with clear overflows being produced.

This testwork was sufficiently successful to warrant taking the next step in the process development, namely the operation of a 'Proof of Concept' pilot plant campaign.

Solvent Extraction Testwork

At the time of the above testwork, the SX Group of the Parker Centre, CSIRO Minerals in Perth, Australia developed a process that allowed for the separation of copper, cobalt and zinc from calcium, magnesium, and manganese by solvent extraction with synergistic mixtures of commercially available organic extractants.

The combination of the organic extractants results in an enhanced chemical effect (essentially a 'synergistic' shift in the pH50 of separation) for the system under consideration, allowing efficient separations of metals in solution to take place.  This system of reagents allows for the separation of copper, cobalt, and zinc from manganese, magnesium, and calcium.  This technology provides a conventional means of dealing with the high levels of manganese in solution and is both fundamental to and appropriate for metals recovery from Boleo solutions.

In preliminary testwork at pH 4.5, 100% of the copper was extracted, cobalt extraction was in the range of 94% to 98%; zinc extraction was in the range of 64% to 80% and manganese extraction in the range of 0.76% to 1.55%.  In almost all cases, no calcium and magnesium were extracted at pH 4.5.  The separation factors of copper and cobalt over manganese ranged in the ten thousands and thousands, respectively, suggesting a complete separation of copper

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and cobalt from manganese.  The separation factors of zinc over manganese ranged in the hundreds, suggesting that a good separation of zinc from manganese can be obtained.

The success of this testwork promoted further investigation and after some additional bench scale testing the concept was included in the flowsheet for 'Proof of Concept' piloting.

Proof of Concept Pilot Campaign Testwork

A 'Proof of Concept' pilot plant campaign was conducted at the SGS Lakefield facility from November 16 to 28th , 2004 treating a bulk sample of Boleo oxide ore from manto 3 grading 1.6% Cu, 0.087% Co, 0.58% Zn, 3.23% Mn and 8.71% Fe.  The pilot plant flowsheet comprised:

· attritioning of the ore with grinding of coarse ore particles in sea water to form an ore slurry

· acid oxidation leaching of the ore with sulphuric acid in seawater

· acid reduction leaching of the ore with sulphur dioxide and sulphuric acid in seawater

· partial neutralization with limestone

The picture below shows the oxidative leach, the reductive leach, and the partial neutralization steps undertaken at the SGS Lakefield facility in Ontario.

· Counter current decantation (CCD) washing of the leach residue in thickeners, to separate the metal rich aqueous solution from the clay waste.  The picture below shows four of the six stage CCD thickeners.

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· Copper solvent extraction and electrowinning. The picture below shows the copper solvent extraction mixers/settlers to the left and the electrowinning process to the right, under a fume extraction hood.  The cathodes can be clearly seen suspended in the electrowinning cell.

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· Iron removal by pH adjustment and oxidation with air (and polish with hydrogen peroxide if required).  The picture below shows the iron removal setup.

· Thickening of the iron residue prior to disposal.

· Direct Solvent Extraction technology for selective recovery of cobalt and zinc and small amounts of residual copper.  As mentioned above DSX® technology is the property of Commonwealth Scientific Industrial Research Organization (CSIRO), Perth, Australia.  The picture below shows the setup of the DSX® circuit.

SGS Lakefield and Bateman Engineering have previously jointly reported the summary findings from the pilot plant.  Highlights of the campaign included the following:

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A total of nearly 2 tonnes of ore were treated through the pilot plant.

The pilot plant operated continuously for a total of 12 days in leaching, 11.5 days in CCD, 9.5 days in Copper SX/EW and 9 days in Cobalt and Zinc SX using the DSX® technology.

The oxidation, reduction leaching circuit gave excellent extractions of copper, cobalt and zinc.  Copper extraction exceeded 90% during pilot operation.  Cobalt extraction varied from 80% to 90%.  Zinc extraction was generally above 70%.  These numbers are indicative of the potential of the Boleo process to extract the three pay metals copper, cobalt, and zinc.

The CCD circuit was set up to simulate the use of the “high-rate” type of thickeners with recirculation of overflow solution to dilute the feed slurry prior to flocculation.  This method of settling and washing was based on recommendations from bench-scale testing by Outokumpu and Pocock Industrial and proved to be highly effective.  The leach residue settled quickly, producing clear overflow solutions to advance to copper, cobalt and zinc recovery.

15.5 kg of copper metal were electrowon from the solvent extraction strip solutions at high efficiency.  A picture of the first copper cathode produced during this campaign appears below.  These cathode samples assayed at better than LME grade.

The iron removal circuit was designed to remove iron, aluminium and other impurities from the solution prior to recovery of cobalt and zinc using DSX® technology.  The iron removal circuit consistently produced very low concentrations of key impurities in solution with negligible losses of cobalt and zinc.

· The DSX® circuit for cobalt and zinc recovery performed very well.  In the Lakefield pilot plant, cobalt and zinc were recovered with high overall extraction efficiency (+95%) to produce a concentrated zinc sulphate solution (for subsequent production of zinc sulphate monohydrate crystals) and a concentrated cobalt solution (for subsequent production of cobalt metal cathode).

Two further metallurgical tests were performed on the product streams from the pilot plant:

· Production of zinc sulphate monohydrate crystals.  Zinc sulphate monohydrate was produced by evaporative crystallization of the zinc strip solution from the DSX® circuit.

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·   Production of cobalt cathode.  The cobalt strip solution from the DSX® circuit containing cobalt along with small amounts of zinc and nickel was treated in a zinc solvent extraction circuit followed by a cobalt SX/EW circuit to purify the cobalt solution for electrolysis as high grade (+99.9%) cobalt cathode.  Conventional solvent extraction reagents and process steps were utilized for this purpose.

In addition certain bench scale tests were conducted to obtain additional data for feasibility study purposes including:

·   Environmental testing of residues and solutions produced in the pilot plant program

·   Characterization of High Acid Consuming (HAC) material from the Boleo site containing limestone and other alkali minerals.  HAC material is intended to be used as a low cost bulk neutralizing agent in the commercial Boleo plant

·   Ore scrubbing and grinding testwork for developing final design for ore preparation circuit, including Bond Work Index determination for both ROM ore and HAC material

·   Testwork on oxidation and precipitation of iron from the DSX® circuit feed solution to ensure maximum removal of iron with minimum treatment time and reagent consumption

·   Leach testing on 24 individual samples of ore that were composited to form the pilot plant feed.  These tests were used to assess leach variability and acid consumption variability of the ore, and the production of an acid consumption model.

The successful completion of this pilot program was an important milestone in moving the Boleo project forward.  For the first time Boleo ores had been treated in a continuous pilot plant program to leach, separate and recover pay metals in final commercial form.

16.1.3 Fully Integrated Pilot Campaign

Refinements to the process flowsheet followed in an attempt to optimize the capital and operating costs for the proposed operation as did targeted testwork initiatives designed to support the proposed flowsheet variations.

Targeted DSX® system testwork conducted by Bateman’s Solvent Extraction Group in late 2005 produced an optimum set of conditions for the operation of the DSX® circuit, relating in particular, to the relative proportions of the two extractants and the operating pH.

Boleo limestone, available on the Boleo Reserve, was found to be effective in bulk tailings neutralization duties and iron precipitation in the Iron Removal Stage during bench scale testing at SGS Lakefield.  Soda ash was found to be effective at pH control.  Milling in acidic raffinate was found to bolster circuit tenors and improve the overall 'water balance' for the circuit.

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These various initiatives gave rise to the need to conduct a second pilot campaign with certain targeted objectives.  The campaign was duly scheduled for July 2006 and the pilot plant constructed and commissioned to meet these dates.

·   Objectives for the 19-day pilot campaign included:

·   Building on and verifying the “Proof of Concept” pilot plant results of 2004

·   Demonstrating the flowsheet feasibility on a larger scale and to a greater extent of integration than in the previous campaign

·   Assessing larger scale CCD settling behaviour and performance

·   Testing the suitability of the DSX® technology for the selective recovery of zinc and cobalt from a Mn matrix using an optimised mixture of the synergistic reagents

·   Confirming key reagent consumptions, including the use of Boleo Limestone (HAC) for pH adjustment

·   Demonstrating the use of soda ash in pH control in the various Solvent Extraction circuits as a lower cost replacement for sodium hydroxide

·   Demonstrating the Cadmium Cementation step

·   Demonstrating Manganese Carbonate production

·   Confirming product quality

·   Verifying existing design criteria and confirming current design assumptions

·   Extracting engineering design data

·   Testing the proposed plant control philosophy.

·   In addition to continuous piloting of the Boleo flowsheet, on-site bench-scale testwork was conducted by the following vendors and industry specialists:

·   Outokumpu Technology – CCD & high rate thickening

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·   Pocock Industrial Inc – CCD & high rate thickening

·   RPA Process – Filtration of iron residues, manganese carbonate product

·   Mixtec – Agitation testing in oxidative and reductive leach, partial neutralization and tailings neutralization

·   SGS Lakefield – Environmental characterization of Boleo Pilot tailings

·   SGS Lakefield – Production of purified cobalt carbonate

·   SGS Lakefield – Cobalt removal from DSX® zinc solution.

·   Samples were also sent off-site for treatment by the following:

·   GLV Pty Ltd – Paste Thickening options for the CCD Circuit

·   Saskatchewan Research Council – Tailings pump loop testing

·   Jenike & Johanson Ltd – Flowability testing of crushed ore, materials handling testwork.

Pilot Campaign Ore Feed

Approximately 10 tonnes of Boleo ore, composited from 143 channel samples of underground oxide Manto 3 ore from the test mining exercise conducted by AMDAD and Agapito in late 2005/early 2006, were homogenised, and crushed to produce the feed to the pilot plant.  A 200 kg composite sample was prepared and used for preliminary grinding testwork.

Of the initial 10 tonnes prepared approximately 8 tonnes were milled during the campaign, and approximately 5 tonnes were consumed in the pilot campaign.  The average grade of the elements of interest for the duration of the campaign is shown in Table 16-1.

Table 16-1: Comparison of As Received, 200 kg Composite and Scrubber Feed Assays

Description	H2O %	Cu %	Co g/t	Zn g/t	Mn %	Fe %
Baja Assays – by Chemex (As Received from Site)
Average	27.39	2.0078	1252	4753	3.9601	8.2383
Length Weighted Average		1.9928	1235	4757	3.9464	8.1712
Boleo Bench Testwork – 200 kg SGS Composite Sample
Calculated	27.81	2.1388	1177	4820	3.6855	8.3434
Assayed		1.8200	1130	4020	4.8200	8.0000
Composite Samples Taken off the Feed Belt to Scrubber During Milling Campaigns
Weighted Average		2.1700	1320	4800	5.0400	8.1400

Grade variation of the prepared feed slurry over the 19-day campaign is shown in the following table.

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Table 16-2: Leach Feed Solids Assays

Sample	Cu %	Co g/t	Zn g/t	Mn %	Si %	Fe %
Feed Ore 1A	2.16	1020	4780	4.14	18.7	8.04
Feed Ore 1B	2.25	1230	4650	5.58	19.4	8.15
Feed Ore 2A	2.26	1410	4810	5.99	18.5	8.22
Feed Ore 2B	2.26	1240	5270	4.27	20.4	8.90
Feed Ore 2C	2.27	1570	4990	5.21	19.0	7.90
Feed Ore 3A	2.18	1370	5130	5.25	20.1	8.59
Feed Ore 3B	2.04	1550	4640	5.45	20.6	7.71
Feed Ore 3C	2.04	1370	4990	4.16	21.2	8.58
Minimum Feed Ore Grade	2.04	1020	4640	4.14	18.5	7.71
Maximum Feed Ore Grade	2.27	1570	5270	5.99	21.2	8.59
Average Feed Ore Grade	2.18	1350	4910	5.01	19.7	8.26

Oxidative, Reductive & Partial Neutralization Leach Stages

Recovery of Cu, Co and Zn from Boleo ore requires three atmospheric leach stages.

The first stage is an oxidative leach with sulphuric acid (pH 1.2) at 70°C to 80°C for three hours.  No reagent addition was required to maintain a redox potential of 600 mV to 950 mV due to the natural occurrence of MnO2 within the Boleo ore body.

A second stage reductive leach with sulphuric acid (pH 1.5) at 70°C to 80°C for three hours ensures dissolution of the manganese and associated minerals.  A redox potential of 400 mV is maintained by addition of SO2 gas.  This reductive step releases additional copper, cobalt, and zinc locked within the manganese mineralization.

A partial neutralization stage, conducted at pH 2.0 at 70°C 80°C for one hour follows prior to copper solvent extraction to ensure high extraction efficiencies in copper SX by subtle pH adjustment.  The pH modification is also carried out to minimize gel formation, especially silica gel.

Impurities leached include calcium, iron, magnesium, and aluminum, contributing to the overall acid consumption of the ore.

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Photo 1       Overview of the Leach Circuit.

Counter Current Decantation (CCD)

Due to the fine nature of the feed slurry, counter current decantation was selected as the residue wash step in order to reduce soluble losses.  The CCD circuit setup included six stages of thickeners, feed dilution to 2% to 3% solids, underflow washing with brine and manganese carbonate precipitation thickener overflow at a wash ratio of 1.75 m3/t and flocculation with Hychem 301.

The CCD circuit was operated at 40°C as per the requirements of the downstream solvent extraction circuit.  The final tailings were drummed, with a portion blended with iron removal circuit residue, neutralized with HAC and lime to pH 7, and sent for pump loop testing by the Saskatchewan Research Council and environmental characterization testwork by SGS Lakefield’s Environmental Group.

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Photo 2          Overview of the piloting Counter Current Decantation (CCD) equipment setup.

Copper Solvent Extraction (CuSX)

The Cu SX process utilized 20%v/v LIX 664N extractant and Orfom SX 80 CT diluent.  The circuit consisted of two extractions, one wash, and two strip stages operated at 40°C.

Pregnant Leach Solution (PLS) from CCD 1 overflow was collected and filtered to remove suspended solids using a 1 µm in-line cartridge filter prior to being pumped to a storage container ahead of the CuSX circuit.  The solution was filtered with a 0.5 µm filter prior to entering the first extraction stage.  No pH adjustment was done in the wash and strip stages.

Organic entrainment removal was via a plastic chip coalescer for the raffinate stream and a combination of a plastic chip coalescer and a multimedia activated carbon filter for the loaded electrolyte stream.

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Photo 3                      Overview of the piloting Copper Solvent Extraction equipment setup.

Copper Electrowinning (CuEW)

The CuEW process utilized a conventional circuit with lead-calcium-tin anodes and stainless steel starter cathode sheets, operated at 40°C and a target current density of 250 A/m2 over a nominal four day strip cycle.  The lead anodes were conditioned prior to the start of the campaign to reduce the potential for lead contamination of the copper product.

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Figure 16-1: Cu Solvent Extraction and Electrowinning Flowsheet, and Operating Parameters

Iron Removal

The objective of this step is to precipitate the iron, aluminium, and residual copper as hydroxides by adjusting the pH of the process feed stream (CuSX raffinate) with HAC to pH4, effectively oxidizing Fe+2 to Fe+3.

Initially, the iron was removed in two stages with HAC and air being fed to the first stage, and lime and oxygen fed to the second stage.  The circuit was modified during the campaign by removing one of the original thickeners giving a circuit configuration consisting of a feed tank and four reaction tanks followed by one thickener.

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The iron precipitate was thickened using Hychem 301.  The thickener overflow proceeded to the Co Zn DSX® circuit and the underflow was filtered.  The filter cake was bagged and drummed.  The filtrate was returned to the second reaction tank.  The circuit was operated at 50°Cduring the campaign to be representative of the operating temperature of the industrial plant.

Photo 4          Pilot Plant Iron Removal Reactors and Iron Removal Thickener

CSIRO Direct Solvent Extraction (DSX®)

The DSX® process utilized a 6.25% Versatic 10 (0.33M) and 13.2% LIX 63 (0.3M) v/v organic mix in Orfom 80 SX CT diluent.  This organic mixture was developed for the selective recovery of Co and Zn in a matrix containing high manganese concentrations (typically 20 g/L to 50 g/L Mn).

Extensive bench-scale testing by the CSIRO had indicated selective recovery of Co and Zn from Boleo solutions and this had also been successfully demonstrated in the previous pilot plant campaign.  At that time however, the process suffered the disadvantage of organic degeneration, thought to be catalyzed by manganese in the loaded organic solution at levels greater than 1 g/L Mn,  acerbated by elevated temperature (>30°C).

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Subsequent testwork by the CSIRO and the Bateman Solvent Extraction Group (also known as Bateman Advanced Technologies or BAT, a specialist solvent extraction division within the Bateman Group) confirmed the effect of manganese on the organic stability and lead to a series of tests by CSIRO to optimize both the circuit conditions and the relative proportion of the synergistic reagents.  Testwork successfully demonstrated that degradation can be all but eliminated by ensuring that manganese does not load onto the LIX 63 extractant.

The key process parameter in this regard is to ensure that the pH is maintained below 4.5 pH units.

The circuit consisted of three extraction, two scrub, two Zn strip and two bulk (Co, Zn) strip stages operated at 30°C.

Photo 5        Overview of the Pilot Plant Direct Solvent Extraction (DSX®) Circuit

Iron free solution was collected and filtered to remove suspended solids with an in-line cartridge filter prior to being fed to the DSX® circuit.  pH in extraction was controlled by dosing a 100 g/L sodium carbonate solution; pH in Zn strip and bulk strip stages was controlled by dosing 200 g/L sulphuric acid.  Zn and Co strip solutions were collected for further processing to zinc sulphate and cobalt metal respectively in the second phase of the testwork campaign.

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The DSX® operating parameters were as follows:

Figure 16-2: Direct Solvent Extraction (DSX®) Flowsheet and Operating Parameters

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Manganese Carbonate

The objective of this circuit is to precipitate manganese from the DSX® raffinate stream.  The raffinate from the Co Zn DSX® circuit is fed to the manganese precipitation tank, which is maintained under an inert atmosphere with nitrogen gas.

Sodium carbonate (Na2CO3) is used to precipitate manganese as its carbonate.  The treated stream is processed in a thickener; the overflow is recycled to the CCD circuit to be used as wash water and the underflow is filtered.

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Pilot Plant Controls

Pilot plant controls included regular flow rate/mass measurements, on-line process variable trending (temperature, redox potential and pH), analytical laboratory assay trending, bench titrations, reagent consumption monitoring (flow metres, load cell measurements, containers) and log sheets.  Actual flow rates were cross-checked with the load cell data to produce a full mass balance with the following accuracies:

· solids within 5 % deviation

· solution within 1 % deviation.

A high level process review was performed daily by the management team on site.

Table 16-3 shows the analytical methods utilized by SGS Minerals Services Analytical Group.

Table 16-3: SGS Minerals Services Analytical Methods Employed

Area	Sample Type	Method
Batch Grinding	Pulp	SG
	Solids	Malvern
	Liquids	SG, ICPLa, Cl, Fe2+, FAT
Leach and Partial Neutralization	Pulp	SG
	Solids	ICPS
	Liquids	SG, ICPLa, Cl, Fe2+, FAT
CCD	Pulp	SG, TSS
	Solids	ICPS
	Liquids	SG, ICPLb
Iron Removal	Pulp	SG, TSS
	Solids	ICPS
	Liquids	SG, ICPLa, Cl, Fe
Cu SXEW	Aqueous	SG, ICPLa, Cu, Ge/In/Ga, Cl, FAT
	Organic	SG, ICPO
	Cathode	ICPC
DSX®	Aqueous	SG, ICPLa, Fe, Ge/In/Ga, Cl
	Organic	SG, ICPO

Special sample preparation for both analytical and SG determination samples included the multiple washing (two-stage re-pulp) of solids to minimize errors from soluble metal contributions.

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Commissioning & Campaign Duration

Pilot plant commissioning commenced on 5th June 2006.  The integrated pilot plant campaign ran until 24th June.  The zinc solvent extraction and cobalt solvent extraction and electrowinning circuits were run during the period of 4th to 15th July in a separate campaign.  The start up times for the various sections are shown in Table 16-4 .

Table 16-4: Boleo Pilot Plant Timeline

Section	Time	Date
Start of Campaign Milling	Batch	5 June
Leach	23h00	6 June
CCD1 (Solids feed)	10h30	7 June
CuSX	10h45	8 June
Fe Removal	19h30	8 June
CuEW	09h00	10 June
DSX®	01h00	10 June
MnCO3	10h30	11 June
First Cu Cathode	15h00	13 June
End of Campaign
Leach	06h00	24 June
CuSX	23h00	25 June
Fe Removal	07h00	27 June
DSX®	07h00	27 June
MnCO3	07h00	27 June

Summary of Piloting Results

Leach and Partial Neutralization

Table 1-1 shows the leach and partial neutralization operational control parameters.

Table 16-5: Leach and Partial Neutralization Operating Parameters

Parameter	Unit	Oxidative Leach		Reductive Leach		Partial Neutralisation
		Tank 1	Tank 2	Tank 1	Tank 2
Temperature	ºC	80	80	80	80	80
pH (Ag/AgCl – sat’d KCL)		1.5	1.4-1.7	1.2-1.5	1.2 -1.5	2.06
Redox Potential	mV	>800	>800	400	400	~400

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The average leach extractions for the various pH conditions of the campaign in Oxidative Leach Tank 2 are shown in Table 16-3.  The figures were calculated using a silica tie method to compare the metals values in the feed with the values in the washed residue from CCD.

Graphs of the metals extractions for the duration of the pilot plant campaign are shown in  Figure 16-3 to Figure 16-6.

Table 16-6: Leach Efficiencies

Condition	OL2 pH	% Extraction Cu	% Extraction Co	%Extraction Zn	% Extraction Mn
OL2 pH = 1.7	1.7	90.5	80.8	52.9	96.6
OL2 pH = 1.5	1.5	90.5	79.4	54.4	95.8
OL2 pH = 1.4	1.4	92.4	82.1	60.1	97.7
Start-up	1.2	94.1	89.3	71.6	94.9
Overall average	1.4	91.8	82.4	59.1	96.5

Figure 16-3: Boleo Metal Extractions – Cu

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Figure 16-4: Boleo Metal Extractions – Co

Figure 16-5: Boleo Metal Extractions – Zn

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Figure 16-6: Boleo Metal Extractions – Mn

Note:  The following abbreviations apply – Leach Feed (LFS), Oxidative Leach (OL2), Reductive Leach (RL2), Partial Neutralization (Neut 1) and CCD6 Underflow (CCD 6 U/F).

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Counter Current Decantation (CCD)

The CCD circuit average operational control parameters were as shown in Table 16-7.

Table 16-7: CCD Operating Results

Parameter	Unit	CCD 1	CCD 2	CCD3	CCD4	CCD5	CCD 6
U/F Density	% solids	10	12	14	15	16	15
Wash Ratio	m3/t	1.75
Wash Efficiency	%	91
Flocculant Addition Total	g/t	470

Figure 16-1 shows the improved operational performance of the CCD circuit as the campaign progressed.

Figure 16-7: Boleo Pilot Plant CCD Thickener Mud Levels

Observations:

·   The CCD circuit produced underflow densities ranging from 10% to 24 % solids as reported on a shift basis.  However, the underflow density calculation is dependent on using a fixed value for the SG of the dry solids and small variations in the SG will affect the calculation.  As such the underflow densities obtained during vendor testwork (shown in Table 16-8) give a more accurate indication of the values attainable.

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·   Wash efficiencies as high as 99% were achieved but only at wash ratios in excess of 4 m3/t.

·   CCD 1 overflow clarities were generally good with clarity wedge readings of 47 on a 0 to 47 scale.

·   There is potential for two-stage flocculation to reduce flocculant consumption.

·   Analysis of the CCD6 underflow assays revealed that additional leaching was taking place in the CCD circuit.

Vendor Testwork

Outokumpu Technology and Pocock Industrial Inc. carried out thickening testwork during the campaign.  In addition, samples were sent to GL&V for paste thickening testwork.  Table 16-3 shows a summary of the vendor results.

Table 16-8: Vendor Testwork – CCD Parameters

Parameter	Outokumpu	Pocock	GLV
Diluted feed, % solids	2-3	2-3	1.5
Max underflow, % solids	16.1	16.1	21-22
Overflow clarity – TSS, ppm	60-85	80-130	-
Optimum rise rate, m/h	1.64	0.87	-
Typical bed height, mm	180-230	1000	300
Flocculant	Hychem NF301	Hychem NF301	Hychem 302

It was noted that underflow densities were as much as 6% (absolute) lower at similar rise rates to those achieved in comparative tests conducted by Outokumpu in 2004.  The reasons for the difference in settling performance are not well understood but it is postulated that they result from the combined effects of milling in acidic raffinate and the use of Boleo limestone for neutralization purposes.

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Copper Solvent Extraction (CuSX)

The CuSX circuit extracted an average of 98.6% Cu.  The average pregnant leach solution and raffinate assay values are shown in Table 16-9.

Only minor amounts of manganese and iron were co-extracted with the copper.  The average solution and organic assays at different stages in the SX process follow :

Table 16-9: Copper Solvent Extraction Solution Assays

Element	Unit	PLS	Raffinate
Cu	mg/L	2515	34.5
Co	mg/L	265	257
Zn	mg/L	1023	976
Mn	mg/L	11710	11601
Fe Total	mg/L	5916	5837
Ca	mg/L	601	589
Mg	mg/L	7573	7442
Al	mg/L	3369	2901
Ni	mg/L	28	28
Si	mg/L	50-170	<140
Cl	mg/L	14535	13570

Table 16-10: Copper Solvent Extraction Organic Assays

Element	Unit	Organic Assays
		Loaded	Scrubbed	Stripped
Cu	mg/L	5620	5417	3200
Mn	mg/L	0.10	0.07	<0.05
Fe Total	mg/L	36	8	0.9
Ni	mg/L	< 2	< 2	< 2

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Copper Electrowinning (CuEW)

Four plating cycles were carried out during the campaign.  The conditions are shown in Table 16-11.

Table 16-11: Plating Cycle Times and Operating Conditions

Cycle Number	1		2		3		4
Start Time	June 9 17:50		June 13 15:33		June 16 18:43		June 21 15:18
End Time	June 13 15:06		June 16 11:51		June 21 15:18		June 25 23:18
Cathode Mass (kg)	A	B	A	B	A	B	A	B
	8.78	8.88	4.41	4.52	5.74	5.88	4.52	4.61
Av. Current Density,A/m2	183		162		155		167
Current Efficiency, %	94.4		97.3		99.5		99.6

Although SGS Lakefield was a certified assayer for the COMEX and NYMEX exchanges only, cathode quality was assessed against COMEX/NYMEX and LME specification.  Three out of four cycles produced copper quality exceeding the LME grade-A specification.

(see Table 16-3).  The third cycle was out of specification due to a lack of control in impurity carryover.  Generally cathode quality was considered excellent when the circuit ran well and impurities were controlled.

Table 16-12: Cathode Quality

Sample Number	LME Purity %	S ppm	Fe ppm	Pb ppm
1A	>99.994	<20	<4	<3
1B	>99.996	<15	2.4	<1
2A	>99.991	<20	17	<3
2B	>99.993	<20	12	<3
3A	>99.995	<15.7	1.4	<1
3B	>99.993	<27.3	13	<1
4A	>99.996	<15.6	1.3	<1
4B	>99.994	<19.3	<4	<1

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Photo 8       Cathode from Electrowinning Cycle No.1 – 13 June  2006

Iron Removal

The occasionally erratic performance of the iron removal circuit did not adversely affect the downstream DSX® operation.  Many of the operational problems resulted from the relatively small size of the pilot plant equipment.  The blockage of spargers and the poor oxygen utilization are not expected to occur in the industrial plant with the frequency that they did in the pilot plant.

The average concentration of iron in the feed stream was 5,454 mg/L and in the thickener overflow (feed to DSX®) it was 17 mg/L giving an average iron rejection for the campaign of 99.7%.

It must be noted that the iron levels in the thickener overflow were generally less than 1 mg/L and the average quoted above is a result of excursions resulting from blockage of oxygen spargers and / or limestone (HAC) addition.

The iron removal thickener in the pilot plant gave an average percent solids in the underflow of 49%.  This was supported by Outokumpu testwork results which reported a maximum underflow density of 46% solids, at a rise rate 3.81 m/h and an overflow clarity of 72 ppm solids.

Observations:

·   In general the iron precipitate settled and filtered easily and gave a clearer overflow at higher temperatures.  The initial operating temperature of 60°C was dropped incrementally to 45°Cover a period of two days from 15th June.  However, this resulted in poorer settling performance, reduced overflow clarity, and poorer filtration characteristics and a decision was taken to run the circuit at 50°Cfor the remainder of the campaign.

·   The occasional use of hydrogen peroxide to “trim” the remaining ferrous iron did not cause problems with the DSX® extractant during the campaign.

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·   A combination of coagulant (Magnafloc 368) and flocculant (Magnafloc 155) was found to best flocculate the iron removal thickener feed stream.

CSIRO Direct Solvent Extraction (DSX®)

The DSX® circuit ran trouble-free for the 425 hours of the campaign.  Crud was formed in the extraction stages but did not adversely affect the physical or chemical behaviour of the circuit.  The crud is likely to have been caused by particulates or precipitates in the feed stream and was not formed by products of organic degradation as was the case in the first pilot plant campaign in November 2004.

None of the black “manganese crud” of 2004 was evident and this is attributed to the excellent control of pH that ensured low manganese loading on the organic at approximately 100 mg/L throughout the campaign.

Figure 16-8: Manganese Loading On DSX® Organic Phase

Metal Extraction

The DSX® extractions for copper, cobalt, and zinc are shown in Table 16-3.

Table 16-13: DSX® Metals Extractions

Description	Unit	Cu	Co	Zn
Extraction	%	99.57	99.48	99.03

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Zinc Selective Strip

Figure 16-9 shows that a zinc tenor in the zinc strip liquor of 40 g/L to 45 g/L can be maintained while keeping a Zn:Co ratio of approximately 50:1.

Figure 16-9: Zinc in Solution vs. DSX® Operating Time

Bulk Strip

Figure 16-10 shows the cobalt and zinc tenors in the bulk strip liquor.  Cobalt concentrations in excess of 10 g/L were attained while maintaining a Zn:Co ratio of 1.6:1.

Figure 16-10: Cobalt in Solution vs. DSX® Operating Time

Observations

·   Excellent pH control in extraction was achieved with 100 g/L sodium carbonate.

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Manganese Carbonate

Manganese carbonate was precipitated from the DSX® raffinate by the addition of 150 g/L sodium carbonate slurry.  The sodium carbonate was added at 80% of stoichiometric requirements to minimize the co-precipitation of calcium and magnesium.  Table 16-14 shows the average results of 19 of the 21 batches produced.  Intermittent overdosing of the sodium carbonate resulted in some low manganese values but in general values of 45% or greater were achieved.

Table 16-14: Manganese Carbonate Product Assay Values

Element	Value	Units
Mn	44.2	%
Ca	1.1	%
Mg	0.2	%
Al	0.4	%
Fe	173	g/t
Co	29	g/t
Zn	201	g/t
Al	0.4	g/t
Ni	365	g/t
Si	443	g/t
Cu	< 5	g/t
Cd	< 5	g/t

Observations:

· The pale pink manganese carbonate precipitate settled, filtered and washed well.

   ·   Seeding of the feed from 1.4% to 5.8% solids with recycled underflow led to better flocculation, improved underflow percent solids (62% to 67%) and clearer overflows (47 on the clarity wedge).

Reagent Consumption

Consumption (kg of reagent per tonne of dry feed) of the key reagents in the leach and CCD circuits for various pH conditions is shown in Table 16-3.

Table 16-15: Summary of Leach and CCD Reagent Consumptions @@RE DO TABLE

Condition	RL2 ORP	H2SO4 Rate (kg/t)	SO2 Rate (kg/t)	HAC Solids Rate (kg/t)	Flocculant Rate (kg/t)
OL2 pH = 1.7	399	228	81	69	0.53
OL2 pH = 1.5	397	228	71	73	0.40
OL2 pH = 1.4	399	330	129	116	0.46
Start up	427	496	144	362	0.55
Overall average	404	311	106	138	0.47

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16.1.4 Variability Testwork

In March and April of 2007 a bench scale testwork programme was initiated at SGS’s Lakefield Research laboratories with a number of objectives:

·   to increase the level of confidence in the design acid consumption values which in turn affect the size of the acid plant needed for the project.

·   to assess leach recovery variability across the orebody.

·   to establish and increase confidence in the grade-recovery relationships for each of the paymetals.

To this end a total of twenty composite samples were produced from approximately 350 individual drill core samples.  The drill core samples originate from across the Boleo Reserve, covering an area of some 10 km2, representing some 25 to 30% of possible resource.   The 350 individual core samples were predominantly manto 3 oxide material with about 15% of the samples being sourced from manto 2.

The samples were composited to reflect the first five years of feed to the plant according to the [then] mining plan.  Each of the samples was therefore designed to represent 4 months worth of plant feed.

Twenty individual bench scale leach tests were carried out, simulating the proposed process flowsheet, to generate 20 data sets of grade-recovery relationships for each of the copper, cobalt, zinc and manganese elements as well as the acid consumption levels required to achieve the leach recovery.

Two additional leach tests were added to the series where leach testing was conducted on feed ore from the Fully Integrated Pilot Plant campaign so as to allow a direct comparison to be made with pilot campaign results.  Intermittent duplicate analyses were also undertaken to provide additional confidence in the results.

The summary metal extractions for the work reported are shown in Table 16-16. The 90% average copper extraction achieved in the variability work compares favourably in the 2006 Demonstration Pilot Plant Results.

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Table 16-16 : Variability Testing Extractions of Principle Elements, wt%

	Copper			Cobalt			Zinc			Manganese
	OL	RL	+48 h	OL	RL	+48 h	OL	RL	+48 h	OL	RL	+48 h
Average	86	89	90	46	58	62	47	54	58	76	92	93
Maximum	95	95	96	63	76	78	70	77	81	97	98	98
Minimum	73	77	80	25	29	34	27	29	34	26	52	55
Std Dev, %	7	5	4	24	21	19	27	22	21	31	11	11

The cobalt extraction was on average lower then that achieved in the Demonstration Pilot Plant. The reasons for this are not clear but are believed to relate to a combination of the ‘aged’ nature of many of the samples and the limitations of the bench-scale test rig.  Specifically with the particular SGS test rig used for this work, it was difficult to stir effectively gas into the pulp due to the use of an unbaffled reactor and as a result, it was difficult to maintain a 400 mV ORP throughout reductive leaching.  This value of 400mV is the value found in the pilot plant campaigns to give effective Co and Mn leaching.

The extraction of zinc was good at 58% and compared well to the pilot plant results.  The Mn extraction on average was slightly lower than the pilot plant results at only 93%.  It is believed that this lower Mn extraction is again symptomatic of the physical limitations of the bench-scale test apparatus with respect to sulphur dioxide mixing and take-up in the leach slurry.

The use of acid averaged 185 kg H2SO4/t for the variability testwork (excluding two particular tests where surplus acid was added due to a control loop malfunction and the two ‘sighter’ tests on pilot plant feed material).  This value compares favourably with historical consumption vales and is firmly in line with the acid consumption levels experienced during the two successful pilot campaigns.

The range of acid use was 29 kg/t to 500kg/t for this particular series of tests.  This is clearly too large a variation to be sustained in an operating plant over any extended period.  It will be necessary to blend the ore on stockpiles to achieve a more consistent use of acid in leaching.

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16.1.5 Flow Sheet Improvements

A number of improvements were incorporated into the design of the Boleo plant during detailed design and engineering of the project in 2008 and 2009.  The modifications are expected to improve the constructability, operability and maintainability of the process plant with corresponding economic benefits for the project.

Sea Water Milling

The DFS flowsheet utilised a recycled copper raffinate solution as the mill dilution.  A fixed quantity of raffinate, from the copper solvent extraction circuit, was added to the scrubbing and milling operations before the final milled product proceeded to the leach circuit.  The combination of high levels of sulphuric acid and chlorides present in the raffinate creates very aggressive process conditions in the comminution operations.  To overcome corrosion concerns and to improve long term maintainability, the milling operations are now conducted in sea water at a neutral pH.

The change in milling solution composition is not expected to have a strong influence on the attrition aspects of the scrubbing operation.  This has been confirmed by washability characteristics testing of a range of seven samples by SGS Lakefield in June 2008.

In order to minimize sea water input in the downstream leaching circuit, gravity thickening of the mill cyclone overflow slurry has been employed.  Due to the significant and variable proportion of montmorillonite clays in the Boleo ore and the nature of the leaching process, solid/liquid separation of leached solids has proven difficult in the Boleo process.  However, prior to subjecting the ore to an aggressive hot acid leach, milled ore settles relatively well.  These effects were confirmed by measuring settling behaviours on four ore samples by Emmett Process Consulting and FLSmidth in 2008.  The improved ability to thicken milled ore further improves the ability to control the water balance in the milling/scrubbing/screening operations.

Utilising sea water for milling has further key water balance implications.  Removing raffinate milling reduces the volume of recycled solution and results in an increased volumetric flow rate to some unit operations downstream of copper solvent extraction – primarily iron removal and direct solvent extraction (DSX®).  These downstream operations have been resized accordingly. In addition, any acid or copper present in raffinate previously bound for milling is no longer recovered through recycling.

Sea water milling of ore was tested in the 2004 Pilot Plant and excellent leach extractions were achieved in the oxidising and reducing leach stages.  The retention times for these stages have therefore not been changed.  In the 2004 Pilot Plant loss of Cu occurred after the reducing leach due to difficulties in controlling pH in the partial neutralisation process.  This was resolved

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in the 2006 Pilot plant and therefore overall leach extents have not been changed.  It is expected that using milled ore for neutralisation of leach discharge solution will contribute to  a certain amount of leaching.

A benefit of the sea water milling is the reduction of the circulating iron tenor in the leach circuit. This allows for greater control of acid addition in the oxidative and reductive leach processes.  In the oxidative leach, ferrous ions are oxidized to ferric ions and in the reductive leach the ferric ions are reduced back to ferrous ions.

Oxidation of Iron

Reduction of Iron

As can be seen in the above reactions, acid is consumed in the oxidative leach and generated in the reductive leach.  A decrease in the iron tenor in these circuits limits the influence of the leach chemistry on the acid balance giving a greater degree of pH control.  This decrease in circulating iron tenor also accounts for a decrease in SO2 consumption in reductive leach.  The capacity of the SO2 production plant has however not been changed.

Counter Current Neutralisation

Partial neutralisation of the leach discharge slurry is required to achieve maximum copper recoveries in the downstream copper solvent extraction circuit.  In the DFS flowsheet this was achieved by addition of Boleo limestone (a local low calcium carbonate grade source) to the leach discharge slurry in a single partial neutralisation stage.  The solution from this neutralised slurry was recovered through a counter-current decantation (CCD) circuit and then forwarded directly to solvent extraction.

In the improved flowsheet, the leach discharge slurry proceeds to the CCD circuit and the solution overflow from the first CCD stage is neutralised with milled ore slurry prior to solvent extraction.  The modification gives the ability to use the natural neutralising capacity of Boleo ores as a replacement for limestone.  A split of 40% of milled/thickened ore is contacted with the CCD overflow solution.  Boleo ore was proven, over a range of six samples, to be an effective neutralising agent, successfully raising the pH of the partially leach solution above 2 (sufficient for high copper extraction) in testwork conducted by SGS Lakefield and Emmett Consulting in 2008.  The neutralising reaction is rapid and the neutralising ore proved incapable of “over neutralising” the solution above pH 2-2.5 even in the presence of excess ore giving improved process control.

Following neutralisation, the slurry is treated in an EIMCO reactor clarifier which primarily acts to clarify the PLS solution advancing to solvent extraction, reducing crud formation and solids

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build-up in the PLS pond.  The reactor clarifier also performs a solid/liquid separation function to recover neutralising ore to the front end of the leach ensuring 100% of the ore passes through the full retention time of the leach circuit.  Only a minor increase in solution ORP was observed in the neutralisation tests.  This indicates that the combination of low reaction temperature and short residence time was insufficient for significant reaction of manganese dioxide, present in the ore, with high levels of ferrous in solution.  This preserves the natural oxidizing capacity of the ore and allows it to be fully utilised in the oxidative leach where it provides the ferric ions necessary to leach secondary copper sulphides that may be present.

Allowing the leach terminal acidity to be effectively consumed by fresh ore reduces the overall acid consumption.  However, the reduced acid consumption in the plant will not result in decreased power production as the acid plant will continuously run at full capacity to meet electricity demand rather than tracking plant acid demands.  Provisions are in place to export acid that exceeds process acid demand.  The reduction in acid consumption has also affected the leach heat balance as the heat of acid dissolution represents a significant heat input.  A steam heat exchanger has been added to the leach circuit which provides the benefit of increased control of leach temperature.

The introduction of the counter-current neutralisation significantly reduces overall limestone demand which is supplied onsite and supplemented with purchases if required.  Instead of realizing the benefit of a smaller limestone milling and handling system, the size of this system is unchanged so that lower grades of calcium carbonate can be milled to meet plant neutralisation demands.  This has extended the amount of Boleo limestone available for operation.

The removal of Boleo limestone neutralisation reduces the solids loading to the CCD circuit.  Furthermore it has been suggested that the addition of Boleo limestone adversely affects the thickening characteristics of the leach residue.  These two effects create the opportunity of reducing the settling area requirements of the CCD circuit.

This improvement is not expected to alter the leaching extraction achieved when using limestone.

Counter Current Decantation Circuit

In March 2008, a variability test campaign was conducted at SGS Lakefield to generate representative samples, over seven ore samples, for solid/liquid separation tests.  Ore preparation and leaching protocols were developed and observed for all ore types.  The leaching parameters were similar to those in the previous variability leach study performed in 2007.  Solid/liquid separation testing was conducted by FLSmidth personnel at the SGS Lakefield site immediately after receiving the hot leach product.  Analysis of settling test work by FLSmidth resulted in CCD thickener unit settling rate of 0.35 m2/tpd at an underflow solids density of 22%.  This resulted in a decrease in thickener diameter to 60m from 75m in the DFS flowsheet.  Furthermore, the CCD circuit now employs high density thickening (HDT) rather high

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rate thickening (HRT).  HDT is similar to HRT but has the advantage of increased underflow densities due to the greater compressive forces gained by employing a deeper mud bed.

In addition to a reduction in thickener sizing, a thickening inter-stage mixing efficiency study was conducted by Emmett Process Plant Consulting that was reported in April 2009.  This study confirmed that a mixing tank residence time of 45-60 seconds is sufficient to yield an efficiency of 98-100%.  This increase in mixing efficiency results in an increased wash efficiency and/or a reduction in wash requirements.  For design purposes a 2 minute residence time yielding 98% mixing efficiency has been selected.  A review of the parallel raffinate wash scheme resulted in a 30% decrease in the PLS flow to copper solvent extraction while still optimising copper recovery.  The decrease in PLS flow has resulted in a corresponding decrease in equipment sizing in solvent extraction.

Heat Recovery

Recovering heat from the hot leach discharge slurry is a key component in the heat balance of the Boleo flowsheet reducing overall energy demand.  In addition, the PLS advancing to solvent extraction must be cooled to avoid significant evaporative organic losses.

In the DFS flowsheet, heat was recovered in five duty spiral heat exchangers from the hot leach discharge slurry by crossing it with recycled raffinate bound for the comminution circuit.  The partially cooled slurry was further cooled in two duty evaporative cooling towers before proceeding to the CCD circuit for dissolved metal recovery.

Due to the corrosive nature of the leach slurries, materials of construction of the spiral heat exchangers were to be titanium – a costly option.  Abrasion was also a concern to these units due to a small fraction of a hard component in the leach slurries.  Indications from comparable commercial installations indicated that these units could have a lifespan as low as 12-18 months.  The cost and longevity concerns surrounding these units were deemed unacceptable and dictated that an alternate heat recovery strategy be pursued.

The spiral exchanger/cooling tower concept has been replaced with a vacuum evaporative  cooling system where steam is flashed off the hot leach discharge slurry and used to heat raffinate and sea water streams bound for the leaching and comminution circuits respectively.  The proposed system comprises four sequential agitated flash tanks designed to cool the slurry from 77°C to 50°C.  Slurry enters the flash tank via bottom centre inlets and cascades to the next tank via an overflow weir.  The vapours generated in each flash tank contain liquor droplets which are removed in order to obtain a high purity condensate stream with low amounts of solids and dissolved salts.  To obtain this purity, high-efficiency vane type entrainment separators are installed in each vessel.  The incoming vapours enter the bottom of the elements in the condensers, where heat is recovered to the raffinate and sea water process streams that flow counter-current to the condensate.  The recovered process condensate reduces the overall plant desalinated water demand by one third.  Accordingly, the design capacity of desalinated

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water production has been reduced to 200t/h.  The quality of the distillate as well as the thermal properties of the slurry have been confirmed by Hazen Research.

IRON REMOVAL

Following copper recovery, the removal of iron from raffinate is required before advancing solution to cobalt/zinc recovery.  This is achieved via a combination of pH adjustment and oxidation leading to precipitation.  Ferrous iron is oxidized and precipitates as a ferric hydroxide. Aluminum will also precipitate in this step.

An integral aspect of this process is near complete oxidation of ferrous to ferric.  This ensures that iron can be precipitated down to the very low levels required for DSX®.  The DFS flowsheet proposed a 3 hour reaction time, sparged air, and a terminal pH of 4.2.  During the 2006 pilot plant, the pilot equipment sizing was necessarily such that these conditions were insufficient for complete oxidation of ferrous, and a stronger oxidant – peroxide – was employed.  However, it was never envisaged that peroxide would be employed during operations as ferrous oxidation rates are enhanced due to the increased uptake of oxygen at full scale.

To make the process more robust, further investigations into iron removal were conducted by FLSmidth Minerals and Emmett Process Consulting in 2008.  This testwork was once again constrained by the lab-scale equipment.  The testwork indicated that complete oxidation can be achieved with sparged air in 3 hours at an increased pH of 5.0-5.5.  This higher pH would introduce the risk of high cobalt and zinc co-precipitation but can be mitigated by a two-stage precipitation circuit.  Under this scenario the majority of iron would be oxidised and precipitated in the first stage and the resultant slurry would undergo thickening and washing (same as in a single-stage).  The thickener overflow advances to second stage precipitation at a higher pH - precipitating the remainder of the iron as well as some cobalt/zinc.  This slurry would be thickened with the underflow reporting back to the first stage allowing the precipitated pay metals to redissolve.

It is expected that a single stage precipitation circuit will be sufficient when the benefits of full scale operation are realized.  As a fallback position, space has been left in the plant layout to include the second stage iron removal precipitation circuit.

Direct Solvent Extraction

The development history of the DSX® system in the Boleo flowsheet has been detailed in the DFS.  In general, cobalt and zinc are separated from calcium, magnesium, and manganese in a process developed by CSIRO in Australia.  This process makes use of a so-called synergistic effect obtained from the use of two different extractants in combination.  This combination produces a result that neither extractant can individually match. The organic phase consists of the two synergistic extractants, namely Versatic Acid and LIX 63 dissolved in an appropriate diluent.

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A review of previous testwork suggested that further opportunity existed for DSX® optimisation in light of the updated flowsheet.  In August 2009, testwork was conducted by the University of British Columbia to optimise the DSX® system through reducing the extraction concentrations in the solvent system while still facilitating maximum loading of cobalt and zinc and maximum rejection of manganese and other impurity elements.  The ratio of Versatic Acid/LIX 63 was fixed at 1.1:1 as per the previous CSIRO optimisation work.

This work shows that it is possible to dramatically reduce the extractant concentrations while maintaining cobalt loading.  However, if zinc loading is to be maintained at a high value, it is essential to maintain high extractant concentrations.  Organic loading of 0.10-0.12 g/L Co+Zn per percent extractant (LIX 63 + Versatic) should be sufficient to maintain high cobalt and zinc recovery while essentially completely rejecting manganese.  This allows for a reduction in extractant first fills which represents a significant proportion of all process first fill costs as well as extractant top up costs.

A further series of DSX® optimisation tests were conducted in September 2009 at UBC.  This testwork investigated the effect of a reduced LIX 63 concentration whilst maintaining a fixed Versatic Acid concentration.  The assumption behind this series of testing is that Versatic Acid accounts for the majority of the metal loading and LIX 63 acts as synergist to improve the overall selectivity.  This work shows that, as the LIX 63 concentration is decreased, the loading of metals is slightly reduced, however even at low concentrations of LIX 63 it is possible to maintain total zinc and cobalt loading while rejecting manganese.  This reduction in LIX 63 concentration represents an opportunity to further reduce first fill quantities.  However the benefits of this work have not yet been incorporated into the process design.  Further testing is scheduled to confirm the design parameters (specifically extraction and stripping isotherms) necessary to implement this process improvement.

Cobalt Ion Exchange Circuit

In order to produce high quality cobalt metal, it is necessary to remove various impurities from the electrolyte solution prior to advancement to electrowinning.  This is typical done in an ion exchange (IX) circuit tailored to the removal of dissolved impurities – namely nickel, copper and zinc.  The DFS flowsheet had provision for two IX circuits, one to remove copper and zinc and a second to remove nickel.  However review of continuous pilot plant operations have shown that nickel protection is not necessary to ensure a high quality cobalt cathode product due to low nickel levels.

The Boleo orebody is naturally low in nickel, only a couple hundred mg/L were observed to advance to cobalt solvent extraction.  This step, which concentrates and purifies the cobalt solution and which precedes cobalt IX, employs a phosphinic acid extractant.  Phosphinic acid extractants offer excellent separation factors for cobalt over nickel and are used in that duty for numerous industrial applications.  The provision for four stages of extraction and two stages of scrubbing ensures that the majority of co-loaded nickel is crowded off the organic by cobalt.  As a result, nickel levels in the cobalt strip solution, advancing to IX, are typically 40-60 mg/L.  Not

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only is this nickel concentration low enough to ensure a high purity cobalt product, but it is also near the lower limit of nickel concentrations achieved in cobalt electrolytes treated by IX with Dow M-4195 and that family of resins

The combination of a low initial nickel concentration and a solvent extraction system that is highly selective for cobalt over nickel means that an IX system for nickel removal is not necessary.  It has therefore been removed from the flowsheet.

Tailings Neutralisation

All streams that report to the tailings storage facility will do so via a tailings neutralisation facility.  The DFS flowsheet had provisions to neutralise excess acid and precipitate metal salts through the addition of blower air, ground carbonate ore, and slaked lime.  The present flowsheet provides only the addition of ground carbonate ore resulting in a decrease in terminal pH from 7 to 6. This still ensures all excess acid will be neutralised, however dissolved metal salts will precipitate to a lesser extent.  The change is expected to meet the operational and environmental criteria that have been defined for management of tailings operations.

Sulphur Supply

Sulphur is a vital reagent to the Boleo process and it provides the acid required for leaching and facilitates the generation of a significant proportion of power.  The sulphur supply was

previously delivered to site as molten sulphur.  Prill sulphur is now the preferred delivery method for primarily logistical reasons.  With a larger range of suppliers, the security of supply has been strengthened.  Furthermore it provides additional freight cost saving by taking advantage of “backhaul” opportunities that exist in nearby shipping lanes.  Additional equipment has been added to the plant design to provide for delivery and storage of prill sulphur as well as appropriate sulphur melting facilities to feed the sulphur burners.

Surface Ore Transportation

An economic trade-off study between two surface ore transportation systems provided the basis for the decision to switch from overland conveyors to using haul trucks.  A key benefit of trucking is that it provides greater flexibility for blending and stockpiling operations.

Limestone supply

The changes to partial neutralisation and tailings neutralisation have reduced the demand of limestone to the extent that purchased limestone is not required.  Further the grade of limestone that the plant will be able to use has been reduced.  Testwork was conducted at SGS Lakefield to characterise the chemistry, mineralogy and performance of different grades of limestone.  The results indicate that the reactivity of lower grade limestone should be sufficient for use on the plant.

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Reagent Consumption

A short experimental batch leaching program was undertaken at SGS Lakefield during 2009 to investigate the absorption of SO2 in the reducing leach and the rates of leaching of Cu, Co, Zn and Mn.  The test work results gave lower extraction efficiencies than obtained in previous pilot, bulk and bench scale leaching tests.  While the indications are that there could be near complete uptake of SO2 during continuous operation of the Boleo reducing leach it was not possible to complete the SO2 mass balance.  Ongoing work on improving experimental procedures is in progress as part of detailed process design activities being undertaken to finalise equipment selection and establish metallurgical operational practises for the site.

16.2 Process Plant Design

16.2.1 Introduction

The proposed treatment route for the Boleo ore consists of a two-stage, whole of ore, sulphuric acid and sulphur dioxide leach followed by solid-liquid separation in a counter current decantation circuit prior to solution purification and metal concentration using solvent extraction technologies.  Cobalt and copper will be electrowon to produce high purity products for export to global metal markets.  Zinc will be recovered in a fluid bed granulator as a zinc sulphate monohydrate salt suitable for incorporation into fertilizer and animal feed applications.

Sulphuric acid is manufactured on the site in a stand-alone acid producing facility, employing sulphur as the main feedstock.  Electricity is produced by harnessing the energy in the steam produced during the acid production process in a so-called 'co-generation' facility consisting of a steam turbine and generator.  Limestone, which is available on the mine site, is crushed in the field and milled on site to provide for plant bulk neutralization duties.

The schematic flow diagrams summarizing the proposed plant processing route are included in Figure 16-11  to Figure 16-13.

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Figure 16-11: Schematic Flow Diagram Sheet 1

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Figure 16-12: Schematic Flow Diagram Sheet 2

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Figure 16-13: Schematic Flow Diagram Sheet 3

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16.2.2 Process and Related Plant Description

The proposed process and related plant consists of the following main areas:

ROM Ore Stockpile - The ore from underground mining will be stockpiled near the underground portals.  Ore from the portal stockpiles will be loaded onto 85 t capacity haul trucks by front-end loaders and transported directly to the Run of Mine (ROM) dump hopper or to intermediate or ROM area stockpiles.  Intermediate stockpiles with lower grade material will be reclaimed at a later date and sent to the ROM hopper via haul truck.  The haul trucks will dump the ore into a 200 t capacity hopper.  Ore will be metered from the hopper by a variable speed apron feeder which will discharge onto a discharge conveyor for transfer onto the ROM stacker conveyor.  A traveling tripper, on the stacker feed conveyor, will discharge the ore onto the partially luffing stacker for placing the ore on the 245,000 t capacity blended stockpile.  One half of the stockpile will be built with the stacker traveling back and forth to create a blended pile.  The other half of the pile will be reclaimed by front-end loaders that will discharge into one of three hoppers, from where the reclaimed ore will be fed by variable speed belt feeders onto the reclaim conveyor at a nominal rate of 395 tonnes per hour.

Secondary Ore Sizing - The reclaim conveyor will feed the ore to a toothed-roll type sizer for size reduction of the ROM ore.  The fines in the ore pass through the sizer.  There will be two sizers with one unit in operation and the other in standby for  maintenance

Scrubbing, Screening, Pebble Crushing & Milling - The ore from the sizer will drop, by gravity, to a rotary scrubber designed to slurry the high clay component of the ore.  The scrubber is a rotating mill without grinding balls.  Seawater will be fed to the scrubber, along with crushed ore, and the resulting mixture will be tumbled such that clay and ore fines will be scrubbed (washed) off the rocks.  The slurry and rocks exiting the scrubber will be discharged through a rotating trommel screen attached to the end of the scrubber.  The coarse ore fraction consisting of manganese nodules and rocks will be discharged from the end of the trommel and will be conveyed, via a belt conveyer, to a pebble crusher.  The trommel under-size slurry of clay and fines, will drop directly into the cyclone feed sump.  From the cyclone feed sump, slurry will be pumped to a cyclone cluster for classification.  The cyclone overflow will be the ground ore product from the grinding circuit with 80% of the particles size passing 206 microns.  The cyclone underflow returns, by gravity, to the ball mill.

The pebble crusher discharge will be conveyed and fed to the ball mill for grinding along with the crushed ore along with cyclone underflow.  Ball mill discharge will be gravity fed to the cyclone feed sump.  In the ball mill, slurrying and grinding will occur in seawater using fifty mm carbon steel balls as grinding media.  Ball handling equipment is included to add make-up balls.

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The cyclone overflow slurry at 15% solids will be screened for removal of trash using  two trash screens with one operating and the other standby.  Screened slurry will then be fed to a high rate thickener for thickening the slurry to 35% solids concentration. Flocculent, from a flocculent makeup package and storage tank, will be added to promote thickening.  Some thickener underflow will be pumped to the leach circuit, the balance to the partial neutralization circuit.  The thickener overflow will be recycled to mill.  Slurrying and milling take place in sea water that has been heated by recovering heat from the slurry discharged from the reducing leach and supplementary steam heating when required.

Oxidative Leach - The ground ore from the mill thickener will be leached in four stages of hot atmospheric agitated tanks with sulfuric acid and recycled copper SX raffinate to dissolve the copper, some manganese, and zinc containing minerals.  A small portion of the cobalt will be leached at this stage as well.  The slurry will flow, by gravity, from tank to tank and then to the reductive leach tanks.  The tanks will be arranged to allow for bypassing.  The heat of solution of sulfuric acid plus heat exchangers will maintain this circuit at 80oC.  The tanks are covered and ventilated to the scrubber serving the Reductive Leach.

Reductive Leach - The slurry will be subjected to a hot atmospheric, agitated, reductive leach where SO2 (sulfur dioxide) gas will be sparged into the  slurry to facilitate the leaching of the remaining copper, cobalt and manganese minerals.  Again, slurry cascades by gravity through the tanks and will then be pumped to the heat recovery circuit.  The tanks will be arranged to allow for bypassing.

The tanks will be closed top and ventilated to a scrubber to capture any residual SO2.  The scrubber liquid will be a soda ash (sodium carbonate) solution.  Blowdown will be to the leach tanks.  The scrubber will be sized for upset condition of full SO2 flow during shutdown.

Partial Neutralization - The slurry leaving the reductive leaching circuit will contain residual acid which is partially neutralised to ensure optimal copper recovery in the downstream copper solvent extraction circuit.  Partial neutralization is achieved via the addition of milled ore by diverting a portion of the oxidative leach feed to neutralisation.  A thickener clarifier separates partially leached ore and pregnant solution termed pregnant leach solution or PLS.  The PLS is transferred to the copper SX circuit for copper extraction and concentration prior to electrowinning.

Slurry Cooling – The slurry discharged from the reductive leach process is cooled in multi stage evaporative coolers and heat exchangers to ensure optimal performance of copper solvent extraction and reduce leaching heating costs.  The system recovers heat from the leach discharge slurry and preheats the sea water and raffinate streams fed to the head of the milling

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and leaching circuits.  Slurry will be flashed under a slight vacuum and the resultant steam will be condensed. The resultant condensate will supplement the high purity process water produced by the desalination plants.

Counter-Current Decantation (CCD) Circuit - Cooled leach residue slurry discharged from the evaporative coolers will be washed in a six-stage CCD circuit to recover soluble copper, zinc, and cobalt.  Deep bed thickeners will be utilized.  Fresh seawater will be introduced into the mixing tank feeding the last thickener, along with slurry underflow from the preceding thickener.  Thickener underflow will be pumped from the last CCD stage to a tailings neutralization facility.  Thickener overflow from the first CCD stage will contain the bulk of the leached copper, zinc, and cobalt and is called pregnant leach solution (PLS).  The PLS will be transferred, by gravity, to the partial neutralization circuit.  The circuit will be arranged to allow bypassing.

Copper Solvent Extraction (SX) – The copper solvent extraction circuit consists of a 2-stage extract, 1-stage wash, 1-stage parallel extract, and 2-stage strip operation.  Copper is recovered from the PLS into an organic liquor known as an extractant.  The PLS aqueous solution, now stripped of copper and known as raffinate is split between the leaching circuit, (where it is used to dilute the milled ore fed to the leach), the iron removal process (where iron and aluminium, are removed prior to further concentration of the zinc and the cobalt in the DSX® operation) and the CCD circuit (where it is used as additional wash solution to improve CCD wash efficiency).  The wash stage removes potentially high levels of species such as iron and chlorides that are deleterious impurities to electrowinning.

The loaded organic stream from the copper extraction circuit is ‘stripped’ with spent copper electrolyte from the copper electrowinning operation, producing loaded copper electrolyte and stripped organic, the latter being recycled to copper SX.  Loaded copper electrolyte is pumped, via multimedia filters to remove carry-over organic and entrained solids, to copper electrowinning where copper is recovered as LME grade cathode.

Copper Electrowinning (EW) – Copper metal is electro-deposited from filtered loaded electrolyte, onto stainless steel blanks, known as cathodes, over a nominal 6-day cycle.  Cathodes are harvested via an automated stripping machine on a semi-continuous basis.

An overhead crane will lift out sets of cathodes from the EW cells.  The cathodes are then washed, and fed on a chain conveyor type system to the automated stripping machine.  Copper cathode is automatically 'stripped' from the stainless steel blanks, sampled, weighed and packaged for sale.  Spent copper electrolyte is eventually returned to the copper SX circuit.

The EW tankhouse will be enclosed in a building with cross-flow ventilation for acid mist control. Acid mist will be suppressed with a layer of small polypropylene balls, and a surfactant, as is industry practice.  Additional support equipment includes the water heaters for indirect electrolyte heating.

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Iron Removal – The copper raffinate from the copper SX circuit will contain residual copper, zinc, cobalt, iron, and manganese.  This step has been designed to remove residual acid and iron and aluminum from solution.  Air is sparged through the solution to convert the ferrous ion to ferric ion and, by increasing the pH using Boleo carbonate, encourages the formation of iron hydroxide precipitates.  The products of the iron removal process, primarily goethite, will be thickened and washed in a small CCD circuit in order to recover as much process liquor and dissolved metal values as possible before being diverted to tailings neutralization.

Thickener overflow will be pumped to a sand filter for final solids removal ahead of the DSX® circuit.  This solution is maintained at a pH of 4.5 in preparation for the DSX® extraction step.

Tailings Neutralization and Disposal – Thickened tailings from the main CCD circuit and thickened underflow from the iron removal CCD circuit will be treated in a neutralization operation where Boleo carbonate is added to neutralise excess acid and precipitate metal salts contained in the tailings solution.  The discharged slurry is expected to meet disposal criteria required by regulatory authorities.

The slurry discharged from this process is pumped to a tailings storage facility situated to the west of the plant site known as the Curuglú area.  Multiple stages of pumping will be used to generate the required pressure to overcome the static and dynamic losses in the pipe.  Initially 6 stages of pumps will be required, with a seventh to be added after 7 years as the dam height increases.

DSX®– Cobalt and zinc contained in solution from the iron removal clarifier overflow, are separated from manganese, magnesium, and calcium into an organic phase via the DSX® process.  The DSX® process employs two extractants known as LIX63 and Versatic 10 to effect this separation.  The DSX® solvent extraction circuit consists of a 2-stage extract, 2-stage scrub and 2-stage strip operation.  The organic phase is scrubbed with aqueous solution from zinc stripping to remove small quantities of impurity from the loaded organic.

The cobalt and the zinc in the loaded organic is then stripped and concentrated into an acidic aqueous stream for further treatment.  This stream is pumped to the zinc solvent extraction circuit after a cadmium purification step removal step.

Cadmium Removal – the DSX® strip solution is fed to a cadmium removal step where zinc dust is used to purify the solution of excess cadmium at a pH of 3.3 using cementation. This enables the final zinc product to meet the quality specifications required for use as an additive to animal feed.  The cadmium free solution is filtered in a plate and frame filter with the filter cake being disposed as hazardous waste and the filtrate being passed to zinc solvent extraction.

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Zinc Solvent Extraction – The purified bulk strip solution from DSX® becomes the feed to a zinc removal solvent extraction operation, prior to cobalt SX and cobalt EW.  The zinc solvent extraction circuit consists of a conventional 4-stage extract, 2-stage scrub, and 2-stage strip operation.  Strip liquor, rich in zinc, reports to two places in the flowsheet.  A small proportion is returned to the DSX® as scrub liquor and the remainder the feed to a facility producing zinc sulphate monohydrate crystals in a fluidized bed granulator.

Zinc Sulphate Monohydrate Production – The zinc strip solution is neutralized using lime and then pumped to a fluidised bed granulator where hot air is used to evaporate water and dry the product.  Particles are 'grown' to the appropriate size and shape, cooled, screened and packaged in a stand-alone production facility.  The hot air used in the process is produced through the combustion of diesel.

Cobalt Solvent Extraction, Electrowinning – Raffinate from the zinc secondary SX circuit, now containing only cobalt and very small quantities of zinc, nickel, and iron, will report to the cobalt SX operation.  The circuit consists of a 4-stage extract, 2-stage scrub and 3-stage strip operation.  The strip liquor from the cobalt SX operation is further purified through ion exchange and passes onto electrowinning to produce high grade cobalt metal.  Electrowon cobalt is crushed, screened and dispatched in drums as a premium metal product.

Sulphuric Acid Generation and SO2 Production – Sulphuric acid and sulphur dioxide are produced in a 'stand alone' facility, which consists of a number of unit operations including:

· prilled sulphur storage

· sulphur melting

· air drying

· sulphur burning

· gas conversion (employing catalysts)

· economizers and heat exchangers

· gas scrubbing and acid storage

· SO2 gas production.

The SO2 is sparged into the slurry in the reductive leach operation.  Sulphuric acid is dosed according to the numerous needs of the operation.  Sulfuric acid will be stored in three large tanks and distributed to various duties throughout the plant.  Excess sulfuric acid will be pumped to barges at the port facility for export.

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Power Production – Steam exported from the sulphuric acid and SO2 gas plants is used to generate electricity via the use of a steam-driven turbine and power generation system.  The power generated via this cogeneration system will produce about 70% of the overall plant requirements, making the plant and mining operation nearly self sufficient in electricity. The power shortfall will be made up by the use of heavy fuel oil (HFO) powered generating sets totaling of 20.44 MW capacity.  Power will be generated and distributed at 13.8 kV.

To maximize turbine output and recycle condensate back to the waste heat boilers, a seawater cooled condenser will be used. Seawater will be pumped from the dock area though the condenser and warm water will be discharged back into the sea.  Sodium carbonate (soda ash) scrubbers on the HFO generator exhaust gas will be used to control sulfur oxide gases, sulfuric acid mist, and particulates.

Power Supply and Distribution

The majority of electrical power will come from the acid plant as co-generation.  Even though acid consumption in the process will vary, the acid plant will be run at full capacity to provide power.  Excess acid produced will be stored for times of higher acid demand, or sold.  To supplement the power requirements, three HFO generators will be provided. Additional generators may need to be rented and hooked into the system if one generator is off-line for maintenance or rebuild.  The generators will be supplied with a black-start diesel powered generator to start the HFO generators and will then switch to HFO.  Site wide power requirements are shown in Table 16-17.

Table 16-17: Site Wide Power Requirements

Summary of Site-wide Power Requirement Power
Description	MW
Total Installed Power - Duty & Standby	105.3
Power Drawn During Normal Operation	52.5
Demand Load	65.8
Cogeneration Power Output	42.3
Power from grid (Demand)	1.5
Supplementary Power Required	23.5
Supplementary Power Installed (HFO Generators)	20.4
Combined Emergency Power Required	TBD
Dedicated Emergency Generators	0

The camp and administration building power will be supplied from the Commission Federal de Electricidad (CFE) grid from the 34.5 kV line which runs right across the front of the MMB property.  The excess capacity of this line is limited (up to 5MW) and will be used for the camp and administration building.  This system will also provide construction power.

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The emergency power for thickeners, pumps, boiler, and other uses have not been estimated.  In case of a total power trip, it will critical to be able to restart or reconnect at least one HFO generator as quickly as possible.

Power Distribution - Both steam turbine and HFO generators will generate power at 13.8 kV in parallel operation. Power to various areas of the plant site will be distributed at 13.8 kV level.  A main plant 13.8KV substation will be provided close to the cogeneration and HFO generation facilities.  Secondary unit substations will then step the voltage down to the utilization voltages of 4160 and 480 volts.

Power to the mine sites will be distributed, after stepping up, via four 34.5 kV overhead power lines to four main load centers, Boleo, Soledad, Purgatorio and Providencia. Power for the surface mine repair facility will be stepped down to the utilization voltages of 4160 and 480 Volts.  Power to mine portals will be stepped down to 12,470 and 480 Volts.

Generally, motors larger than 400 hp will be served at 4.16 kV and motors less than 400 hp will be served at 480 V.

Limestone Milling – A limestone milling circuit is required to produce ground limestone slurry for various bulk neutralization duties throughout the circuit namely Partial Neutralization, Tailings Neutralization and Iron Removal.  Limestone (or Boleo carbonate) is mined on site, transported using haul trucks, crushed in the field to approximately 19 mm, stockpiled and ball milled in sea water.  A set of cyclones is used to correctly size the particle size distribution of the slurry which is then circulated through the plant on a ring main system.

Process Water – Seawater will be the main source for process water throughout the circuit, but not all process water will be seawater.

Desalinated Water – Desalinated water, required for particular duties throughout the plant, will be produced in one reverse osmosis and one thermal plant using waste heat.  A small proportion of this water production will be diverted for use throughout the plant and mining operations as potable water.  A third smaller reverse osmosis plant has already been installed to provide water for construction and the camp.  A small chlorination plant will treat desalinated water to ensure a steady supply of potable water to both the plant and the mine operation.  Reject brine reports to the seawater discharge system.

Steam - The Acid Plant, while supplying project use and excess sulfuric acid, will supply high pressure (HP) steam at 62 barg and medium pressure (MP) steam at 10 barg. The sulfur dioxide plant, while supplying project SO2, will also supply HP steam.  High pressure steam will be used to generate power; steam extraction from the turbo generator will be low pressure steam at 5 barg.

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The HFO power plant will produce MP steam with waste heat from the generator combustion gasses.

A MP diesel fired boiler  will be employed for make up steam and start up conditions.

Tailings Dam – Tailings from the processing plant will be contained in a zero discharge facility located in the Curuglú Arroyo by a series of rock-fill dams.  There will be two significant dams, identified as the Main Dam and the North West Saddle Dam, as well as a series of small levies along the ridge above the Soledad Arroyo (the East Saddle Dam). The locations of these features are shown in Figure 16-14.

Figure 16-14: Location of Tailings Storage Facilities

Construction of the Main Dam will be accomplished in phases.  A starter dam will be completed by a contractor prior to plant startup.  This starter dam will be constructed to the 173-m elevation, with the downstream arroyo being filled to the 150-m elevation. Constructing this downstream fill will facilitate downstream access to the dam.  The dam will be raised in seven phases as shown in Figure 16-15.  The material for the first 3 years of construction will be overburden, stockpiled from the mining of the manto reserves within the footprint of the TSF.

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Figure 16-15:              Tailings Storage Facility Wall Design

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Dock - The first marine structure will be a construction dock consisting of a rock-fill jetty with sheet piling at the end to receive oversized loads that cannot be trucked on Highway 1.  Barges will come from Guaymas with oversize loads.  The permanent dock will then be constructed by adding a similar jetty to the south and connecting the two at the seaward side with additional fill forming a settling basin.  This endwall will have large diameter pipes in the bottom to allow water from 3 m depth or more to flow into the settling basin formed by the rock fill.

After receiving the oversized loads, a pier consisting of a steel trestle sitting on steel piles will extend 150 m into the sea to accept heavy fuel oil and lightering barges.  The lightering barges will unload sulfur prills and soda ash into a hopper from where the prills and soda ash will be conveyed to a storage facility in the plant area.  Excess sulfuric acid will also be shipped, by barge, to market.

Sulfur Prills - Solid sulfur will be shipped to site in the form of prills and will be unloaded from Handymax sized self unloading ships onto lightering barges.  Rented tug boats will move the lightering barges from the ship to the barge pier.  The self unloading barges will feed a hopper over the unloading conveyor.  The unloading conveyor will convey, via two transfer conveyors, onto the sulfur stockpile stacking conveyor which in turn will discharge onto the 50,000-ton stockpile.  The stacking conveyor will be a radial stacker and produce a kidney shaped pile.  The conveyor transfer points will have foggers and dry dust collection dependent on the product being conveyed.  Desalinated water will be used for dust control and to keep the pile wet to prevent fires.  The unloading conveyor and transfer conveyors will be common for sulfur prills and soda ash.  Sulfur prills will be reclaimed from the stockpile using front end loaders, and loaded into a hopper feeding the sulfur melter feed conveyor.

Soda Ash - Soda ash will be unloaded from Handymax ships in 5,000 tonne lots onto the lightering barges that will then be positioned by tug at the dock.  The self unloading barges will transfer soda ash to the same hopper feeding the same unloading conveyor used for sulfur prills.  A wash down system will be provided to clean the belts periodically. In addition, when over water, the conveyors will be in a gallery.  The conveying system will convey at 1,000 tons per hour through two transfer conveyors onto the silo feed conveyor. The silo feed conveyor will discharge into a 10,000 tonne storage dome.  Soda ash will be reclaimed out of the storage dome by a front end loader to a bucket elevator that discharges into the soda ash solution preparation facility. Soda ash will be dissolved in water in an agitated tank before being pumped to day tanks in the plant.

Diesel - Bulk imports of fuel from tanker trucks will be held in a single 2,000 m³ tank situated at the plant site within a synthetically lined berm with an appropriate spill capacity.  Diesel will be used in a number of operations including the mining fleet and other mobile equipment, the zinc

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sulfate plant (as a heat source), the low pressure steam boiler, and the acid plant start-up pre-heater.

Heavy Fuel Oil (HFO) - The HFO for the HFO power plant will be unloaded from the barge by unloading pumps to one of two 1,500 m3 HFO storage tanks.  Two tanks will be provided to allow clean up of one tank to remove any solids.  Filters will be also provided to remove unsettled solids as the HFO is pumped to the generators.  The storage tanks will be inside a lined bermed area.  The fuel oil unloading line will be steam traced up to the storage tank.  HFO needs to be kept at an elevated temperature to be pumped and the storage tanks are equipped with low pressure steam heating coils.

Seawater - Seawater will be drawn from an intake structure located at the west end of the settling basin formed by the dock rock filled levees.  The intake structure houses screens and a chlorinator.  Large debris will settle at the bottom of the basin in the calm waters. The seawater then passes through two screens to prevent trash, debris, or marine life from entering the vertical centrifugal seawater pumps.  Hypochlorite will be added to reduce algae growth throughout the process plant.

Seawater will be drawn from a depth of 3 to 4 m at a temperature varying from 20°C to 30°C to supply the various needs of the plant.  The seawater will be pumped primarily to the condenser on the cogeneration plant at approximately 10,000 m3/hr, which will be 76% of the seawater demand of the plant.  In order to minimize power for seawater pumping, the pumps will only be designed to pump to the top of the condenser when the condenser is under almost a full vacuum.  To provide the vacuum, the discharge out the top of the condenser will flow to barometric leg and a seawater return sump located north of the sewage lagoons.  A vacuum pump will be required to remove non condensable gasses from the return line.

Sea water will also be used in coolers/heat exchangers, feed to the desalination plant, and for the water in the process.  Seawater booster pumps will be provided for the pressure to be pumped to the various areas.

During construction, the existing seawater pump will pump to a lined pond on the Soledad Plateau for soil compaction water. A series of ponds and pumps provide water for road and dam construction. The lines and pumps will be left in place for further dam lifts and road watering for dust control

Seawater that will be used for cooling water and for desalination plant brine will flow via a separate line to the seawater return sump.  The seawater return sump will then flow by gravity to be discharged back into the sea near the mouth of the Boleo Arroyo.  An allowance is included for a diffuser on the seawater discharge to minimize the hot spots and promote mixing.

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Fire Protection – Fire protection will be provided through a combination of a high-pressure water main and hydrant system, automatic sprinkler systems, with or without foam (as appropriate), and fire detection and alarm systems.  The fire alarm systems will consist of smoke or heat detectors, horns, bells, strobe lights, and manual pull stations.  A high-pressure firewater ring main will be provided throughout the plant and associated buildings.  Fire hydrants and hoses will be provided.  Buried water mains will be HDPE; above ground piping will be carbon steel.  Hose cabinets, internal standpipes, and fire extinguishers will be strategically located throughout the plant site to provide suitable coverage.  Fire water will be desalinated water with backup being seawater.

The firewater will be desalinated water and will be stored in the lower dedicated portion of the storage tank located on the hill south of the plant site.  As required pressure will be 7 to 8 bar, an electric firewater pump, backed up by a diesel driven pump will be used to supplement the gravity system.  As a backup, the seawater supply system will be piped to the firewater pump system.

Storm Water - All surface run off associated with the plant area will be directed throughout the plant area to two collection ponds, one situated within the main hydrometallurgical plant area to the south of Soledad Creek, and the other serving the CCD area, north of Soledad Creek.

The ponds are intended to collect rainfall run-off or spills at the plant to prevent contamination of the local watercourses. The ponds are fed by a series of runoff diversion channels, spoon drains, and berms.

The quality of the water in the catchment ponds will be monitored and will either be used as process water (if not contaminated) or diverted to tailings neutralization (if contaminated) as required. Diversion bunds around the perimeter of the facility will prevent clean rainwater, which has not passed through the plant site, from entering the pond.

Sewage - A package sewage treatment plant will be provided, suitably located away from the populated plant areas and across the normal wind flow direction, to limit any odor carryover. The plant will be sized to cater for up to 3,000 persons.

The largest mine population will be to be found in the area of the administration offices. Wastewater from all human sources in the administration and ablutions buildings will be collected in a small underground concrete tank adjacent to the ablutions building. It will be pumped automatically at frequent intervals to the sewage plant for treatment by biodegradation.

After treatment, water quality levels will have a pH of 6.5 to 8.5, with 30 mg/L of suspended solids, 20 mg/L of biochemical oxygen demand, and total nitrogen of 15 mg/L.

The water, once treated, will be used for dust suppression, gardening etc. Sludge solids will be removed and stored in lined ponds until suitable for fertilizer

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Other Waste Products - A sanitary land fill will be located south of the plant site off of the Purgatorio Road.  An aggressive recycling program will be implemented.  Waste not recyclable will be trucked to the land fill for burial.  The proposed site is above the flood and runoff areas.

Flammable wastes will be impounded and sent to an appropriate facility.

Fuel-derivative wastes will be trucked away from the site by the fuel company which will supply the site oils and lubricants.

Certain hazardous process waste products, generated on an irregular basis will also require handling by a specialist waste disposal company or service.

Communications- Baja California telephone systems are fiber optic, eliminating the need for satellite communications.  Communications facilities at the project site will include a 100-line PABX telephone system, a mobile radio system, a telemetry control system and a mine vehicle dispatch system.

The telephone system will service the various buildings in the plant site area including administration buildings, warehouses, various control rooms within the plant (e.g., grinding, leaching, SX, EW), maintenance workshops, laboratory, etc. Phone stations may be provided in such geographically distant areas such as the CCD area, the tailings neutralization area, and the ROM pad. The system will feature the various “intelligent” telephone options that are currently on offer such as voice mail, call forwarding etc.

A number of national and international connections will be required for administrative purposes.

Mobile radio communication system will consist of base stations and a mixture of hand-held portable radios and vehicle mounted radios. The radio will be required to support up to 100 users.

Buildings -Buildings specified for the project include:

· Security and Access Control

· Administration

· Laboratory

· Medical Clinic

· Workshops

· Warehousing

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· Change Houses

· Construction (then subsequently the Operations) Camp

· Reagent Storage Shed

· MCC Buildings, Control Rooms, and Substations.

16.2.3 Process Design and Operating Assumptions

The design and expected operational performance of the Boleo process plant has been defined and recorded as follows:

·   Metallurgical and engineering reports.

·   Design criteria, schedules, datasheets, drawings and specifications of equipment.

·   An electronic model of the process has been established using Metsim®, a proprietary process simulation software package that produces a mass balance of all process streams.

·   The mass balance is used as input to an Excel® based operating cost model.  This model is time based and uses the mine plan to simulate and predict the ramp-up and phased implementation of production as well as all the economic factors of operating cost.

·   The following assumptions have been made regarding metal recovery.

·   Demonstration Pilot Plant recoveries for Cu, Co and Zn have been used as the basis of estimation.

·   Cu recovery is assumed to be constant over the mine plan.

·   Co and Zn recoveries are lower during ramp-up.

·   Zn recoveries reduces after year 12 of the Mine Plan due to a constraint in extraction capacity when processing ore with high Zn grade.

·   No Mn recovery is assumed for this phase of estimation.

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17 MINERAL RESOURCE & MINERAL RESERVE ESTIMATES

17.1 Summary

Wardrop was commissioned to update the Boleo Mineral Resource Estimate by taking a slightly different approach to that chosen in the Feasibility Study and all resource figures quoted in this section are from their 2009 report2.  For the Feasibility Study a 3D block model and a 2D “Gridded Seam” model were generated for use in open pit and underground mine planning, respectively.  For the update, it was decided to generate the 2D model from the 3D block model by compositing vertically adjacent blocks according to set mining and economic parameters.  This was done to allow greater flexibility in selecting the local mining thickness within the limits of the anticipated mining method.  The process consisted of:

· generating 3D block models of the seven mantos of interest.

· flattening each manto model and corresponding drillhole dataset to a common basal elevation for geostatistical analysis and grade interpolation.

· compositing vertically adjacent blocks to generate 2D “Pseudo-seam” models

· translating both models back to real space for use in mine design and scheduling.

17.2 3 D Geological Interpretation

The geology of the Boleo deposit is well known (see Sections 6.0, 7.0, and 9.0) and has been extensively studied.  Mineralization is effectively confined to seven shallowly dipping layers, referred to as “mantos.”  A large number of faults have been identified from either surface mapping or mapping of historical workings.  These faults play an important role with respect to local continuity of both geology and grade.

Faults were modelled as steeply dipping planes based on historic mapping and verified by MMB geologists.  The fault wireframes were supplied to Wardrop as 3D DXF files.

Baja supplied three-dimensional surface models of the footwall and hanging wall contacts for each of the seven principal manto deposits (mantos 0, 1, 2, 3aa, 3a, 3, and 4) as 3D DXF files.  The surfaces were updated versions of those generated during the Feasibility Study.  They honoured the geological logging of the upper and lower manto contacts in the diamond drill holes, regardless of grade, and took into consideration the many north-south trending faults

  2 Stubens,. T.C., Wardrup Engineering Inc (2009) “Mineral Resource Update on the El Boleo Property, Baja California Sur, Mexico.” November 2009.

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which offset the mantos.  Wardrop checked to ensure that the surfaces honoured the drilling and the faults and that the relative locations of the mantos were maintained.

A detailed description of the generation of manto and Fault wireframes can be found in Section 2.21 of the Boleo Feasibility Study.

Figure 17-1: Final Triangulation of Manto 3 Surface with Post-Ore Faults (Oblique View)

The same process was followed for all seven manto footwall surfaces.

17.3 Resource Estimation Technique

17.3.1 Block Model vs. Gridded Seam Model

The Boleo copper-cobalt-zinc-manganese resource will be mined by a combination of open pit and underground methods.  Consequently, different types of block model must be constructed that reflect the different requirements of open pit and underground mine planning.

During the Feasibility Study, two separate models were built: a 3D block model for open pit mine design and a 2D “Gridded Seam” Model for underground mine design.  It is acknowledged that 2D models are appropriate for certain types of deposits, like coal seams and precious metal vein deposits, which have clearly defined geological limits and where the geological limits will also be the mining limits.  However, in situations where one or both of the mining limits will be determined based on grade or where there is a recognized variability or trend in metal grades across the deposit, a 2D approach can be overly restrictive and inflexible.  A 2D approach

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requires creating a single composite from each drill intersection through the deposit and the subsequent treatment of deposit thickness as a variable in the resource estimation process.  Implicit in the method is the assumption that the grades are homogeneous over the deposit thickness when a single composite is made.  This is clearly not the case in the Boleo manto deposits, as discussed below.

The manto deposits at El Boleo vary greatly in thickness and display well recognized trends in grade between their upper and lower contacts.  Generally-speaking the mantos show high Cu grades adjacent to their basal contacts which tend to decrease upwards.  The opposite trend is observed with zinc grade.  So using the basal manto contact as the lower mining limit is quite logical but the determination of the upper mining limit and 2D composite boundary will be based on economics and the limits of the planned mining method.  It was therefore, decided to generate a single 3D block model for each manto and to build the 2D models from the 3D models.  This would preserve the variability of grades within the seam and allow better optimization of mining height in each manto deposit.

17.3.2 Data Domains and Flat Models

The definition of data domains is required to limit data used for grade estimation to the data within the area being considered.  The most obvious control that needs to be applied at El Boleo is to keep separate the data and estimation process for each manto.  Data for each manto was loaded into different data files and individual models were created and the estimation process was carried out separately for each manto.

The second controlling influence on the data that needs to be considered is the effect of the faulting.  Faulting is predominantly post mineralization and as such, has resulted in fragmentation of once continuous manto units into numerous blocks, each bounded by faults.  In plan, the continuity of the mantos remains.  In section, continuity before a fault termination is significantly reduced to distances as short as 100 m.

In Figure 17-2, a search ellipse is shown in plan and in section.  In plan, the ellipse centred over a highlighted block appears to capture three drill holes that fall within it (yellow points).  However, when viewed in section, the block, which is adjacent to but on the up-throw side of a fault, has its search ellipse well above the data on the down-throw side and consequently no data will be captured from that side of the fault.

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Figure 17-2: Search Ellipses – Plan and Section Views

Plan View:

Section View:

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Because the faulting at El Boleo is predominantly post mineralization, the two segments of manto now offset by faulting would originally have been juxtaposed.  This being the case, the inclusion of the holes on the down-throw side of the fault would be justified.  It is not practical to capture data significantly offset by faults simply by means of modifying the search ellipse parameters (i.e., increasing the vertical search distance) as this would require very large vertical searches that may also result in too much data from higher levels in the manto being captured as well.

To achieve the desired effect the base of each manto intersection in each drill hole was set at the same (fixed) elevation.  The base of the block model was also set at the same elevation.  This step removes the effect of the faulting and returns drill hole intersections to their pre-faulting position, relative to adjacent holes.  It is then possible to consider all the data for each manto as a single mineralized domain.

The effect of this is to create a resource model that has the same EW and NS lateral extents but exists in a reduced or ‘flat’ vertical space (Figure 17-3).  Aligning each hole at the same level not only removes the faulting offsets, but also eliminates the easterly dip of the mantos.  This has an additional advantage in that each block model has a significantly smaller vertical extent as it only has to be slightly thicker than the maximum manto thickness.  These models are referred to as ‘flat’ models.

Figure 17-3: Alignment of Data in ‘Flat’ Space

All grade modelling steps for 3D block models were carried out using the flat models.  Partial Blocking was used to capture the entire manto within the block model (Figure 17-4).

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Figure 17-4: Grade Estimation Block Models

All blocks retained both flat and true mid-point elevations (Z_Flat and Z_Org) so returning the blocks to their original positions in real space is accomplished by having the block elevation equal to Z_Org.

17.4 Block Models

Datamine Studio 3 (Version 3.16.2299.1) was used for all 3D modelling and block model estimation.

The 3D Manto Block Models were built using blocks measuring 50 m by 50 m by 0.9 m in the X, Y and Z dimensions respectively.   The individual manto deposits are very thin relative to their great lateral extents, so the small block height was used to attempt to capture the grade variability in the vertical direction.  Furthermore, 1.8 m is planned to be the minimum mining height and it is evenly divisible by 0.9.  The maximum mining height is 4.2 m.

Table 17-1: Block Model Parameters

	UTM_X	UTM_Y	Z
Minimum	364,000	3,019,000	-217.8
Maximum	376,500	3,029,200	288.0
Model Size	12,500	10,200	505.8
Block Size	50	50	0.9
Blocks	250	204	562

The manto block models were built in “real space” and then flattened using the footwall elevations shown in Table 17-2.

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Table 17-2:                      Manto Names, Numbers, and Flattened Footwall Elevations

Manto		Flat-Z
Name	Number	FW Elev.
m0	11	225
m1	10	189
m2	20	153
m3aa	32	135
m3a	31	117
m3	30	90
m4	40	9

17.4.1 Accounting for Past Mining Activities

Mantos 1 and 3 have been extensively mined in the past.  The locations of the old mines are well known (Wilson 1955); however, due to the method of mining adopted the precise location and extent of voids, pillars and back-filled excavations are not known.

Ore was mined using a short wall mining method; however, only high-grade ore (> 3.5% Cu) was taken to the surface for processing.  It is estimated that the ore processed amounted to only 40% of mined material, the remaining 60%, referred to as “Retaque,” was side-cast (backfilled) into old excavations as mining progressed.  Exploration drill holes have intersected voids in about 67 instances and a similar number of Retaque intervals have been identified.  The estimation process has been carried out effectively ignoring this issue other than treating voids as ‘missing’ or ‘no data’.

The resources were however modified to account for the previous mining activities.  This could be achieved by simply removing an amount of ore equal to the known treated ore tonnage, 13.6 Mt.  However, this may be overly conservative since many of the workings have collapsed and the void volume is now lower than was originally the case.

Over time, the lower grade clayey breccias, that typically overlie the higher-grade laminated manto material that was mined, have converged or sagged into the original voids and areas of back-fill.  This has compressed the backfill and reduced or eliminated the voids associated with the previously mined areas.  The higher-grade ore mined and processed has effectively been ‘replaced’ by lower grade material, firstly due to backfilling as mining progressed and later by the collapse of the workings where voids were left.  This replaced material has been sampled and can be considered in situ.  Voids still remain though and these have been accounted for by:

· digitizing a simplified footprint shape around the area of historic workings

· extending the footprint area up and down to form a 3D object that enveloped the workings and the entire thickness of the manto

· flagging all model blocks within this ‘mined envelope’ as ‘mined’

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· determining the proportion of sample intervals within the area affected by past mining that intersected voids and factoring the tonnes by the proportion of sample intervals that were voids.

For Manto 3 a total of 443 m of drill hole intersection occur within the mined volume shape of which 12% are voids (Figure 17-5).  It is appropriate therefore, to factor the tonnes down by 12%.  To accomplish this, the SG was reduced by 12% from 1.41 to 1.24 t/m3 for the previously mined area.

Figure 17-5: Retaque (Stope Backfill) on Left Side of Photo, Texcoco Mine

Note: over time, hanging wall has compressed Retaque, leaving no void space

Drill hole information showed a smaller percentage of voids encountered in Manto 1 than Manto 3, but it was decided to reduce Manto 1 tonnage by the same 12% factor for the historically mined area.

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Figure 17-6: Manto 3 Old Mine Areas – Location Drill Holes which Encountered Voids (Solid) and Retaque (Open)

17.5 Data

17.5.1 Updated Database

Ninety diamond drill holes totalling 5,062.75 m were completed after the cut-off date for the Feasibility Study and added to the database for this resource update.  Thirty-seven surface trenches from 2006 totalling 106.4 m were also added.  The current database contains 1,364 drill holes and 37 trenches totaling 127,879 m and 106.4 m, respectively.  Sample intervals were flagged with a code indicating the manto each one represented.  In total 12,533 samples

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totalling 11,882.9 m of core from 1,284 drill holes and 36 trenches were used in grade estimation and modelling.

17.5.2 UTM Transfer

All drill hole collars and 3D models of manto surfaces, faults, mined out areas and topographic surface were converted from a local grid system (“Wilson”) to UTM coordinates (Zone 12 NAD83 Datum).

17.5.3 Secondary Elements

During the Feasibility Study, it was determined that a better understanding of the concentration of various elements would aid in the operation of the processing plant and in the estimation of plant operating costs.  To that end, analytical data for the following elements was added to the geological database constructed for the Mineral Resource Update:  Calcium (Ca%), Iron (Fe%), Aluminum (Al%), Magnesium (Mg%), Strontium (Sr ppm), Uranium (U ppm), Vanadium (V ppm), total Sulphur (S%) and sulphide Sulphur (S_fide %).  Of these elements, only the iron data are reasonably complete.  Analysis of the samples for the other elements has only recently become standard practice, so approximately 2/3 of the intervals do not have data for these elements.  Therefore, grade estimates for these elements will be dependent upon the sample locations and have lower confidences.

17.5.4 Compositing

An average 0.9 m composite length was selected to better fit the block height and because the average sample length of all the manto samples is 0.95 m.  To insure that all manto samples were included in the composites, remnants less than 0.45 m in length were added to the previous composite. Thus the final composite in a given manto intersection was between 0.45 m and 1.35 m in length.

17.5.5 Flattening

All spatial analyses and resource grade estimates were performed in “flattened space”.  The drill data and 3D block models of each manto were flattened to a nominal basal elevation (Table 17-4) to undo the effects of the post-ore faulting and allow the estimation of blocks across the faults (see Section 17.3 for details).

17.5.6 Density

For the purposes of resource calculations, a global in situ dry bulk density of 1.41 t/m3 has been used for all mantos.  It is based on an average calculated from 993 wet density measurements and the corresponding analysed water contents of the samples (Section 12.3).  A density of 1.24 t/m3 was applied to previously mined areas of mantos 1 and 3.

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17.6           Exploratory Data Analysis

The summary statistics of the composited assay data for each Manto are shown below, in Table 17-3.

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Table 17-3:           Summary Statistics of Composited Assay Data

M0	Length	Cu (%)	Co (%)	Zn (%)	Mn (%)	Ca (%)	Fe (%)	Al (%)	Mg (%)	Sr (ppm)	U (ppm)	V (ppm)	S (%)	S_fate (%)	S_fide (%)
Mean	0.91	0.02	0.01	0.48	1.07	7.89	3.73	4.85	4.83	442.22	0	149.90	0.51	0.06	0.44
Max	1.35	0.45	0.08	3.32	6.63	19.90	11.34	8.29	8.65	2360	0	409	3.18	0.86	3.04
Q3	0.95	0.02	0.01	0.69	1.31	10.62	4.15	5.71	5.74	530	0	190	0.85	0.07	0.73
Median	0.90	0.01	0.01	0.39	0.81	7.01	3.57	5.03	4.98	420	0	126	0.13	0	0.10
Q1	0.85	0.01	0.00	0.17	0.49	4.32	3.04	3.98	3.97	325	0	90	0.03	0	0.03
Min	0.57	0.002	0.001	0.01	0.02	0.68	1.73	1.88	1.19	108	0	47	0	0	0
IQR	0.10	0.01	0.01	0.52	0.82	6.30	1.11	1.73	1.76	205	0	100	0.82	0.07	0.69
CV	0.09	1.79	0.89	0.86	0.89	0.57	0.28	0.26	0.28	0.50	0	0.52	1.26	2.05	1.32
Number	484	484	484	484	484	250	481	250	250	250	26	250	248	248	248
ns*		0	0	0	0	234	3	234	234	234	458	234	236	236	236
M1	Length	Cu (%)	Co (%)	Zn (%)	Mn (%)	Ca (%)	Fe (%)	Al (%)	Mg (%)	Sr (ppm)	U (ppm)	V (ppm)	S (%)	S_fate (%)	S_fide (%)
Mean	0.90	0.60	0.04	0.81	2.59	3.74	4.50	4.95	3.77	536.92	5.52	170.01	0.85	0.14	0.72
Max	1.31	14.95	0.48	20	21.8	23.80	23.24	9.27	10.95	3982	50	1050	7.55	2.05	7.33
Q3	0.94	0.52	0.05	0.86	3.42	4.56	5.19	6.31	4.66	670	0	190	1.51	0.16	1.20
Median	0.90	0.05	0.02	0.49	1.87	2.64	4.14	5.29	3.48	490	0	129	0.10	0	0.08
Q1	0.85	0.01	0.01	0.25	0.86	1.94	3.37	3.60	2.73	288	0	84	0.03	0	0.03
Min	0.48	0	0	0	0	0.39	0.81	0.52	0.76	50	0	14	0	0	0
IQR	0.09	0.51	0.04	0.60	2.56	2.61	1.82	2.72	1.93	382	0	106	1.48	0.16	1.17
CV	0.09	2.20	1.29	1.78	0.99	0.86	0.44	0.38	0.41	0.71	0	0.86	1.47	2.09	1.55
Number	836	836	836	836	833	453	808	453	453	453	45	453	379	379	379
ns		0	0	0	3	383	28	383	383	383	791	383	457	457	457
table continues…

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M2	Length	Cu (%)	Co (%)	Zn (%)	Mn (%)	Ca (%)	Fe (%)	Al (%)	Mg (%)	Sr (ppm)	U (ppm)	V (ppm)	S (%)	S_fate (%)	S_fide (%)
Mean	0.90	0.30	0.04	0.89	3.71	3.24	5.72	4.08	3.48	627.99	1.50	199.57	0.48	0.06	0.42
Max	1.31	12.91	2.84	9.28	22.54	16.50	19.94	11.50	8.51	5000	81	2580	7.89	1.06	7.51
Q3	0.95	0.27	0.06	1.03	5.33	3.38	6.94	5.52	4.33	750	0	250	0.51	0.07	0.42
Median	0.90	0.09	0.03	0.58	2.87	2.12	5.29	4.38	3.22	535	0	157	0.08	0	0.07
Q1	0.85	0.02	0.01	0.30	1.15	1.63	4.06	2.41	2.44	350	0	96	0.03	0	0.03
Min	0.46	0	0	0	0	0.39	1.45	0.16	0.31	44	0	20	0	0	0
IQR	0.1	0.25	0.05	0.73	4.18	1.75	2.88	3.11	1.89	400	0	154	0.48	0.07	0.39
CV	0.09	2.48	1.76	1.21	0.89	0.91	0.42	0.47	0.40	0.89	0	0.89	1.91	2.15	2.01
Number	2,132	2,132	2,132	2,132	2,118	863	1,991	857	857	857	202	857	645	645	645
ns		0	0	0	14	1,269	141	1,275	1,275	1,275	1,930	1,275	1,487	1,487	1,487
M3AA	Length	Cu (%)	Co (%)	Zn (%)	Mn (%)	Ca (%)	Fe (%)	Al (%)	Mg (%)	Sr (ppm)	U (ppm)	V (ppm)	S (%)	S_fate (%)	S_fide (%)
Mean	0.91	0.51	0.07	0.70	4.94	1.82	6.31	2.76	2.87	724.18	7.11	145.76	0.62	0.16	0.46
Max	1.35	5.20	0.28	5.62	18.02	4.11	19.50	6.43	5.79	2750	35.5	556	2.75	0.82	2.75
Q3	0.95	0.70	0.11	0.89	7.60	2.14	8.34	3.88	3.48	1005	0	176	0.87	0	0.87
Median	0.91	0.19	0.05	0.58	3.88	1.73	4.86	1.99	2.76	599	0	96	0.59	0	0.17
Q1	0.83	0.08	0.03	0.28	1.39	1.27	3.32	1.42	2.08	300	0	65	0.17	0	0.08
Min	0.71	0	0	0.05	0.09	0.69	2.23	0.75	1.32	124	0	42	0.08	0	0
IQR	0.12	0.61	0.08	0.61	6.21	0.87	5.02	2.46	1.40	705	0	111	0.70	0	0.79
CV	0.12	1.50	0.84	0.96	0.84	0.39	0.61	0.61	0.36	0.74	0	0.82	0.99	1.85	1.46
Number	162	162	162	162	158	54	142	54	54	54	5	54	21	21	21
ns		0	0	0	4	108	20	108	108	108	157	108	141	141	141
table continues…

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M3A	Length	Cu (%)	Co (%)	Zn (%)	Mn (%)	Ca (%)	Fe (%)	Al (%)	Mg (%)	Sr (ppm)	U (ppm)	V (ppm)	S (%)	S_fate (%)	S_fide (%)
Mean	0.90	0.35	0.06	0.56	3.84	2.40	8.37	3.39	2.72	1339.24	4.54	122.22	0.35	0.07	0.28
Max	1.35	7.68	0.63	5.91	21.71	12.91	33.32	9.08	8.48	82545	107.4	1220	7.49	2.43	7.49
Q3	0.94	0.32	0.08	0.72	5.67	2.69	9.93	5.11	3.39	780	0	143	0.31	0.02	0.23
Median	0.89	0.08	0.04	0.43	2.73	1.90	7.26	2.99	2.50	450	0	99	0.09	0	0.07
Q1	0.84	0.02	0.02	0.26	0.90	1.37	5.47	1.62	1.82	245	0	60	0.03	0.00	0.03
Min	0.47	0.00	0.00	0.00	0.00	0.43	1.04	0.71	0.82	76	0	0	0	0	0
IQR	0.10	0.30	0.06	0.47	4.77	1.32	4.46	3.49	1.57	535	0	83	0.28	0.02	0.20
CV	0.12	2.02	1.07	0.81	0.95	0.75	0.52	0.58	0.43	4.07	0	0.84	2.01	3.84	2.25
Number	2,115	2,115	2,115	2,115	2,070	604	1,724	598	598	598	133	598	297	297	297
ns		0	0	0	45	1,511	391	1,517	1,517	1,517	1,982	1,517	1,818	1,818	1,818
M3	Length	Cu (%)	Co (%)	Zn (%)	Mn (%)	Ca (%)	Fe (%)	Al (%)	Mg (%)	Sr (ppm)	U (ppm)	V (ppm)	S (%)	S_fate (%)	S_fide (%)
Mean	0.90	0.86	0.07	0.36	2.12	2.14	7.29	3.65	2.39	2319.49	13.64	123.12	0.43	0.11	0.32
Max	1.35	16.20	1.47	14.03	34.00	18.30	55.20	10.15	7.93	99444	240	1088	7.35	5.68	7.35
Q3	0.93	1.16	0.09	0.42	2.82	2.39	8.25	5.54	2.86	896	0	157	0.57	0.10	0.34
Median	0.90	0.34	0.05	0.27	1.51	1.46	6.20	2.65	2.18	480	0	90	0.16	0	0.08
Q1	0.87	0.08	0.02	0.18	0.60	0.93	4.63	1.89	1.70	280	0	52	0.04	0	0.03
Min	0.47	0	0	0	0	0.28	0.92	0.27	0.23	0	0	0	0	0	0
IQR	0.06	1.08	0.06	0.24	2.22	1.46	3.62	3.65	1.16	616	0	106	0.53	0.10	0.31
CV	0.07	1.49	1.02	1.35	1.11	0.98	0.63	0.57	0.43	3.73	0	0.87	1.64	2.94	1.93
Number	5,296	5,296	5,296	5,296	5,111	1,957	4,165	1,936	1,936	1,936	299	1,936	850	850	850
ns		0	0	0	185	3,339	1,131	3,360	3,360	3,360	4,997	3,360	4,446	4,446	4,446
table continues…

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M4	Length	Cu (%)	Co (%)	Zn (%)	Mn (%)	Ca (%)	Fe (%)	Al (%)	Mg (%)	Sr (ppm)	U (ppm)	V (ppm)	S (%)	S_fate (%)	S_fide (%)
Mean	0.90	0.31	0.03	0.23	1.66	5.06	7.47	4.30	3.20	584.42	4.13	136.45	0.35	0.05	0.30
Max	1.30	11.05	0.46	7.32	30.97	31	41.01	8.96	10.72	5170	70	1480	6	3	6
Q3	0.91	0.32	0.04	0.28	1.61	7.33	8.60	5.69	4.03	540	0	151	0.40	0.03	0.22
Median	0.90	0.12	0.02	0.17	0.59	3.11	5.91	4.57	2.56	400	0	110	0.10	0	0.07
Q1	0.88	0.04	0.01	0.11	0.25	1.60	4.33	2.88	1.99	278	0	80	0.04	0	0.03
Min	0.56	0	0	0	0	0.40	1.22	0.17	0.27	82	0	10	0	0	0
IQR	0.03	0.28	0.02	0.17	1.36	5.72	4.27	2.82	2.04	262	0	71	0.36	0.03	0.19
CV	0.04	2.07	0.97	1.18	1.82	0.94	0.70	0.43	0.58	1.36	0	0.86	1.78	3.60	2.01
Number	2,160	2,160	2,160	2,160	2,137	831	1,568	660	660	660	132	660	579	579	579
ns		0	0	0	23	1,329	592	1,500	1,500	1,500	2,028	1,500	1,581	1,581	1,581
All	Length	Cu (%)	Co (%)	Zn (%)	Mn (%)	Ca (%)	Fe (%)	Al (%)	Mg (%)	Sr (ppm)	U (ppm)	V (ppm)	S (%)	S_fate (%)	S_fide (%)
Mean	0.90	0.55	0.05	0.50	2.60	3.27	6.82	3.96	3.00	1374.41	6.90	144.53	0.48	0.09	0.39
Max	1.35	16.20	2.84	20	34.00	31	55.2	11.5	10.95	99444	240	2580	7.89	5.68	7.51
Q3	0.93	0.57	0.07	0.56	3.51	3.68	7.95	5.62	3.76	730	0	180	0.58	0.07	0.43
Median	0.90	0.14	0.03	0.31	1.59	2.03	5.78	4.11	2.61	468	0	110	0.11	0	0.08
Q1	0.86	0.03	0.02	0.18	0.57	1.28	4.23	2.02	1.94	290	0	68	0.04	0	0.03
Min	0.46	0	0	0	0	0.28	0.81	0.16	0.23	0	0	0	0	0	0
IQR	0.07	0.54	0.05	0.39	2.94	2.40	3.72	3.60	1.82	440	0	112	0.54	0.07	0.40
CV	0.08	1.91	1.23	1.46	1.16	1.04	0.62	0.51	0.49	4.29	0	0.89	1.75	2.90	1.92
Number	13,185	13,185	13,185	13,185	12,911	5,012	10,879	4,808	4,808	4,808	842	4,808	3,019	3,019	3,019
ns		0	0	0	274	8,173	2,306	8,377	8,377	8,377	12,343	8,377	10,166	10,166	10,166

*ns=not sampled

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17.7 Variography

Semi-variograms were calculated for Cu, Co, Zn and Mn using the flattened composite data.  Semi-variogram maps in the plane of each manto were generated to help determine the major and minor directions of continuity.  Then the semi-variograms in those directions were modelled.  Down-hole semi-variograms were also calculated to determine the continuity of metal grades in the vertical direction.  The nugget effect from the down-hole semi-variogram was also used for the horizontal models.  The Boleo semi-variograms have been plotted and the parameters are summarized in Table 17-4 through to Table 17-7.  The clock-wise rotation of the X and Y variogram axes about the Z axis is shown.

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Table17-4: Variogram Parameters (Manto 0 and 1)

Manto 0						Manto 1
Cu	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.					sph.	sph.	sph.
variance	0.150	0.493			0.643	0.057	0.650	0.203	0.197	1.107
range-X	-	395				-	171	375	2748
range-Y	-	395				-	196	424	1484
range-Z	-	4.3				-	2.0	3.5	7.0
Rot.(cw)	0					35
Co	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.				sph.	sph.
variance	0.000	0.187	0.134		0.321	0.039	0.483	0.209		0.731
range-X	-	95	350			-	139	2609
range-Y	-	95	350			-	257	3170
range-Z	-	2.0	4.0			-	2.5	10.0
Rot.(cw)	0					35
Zn	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.				sph.	sph.
variance	0.070	0.112	0.355		0.537	0.069	0.409	0.099		0.576
range-X	-	43	250			-	189	3510
range-Y	-	43	250			-	199	2192
range-Z	-	2.0	4.0			-	3.8	10.5
Rot.(cw)	0					35
Mn	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.				sph.	sph.
variance	0.058	0.196	0.182		0.436	0.077	0.480	0.059		0.616
range-X	-	230	415			-	237	3007
range-Y	-	230	415			-	232	2446
range-Z	-	2.0	4.0			-	2.9	10.5
Rot.(cw)	0					35

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Table 17-5: Variogram Parameters (Manto 2 and 3aa)

Manto 2						Manto 3aa
Cu	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.	sph.			sph.	sph.	sph.
variance	0.061	0.734	0.090		0.885	0.172	0.443	0.261		0.876
range-X	-	40	1000			-	157	484
range-Y	-	45	620			-	105	377
range-Z	-	3.0	6.0			-	2.2	4.0
Rot.(cw)	0					30
Co	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.				sph.	sph.
variance	0.042	0.350	0.150		0.542	0.006	0.354	0.132		0.492
range-X	-	95	1130			-	109	372
range-Y	-	50	790			-	118	297
range-Z	-	3.0	5.0			-	2.9	7.6
Rot.(cw)	0					30
Zn	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.				sph.	sph.
variance	0.075	0.239	0.067	0.164	0.545	0.001	0.300	0.190		0.492
range-X	-	155	495	1520		-	238	632
range-Y	-	90	415	690		-	254	462
range-Z	-	1.5	6.2	7.0		-	2.2	3.0
Rot.(cw)	15					30
Mn	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.				sph.	sph.
variance	0.049	0.308	0.088	0.147	0.592	0.001	0.436	0.240		0.677
range-X	-	120	330	1200		-	115	323
range-Y	-	130	235	400		-	206	307
range-Z	-	1.5	6.2	7.0		-	2.7	5.0
Rot.(cw)	0					30

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Table 17-6: Variogram Parameters (Manto 3a and 3)

Manto 3a						Manto 3
Cu	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.	sph.			sph.	sph.	sph.
variance	0.173	0.497	0.158	0.172	1.001	0.090	0.518	0.201	0.099	0.908
range-X	-	88	636	2277		-	50	160	935
range-Y	-	60	866	1110		-	55	130	965
range-Z	-	1.9	5.0	9.8		-	3.0	5.8	7.5
Rot.(cw)	0					15
Co	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.				sph.	sph.	sph.
variance	0.030	0.294	0.123	0.113	0.560	0.050	0.279	0.071	0.080	0.480
range-X	-	50	111	2371		-	25	315	1060
range-Y	-	50	139	1143		-	35	185	565
range-Z	-	2.6	3.0	7.6		-	4.0	5.0	6.8
Rot.(cw)	0					15
Zn	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.				sph.	sph.	sph.
variance	0.030	0.193	0.123	0.045	0.392	0.020	0.107	0.081	0.083	0.291
range-X	-	50	158	1133		-	135	260	810
range-Y	-	50	168	603		-	40	115	535
range-Z	-	3.0	3.0	6.0		-	1.0	2.0	5.0
Rot.(cw)	0					15
Mn	c0	c1	c2	c3	Sill	c0	c1	c2	c3	Sill
type		sph.	sph.				sph.	sph.	sph.
variance	0.090	0.542	0.140	0.071	0.842	0.090	0.516	0.041	0.078	0.724
range-X	-	65	407	2647		-	65	555	940
range-Y	-	70	430	838		-	70	220	440
range-Z	-	3.0	4.5	5.5		-	5.0	6.0	7.5
Rot.(cw)	0					15

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Table 17-7:             Variogram Parameters (Manto 4)

Manto 4
Cu	c0	c1	c2	c3	Sill
type		sph.	sph.	sph.
variance	0.045	0.601	0.137		0.783
range-X	-	163	2072
range-Y	-	262	1078
range-Z	-	3.0	12.6
Rot.(cw)	0
Co	c0	c1	c2	c3	Sill
type		sph.	sph.
variance	0.014	0.236	0.146		0.397
range-X	-	205	407
range-Y	-	163	674
range-Z	-	3.0	12.4
Rot.(cw)	0
Zn	c0	c1	c2	c3	Sill
type		sph.	sph.
variance	0.010	0.173	0.208		0.392
range-X	-	182	440
range-Y	-	355	557
range-Z	-	2.2	10.6
Rot.(cw)	0
Mn	c0	c1	c2	c3	Sill
type		sph.	sph.
variance	0.014	0.501	0.215		0.730
range-X	-	191	1054
range-Y	-	332	454
range-Z	-	2.4	5.1
Rot.(cw)	0

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17.8 Grade Interpolation and Resource Classification

The grades of the four principle elements (Cu, Co, Zn and Mn) were estimated using ordinary kriging and the semi-variogram parameters shown in Table 17-4 to Table 17-7 (above).  The grades of the secondary elements were estimated using the Inverse Distance Squared (ID2) method.

A three step search strategy was employed to estimate the block grades and simultaneously assign a resource class number to each block.  The first pass search radius was the semi-variogram range at 57.5% of the sill and was used to estimate Measured Resource blocks.  The second pass search radius was the semi-variogram range at 80% of the sill.  Blocks estimated in the second pass and not in the first pass, were classified as Indicated.  The third pass search radius was used to estimate Inferred blocks and was approximately 2.5 times the second pass search radius.  Limits were placed on the number of composites from a single drillhole that could be used to estimate the grade of an individual block.  This was done to force the contractor to use data from more than one hole thus maintaining grade variability in the vertical direction.  The details of the search parameters used to estimate the block grades are shown in Table 17-8.

Table 17-8: Estimate Block Grades Search Parameters for Copper

Manto	Rotation (Clockwise)	Search Radii			Composites			Comments
		X	Y	Z	Min	Max	Max/ddh
Measured Resources
0	0	169	169	2.2	9	20	4	Estimated in Flattened Space
1	35	176	216	2.9	9	20	4	Search = SV Range at 57.5% C
2	0	216	137	2.9	9	20	4	Minimum 3 holes
3	15	189	118	2.9	9	20	4
3a	0	173	216	2.9	9	20	4
3aa	30	208	171	2.2	7	15	3
4	0	151	216	4.3	13	25	6
Indicated Resources
0	0	235	235	3.0	5	20	4	Search = SV Range at 80% C
1	35	245	300	4.0	5	20	4	Minimum 2 holes
2	0	300	190	4.0	5	20	4
3	15	263	164	4.0	5	20	4
3a	0	240	300	4.0	5	20	4
3aa	30	290	238	3.0	4	15	3
4	0	210	300	6.0	7	25	6
Inferred Resources
0	0	625	625	8.0	3	20	4
1	35	650	800	10.6	3	20	4
2	0	650	410	8.6	3	20	4
3	15	450	275	6.9	3	20	4
3a	0	640	800	10.6	3	20	4
3aa	30	770	630	8.0	3	15	3
4	0	450	650	12.9	3	25	6

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17.9 Resource Tabulation

17.9.1 Copper Equivalent Calculation

The recovered copper equivalent (CuEq) is calculated for each cell to be used for cut-off grade purposes, only.  This allows the cut-off to capture the contribution of each metal to the value of the block.  It is standard procedure in poly-metallic deposits.  The formula used to calculate CuEq (%) is below and the metal price and recovery assumptions used in its calculation are summarized in Table 17-9.  To provide consistency with the Feasibility Study, the same metal prices were used.

Table 17-9 Copper Equivalent Assumptions

Metal	Price ($/lb)	Recovery
Cu	1.50	91.2%
Co	15.00	78.5%
Zn	1.20	65.5%

Where:

CuEq = (Cu% x Cu-Pr. x Cu-Rec. + Co% x Co-Pr. xCo-Rec. + Zn% x Zn-Pr. x Zn-Rec.) / Cu-Pr.

· Cu% = Copper Grade, Cu-Pr = Copper Price, Cu-Rec = Copper Recovery

· Co% = Cobalt Grade, Co-Pr = Cobalt Price, Co-Rec = Cobalt Recovery

· Zn% = Zinc Grade, Zn-Pr= Zinc Price, Zn-Rec = Zinc Recovery

The Boleo Block Mineral Resource Estimate is summarized in Table 17-10 to Table 17-12 using a series of CuEq (%) cut-off grades.  A break-even CuEq cut-off grade of 0.5% will be used for reporting purposes.

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Table 17-10:               Block Model Mineral Resource Estimate at 0.5% CuEq Cut-off

Manto		Tonnes (106)	Cu (%)	Co (%)	Zn (%)	Mn (%)	CuEq (%)
0	Measured	0.210	0.02	0.013	0.87	1.36	0.57
	Indicated	2.844	0.04	0.011	0.99	2.12	0.64
	M+I	3.055	0.04	0.011	0.99	2.07	0.64
	Inferred	0.606	0.03	0.008	0.97	2.41	0.60
1	Measured	5.677	1.17	0.054	0.77	2.69	1.90
	Indicated	18.448	0.96	0.057	1.14	3.38	1.93
	M+I	24.124	1.01	0.057	1.05	3.22	1.92
	Inferred	25.343	0.71	0.060	0.97	3.42	1.63
2	Measured	7.122	0.46	0.058	1.00	4.26	1.40
	Indicated	44.626	0.42	0.053	1.05	4.60	1.35
	M+I	51.748	0.43	0.054	1.04	4.56	1.36
	Inferred	39.067	0.30	0.048	0.95	4.45	1.14
3aa	Measured	1.692	0.59	0.088	0.92	6.06	1.71
	Indicated	1.243	0.64	0.085	0.81	5.29	1.67
	M+I	2.936	0.61	0.087	0.87	5.73	1.69
	Inferred	0.196	0.57	0.066	0.50	5.66	1.29
3a	Measured	19.774	0.42	0.080	0.65	4.62	1.35
	Indicated	24.274	0.42	0.051	0.63	4.15	1.12
	M+I	44.048	0.42	0.064	0.64	4.36	1.22
	Inferred	4.676	0.35	0.055	0.58	4.87	1.06
3	Measured	43.768	1.02	0.081	0.32	2.06	1.73
	Indicated	62.680	1.14	0.060	0.51	2.50	1.77
	M+I	106.448	1.09	0.068	0.43	2.32	1.76
	Inferred	35.094	0.64	0.046	0.90	2.54	1.41
4	Measured	7.574	0.80	0.050	0.37	3.01	1.32
	Indicated	24.740	0.53	0.038	0.30	2.27	0.94
	M+I	32.314	0.59	0.041	0.32	2.44	1.03
	Inferred	54.870	0.38	0.034	0.29	1.71	0.77
TOTAL	Measured	85.818	0.82	0.074	0.50	3.04	1.59
	Indicated	178.855	0.74	0.053	0.71	3.32	1.46
	M+I	264.673	0.76	0.060	0.64	3.23	1.50
	Inferred	159.852	0.47	0.045	0.70	2.93	1.15

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Table 17-11:                   Block Model Mineral Resource Estimate at 0.75% CuEq Cut-off

Manto		Tonnes (106)	Cu (%)	Co (%)	Zn (%)	Mn (%)	CuEq (%)
0	Measured	0.003	0.04	0.013	1.19	2.19	0.77
	Indicated	0.426	0.04	0.010	1.35	2.83	0.83
	M+I	0.429	0.04	0.010	1.35	2.82	0.83
	Inferred	0.032	0.04	0.009	1.29	2.55	0.78
1	Measured	5.060	1.30	0.059	0.78	2.73	2.05
	Indicated	15.825	1.11	0.064	1.20	3.40	2.14
	M+I	20.885	1.16	0.063	1.10	3.24	2.12
	Inferred	22.147	0.80	0.066	0.99	3.58	1.77
2	Measured	5.963	0.52	0.064	1.08	4.59	1.54
	Indicated	38.324	0.47	0.058	1.13	4.90	1.47
	M+I	44.287	0.47	0.059	1.12	4.86	1.48
	Inferred	32.221	0.32	0.053	1.05	4.86	1.26
3aa	Measured	1.585	0.62	0.091	0.96	6.30	1.78
	Indicated	1.177	0.66	0.087	0.84	5.42	1.72
	M+I	2.762	0.64	0.089	0.91	5.92	1.76
	Inferred	0.159	0.65	0.072	0.57	6.26	1.45
3a	Measured	16.144	0.49	0.089	0.69	4.99	1.51
	Indicated	18.335	0.51	0.057	0.68	4.50	1.27
	M+I	34.479	0.50	0.072	0.68	4.73	1.38
	Inferred	3.476	0.44	0.062	0.60	5.47	1.20
3	Measured	39.722	1.11	0.085	0.32	2.12	1.84
	Indicated	56.217	1.24	0.062	0.53	2.55	1.90
	M+I	95.939	1.19	0.072	0.45	2.37	1.88
	Inferred	30.195	0.71	0.049	0.97	2.54	1.54
4	Measured	4.838	1.09	0.061	0.46	3.81	1.72
	Indicated	13.577	0.72	0.046	0.37	2.63	1.21
	M+I	18.414	0.82	0.050	0.39	2.94	1.35
	Inferred	21.253	0.62	0.037	0.37	1.96	1.05
TOTAL	Measured	73.314	0.93	0.081	0.52	3.20	1.75
	Indicated	143.881	0.87	0.059	0.77	3.55	1.66
	M+I	217.196	0.89	0.067	0.69	3.43	1.69
	Inferred	109.484	0.59	0.052	0.87	3.42	1.39

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Table 17-12:                 Block Model Mineral Resource Estimate at 1.0% CuEq Cut-off

Manto		Tonnes (106)	Cu (%)	Co (%)	Zn (%)	Mn (%)	CuEq (%)
0	Measured	0	0	0	0	0	0
	Indicated	0.008	0.03	0.010	1.81	3.45	1.06
	M+I	0.008	0.03	0.010	1.81	3.45	1.06
	Inferred	0	0	0	0	0	0
1	Measured	4.378	1.46	0.063	0.78	2.74	2.24
	Indicated	13.835	1.24	0.070	1.24	3.41	2.33
	M+I	18.213	1.29	0.068	1.13	3.25	2.30
	Inferred	20.563	0.85	0.069	0.99	3.64	1.84
2	Measured	4.569	0.63	0.072	1.17	4.94	1.75
	Indicated	29.346	0.54	0.064	1.25	5.15	1.65
	M+I	33.915	0.55	0.065	1.24	5.12	1.66
	Inferred	21.463	0.39	0.059	1.19	5.31	1.44
3aa	Measured	1.419	0.67	0.096	1.00	6.64	1.89
	Indicated	1.039	0.71	0.091	0.90	5.67	1.83
	M+I	2.459	0.69	0.094	0.96	6.23	1.87
	Inferred	0.133	0.71	0.075	0.63	6.76	1.57
3a	Measured	11.841	0.61	0.104	0.71	5.44	1.74
	Indicated	11.627	0.66	0.068	0.70	4.92	1.50
	M+I	23.467	0.64	0.086	0.70	5.18	1.62
	Inferred	2.174	0.54	0.072	0.66	5.92	1.40
3	Measured	34.259	1.23	0.090	0.33	2.17	2.00
	Indicated	48.797	1.38	0.065	0.55	2.59	2.06
	M+I	83.056	1.32	0.075	0.46	2.41	2.03
	Inferred	23.566	0.83	0.051	1.08	2.57	1.72
4	Measured	3.365	1.40	0.070	0.51	4.13	2.09
	Indicated	7.893	0.90	0.052	0.44	3.02	1.46
	M+I	11.258	1.05	0.058	0.46	3.35	1.65
	Inferred	9.624	0.80	0.043	0.41	2.11	1.27
TOTAL	Measured	59.831	1.07	0.088	0.53	3.29	1.95
	Indicated	112.545	1.03	0.065	0.83	3.65	1.88
	M+I	172.376	1.04	0.073	0.72	3.53	1.91
	Inferred	77.522	0.70	0.058	0.99	3.66	1.61

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17.10 Pseudo-seam Models

In a final post-estimation process, each manto model was composited to produce a 2D “Pseudo-seam” model to be used by mining engineers for room and pillar (stope) design, scheduling and for the calculation of a Mineral Reserve estimate.  The anticipated minimum and maximum mining heights for El Boleo underground mining (1.8 m and 4.2 m respectively) were used to composite the blocks along with a Cu cut-off grade of 0.5%.

This process is performed in an Excel spreadsheet where vertically contiguous blocks are composited from the footwall upwards.  The blocks are automatically composited to reach the minimum mining width or the manto hanging wall contact, regardless of grade.  If the hanging wall contact is encountered before the minimum mining width is reached, the composite is extended upwards by adding sufficient “dilution” of zero grade to reach the minimum mining width.  Overlying blocks are added, assuming they satisfy the cut-off grade requirement, until either: a “sub-ore grade” block is encountered, the manto hanging wall contact is encountered, or the maximum mining width is reached.

In Figure several possible situations are shown that could occur during the block compositing operation.  Blue blocks have a grade that is below the cut-off used for compositing and the red block grades are greater than or equal to the cut-off.  The resultant Pseudo-seam model heights are shown below each column of blocks.

Figure 17-7: Block Compositing to Generate Pseudo-seam

The Pseudo-seam Model Manto Mineral Resource estimates are reported in Tables 17.13, 17.14 and 17.15 using a series of CuEq cut-off grades.

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Table 17-13:                    Pseudo-seam Model Resource Estimate at 0.5% CuEq Cut-off

Manto		Tonnes (106)	Thick (m)	Cu (%)	Co (%)	Zn (%)	Mn (%)	CuEq (%)
0	Measured	0.082	1.80	0.02	0.014	0.79	1.04	0.54
	Indicated	1.891	1.80	0.04	0.012	1.02	2.08	0.66
	M+I	1.973	1.80	0.04	0.012	1.01	2.04	0.66
	Inferred	0.355	1.80	0.03	0.008	0.99	2.41	0.61
1	Measured	4.589	2.77	1.45	0.058	0.73	2.54	2.17
	Indicated	15.140	2.34	1.13	0.064	1.07	3.32	2.09
	M+I	19.729	2.44	1.21	0.063	0.99	3.14	2.11
	Inferred	20.607	2.36	0.83	0.068	0.85	3.25	1.74
2	Measured	4.531	2.39	0.64	0.067	0.99	4.19	1.63
	Indicated	31.509	2.16	0.50	0.059	1.05	4.64	1.47
	M+I	36.041	2.19	0.52	0.060	1.04	4.58	1.49
	Inferred	25.063	2.02	0.34	0.053	0.96	4.58	1.23
3aa	Measured	1.592	2.00	0.66	0.093	0.97	6.21	1.83
	Indicated	1.022	1.87	0.62	0.080	0.73	4.98	1.57
	M+I	2.613	1.95	0.64	0.088	0.88	5.73	1.73
	Inferred	0.163	1.94	0.60	0.067	0.53	5.93	1.36
3a	Measured	15.547	2.05	0.50	0.084	0.67	4.87	1.47
	Indicated	20.841	1.96	0.44	0.049	0.62	4.10	1.11
	M+I	36.388	2.00	0.47	0.064	0.64	4.43	1.26
	Inferred	3.849	1.99	0.36	0.053	0.56	4.75	1.04
3	Measured	31.617	3.42	1.34	0.090	0.33	2.12	2.09
	Indicated	48.834	2.92	1.34	0.060	0.54	2.55	1.97
	M+I	80.451	3.12	1.34	0.072	0.46	2.38	2.02
	Inferred	24.670	2.58	0.77	0.048	0.88	2.52	1.55
4	Measured	2.708	3.41	1.47	0.065	0.46	3.70	2.10
	Indicated	10.131	2.91	0.64	0.039	0.30	2.17	1.05
	M+I	12.839	3.01	0.81	0.045	0.34	2.49	1.27
	Inferred	20.639	2.77	0.51	0.031	0.30	1.66	0.87
TOTAL	Measured	60.666	2.90	1.07	0.083	0.52	3.19	1.89
	Indicated	129.369	2.49	0.89	0.056	0.73	3.38	1.63
	M+I	190.035	2.62	0.94	0.065	0.66	3.32	1.71
	Inferred	95.348	2.40	0.60	0.050	0.76	3.13	1.33

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Table 17-14:                      Pseudo-seam Model Resource Estimate at 0.75% CuEq Cut-off

Manto		Tonnes (106)	Thick (m)	Cu (%)	Co (%)	Zn (%)	Mn (%)	CuEq (%)
0	Measured	0.000	0.00	0.00	0.000	0.00	0.00	0.00
	Indicated	0.463	1.80	0.04	0.010	1.32	2.80	0.81
	M+I	0.463	1.80	0.04	0.010	1.32	2.80	0.81
	Inferred	0.019	1.80	0.03	0.009	1.29	2.63	0.77
1	Measured	4.417	2.81	1.51	0.060	0.73	2.55	2.22
	Indicated	13.544	2.40	1.25	0.070	1.10	3.32	2.27
	M+I	17.961	2.50	1.31	0.067	1.01	3.13	2.26
	Inferred	18.418	2.43	0.92	0.074	0.86	3.37	1.87
2	Measured	4.007	2.47	0.70	0.073	1.04	4.48	1.76
	Indicated	28.318	2.20	0.54	0.062	1.10	4.88	1.56
	M+I	32.325	2.23	0.56	0.064	1.09	4.83	1.58
	Inferred	22.297	2.05	0.37	0.057	1.01	4.85	1.31
3aa	Measured	1.547	2.01	0.67	0.094	0.99	6.33	1.87
	Indicated	0.996	1.87	0.63	0.081	0.74	5.04	1.60
	M+I	2.544	1.96	0.65	0.089	0.89	5.83	1.76
	Inferred	0.144	1.96	0.65	0.070	0.58	6.27	1.44
3a	Measured	13.757	2.08	0.55	0.091	0.69	5.07	1.57
	Indicated	15.512	2.01	0.54	0.056	0.66	4.44	1.28
	M+I	29.269	2.04	0.55	0.073	0.67	4.73	1.42
	Inferred	2.720	2.07	0.46	0.062	0.59	5.49	1.22
3	Measured	31.380	3.43	1.34	0.090	0.33	2.13	2.10
	Indicated	47.197	2.96	1.37	0.062	0.54	2.56	2.02
	M+I	78.576	3.15	1.36	0.073	0.46	2.39	2.05
	Inferred	22.699	2.64	0.82	0.050	0.92	2.49	1.62
4	Measured	2.328	3.67	1.66	0.071	0.50	3.98	2.34
	Indicated	6.744	3.40	0.81	0.045	0.35	2.46	1.27
	M+I	9.071	3.47	1.03	0.051	0.39	2.85	1.54
	Inferred	11.555	3.45	0.70	0.033	0.36	1.80	1.08
TOTAL	Measured	57.437	2.96	1.12	0.086	0.52	3.22	1.96
	Indicated	112.772	2.58	0.99	0.061	0.76	3.51	1.78
	M+I	170.209	2.71	1.03	0.069	0.68	3.41	1.84
	Inferred	77.852	2.52	0.68	0.055	0.83	3.38	1.50

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Table 17-15:                        Pseudo-seam Model Resource Estimate at 1.0% CuEq Cut-off

Manto		Tonnes (106)	Thick (m)	Cu (%)	Co (%)	Zn (%)	Mn (%)	CuEq (%)
0	Measured	0	0	0	0	0	0	0
	Indicated	0	0	0	0	0	0	0
	M+I	0	0	0	0	0	0	0
	Inferred	0	0	0	0	0	0	0
1	Measured	4.230	2.84	1.56	0.062	0.73	2.57	2.28
	Indicated	12.385	2.46	1.35	0.074	1.12	3.32	2.40
	M+I	16.615	2.55	1.40	0.071	1.02	3.13	2.37
	Inferred	17.518	2.45	0.96	0.077	0.85	3.39	1.92
2	Measured	3.464	2.57	0.78	0.079	1.08	4.77	1.90
	Indicated	23.619	2.28	0.60	0.067	1.17	5.08	1.69
	M+I	27.083	2.31	0.63	0.069	1.16	5.04	1.72
	Inferred	16.142	2.14	0.43	0.062	1.12	5.27	1.47
3aa	Measured	1.446	2.02	0.70	0.097	1.02	6.57	1.94
	Indicated	0.844	1.89	0.69	0.085	0.83	5.34	1.73
	M+I	2.290	1.97	0.70	0.093	0.95	6.12	1.86
	Inferred	0.138	1.96	0.66	0.071	0.59	6.41	1.47
3a	Measured	10.510	2.17	0.66	0.104	0.71	5.44	1.79
	Indicated	9.988	2.12	0.70	0.066	0.67	4.78	1.50
	M+I	20.498	2.14	0.68	0.085	0.69	5.12	1.65
	Inferred	1.762	2.22	0.56	0.071	0.64	5.87	1.41
3	Measured	30.669	3.46	1.36	0.091	0.33	2.14	2.13
	Indicated	44.193	3.01	1.44	0.063	0.56	2.60	2.10
	M+I	74.862	3.19	1.41	0.075	0.46	2.41	2.11
	Inferred	19.476	2.72	0.91	0.052	0.98	2.54	1.75
4	Measured	1.985	3.90	1.88	0.076	0.54	4.15	2.59
	Indicated	4.634	3.65	0.94	0.049	0.40	2.77	1.45
	M+I	6.619	3.72	1.22	0.058	0.44	3.19	1.79
	Inferred	5.846	3.64	0.85	0.039	0.41	1.88	1.29
TOTAL	Measured	52.303	3.07	1.20	0.090	0.51	3.21	2.07
	Indicated	95.664	2.68	1.11	0.065	0.79	3.56	1.94
	M+I	147.967	2.82	1.14	0.074	0.69	3.44	1.99
	Inferred	60.882	2.56	0.78	0.061	0.91	3.55	1.67

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17.11           Mineral Reserves

Underground and open-cut mine designs, completed by Agapito using a 0.5% CuEq cut-off and the Wardrop Block and Pseudoseam models, cover the current 23 year mine plan of 70.1Mt grading 1.38% Cu, 0.07% Co and 0.58% Zn.  A minor quantity of Inferred resource is also included within the 70.1 Mt planned as this material occurs in areas through which principle accesses must proceed.

The above represents a minor change from the 2007 DFS where mine designs covered a 25 year mine plan totalling 74.4 Mt at 1.47% Cu, 0.08% Co and 0.57% Zn.  A minor quantity of Inferred resource was also included within the 74.4 Mt plan.

Proven and Probable reserves as published in the 2007 DFS are sufficient to cover the current underground and open-cut mine plans.  Underground Proven and Probable reserves are noted separately from open-cut reserves as there are slight differences in the models used for each.

Underground mining will take place in Mantos 1, 2, 3 and 4 in areas where drilling has shown grade continuity and thicknesses to be sufficiently established to classify these areas as Proven and Probable Reserves (17.16).

Table 17-16: Underground Proven and Probable Reserves

Underground Reserves		Cu%	Co%	Zn%	Mn%
Class	tonnes x106
Proven	29.8	1.69	0.08	0.46	2.38
Probable	37.6	1.34	0.07	0.69	3.36
Total	67.4	1.49	0.07	0.59	2.93

Open-cut mining will take place in Mantos 2, 3a, 3aa, and 3 in areas where drilling has shown grade continuity and thicknesses to be sufficiently established to classify these areas as Proven and Probable Reserves (17.17).

Table 17-17: Open Cut Proven and Probable Reserves

Open Cut Reserves		Cu%	Co%	Zn%	Mn%
Class	tonnes x106
Proven	11.1	0.75	0.10	0.37	2.74
Probable	6.5	0.70	0.07	0.48	3.08
Total	17.6	0.73	0.09	0.41	2.87

Proven and Probable reserves shown are contained within the Measured and Indicated resources defined at the 0.5% CuEq Cutoff and are not additional to them

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18 OTHER RELEVANT DATA & INFORMATION

18.1 Mining

The mining plan supporting the economics of El Boleo has been projected for an initial 23 years of plant operation.  The plan consists of a series of underground and surface mines scheduled in such a manner as to provide the highest grades of copper ore for the first 12 years of operation and then the highest grades of copper equivalent ore for the remaining years of the plan.

The nominal production rates for all mining operations is 3.1 Mdt/a.  At full production, the mining workforce totals an estimated 400 salaried and hourly employees.  Skilled and experienced mentors, used in the start-up years for training and skill development needs, are essentially phased out by Project Year 3.

The surface mining plan incorporates all the surface mining and surface infrastructure activities required for executing both the underground mine plan (Section 18.2), the surface production schedule, the calcium carbonate quarry, tailings dam lifts, haul roads, utilities, and ROM ore haulage.  The surface plans were prepared primarily by Agapito Associates, Inc (AAI), with contributions from Baja Mining in regards to designs and operating plans and certain costs and by Wardrop Engineering (Closure plans).

The underground plan was prepared by AAI and schedules mining activities for the initial 23 years of plant operation.  The underground plan is based on detailed geotechnical analysis of:

· the ground conditions observed and measured in the test mine

· the geomechanics testing of core samples

· the experience of AAI and Baja Mining staff in room and pillar mining operations and the predicted behaviour of the ground.

18.1.1 Surface Mining & Mine Design

Surface mining activities at the Boleo Project will consist of:

· mining of manto ore from a series of shallow pits/contour outcrop cuts throughout the mine life as scheduled by estimated delivered grade.

· a quarry to supply calcium carbonate in the form of limestone for use in the ore processing plant.  Although these activities are included in the surface mining operations, the cost of mining the calcium carbonate is included in plant costs.

· a quarry to supply gypsum rock for road surfacing material.

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· The mining fleet and personnel will also be used to assist other areas of the operation including:

· excavation and haulage of rock fill for use in the starter tailings dam wall construction and to construct all the subsequent lifts of the dam and other TSF structures.

· haulage of ore from underground portal and surface mining sites and general earthworks such as road maintenance, road extensions and stockpile facilities.

Although nearly all of the ore for the first 23 years will be sourced from the underground mines, the surface mine fleet is an essential part of the project because of the ongoing infrastructure construction needs and it allows calcium carbonate, which is major input cost for ore processing, to be sourced at less than one-quarter the cost of purchasing limestone from the nearest supplier.

The open cut copper pits provide 5% of the processed amount of ore to the plant and support the mining of underground ore reserves.  The shallow pits and the quarry mining fleet also provide an easily accessible alternative ore source if the underground mine suffers any disruption during the project years.

Open Cut Mining

Open cut operations at Boleo will consist of ore mining from the shallow sections of Manto 3 and overlying mantos in the northern half of the deposit and quarrying of calcium carbonate from the Dos de Abril area in the north western corner of the project lease and later at the Santa Marta area just west of the plant ROM stockpiling area.

The open cuts will be phased throughout the project years to compliment the declining underground mining operations, allowing full production feed rate to the plant to be maintained.  The current mine plan shows a total of 5% of total ore production will be from surface open cuts.

Open Cut Mining Analyses

All of the open cut targets mine to Manto 3 as the base unit.  Some pits include areas of Mantos 3A and 3AA above Manto 3.  Currently, no significant amount of Manto 1, 2 or 4 ore is included in the open cut targets.  Open cut modeling was performed using the commercial software, TechBASE® and is readily used for this type of mining and civil earthworks design.

Resource Model

Since open cut operations allow selective mining based on close spaced grade control sampling in the cuts, a block model was used as opposed to the single layer seam models used for underground mining.  Block grades were estimated as whole block values using ordinary kriging with a block size of:

50 m north south

50 m east west

0.9 m vertically.

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This allows for selecting a particular mining block vertically as well as laterally.

Mining Loss & Dilution

Grade control drilling is allowed for on a 20 m x 20 m pattern with samples every 0.5 m vertically through the mantos.  The open cut mine plan in this report uses whole block grade estimates.  The increased definition provided by the grade control should allow exclusion of lower grade portions of the large resource blocks.  For this reason, it is assumed that the whole block estimates represent a diluted resource and no further dilution adjustments are made.

Pit Wall Slopes

A series of geotechnical core holes were drilled during the mid 1990s.  Geotechnical analysis of the data at that time suggested an average pit wall slope of 45°.  This slope was adopted for current open cut mine planning.  Geotechnical analysis undertaken as part of the feasibility study is considered adequate for initial level of confidence slope design purposes but an on-going process of data collection and slope monitoring during operations will be required to allow pit slopes to be optimized.

Mining Method

The open cut ore pits will use the same fleet as the calcium carbonate quarry but the mining method places more emphasis on the bulldozers as primary waste mining units.  All of the pits start at outcrop and in many cases; this outcrop line is in an arroyo wall so that there is open space below the pit start position.  Almost all of the overburden consists of weak sandstones, siltstones and clay and the mantos themselves are free digging.  Overburden is 0 m to 40 m thick and the economic manto thickness is mostly 1 m to 4 m.  The floor of Manto 3 in the open cut areas is hard conglomerate.

The overburden stripping strategy is illustrated in Figure 18-1.

The general mining sequence for mining in the calcium carbonate quarry (and ore) is:

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· Drill and blast all overburden.  It could be ripped but light blasting at an average powder factor of 0.33 kg of explosive per cubic metre will greatly improve productivity of the bulldozers and excavators.  No drill or blast is needed for ore mining.

· In the first block of each pit, the overburden is moved to an out of pit waste dump adjacent to the pit.  Where the elevation of the pit start point allows, this initial block overburden can be moved by dozer push.  In flatter areas it will be mined by a 95 tonne hydraulic excavator loading 85 tonne trucks.

· In subsequent blocks the overburden will be mostly dozer pushed over the mined out area in the previous block.  In some cases, the pit geometry or manto dip will require some portion of the overburden to be mined by the excavators.  Overburden for each block was assigned dozer and excavator percentages based on an assessment of the block geometry in cross-section.

Figure 18-1: Overburden Stripping Strategy

· In all blocks overburden removal with the dozer will cease a little above the expected roof position of the economic manto or target CaCO3 zone as defined from the previous block and the resource model.  The manto will be grade control drilled from this bench to determine the ore horizon and average grades for the block.

· The excavator will mine the last waste and the trucks will place the waste on the dozer push waste in the previous block.

· The excavators will mine the manto ore and the haul trucks will haul it to the nearest intermediate stockpile or directly to the plant ROM.  In the case of the pits in the east end of Soledad Arroyo, the trucks will haul directly to the process plant stockpiles.

· Light blasting will be used as needed in the manto overburden to promote productivity of the excavator and dozer fleets.

The detailed shape of the mining blocks within each pit will depend on the thickness and dip of the manto and the overburden depth and topography across the pit.  However, in most cases each block will be mined a series of strips no more than 70 m wide to accommodate efficient dozer stripping.

Pit Selection

Prior to settling on the combination of underground and small scale, open pit mining adopted for this analysis, the potential for a large scale, low grade open pit operation with limited selectivity was

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assessed as a possible method of exploiting multiple mantos. This initial analysis of open cut potential was performed in 2005 using standard Whittle pit and Minex® seam optimization software. The resultant open cut operation was extensive, large scale and low grade and characterized by high stripping ratios.  While the potential for a very large scale, low grade operation was demonstrated from a technical point of view without regard to cost, this proposed mining method created a number of issues:

· Large scale open cut mining would deliver average mill head grades of around 0.9% Cu.  To achieve the desired level of copper metal production and sufficient cashflow for economic viability the metallurgical process plant would need to operate at between 6 and 10 Mdt/a.

· The extensive mining operation would create major changes to the landscape over a large area with correspondingly significant environmental impacts which MMB wished to avoid, if possible.

· To access the relatively narrow but extensive mineralized mantos large volumes of overlying waste would have to be removed resulting in waste to ore stripping ratios of up to 20 :1

· The initial capital cost in terms of pre-production waste stripping and mining equipment fleet acquisition combined with the capital cost of a metallurgical processing plant with a capacity of 6 to 10 Mdt/a was thought to be beyond Baja Mining’s capability to finance.

· The distribution of relatively thin mineralized mantos and significant thicknesses of totally barren overburden resulted in an uneven production schedule.  Attempts to maintain a steady, year on year plant feed rate resulted in significant spikes in overburden stripping requirements with corresponding inefficient equipment fleet utilization.

Work to date on open cut and underground operating costs shows that underground mining can deliver ore at an estimated range of US$9.00 to US$12.00/dt.  Open cut ore mining costs are estimated to deliver ore in the range of US$7.50 to US$10.00/t.  This incorporates a waste to ore ratio below an estimated 9 to 1.

This finding caused a reassessment of underground mining potential and the eventual development of the combined underground and small scale open pit mining method that forms the basis of this Technical Report which assumes open cut mining will supplement underground mining though out the project life.  The variation in underground mine production grades will change the amount of calcium carbonate required for the process.  It is expected that the rate of quarrying will be adjusted to meet to plant requirements and occasional turndown of quarry operations will be required.  This makes it possible to transfer the calcium carbonate quarrying fleet to the open cuts and mine sufficient ore to maintain 3.1 Mdt/a feed for the process plant.

The open cut mining areas shown in yellow and outlined in blue in are scheduled to be mined from highest to lowest copper equivalent grade while maintaining efficient access from the pits to the truck loading points developed for the underground mines.

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Figure 18-2: Open Cut Mining Locations

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18.1.2 CALCIUM CARBONATE QUARRY

The ore treatment process requires 870,000 tonnes of calcium carbonate grading 45% CaCO3 per year for various neutralisation duties.  This tonnage requirement varies according to the ore grade and pay metal mix.  This calcium carbonate will be sourced initially from a section of the coquina fossiliferous limestone bed in the north western corner of the mining area.  The coquina bed is part of the Gloria formation which occurs high in the Boleo sequence and can be seen outcropping along many ridge lines.  Locally this section of the coquina bed can average 60%- 65% CaCO3.  In later project years, other sites located on the property will supplement the initial quarry site.

The quarry mining method is based around one 95 tonne hydraulic excavator loading 55 tonne rigid body dump trucks.  All material is blasted using an average powder factor of 0.33 kg of explosives per cubic metre of rock.

As the quarry advances into the arroyo walls, some parts of the pit can reach up to 100 m height from the pit floor to the pit crest.  There is limited room to develop haul roads over this height and the cost of hauling down large elevation changes is prohibitive.  To overcome the elevation differences, two large bulldozers will be used to push the blasted waste rock down to the excavator working level, which will generally be no more than 30 m above the final pit floor.

The hydraulic excavator is in backhoe configuration to allow greater flexibility in mining irregular bench shapes and forming access ramps onto the benches and across the waste dumps.  The dump trucks haul ore to the crushing station at the head of the Boleo haul road or the low grade stockpile and waste either to the out of pit waste dump or to a final in pit waste position above one of the mined out areas.

Calcium carbonate hauled to the crusher will be held in stockpile and rehandled into the crusher by a wheeled loader.  The haul road also provides a route for copper ore from several of the underground mines scheduled for the same area, so both calcium carbonate and ore will be trucked from the area.

Part of the calcium carbonate deposit overlies potential ore grade Manto 3.  This area was considered too deep to surface mine initially but after the southern portion of the quarry is removed, the reduced overburden depth makes it possible to extend copper ore mining farther north.

Purchased Calcium carbonate

The hydrometallurgical copper cobalt recovery process requires approximately 145 kg of pure CaCO3/dt of ore.  As currently planned, the onsite quarry operations can supply all the plant needs during the projected operating life.  As the grade of calcium carbonate from the quarry rises and falls over the LOM, more tonnes of calcium carbonate are required during periods of lower grade to maintain the CaCO3 tonnage fed to the plant.  The quarry schedule assumes that calcium carbonate feed production rate will be capped at 1 Mt/a.  At full ore production of 3.1 Mdt/a, the 1 Mt/a calcium carbonate limit is reached when the grade falls below 45% CaCO3.  Imported calcium carbonate at 90% CaCO3 can be purchased to maintain feed grade at no less than 45% CaCO3 and the total calcium carbonate feed rate at no more than 1 Mt/a.

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High-grade calcium carbonate can be purchased from commercial quarry operations in mainland Mexico and barged across Sea of Cortez from Guaymas.

18.1.3 Other Surface Mining Tasks

In addition to quarrying calcium carbonate and mining copper ore, the open cut mining fleet will be used for other functions requiring bulk excavation or haulage capability.  These tasks will be achieved by either using spare time in the fleet work roster or by providing additional equipment, which can be used interchangeably with the quarry or mining operations.

Rock Fill for Tailings Dam Wall Construction

During the pre-production year the starter dam wall will require an estimated 1.95 million bank cubic metres of earthen fill to be quarried from within the tailings dam impoundment area to form the bulk of the initial wall structure.  An additional 1.8 million bcm of earthen fill will be placed to form a “tail” up to the 150m level to provide a working surface for the remaining dam lifts.  The surface mining group will excavate this rock as part of mining work to remove overburden and expose valuable manto ore within the footprint of the tailings depositional area.  This is partially due to avoid the contractor’s cost that would include mobilization, accommodation for a significant workforce, amortization of the contractor’s equipment and the contractor’s margin.  In order to avoid these additional costs part of the calcium carbonate quarry fleet will be mobilized in Month 4 of the pre-production year -2.  The excavator, two 55 tonne trucks, a blast hole drill and several support units including a bulldozer, a grader and a water truck will work 2 x 12 hour shifts per day, six days per week for an estimated seven months delivering rock to the dam wall site.  At the conclusion of this work, the fleet will move to other required earthworks and then to the calcium carbonate quarry to build stocks ready for commissioning and operation of the process plant.

Ore Haulage from Underground Portal Sites

Most of the underground mine entry points will be serviced by an 85 tonne mine haul truck fleet to take mined ore to the process plant stockpiles.  Ore from portal sites will be re-handled off mine stockpiles by a 6.5 m3 wheeled loader into 85 tonne articulated haul trucks for transport to the nearest intermediate stockpile location or directly to the plant stockpile.  The truck and loader size selection is productive and is large enough to handle all the ore from active portal sites.  It is not necessary to have a loading and haulage fleet dedicated to each portal site.

General Earthworks and mine operations’ infrastructure

The open cut mining fleet will carry out other support roles for the operation including lifts of the main and supporting tailings dam(s), road surfacing with gypsum mined from outcrops in Boleo Arroyo, road watering for dust suppression and general site earthworks as required.  Extending hauling roads, creating new underground portal sites, tailings dispersion changes, contemporaneous reclamation, garbage collection, haul road watering for dust control are just a few of the examples of on going work assigned to the surface mining group.

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18.1.4 Surface Mining Costs

Capital Costs

Surface capital mining costs (as set out in Section 18.1.4.)  are expected to be incurred within 24 months of re-commencement of construction and includes the cost of surface mining equipment as part of the capital cost estimate in years 2010 to 2012, prior to plant startup.  The leasing of major surface mining equipment is estimated for both initial and sustaining capital projections.  The estimate of capital costs excludes the cost of assets which are expected to be leased.

Beginning in year 2012, there are sustaining capital needs to

· extend haulroads, utilities and beltlines

· continue with the tailings dam lifts

· open additional underground mine portal sites.

· Increase the mining equipment fleet to meet project production levels

Operating Costs

Average surface mining costs, excluding costs of equipment leases, for each activity are:

· Mining overburden rock for the starter dam of the tailings dam wall:

1.9 Mbcm at US$5.41/bcm

· Calcium carbonate quarry operations:

average mining cost per BCM waste and calcium carbonate at US$3.51/t

Table 18-1: Average Cost per Tonne for calcium carbonate

Life of Project	Mined	CaCO3	US$/t	US$/t
	Tonnes	Tonnes	Mined	CaCO3
Quarried CaCO3	18,453,278	9,337,359	3.25	6.42
Purchased CaCO3	0	0	24.00	26.67
Total	18,453,278	9,337,359	3.25	6.42

· Loading and hauling mined ore to stockpiles and ROM loading points:

92.4 Mt (wet) at US$0.93/t

· Open cut mining operations:

Table 18-2: Average LOM Costs

Life of Project	US$/m3 Ore and Waste	US$/wt Ore and Waste	US$/wt Ore	US$/dt Ore
Average Cost	1.09	0.55	4.03	5.41

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18.1.5 Underground Mining & Mine Design

Underground mining operations at Boleo will consist of mechanised room-and-pillar mining with pillar extraction.  The initial mining approach is to target and sequence mining extraction to take advantage of the higher copper (Cu) grade concentrations, or “bull’s eyes” that exist in each of the mantos during the early production years.  The project life is a minimum of 23 years.  Mine planning by Agapito has been based on the geologic model provided by Wardrop.

Underground mining will take place in Mantos 1, 2, 3 and 4 where (1) there is reasonably sized, contiguous mineralized zones classified as either measured or indicated resources in the current resource model, (2) the mantos have a Cu content greater than 0.2% at a minimum mining height of 1.8 m, and (3) the mantos have an overburden thickness of at least 20 m.  Approximately 52.8% of the underground mining is in Manto 3, 25.6% in Manto 2, 18.2% in Manto 1 and 3.5% in Manto 4.

The underground mining resource was divided into independent mines, each of which is planned for access from logical portal locations (manto outcrops, box cuts, and/or shallow slopes).  To the extent practical, the mines segregate the resource by copper grade.  The primary objectives for sequencing the mines were winning the higher copper grade ore first, while optimizing the initial plant feed grade higher than 2.0% Cu and then maintaining a higher feed grade for as many years as practical.  The copper grade gradually decreases by year for the life of the project.

Production and mined ore grades for the underground mine plan were modelled by Agapito using SurvCADD® software, a mining design program popular with seam type of underground mining.

Underground mine production starts with one mining unit approximately 18 months prior to plant commissioning.  Production is systematically ramped up over the next 24 months to six mining units operating on a 3 crew, 24/6 rotation producing a targeted annual production rate of approximately 3.1 Mdt/a, coinciding with the start of the second year of plant operation.  Underground production is ramped up to 3.1 Mt by the end of project year 1.  The sixth underground mining unit is added to compensate for declining ore grade and production fluctuations brought about by the scheduling of advance and retreat mining.

Underground Mining Method

Because the Boleo mantos display similar depositional and geotechnical similarities with seam deposits (bedded and relatively soft ground), the recommended underground mining method is similar to coal, trona, potash, and salt seam room-and-pillar mining with pillar removal as successfully used in North America, Australia, and South Africa.

Underground mining will utilize (1) a combination of continuous miners, rubber-tired batch haulage, and mobile roof-bolting equipment for access drift and pillaring crosscut development and (2) remote-controlled continuous miners, continuous haulage, and mobile roof supports (MRS) for pillar extraction.  Development mining and pillar extraction may be independent activities with different equipment and crews; however, mine scheduling and synchronizing mine production with plant feed

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requirements dictates some units conducting both development and pillar mining operations.  For both development and pillar removal, ore haulage from the producing units to the portal stockpile will be by belt conveyor.

Ore from the underground mines will be transported to the plant by haulage truck from the surface stockpiles at the mine portal.  Hauling to the plant ROM stockpile is expected to work on a 24/6 basis.

Depending on mine size, three or four parallel drifts (access mains), connected at regular intervals by crosscuts, will be excavated from the mine portal through the manto to access mining districts.  Most mine portals can be established above the typical storm runoff levels but where this is not possible; flood protection measures will be taken to minimize the possibility of flooding the workings.  Emergency pumping equipment is included in the capital cost estimate for the underground mines.

Where production scheduling and mine conditions allow, development work for an entire district will be completed prior to initiating pillar removal.  Multiple, independent blocks of ore, called “districts” will be simultaneously mined in various stages of development and pillar removal.  Equipment will be moved between districts as needed.  A conceptual mine plan for a typical mining district is shown in Foigure 18-3.  A preliminary layout for proposed Mining Area, M310 is shown in Figure 18-4.

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Figure 18-3:              Conceptual Mine Layout for a Typical District

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Figure 18-4: Mining Area M310 Preliminary Layout

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Resource Recovery

Underground resource recovery will be impacted by extensive faulting, steep manto dip, mining method/equipment limitations, and historical mine workings.  Except where access drifts cross faults, a 10 m buffer is left unmined on both sides of faults.  Areas where the manto dip exceeds 25% are not mined.  In general, during advance mining operations, areal ore recovery is limited to 25% of the available in place material.  For retreat mining operations, areal recovery is estimated to be 75% of the available in place ore, providing an estimated overall recovery of available in place material averaging 65%.

Resource Grade Dilution

The underground resource model for each manto initially calculates ore grade for a 1.8 m interval.  Where total manto thickness is less than 1.8 m, the grade is diluted to account for a minimum 1.8 m mining height.  Where the manto thickness is greater than 1.8 m, the Cu grade of each adjacent 0.3 m intervals are sequentially checked.  If the immediately adjacent 0.3 m interval exceeds 0.5% Cu , 0.3 m is added to the resource height and average grade is adjusted to include that additional increment, and only then is the Cu grade of the next 0.3 m interval evaluated for inclusion in the resource model.

The mining equipment selected for Boleo can adapt to variable mining heights between 1.8 m and 4.2 m.  Within those mining limitations, the excavation height will approximate the grade interval, not the 0.3 m increments of the resource model.  Consequently as the maximum height is determined by Cu grade only and not the CuEq grade plus having equipment capable of mining to this grade interval, the resource model already accounts for reasonable grade dilution and no additional dilution adjustments are appropriate.

 Geotechnical Mine Design

All underground pillar and excavated opening dimensions used for mine layout and design are based on (1) observations made and documented3 during test mining at Boleo’s Texcoco mine during the last quarter of 2005 and 2006, and (2) a geotechnical study conducted by AAI.4

 Productivity

Productivity at the underground mines is based on observations made during test mining at the Texcoco Mine including a roof-bolting productivity test conducted in November of 2006 and AAI’s extensive experience with similar production equipment in coal, potash, and trona mines.

3 Agapito Associates, Inc. (2006), Geotechnical Performance Study for Underground Mining of El Boleo Copper Cobalt Project, Texcoco Test Mine Including Operations Observations and Recommendations, draft report to Baja Mining Corp, July.

4 Agapito Associates, Inc. (2007), “Geotechnical Evaluation for Underground Mine Design,” report to Baja Mining Corp, February, 2007.

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The shift production was further refined by Baja Mining as a result of their knowledge of the specific site conditions. It is assumed that developing mining units will produce 545 dt per unit shift and pillaring units will produce up to 2,536 dt per unit shift.

Worker Safety

Underground mining inherently contains hazardous working conditions.  The designed mining systems, including fire prevention and fire fighting, emergency transportation, ventilation, power, communications, first aid and production equipment specifications have incorporated not only applicable Mexican mining laws, but also North American, specifically, US Mine Safety and Health Administration (MSHA) standards as promulgated in Chapter 30, Part 57 where applicable.

Worker health standards for hazards such as noise and dust were incorporated into the basic mining plans where applicable in accordance with US Mine Safety and Health (MSHA) standards as promulgated in Chapter 30, Part 56.

Worker safety training has been incorporated into schedules and plans as both newly hired experienced and newly hired inexperienced programs.  Annual refresher training has been allowed for in the basic mine production schedules.

Mine layouts

Figure 18-12 shows the overall mine working layouts.  Because there are mine workings scheduled for production above (or below) another target manto, there is a time delay of at least two years to allow subsidence and settling to occur where necessary.

18.1.6 Underground Mining Costs

Underground mine costs were based on the following:

· Capital (initial and sustaining)/leasing costs were estimated on an annual basis for the equipment and infrastructure required to produce the required ore tonnage and grade from the proposed underground mine sequencing plan.

· Labour costs were estimated by Baja Mining for each employee classification and applied against a detailed workforce schedule developed to produce the required ore tonnage and grade from the proposed underground mine sequencing plan.

· Maintenance cost per dry tonne of ore produced was estimated by averaging actual maintenance costs from five US underground coal mines utilizing similar equipment as proposed for Boleo.  These costs were then adjusted in specific categories by Baja Mining in an effort to improve the accuracy of the estimate given the knowledge of the deposit and the robust nature of the specified equipment.

· Supply cost per dry tonne of ore produced was estimated by calculating the anticipated installed roof support cost per wet ton of ore produced and adding an estimated cost per wet ton of other supplies based on experience from US underground coal mines utilizing similar mining methods as proposed for Boleo.

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Figure 18-5: Overall Mine Work Layout

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Capital Costs

Capital costs for underground equipment (as set out in Section 19.1.3) are expected to be incurred within 24 months of re-commencement of construction are estimated for years 2010 (6 months), 2011 and 6 months of 2012 for a total period of 24 months.  Underground mining equipment, portal conveyors, mine fans, chillers, mine electrical equipment, and speciality surface equipment capital and leasing costs were developed for the initial capital expenditure as well as for the entire the life of the project.  The estimate of capital costs excludes the cost of assets which are expected to be leased.

Leasing costs include an estimate of lease financing and administrative costs.  Operating leasing terms have been developed from active proposals received from third party leasing organizations.  At least one potential large equipment supplier to the project has stepped forward with a detailed proposal to lease all their equipment and to lease a substantial amount of other project related equipment not of their manufacture.  Downpayments for leasing the selected equipment from the supplier are estimated at 15% of the equipment cost and 20% of the equipment cost for machinery supplied by a third party respectively.

Operating Costs

As stated earlier, operating costs for the projected mine plan were developed from experience and data from operating mines in similar mining conditions.  AAI had developed initial operating costs which were subsequently modified by Baja to reflect management’s estimate of costs based on their experience and knowledge of the unique Boleo site conditions.  These were primarily related to updated maintenance cost estimates and the projected rebuilding and replacement schedules for primary mining equipment.  Operating costs vary with the production schedule estimated for each year of mining.  Operating lease costs are not included in the estimates for mine operating costs.

Table 18-3: Projected Underground Mine Costs@@redo

Year	Labor	Roof Support	Operating Supplies	Maintenance	Total $    (Wet mined tonne)	Total $    (Dry mined tonne)
-2 Ramp up	$12.72	$1.63	$0.80	$0.75	$15.91	$21.32
-1 Ramp up	$9.48	$1.80	$0.80	$1.13	$13.21	$17.70
1 Ramp up	$1.41	$0.85	$0.80	$1.51	$4.57	$6.13
Typical Year	$1.14	$0.85	$0.80	$1.51	$4.30	$5.76

Note: *Excludes administrative and equipment leasing costs.

18.1.7              Surface Infrastructure Development

The property will be developed to support the access, utilities and movement of ROM materials from the various underground mining portals, surface quarry site and tailings dam construction site using design plans developed by Baja Mining, AAI and Amec E&E companies.

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Mine Dry Facilities

The mine plan will be developed by building a central operations and changing facility (CF) strategically located at the plant site.  This will be supported by mobile changing facilities located at each of the active portal sites (PF).  Mining crews assigned to work specific mining areas will use these changing facilities at the portal to reduce the needed transportation time each shift.  These mobile facilities will each contain small changing rooms, an office, stockroom and other mine related facilities.  The facilities are scheduled to be moved as mining is scheduled across the property.

Transporting mine labour is planned as a two-step process.  First, all mine labour would assembled at the plant changing facility prior to the start of shift for transport to the active portal facility changing room.  After being transported to their assigned mining portal, the miners would change clothing and obtain their mine safety and work equipment.  The start of shift will commence at the boarding of mine transports from the portal facility for the ride to specific work assignments in active mining panels.

Management, engineering, surveying, and other support activities are expected to report to the central facility at the plant site.  Offices and storage have been allowed for.

Power

Mine power will be supplied from HFO fuelled generators as part of the co-generation plant located near the processing plant using overhead high voltage lines to the central facilities where it will then be split to the active mining operations.  The power will be stepped down to supply the needed voltages at the portal areas.

Water

Processed water will be trucked to each active mining portal and pumped and stored in larger tanks to provide for changes in consumption at the PF.  Fire water for each underground mine will be drawn from the portal facility tanks of sufficient size to meet fire-fighting requirements based on mine size.

18.2           Environmental

The project is located within the Vizcaíno Biosphere Reserve, which is centered on the Desierto de Vizcaino on the west central coast of the Baja California Peninsula.  Originally, the area included 1.5 Mha dedicated to the protection of the Western Coast as a refuge zone for the grey whale and the distribution zone of the Mexican antelope (Berrendo).  In 1986, the area was extended south to protect the ancient paintings at the Sierra de San Francisco; the historic buildings, dating from the late 1800s, in the town of Santa Rosalía which are associated with the early mining of the Boleo district and the coast of the Sea of Cortes.  Currently, the El Vizcaino Biosphere Reserve covers a total surface area of 2,546,790 ha as stipulated in the official decree dated November 30, 1988.

Although the area of the project is recognized as being significantly affected by historical mining activity, the fact that it is located within the boundaries of the biggest natural protected area in Mexico places considerable attention on its evaluation.  Final approval is, jointly, decided by two independent

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branches within the Secretariat for the Environment and Natural Resources (SEMARNAT): the National Commission for Natural Protected Areas (CONANP) and the General Direction for Environmental Impact and Risk (DGIRA).  Mention must be made of the fact that the surface property of MMB, where the project will be developed, has no significant archaeological value as determined by the National Institute for Anthropology and History (INAH).  Additionally, all the buildings which are considered a cultural heritage are outside of the Boleo project and study area boundary and will not be affected by project development.

During 2006, MMB successfully completed a full Environmental Impact Assessment that covers the construction, operation, and closure phases of the Boleo project.  Given the complexities of the project itself and the environmental sensitivity surrounding the project location, the Mexican Federal environmental agency, Secretaria de Medio Ambiente y Recursos Naturales, (SEMARNAT) requested the submittal of an Environmental Impact Manifest with a regional scope.  The change of scope required additional fieldwork to characterize fully the regional area of influence of the project.  This, in turn, caused a delay in the EIM submission date from February to May 2006.  The evaluation process also included a request from SEMARNAT to submit additional information relating to the project to better clarify the identified environmental impacts.  This request was given to MMB on July 5th, 2006.  The information was formally filed on October 2nd, 2006.

Finally, after incorporating the observations and recommendations from the National Commission for Natural Protected Areas, the Secretariat for Urban Planning, Infrastructure, and Ecology of the State Government of Baja California Sur and the Municipal Presidency of Mulegé at Baja California Sur, the environmental impact resolution was issued on November 27th, 2006 and delivered to MMB on December 7th, 2006.  This resolution authorizes the construction, operation and closure of El Boleo Mining Project.  The official document number containing this resolution is S.G.P.A.-DGIRA.-DDT.-2395.06 and is signed by the General Director for Environmental Impact and Risk (DGIRA).

This authorization allows MMB to initiate the procedures to obtain permits that are more specific.  In 2007, MMB will concentrate its efforts in securing these additional permits and in managing the terms and conditions that were established in the environmental impact authorization.

Table 58 is a summary of environmental activity showing the various project phases and the required mitigation or compensation needed to address such activity.

As with the exploration permits, MMB has setup a permit management and compliance plan.  The main actions taken to date are:

· A covenant agreement with CONANP was set up and signed on January of 2007.  Through this mechanism, the agreed compensation funds will be made available to the El Vizcaino Biosphere Reserve as established in this agreement.  As mandated by this legal instrument, a technical subcommittee that will oversee the use of these funds was set up in March 2007.  MMB holds a seat at this subcommittee with full voting powers.

· An application for the authorization to mine in a natural protected area was submitted to CONANP.

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· MMB has requested for an extension of the terms and conditions that had early deadlines (one to three months after permit issuance).  The extension has been granted until August of 2007.

· MMB has initiated with the integration of the technical report to request a change in land use as mandated by the General Law of Sustainable Forestry Development.  Additionally, it has engaged in ongoing negotiations with the Autonomous University of Baja California Sur to act as its lead educational institution for the integration, follow up and supervision of the main mitigation measures contained in the Environmental Impact Manifest and in the permit.

After the construction engineering of the project is finished, the final design will be compared with that submitted in the EIM and approved in the permit.  As allowed by the Regulations of the LGEEPA on Environmental Impact, a modification of the original permit will be applied for if needed.  Significant and environmentally adverse modifications to the original design could require a full impact assessment and the issuance of a supplementary permit.

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Table 18-4: Environmental Activities

Phase of the Project	Activity	Environmental Parameter Affected	Mitgation Measure	Compensation Measure
Construction	Construction of facilities (camp, concrete plant, contention dyke, etc.)	Vegetation abundance and diversity as well as affectation to endemic and protected species	Vegetation Rescue and Transplant Program	Creation of a Trust Fund for CONAMP aimed at supporting conservation activities within the Biosphere Reserve of  El Vizcaíno
	Construction and rehabilitation of access roads	Disturbance of passages, creating barriers for natural movement of fauna	Fauna Protection Plan Road design considering communicating passages (for fauna)
Operation	Open Cuts	Changes in terrain slope	Protection against Erosion Program (program for soil recovery)
		Changes in runoff patterns
		Changes in soil permeability by the displacement of soil layer
		Soil removal  (change of type)
		Abundance: removal of vegetation layer (abundance)	Vegetation Rescue and Transplant Program
		Diversity: selective removal of vegetation species
		Removal of protected and endemic vegetation species
		Abundance: Fauna’s driving away and/or elimination	Fauna Protection Plan
		Interruption of fauna transit
		Some  species of protected or endangered species might be eliminated or driven away
Operation	Tailings Dam	Changes in runoff patterns	Protection against Erosion Program (program for soil recovery)	Creation of a Trust Fund for CONAMP aimed at supporting conservation activities within the Biosphere Reserve of  El Vizcaíno
		Soil removal (changes in permeability, composition and its characteristics)
		Abundance: the flooding of the tailings dam will prevent vegetation from re-colonizing the area thus decreasing abundance	Vegetation Rescue and Transplant Program
		Flooding of the tailings dam will cause some habitats to disappear hence decreasing vegetation and disappearance of endemic and protected flora
		Abundance: Fauna’s driving away and/or elimination	Plan de protección de fauna
		Interruption of fauna transit
		Some  species of protected or endangered species might be eliminated or driven away

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18.3 Community Development

In order to comply with IFC standards and ensure the acceptance of the project from the population of Santa Rosalía, a Public Consultation and Disclosure Plan was integrated (PCDP).  This plan included the following implementation stages:

· Speaker Training: that included a strength – weakness – threat – opportunity analysis

· Stakeholder Identification: derived from a fieldwork conducted at the State and Municipal levels

· Design of Communications Strategy

· Execution of Public Presentations of the project.

A detailed description of this process and its results is contained in the report of the PCDP.  Salient points are summarized in the following paragraphs:

As a result of the field investigation to map out the stakeholders of the project, a communication strategy was set up for the public presentations of the project.  Its main characteristics are:

  · Objectives:  to improve the community’s knowledge of the project; to assess its acceptance and to address any concerns or opposition that may arise to its development.

  · Scope: given the size and area of influence of the project, the consultation is to be directed to all persons or organizations, with a current and active presence in Santa Rosalía; the Municipality of Mulegé or the State of Baja California Sur, which show an interest to participate in the process.

    ·   Communication Contents: background of the project and its sponsors; company vision; project description; legal compliance of the project; description of the environment and socio-economic factors in its area of influence; expected impacts of the project; prevention, mitigation, control and compensation measures; social responsibility and community wellbeing.

   ·   Communication Approach: verbal communication aided by visual and written material designed for the needs of each selected audience.

   ·   Timing: in the interval between the initial authorization of the project and the conclusion of the development of its construction engineering to include pertinent changes that could arise from the presentation process.

    ·   Speakers: management team and support staff from the Mexican company that is promoting the project (MMB).

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18.3.1 Public Presentation Events

The calendar of events is enlisted below:

· Participation of MMB with El Boleo Project during the 13th Annual Science and Technology Week celebrated at Santa Rosalía from October 23 to October 27, 2006.

· Presentation of El Boleo Project at the Hotel Fiesta Inn in La Paz by invitation and for selected stakeholders on February 1, 2007.  The invitation included a CD that contained all files relative to the environmental permitting process of the project.  A balanced mix of attendants, which included government officials; academia; entrepreneurs and civil society organizations (NGO’s), was sought and achieved.

· Meetings with focus groups from Santa Rosalía from March 7 to March 8, 2007.

· Public Presentation of the project at Santa Rosalía on April 19, 2007.

There is strong support from the community who perceive it as an opportunity to promote economic growth for the region.  No opposition to the project was detected during the celebration of these events.

18.3.2 Grievance Mechanism

MMB has pledged to establish a formal office to manage the relationship with the community once the availability of financing for the project is confirmed.  Its initial project is to coordinate the integration of the Land Ordinance Plan that has the objective to identify the most pressing present and future needs of the community.  This identification will enable all involved to set up activities to resolve these needs under a shared responsibility scheme.  The liaison officer will have the task to establish a grievance mechanism to answer the concerns of all valid stakeholders.

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19 ECONOMIC ASSESSMENT

19.1.1 Summary

This economic assessment of El Boleo project is based upon:

· The mineral resource estimate for copper, cobalt and zinc prepared by Wardrop and Baja Mining as discussed in Section 17

· The open cut mine design developed by Agapito Associates Inc, as discussed in Section 17.1.1

· The underground mine design developed by Agapito Associates, Inc., as discussed in Section 17.1.5.

· The process plant design developed by Bateman Engineering Canada Corp.and Baja Mining, using recoveries of copper, cobalt and zinc achieved during the Fully Integrated Pilot Plant testing program at SGS Lakefield Research Ltd., Lakefield, Ontario, conducted under the joint guidance of Baja Mining Corp. & Bateman, as discussed in Section 16.

US$ and metric measures are used throughout this section.

The current base-case is for annual mine production to deliver up to 3.1 Mdt/a to the process facility; with maximum annual metal production of 61,200 tonnes of copper, 2,200 tonnes of cobalt and 37,000 tonnes of zinc sulphate monohydrate.

The remaining project capital cost, at December 31, 2009, for the construction of the mine and process plant complex is currently estimated at US$899.474 million (including Engineering, Procurement, Indirects, Construction Management, Owner’s Costs, and an 11.6 % overall Contingency, but excluding estimated leased equipment).  Total operating costs are estimated at US$35.71/dt of ore treated, averaged over the 23 year project life being modelled.  During the initial 23 years project life a total of 70 Mdt of ore will be processed.  The Total Measured and Indicated Resource is 264.7 dry metric tonnes, potentially leaving a large amount of ore to be processed after the initial 23 years.

Process recoveries used in the economic analysis are as follows:

Copper recovery –        90.65%

Cobalt recovery –          78.76%

Zinc recovery –              64.44%.

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Case 1 financial modelling of the project was done using the weighted average three-year trailing metal price (in accordance with SEC guidelines) at December 31, 2009 of $2.91/lb for copper; $26.85/lb for cobalt; and $1,175/tonne for zinc sulphate monohydrate flat over the life of the project.

Expressions of interest have been received from potential metal off-take partners indicating that net-back price of copper would include a slight premium (assumed to be $0.02/lb) above the LME, or COMEX, price after adjusting for freight.  The financial analysis additionally makes provision for delivery of the products to the end markets, including packaging, freight, and freight insurance.

The modelling, based on the current mine schedule, indicates that the project is financially very attractive at Case 1 metal prices.  Financial modelling, using the Case 1 prices and 23 years for the project life, shows that the project could generate a net after-tax Internal Rate of Return (IRR) of 25.6% with a discounted net present value (NPV), at an 8% discount rate, of US$1,306 million.

Using a 5% discount rate generates an NPV, after tax, of $US1,922 million.

Cash flow analysis was also conducted using the following metal price assumptions to test the project's economic robustness and sensitivity to changes in those prices, as shown in Table 19-1:

· Conservative case - $2.25/lb for copper; $20/lb for cobalt; and $1,100/tonne for zinc sulphate monohydrate.

· Current January 2010 prices - As of January 8, 2010 were $3.40/lb for copper; $21.17/lb for cobalt; and $1,000/tonne for zinc sulphate monohydrate.  

The project is sensitive to four key variables:  Copper price; cobalt price; the capital cost and to operating costs.  The sensitivity of the After-Tax IRR and NPV (at 8% Discount Rate) relative to the Case 1 is shown in the table below to indicate the effect of plus or minus 10% changes in the key variables.

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Table 19-1:               Economic Assessments – Base Case Highlights

Description	Quantity or Value
Mine Production	3,100,000 dt/a
Metal Production	Up to 61,200 t/a Cu cathode Up to 2,200 t/a Co cathode Up to 37,000 t/a ZnSO4 salt
Capital Cost to complete	US$889 million
23 year average Operating Cost,  excluding start-up year	US$35.71/t of ore ($35.71)
Weighted average three-year trailing metal prices	Copper – US$2.91/lb Cobalt  – US$26.85/lb ZSM – US$1,175/t
(After tax) Internal rate of return (IRR)	25.6 %

Table 19-2: Sensitivity to Key Variables

	After Tax IRR			After Tax NPV at 8% ($ millions)
Variable	-10%	Base Case	+10%	-10%	Base Case	+10%
Copper Price	23.2%	25.6%	27.9%	$1,111	$1,306	$1,505
Cobalt Price	25.0%	25.6%	26.1%	$1,246	$1,306	$1,367
Capital Cost	27.7%	25.6%	23.7%	$1,361	$1,306	$1,250
Operating Cost	26.3%	25.6%	24.8%	$1,379	$1,306	$1,234

19.1.2 Product Marketing

The processing facility at El Boleo property will produce London Metal Exchange (LME) grade copper cathode (metal), and high purity cobalt metal.  It is not proposed to produce zinc metal but rather to evaporate the zinc sulphate stripped in the zinc solvent extraction circuit to produce zinc sulphate monohydrate for subsequent sale into the soil micronutrient market, animal feed market, or delivery to a zinc refinery.

Off-take agreements with respect to Copper and Cobalt are in the process of being negotiated, but “Expressions of Interest” have been received and Off-take agreements will be finalized prior to construction.

The copper cathode that is produced on site is expected to meet or exceed LME Grade A purity specifications and should command typical benchmark premiums above LME pricing.  Costs for freight insurance, and freight to market will be deducted from the indicated selling prices

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The current process flowsheet provides for the production of cobalt metal on-site. Costs for packaging (drums), freight insurance, and freight to market will be deducted from the indicated selling prices.

The zinc sulphate monohydrate is considered to be a “value added” product and should command a price equivalent to the value of contained zinc metal, or higher.  It is assumed that the material would be sold through an agent not through an off-take agreement.  The net price received by Baja is also adjusted for costs of packaging (jumbo bags), freight insurance, freight to market, and the agent’s selling commission.

19.1.3 Capital Cost Estimate

The Boleo Feasibility Study (DFS) was issued in July, 2007.  Within this study was an estimate of the capital costs.  In January 2010 an update was issued.  The current update incorporates estimated plant capital , operating and owner’s costs prepared by the Company,.  A revised geological model, in accordance with NI 43-101, has been prepared by Wardrop, a Tetra Tech Company (Wardrop) and used by Agapito Associates, Inc. (AAI) to prepare the current mine plan.  The initial capital cost for the mine was prepared by AAI, however mine sustaining capital and operating costs were prepared by the Company.  The Company utilized these cost estimates to prepare the financial projections.  In doing so, the capital and operating costs were adjusted to reflect leasing of certain equipment, and to capture costs incurred in the fourth quarter of 2009.

As at December 31, 2009, the total project cost remaining, including Engineering, Procurement, Indirects, Construction Management, Owner’s Costs, and an 11.6 % overall Contingency, but excluding leased equipment, is $889 Million.

Table 19-3 provides the updated cost summary of the estimated total project capital costs to complete.  The base date of the cost estimate is third quarter 2009 dollars.  All costs are listed in US$.  The numbers are rounded to reflect that they are approximations.

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Table 19-3:  Capital Cost Estimate to Complete Summary

Project Area	Estimate Capital Cost (US$’000)
Mining and Tailings	68,914
Process Plant	321,033
Services & Infrastructure	84,022
Buildings	9,072
Total Direct Costs	483,041
Construction Indirects & Freight	70,029
EPCM	52,391
Contingency	92,297
Total Construction Costs	697,758
Owner’s Costs and Pre-Development Costs	141,101
Total cost before W/C and Financing	838,859
First Fills, Reagents, Spares and Working Capital	50,616
Total Estimated Capital Cost to complete	889,474

The overall capital cost to complete is estimated at US$889.474 million including a 11.6% contingency.

This number includes the capital cost components of all infrastructures, the process plant, and the mining operation, including the mining fleet, mining infrastructure, tailings dam, haul road construction, electrical power and water reticulation to the various mine sites, waste disposal, construction camp, pre-development of the underground and surface mines and other owners’ costs including various community initiatives.

The Capital Costs set out in the above table exclude:

· Costs of assets which are expected to be leased

· Mine rehabilitation costs, which are funded through a sinking fund

· Mine closure and environmental remediation costs, which are funded through a sinking fund

· Sustaining capital

· Funds expended on the project to December 31, 2009

Provision for the above costs – with the exception of escalation – has been made in Baja Mining Corp’s economic model.  Working Capital requirements will be minimized through the

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involvement of off-take agreements offering prompt payment for metals following transfer of title as the products are loaded onto transportation ships on site for delivery to designated port facilities in Long Beach, California.

The project evaluation has been presented on the basis of a 25-year mine life, but it is expected that the life of the project will be considerably longer.  There is no credit given for recapture of the salvage value of the equipment after the initial 23 years of plant operation.  Reclamation of the open cut mines is expensed as it is incurred during the initial 25 years of mine life on an ongoing basis, so that there will not be a significant closure and reclamation cost associated with mining.

Sources of Costing Information

The Estimate Update used the original Feasibility Study estimate as a basis for the outline structure.  Updated information was input everywhere possible to result in the updated estimate.  Specifically, the following was revised:

· “Tagged Equipment”

Firm and budgetary quotes were obtained for the majority of the tagged equipment.  Any other equipment item costs were reviewed and updated as required.

· Bulk Material

Quantities were  taken offs using the current available drawings and sketches that reflect the current design layouts.  Engineering consulation was obtained as required for added clarity to the available design.

Material purchase pricing was based on the result of a Request for Proposal (RFP) process to acquire current material prices.

· Labor

Mexican craft labor rates were built up. .  Allowances are included in the built-up rates for specialty craftsmen from the USA as required.

Labor costs were developed through the application and extension of unit manhour rates to quantities in each discipline based on US productivity.  Manhours are factored upward from 1.2 to 1.8 (depending on the discipline) to account for Mexican labor productivity.

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· Indirect Costs

Engineering/Procurement costs and Construction Management costs were completely re-estimated based on current manhour estimates for the various tasks involved.

Labor rates were reviewed and updated to current values.

Indirect expenses were re-estimated based on current planning.

The cost of the construction camp was re-estimated.

Catering costs were based on the result of a previously negotiated and updated contract.

Estimating Methodology

· Earthworks

The earthworks quantity was developed from drawings approximating final layouts and elevations.  Earthworks unit pricing was developed from local quotes in Mexico and a build-up of unit rates using equipment fleet requirements.

· Concrete

Outline sketches of concrete requirements were developed from assumed equipment/building loadings. Quantities were taken off from these sketches.  Concrete unit pricing was developed from a build-up of material/labor requirements by type of concrete pour (Slab on Grade, Footings, Walls, etc.).  This enabled the preparation of estimated costs from the quantities and unit costs.

Provision is made in the estimate for the operation of a batch plant at site.

· Structural Steel

Structural steel quantities were estimated from sketches and preliminary drawings.  Steel material pricing was from budget quotes.  Installation manhours and resultant costs were calculated using manhour installation rates and hourly rates appropriate for this part of Mexico.

· Buildings

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The updated building and architectural estimate used existing building layouts to measure the area.  Average Mexican prices per square meter were used to determine building costs.

HVAC and other building services were allowed using a cost per square meter.

· Mechanical Equipment

The majority of mechanical equipment was re-estimated from firm or budget quotes.  Where quotes were not available, in-house data was used for pricing.

Agreements have been reached with major equipment suppliers to fix equipment pricing and deliveries.  Pricing of equipment for major circuits and systems were updated by the following suppliers:

Bateman Litwin	SX-IX-EW
FL Smidth	Crushing & Milling Circuit Iron Removal System Copper CCD Circuit Limestone Crushing/Grinding and Ore Storage System
Fenco	Acid Plant

Piping and Valves

Piping quantities were estimated using current layouts, sketches, and P&ID’s..  Where process throughput has been increased, but P&ID’s not yet updated to reflect the increase, piping sizes have been increased to allow for increased flow.  Piping lengths were estimated using the P&ID’s for pipe sizing and layouts, and sketches to determine the required length of the pipe.

Valves were estimated using the P&ID’s and valve lists.  Special valves were priced from quotes, while more common valves were priced through the RFQ process.

Electrical

Electrical quantities were estimated using single line drawings, current layouts and sketches.

RFQ’s with bills of materials were sent to vendors to obtain electrical material pricing.

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Instrumentation

The instrumentation estimate was performed based primarily on the Instrument List, data sheets and support details as well as GA’s and P&ID’s.  Instrument wiring was quantified with the electrical estimate.

Instrumentation supplied directly was separated from instrumentation supplied in vendor equipment packages.

Laboratory

A consultant provided estimates of laboratory equipment costs and plans of requirements for the site laboratory that will be managed by MMB.

Pre-Engineered Buildings

Budget pricing was obtained for pre-engineered buildings.

Labor Cost Development

Labor rates were developed for composite crews by discipline.  Allowances were made for expat assistance necessary in some trades.  Additionally, subcontract labor rates were obtained for certain skilled work that will be subcontracted out.

The composite construction crew rates are based on a 10 hour day X 6 days per week with costs in accordance with Mexican pay requirements.

Productivity expectations were reviewed based on typical factors expected for Mexico resulting in a productivity factor being applied of between 1.2 and 1.8, depending on the discipline.

Construction Equipment

Construction equipment needs were determined and a schedule of the duration requirements for each equipment piece was prepared.  Fuel, oil and maintenance costs were determined along with the mobilization and demobilization expenses.

Owners Costs

Owner’s costs have been reviewed in detail, by department and total labour needs assessed and updated to include the latest project implementation plans and updated costs.

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19.1.4          Operating Costs

Operating costs were developed for the update from the following sources:

· Extensive bench scale metallurgical and pilot plant testwork data

· Quoted budget prices for reagents and consumables, typically from North American suppliers

· Appropriate labour costs for Expatriates and Mexican nationals for project development in the Baja California area of Mexico, drawing on remuneration experience from the local Gypsum Project, adjacent to the Boleo Resource on the Baja Peninsula and other Mexican Operations

· Maintenance costs based on other plant operations of a similar nature

· Estimates of open cut mining, tailings dam construction, site earthworks and quarry costs from Agapito Associates, Inc.

· Estimates of underground mining costs from Agapito Associates, Inc based on the Boleo test mine experience as well their operating costing data base of similar operations in North America in coal, potash and trona.  These costs were further updated by Baja Mining staff.

· A factored approach to product marketing and product freight costs based on discussions with freight forwarders, shipping agents and interested off take parties.

Operating Costs have been updated from the Feasibility Study to reflect pricing of consumables in mid-2009.  Additionally, any costs that were factored as a percentage of Capital Costs, such as plant maintenance supplies and insurance have been updated to reflect the updated Capex. Prices to minor consumables such as reagents have been updated to September, 2009 prices based on price quotes from the suppliers.

Table 19-4 shows the operating costs for year 5 of operation.

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Table 19-4: Operating Cost Summary Estimate

Operating Cost Category	Annual Operating Cost (US$ ‘000s) *
Mining Costs
Mining Costs
Labour	4,404
Fuel	428
Operating Leases	23,966
Consumables	20,258
Generator Power (diesel)	1,746
Limestone credit	(2,225)
Total Mining	48,577
Process Plant
Labour	3,857
Sulphur	13,705
Soda Ash	13,923
Flocculant	5,114
Limestone	2,225
LIX 84	331
LIX 63	822
Other Supplies	4,514
Maintenance Supplies	7,095
Lab Operation	225
Generator Power	4,264
Operating Leases	4,254
Total Process Plant	60,330
Camp and security	4,606
Office & Insurance	6,541
Total Operating Cost	120,054
Total Operating Cost per ROM Tonne Treated	38.73
Total Freight, Selling & Distribution Costs	12,894
Total Cash Cost Before Taxes	132,948
Cash Cost per Tonne of Ore (no credits)	42.89

Note: * Year 5 of the LOM, 3.1 Mdt/a plant feed.

19.1.5 Summary

The financial model utilizes the updated mine production schedule, showing a 23 year mine life, the associated diluted metal grades based on the Wardrop resource and Agapito mine plans, the updated capital with the Company providing the operating costs, and projected metal prices.

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Financial modelling of the project was done using the weighted average three-year trailing metal price (in accordance with SEC guidelines) at December 31, 2009 of $2.91/lb for copper; $26.85/lb for cobalt; and $1,175/tonne for zinc sulphate monohydrate flat over the life of the project.

Since publication of the DFS the financial model has been refined to incorporate the current expected expenses and revenue on a quarterly basis until the end of construction, first year operations on a monthly basis and by half-year increments during the next four years.  The model has annual periods for the balance of the life of mine.

Since publication of the DFS in June, 2007, the construction schedule and expenditure profile have been further refined.  As a result, this update reflects start of copper production in the 28 months after the beginning of construction, the second half of 2012.  Capital expenditures are now broken down by quarter in the financial model.  In addition, since the financial model is an analysis of the project, starting in 2010, the portion of capital already spent to the end of 2009 ($144.995 Million) is not included in the expenditure profile.

The economic analysis of the Boleo project has been conducted by Baja Mining Corp. on an all equity basis (with no financial leveraging) using the discounted cash flows, after taxes, for the construction period and first 23 years of the project life.  The project is not limited to 23 years by the ore body, and it can be expected that the project will continue beyond that length of time, adding to the 23-year value of the project.  The projected cash flows allow for all capital expenditures, including construction, working capital and sustaining capital.  Since this is an unleveraged model no financing costs or interest were accounted for in the Project’s costs.

On an all-equity basis the financial analysis indicates that the Boleo project has a payback period for invested capital of approximately 40 months from start-up of commercial production of copper (on an after-tax basis).  The potential net present value of the project utilizing Case 1 prices at a 5% discount rate is US$1.922 billion, or $1.306 billion at 8%, and the internal rate of return on investment is 25.6% on an after tax basis.

Using the assumed Case 1 metal prices the annual revenue from cobalt and zinc sulphate sales is potentially adequate to cover more than all of the operating costs.  This will result in a negative cash cost of copper production, net of by-product credits, of about negative 29 cents per pound of copper averaged over the 23 years of operation.  This also does not account for any addition of manganese revenue which could further drop the net cash cost of copper.

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19.2           Boleo Project Economic Analysis Assumptions and Discussion

19.2.1 Capital Cost & Expenditure Timing

The total project capital costs for the construction of the mine and process plant are summarised in Table 19-3.  The timing of expenditures is based on an accelerated project schedule that enables the start up of the copper production circuit in second quarter, 2012, with commissioning continuing after for an additional four months on the DSX® circuit for the recovery of cobalt and zinc.

At this time, there is no allowance for the capital cost for the construction of the manganese carbonate circuit, and it remains an opportunity for future production.

It is expected that there will be an offset of approximately 4 months between the start-up of copper production, and that of cobalt and zinc.  This philosophy was adopted to allow the operating team to focus on the start-up of the main revenue generator (copper) without the distraction of simultaneously starting up the more complex DSX® circuit.  This should result in a faster ramp up rate for copper, which will more than offset the loss of revenue from the other two products, and make the most efficient use of the technical resources.

This schedule will also enable more efficient use of a smaller construction crew and minimize the size of construction camp required, along with potential impact on the infrastructure of the neighbouring town of Santa Rosalía and its residents.

A summary of the Boleo Project development timing is shown in Table 19-2.  The timing of these activities forms the basis of timing of the cash flows in the economic model.

Table 19-5: Summary of Boleo Project Development Timing

Year	Activity	Percentage Expenditure
2007-2009	Feasibility Study and Basic Engineering	13.9%
2010	Mine and plant pre-development Detailed engineering for process plan Site preparation for process plant Commencement of construction of acid plant Surface earthworks development of roads, dam, etc.	19.3%
2011	Construction of Site and Process Plant Underground mining development Construction of permanent wharf	50.5%
2012	Start-up of copper production Construction of DSX® circuit, cobalt EW, ZSM granulation	16.3%

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2011	Construction of Site and Process Plant Underground mining development Construction of permanent wharf	50.5%
2012	Start-up of copper production Construction of DSX® circuit, cobalt EW, ZSM granulation	16.3%

19.2.2 Sustaining Capital

In the economic analysis, provision is made each year for sustaining capital.  This allowance covers cost associated with capitalized rebuilds, refurbishment rebuilds, and replacement of equipment and major spares to maintain the operation at the design capacity.

The provision for sustaining capital includes:

· general allowance based on an average 3% of the direct capital costs of the process plant and infrastructure (each area has a specific rate).

· mine mobile equipment  rebuilds and replacements from Year 1 onwards

· development of underground mine portals

· bi-annual lifts of the tailings dam

· closure costs for the tailings facility and plant demolition

19.2.3 Working Capital

In the cash flow analysis, an estimate of annual working capital requirements is made and includes the following:

· the net balance of accounts payable and receivable

· first fills of reagents – US$27.2 million in year 1

· capital spares and warehouse inventory – US$16.3 million in year 1, equivalent to 1.8% of total capital expenditure

· in process inventories

· finished product inventories up to the point of transfer to the purchaser; this is expected to be once loaded onto the transport ship, destined for the port of Long Beach as has been discussed with interested off-take parties.

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19.2.4 Revenue

Net sales revenue from the sale of copper is calculated based on the LME (or COMEX) price, plus a premium.  It is assumed that copper will be delivered FOB to a warehouse at the port of Long Beach and these costs are part of the selling costs.  The premium received depends on quality and market conditions, which may vary.  It is assumed that the quality will be LME Grade A.  An average premium of $0.02/lb is assumed.  Financial modelling of the project was done using the weighted average three-year trailing metal price (in accordance with SEC guidelines) at December 31, 2009 of $2.91/lb for copper flat over the life of the project

Financial modelling of the project was done using the weighted average three-year trailing metal price (in accordance with SEC guidelines) at December 31, 2009 of $26.85/lb for cobalt flat over the life of the project   Included in the selling costs are packaging (drums) and freight costs.  The quality of cobalt cathode produced during the pilot plant campaign was very high and would indicate that a premium may be available, but this is not assumed in this analysis.  Cobalt is also about to be traded on the London LME which will greatly enhance the long term marketing of cobalt.

Zinc Sulphate Monohydrate is assumed to be sold into the US fertilizer and animal feed markets, using agents.  Provision is made for freight into the core area of use and for a selling commission to the agents.  Financial modelling of the project was done using the weighted average three-year trailing metal price (in accordance with SEC guidelines) at December 31, 2009 of $1,175/tonne for zinc sulphate monohydrate flat over the life of the project is assumed, before the deductions for freight and commissions.

19.2.5 Depreciation & Taxation

Mexican tax laws provide for two methods of calculation of depreciation.  Assets can either be depreciated at a rate of 12% for 8 years, with the remainder taken in the ninth year or at the rate of 87% in the first year of operation with the remaining 13% taken at the end of the project life.  For the purposes of evaluation of this project, under the all equity funding assumption, it has been determined that it is tax effective to claim the 87% depreciation in the first year, and the remainder in the final year.  This assumption could change when the project financing is put into place.

Mexican federal income tax rates are currently 30%, and will be reduced to 29% in 2013 and 28% in 2014.  There is no state income tax.

There is a value added tax in Mexico and the applicable rate is 11% for Baja California., 16% for purchases outside the border zone.  The tax is paid on purchases, but the end product sales will be zero rated as it will be targeted for export. However, the full amount paid is recoverable by mining projects, which has been the case for the past several years of operations and continues to be assumed for this project.

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19.2.6 Manganese Opportunity

Estimates of the economic benefits of recovering manganese at Boleo have not been updated in this report.  Marketing, product and production investigations have continued since the issue of the Feasibility Study.  The following is a summary of the status of development regarding manganese, recovery, production of value added products and the markets for these products:

· The Boleo orebodies contain substantial amounts of manganese.  Preliminary estimates indicate that up to 220,000 tonnes per year of manganese carbonate containing up to 100,000 tonnes per year of manganese as metal can be recovered from the Boleo refinery with a minor capital investment.

· The technology for manganese carbonate production at Boleo has been tested and proven during the fully integrated pilot plant campaign conducted at the SGS Lakefield facility in June 2006.

· Although there is a substantial operating cost increase due to the consumption of soda ash to produce manganese carbonate, estimates of the cost of the product do not exceed the value of manganese ore with a similar manganese content.  Market studies have indicated that the market for Boleo’s manganese carbonate is a limited to a portion of production.  The quality and purity of manganese carbonate produced at Lakefield was high.  There is therefore an opportunity to market a portion of production into certain manganese market segments.  Boleo manganese carbonate is a versatile intermediate product and investigations have been conducted by Baja Mining into the production of higher value added manganese products.

· The production of manganese sulphate monohydrate has been investigated with test work being conducted by SGS Lakefield.  The production technology for manganese sulphate monohydrate is similar to zinc sulphate monohydrate.  The favourable location of the Boleo project to North and South American markets suggests that a portion of manganese carbonate production could be directed to manganese sulphate monohydrate production.

· The largest market for manganese is in the form of ore and ferroalloys used in steel making.  One of the manganese ferroalloys, electrolytic manganese metal, is marketed into steel making applications as well as other non ferrous industries.  Laboratory scale testwork by UBC has shown that it is possible to convert Boleo manganese carbonate into electrolytic manganese metal.  The production of electrolytic manganese metal requires large amounts of electric power and sustainable producers are generally located close to sources of low cost power, ore and labour.  Suitable locations for a refinery as well as the marketing and economics of producing manganese metal are currently being investigated.

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19.3 Economic Assessment Outputs

19.3.1 NPV & IRR Determinations

Table 19-9 contains NPV and IRR outputs from detailed calculations of the project economics at various metal prices.  Three cases are presented:

1) Case 1:

Project economics use the weighted average three-year trailing price (in accordance with SEC guidelines) of $2.91/lb for copper; $26.85/lb for cobalt; and $1,175/tonne for zinc sulphate monohydrate.

2) Case 2:

$2.25/lb for copper; $20/lb for cobalt; and $1,100/tonne for zinc sulphate monohydrate.

3) January 2010 Price Case:

Prices as of January 8, 2010 were $3.40/lb for copper; $21.17/lb for cobalt; and $1,000/tonne for zinc sulphate monohydrate.

Production and Operating Costs

Start-up of the process plant is scheduled for the second half of 2012, based on funding the project during the first half of 2010, in accordance with the following production schedule:

Table 19-6: Production summary

Production Summary
Parameter	Yrs 11	Yrs 2-7	Yrs 8-10	Yrs 11-13	Yrs 14-20	Yrs 21-23
Ore Treated (kt/y)	2,174	3,100	3,100	3,100	3,100	3,021
Grade:    % Cu	2.04	2.02	1.74	1.39	0.91	0.61
% Co	0.074	0.071	0.061	0.070	0.074	0.059
% Zn	0.40	0.47	0.52	0.58	0.62	0.81
% Mn	2.45	2.55	2.58	2.93	3.08	4.95
Production (t/y):
Copper	40,221	56,697	49,009	39,161	25,701	17,249
Cobalt	710	1,708	1,481	1,699	1,802	1,398
Zinc Sulphate Monohydrate	9,027	25,364	28,482	30,469	33,054	36,760

1. Note: year 1 is a ramp-up year and partial year production for cobalt and zinc sulphate monohydrate estimated at 8 months

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Table 19-7: Unit Operating costs and cash flow summary

Unit Operating Costs   (expressed in $/tonne of ore treated)
	Yrs 1	Yrs  2-7	Yrs  8-10	Yrs  11-13	Yrs  14-20	Yrs 21-23
Mining	15.36	14.91	11.89	7.90	8.62	7.47
Process	19.36	18.96	17.87	17.63	17.60	18.45
G&A, Sales	9.70	7.54	7.36	7.07	6.66	6.64
Total ($/t)	44.37	41.41	37.12	32.60	32.87	32.56
Cash cost1 $/lb	0.49	(0.04)	(0.07)	(0.43)	(0.80)	(0.78)
of copper
$000/yr
Cash Flow1	216,002	316,432	233,539	204,462	148,931	97,749
Cash Flow2	146,427	263,258	165,617	145,063	101,867	63,871
Cash Flow3	248,904	344,143	254,629	213,990	148,712	93,037

1Cash cost/lb of Cu is net of cobalt, zinc and sulphuric acid credits. Cash flows are after-tax, using SEC guideline prices of $2.91/lb Cu, $26.85/lb Co and $1,175/tonne ZnSO4·H2O.

2Cash flows are after-tax, using case 2 prices - $2.25/lb Cu, $20/lb Co and $1,100/tonne ZnSO4·H2O.

3Cash flows are after-tax, using January 2010 prices - $3.40/lb Cu, $21.17/lb Co and $1,000/tonne ZnSO4·H2O.

Project Economics

Project economics are presented for three cases:

Table 19-8: Economic analysis

	Case 1 Prices	Case 1 Prices	Jan '10 Prices
IRR – Pre-tax	28.7%	23.0%	31.2%
IRR – After tax	25.6%	20.4%	27.9%
NPV* @ 0%	$3,749	$2,547	$4,024
NPV @ 5%	$1,922	$1,230	$2,122
NPV @ 8%	$1,306	$815	$1,473

*Note: all NPVs are after-tax

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Sensitivities

The project is most sensitive to four key variables: the copper price, cobalt price, capital costs, and operating costs.  The sensitivity of the after-tax IRR and NPV (at an 8% discount rate) relative to the SEC guideline price case (without manganese) are shown in the table and figure below to indicate the effect of + or – 10% changes in the key variables.

Table 19-9: Sensitivity analysis

	After Tax IRR			NPV @ 8% (Millions)
	-10%	Case 1	+10%	-10%	Case 1	+10%
Copper price	23.2%	25.6%	27.9%	$1,111	$1,306	$1,505
Cobalt price	25.0%	25.6%	26.1%	$1,246	$1,306	$1,367
Capital cost	27.7%	25.6%	23.7%	$1,361	$1,306	$1,250
Operating cost	26.3%	25.6%	24.8%	$1,379	$1,306	$1,234

Figure 19-1 Sensitivity to Copper Price, Cobalt Price, Capital and Operating Costs

19.3.2 Annual Cash Flow Determinations

Table 19-10  contains cash flow outputs from the economic models on a year by year basis.

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Table 19-10:                23 Year Detailed Cash Flow – Using SEC Guideline pricing

		Construction costs  to complete	Y1	Y2	Y3	Y4	Y5	Y6	Y7
Throughput:	dt/a		2,174	3,100	3,100	3,100	3,100	3,100	3,100
Grades:	%Cu		2.04%	2.07%	2.18%	1.97%	2.06%	1.98%	1.85%
	%Co		0.074%	0.067%	0.068%	0.073%	0.075%	0.071%	0.070%
	%Zn		0.40%	0.35%	0.45%	0.52%	0.52%	0.47%	0.49%
	%Mn		2.45%	1.81%	2.35%	2.75%	2.96%	2.68%	2.75%
Recoveries:	Cu		90.65%	90.65%	90.65%	90.65%	90.65%	90.65%	90.65%
	Co		62.33%	72.75%	78.76%	78.76%	78.76%	78.76%	78.76%
	Zn		51.00%	59.55%	64.44%	64.44%	64.44%	64.44%	64.44%
Production:	Cu		40,221	58,030	61,206	55,497	57,924	55,676	51,848
(tonnes)	Co		710	1,505	1,662	1,786	1,839	1,739	1,715
	ZSM		9,027	17,955	24,421	28,451	28,284	25,941	27,133
	Excess Sulphuric acid		0	196,125	196,125	196,125	196,125	196,125	196,125
Revenue:	Cu		$259,796	$374,828	$395,341	$358,468	$374,142	$359,622	$334,897
(000's)	Co		$42,034	$89,104	$98,348	$105,713	$108,874	$102,941	$101,488
	Zn		$10,607	$21,097	$28,695	$33,430	$33,234	$30,481	$31,882
	Sulphuric acid		$0	$1,961	$1,961	$1,961	$1,961	$1,961	$1,961
	Total		$312,436	$486,991	$524,345	$499,572	$518,211	$495,005	$470,227
(000’s)
Total Op Cos			$96,435	$118,375	$130,094	$132,750	$132,948	$130,200	$125,899
Initial Capital		$876,278	$6,095
Sustaining Capital			$20,477	$199	$2,616	$5,339	$9,125	$12,518	$9,876
Reclamation funding			$0	$775	$1,550	$1,550	$1,550	$1,550	$1,550
Working Capital			$7,101	$0	$0	$0	$0	$0	$0
Income Taxes			$0	$77	$69	$19,042	$101,459	$105,700	$99,145
Net Cash Flow $000’s		($876,278)	$182,329	$367,565	$390,016	$340,891	$273,129	$245,037	$233,757
		Y8	Y9	Y10	Y11	Y12	Y13	Y14	Y15
Throughput:	dt/a	3,100	3,100	3,100	3,100	3,100	3,100	3,100	3,100
Grades:	%Cu	1.77%	1.74%	1.72%	1.58%	1.39%	1.21%	1.13%	0.97%
	%Co	0.063%	0.055%	0.064%	0.065%	0.065%	0.079%	0.090%	0.090%
	%Zn	0.50%	0.51%	0.55%	0.49%	0.53%	0.73%	0.75%	0.69%
	%Mn	2.70%	2.39%	2.66%	2.58%	2.71%	3.52%	3.62%	3.60%
Recoveries:	Cu	90.65%	90.65%	90.65%	90.65%	90.65%	90.65%	90.65%	90.65%
	Co	78.76%	78.76%	78.76%	78.76%	78.76%	78.76%	78.76%	78.76%
	Zn	64.44%	64.44%	64.44%	64.44%	64.44%	57.31%	57.49%	62.25%
Production:	Cu	49,716	48,918	48,392	44,392	39,065	34,026	31,631	27,158
(tonnes)	Co	1,540	1,346	1,557	1,595	1,584	1,918	2,189	2,199
	ZSM	27,392	28,089	29,965	26,600	29,096	35,712	36,776	36,776
	Excess Sulphuric acid	196,125	196,125	196,125	196,125	196,125	196,125	196,125	196,125
Revenue:	Cu	$321,122	$315,972	$312,572	$286,733	$252,328	$219,780	$204,309	$175,421
(000's)	Co	$91,156	$79,666	$92,143	$94,410	$93,738	$113,554	$129,591	$130,183
	Zn	$32,186	$33,005	$35,209	$31,255	$34,188	$41,961	$43,212	$43,212
	Sulphuric acid	$1,961	$1,961	$1,961	$1,961	$1,961	$1,961	$1,961	$1,961
	Total	$446,425	$430,603	$441,885	$414,359	$382,215	$377,256	$379,074	$350,777
(000’s)
Total Op Cos		$118,895	$115,100	$111,189	$102,502	$99,366	$101,288	$103,211	$104,165
Initial Capital
Sustaining Capital		$2,644	$1,829	$4,006	$1,309	$2,776	$2,926	$3,736	$2,956
Reclamation funding		$1,550	$1,550	$1,550	$1,550	$1,550	$1,550	$1,550	$1,550
Working Capital			$0	$0	$0	$0	$0	$0	$0
Income Taxes		$94,055	$91,113	$87,945	$91,668	$87,050	$78,571	$76,607	$76,380
Net Cash Flow $000’s		$229,280	$221,012	$237,195	$217,330	$191,473	$192,922	$193,969	$165,726

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		Y16	Y17	Y18	Y19	Y20	Y21	Y22	Y23
Throughput:	dt/a	.	3,100	3,100	3,100	3,100	3,100	3,100	2,863
Grades:	%Cu	0.78%	0.83%	0.91%	0.91%	0.88%	0.75%	0.57%	0.51%
	%Co	0.081%	0.077%	0.070%	0.057%	0.052%	0.054%	0.061%	0.062%
	%Zn	0.69%	0.61%	0.52%	0.50%	0.58%	0.71%	0.84%	0.87%
	%Mn	4.02%	3.26%	2.11%	2.18%	2.79%	3.94%	5.20%	5.77%
Recoveries:	Cu	90.65%	90.65%	90.65%	90.65%	90.65%	90.65%	90.65%	101.67%
	Co	78.76%	78.76%	78.76%	78.76%	78.76%	78.76%	78.76%	78.76%
	Zn	62.24%	64.44%	64.44%	64.44%	64.44%	59.64%	51.76%	53.94%
Production:	Cu	22,019	23,439	25,486	25,510	24,667	21,015	15,881	14,850
(tonnes)	Co	1,969	1,875	1,718	1,400	1,266	1,325	1,478	1,391
	ZSM	36,657	33,464	28,681	27,310	31,716	36,279	37,000	37,000
	Excess Sulphuric acid	196,125	196,125	196,125	196,125	196,125	196,125	196,125	196,125
Revenue:	Cu	$142,227	$151,399	$164,616	$164,771	$159,326	$135,740	$102,579	$95,918
(000's)	Co	$116,535	$110,976	$101,667	$82,862	$74,927	$78,410	$87,490	$82,362
	Zn	$43,072	$39,321	$33,700	$32,090	$37,267	$42,628	$43,475	$43,475
	Sulphuric acid	$1,961	$1,961	$1,961	$1,961	$1,961	$1,961	$1,961	$1,961
	Total	$303,796	$303,657	$301,944	$281,684	$273,482	$258,739	$235,506	$223,716
(000’s)
Total Op Cos		$103,448	$102,988	$100,899	$99,217	$99,459	$101,161	$100,259	$93,620
Initial Capital
Sustaining Capital		$2,207	$3,003	$3,111	$2,835	$2,712	$1,516	$638	$0
Reclamation funding		$1,550	$1,550	$1,550	$1,550	$1,550	$1,550	$1,550	$2,925
Working Capital			$0	$0	$0	$0	$0	$0	($7,101)
Income Taxes		$68,379	$55,607	$55,503	$55,582	$50,448	$48,113	$43,800	$37,761
Net Cash Flow 000’s		$128,212	$140,509	$140,881	$122,499	$119,314	$106,399	$89,258	$96,511

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20 CONCLUSIONS

20.1 Geology & Mineral Resource Modelling

A Measured and Indicated Resource has been defined, totalling:

·   265 Mt at 1.50% CuEq. at a cut-off grade of 0.5% CuEq.

·   172 Mt at 1.91% CuEq. at a cut-off grade of 1.0% CuEq.

In addition an Inferred Resource has been defined, totalling:

·   160 Mt at 1.15% CuEq. at a cut-off grade of 0.5% CuEq.

·   78 Mt at 1.61% CuEq. at a cut-off grade of 1.0% CuEq.

20.2 Metallurgy & Process Design

After conducting extensive testwork of both a bench scale and a pilot scale nature, the testwork outputs have been studied, assimilated and incorporated into the process design of the metallurgical plant

Boleo ore has been successfully processed using the proposed metallurgical treatment route in two continuous pilot plant programs to leach, separate and recover pay metals in final commercial form.

Design criteria have been confirmed for the purposes of the metallurgical plant design.  Data for the purposes of formulating process guarantees for the Boleo process plant have been gathered.

Detailed design of the Boleo process plant has been progressed retaining the flowsheet proposed in the DFS.  A number of improvements have been included that optimise process equipment selection, reagent selection and operating parameters.  These improvements have been tested and are expected to result in a more robust plant with improved constructability, operability and maintainability.

20.3 Mining

The potential for surface and underground mining were assessed for the Wardrop resource model.  A series of underground mining operations, supported by small open-cut surface mines, delivering targeted high-grade copper-cobalt-zinc- manganese ore (0.5% to 2.5% Cu) to the process facility was selected as the best alternative for the scale of operation envisaged by Baja Mining Corp.

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The mine plan has scheduled 23 years of underground and surface reserves.

A suitable limestone source was located on the property and will be mined by surface methods to provide calcium carbonate for plant process needs.

A tailings dam and associated tailings disposal facility has been designed with the capacity to support the LOM projections.

20.4 Environmental

An environmental plan was submitted to Mexican Federal authorities and the basic Environmental Impact Manifest (EIM) permit was approved and issued to the Company in December, 2006.

·   This allows the company to begin submitting additional specific permitting requests for construction and operational activities in accordance with the EIM provisions.  This work is underway.

20.5 Economic Assessment

Financial modelling based on the current mine schedule indicates that the project is attractive using all three case metal prices while recovering copper, cobalt and zinc.

Modelling of the case 1 SEC guideline metal prices over a projected 23 year mine life shows that the project could generate a cumulative after tax cash flow of US$3.749 billion with a discounted present value of US$1,306 million at an 8% discount rate, or US$1,922 million at a 5% discount rate.

The cash cost of production of copper, net of by-product prices, is below zero using all three metal case pricing.

·   The addition of manganese production could add extra value to the project.

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21 RECOMMENDATIONS

21.1 Geology & Mineral Resource Modelling

At this stage of the project development the following recommendations are made with respect to Geology and Mineral Resource Modelling:

· Resource definition for the Soledad limestone should be upgraded to measured and indicated status.  This will require further drilling, sampling, assaying and modelling.

· Resource definition for Manto 4 should be upgraded to measured and indicated status.  This will require further drilling, sampling, assaying and modelling.

21.2 Metallurgy & Process Design

At this stage of the project development the following recommendations are made with respect to Metallurgy and Process Design:

· Further development is undertaken of geological, mining and stockpiling processes to ensure that the optimum ore quality is delivered to the process refinery.  This includes further developing methods of assessing the impact of variation in ore quality on metallurgical performance.

· Detailed design of the process needs to be progressed to fully define process control and metallurgical accounting systems that will support measurement and verification of process performance guarantees.

· Further studies are undertaken to define requirements for manganese recovery and processes for producing value added products.

21.3 Mining

At this stage of the project development the following recommendations are made with respect to Mining:

· Geotechnical investigation needs to be conducted in the first mines in Mantos 1, 2 and 4.

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21.4 Economic Assessment

At this stage of the project development the following recommendations are made with respect to the economic assessment.

· Further studies are undertaken to improve the confidence in and accuracy of the economics of recovering manganese from Boleo ores and producing added value manganese products.

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22 REFERENCES

Agapito Associates, Inc. (2006), “Preliminary Geotechnical Performance Study for Underground Mining of El Boleo Copper Cobalt Project, Texcoco Test Mine Including Operations Observations and Recommendations”, draft report to Baja Mining Corp, July 2006.

Agapito Associates, Inc. (2007), “Geotechnical Evaluation for Underground Mine Design,” draft report to Baja Mining Corp, February 2006.

Bailes, R.J., Christoffersen, J.E., Escandon, F., and Peatfield, G.R. (2001).  “Sediment–Hosted Deposits Of the Boleo Copper-Cobalt-Zinc District, Baja California Sur, Mexico”. In Society of Economic Geologists, SP8, p. 291-306, 2001

Brown, J., Todd, I., SGS Mineral Services (2010), “ An Investigation into Sulphur Dioxide Adsorption during the Reductive Leach of El Boleo Ore”,  Jan 2010

Emmett Process Consulting Company (2008), Report on Neutralisation and Thickening Testwork, April 2008.

Emmett Process Consulting Company, (2008), “Raffinate Iron Removal Test Results”, July 2008

Felix, F.C., (1996).  “Specific Gravity.”  Unpublished Internal Company Activity Report, November 1996

FLSmidth Mineral (2008).  “Report of Investigation Thickener Sizing Confirmation Testing”, April 2008.

Hazen Research Inc.(2008),  “Thermal Properties of Leach Liquor”, July 2008.

Hellman &Schofield Pt Ltd (2005),  “Resource Estimate Study, The Boleo Copper - Cobalt - Zinc Deposit, Baja California, Mexico”, April 2005.

Morton, N., Prout, S., Marion, R., SGS Lakefield (2009), “An Investigation by High Definition Mineralogy into the Mineralogical Characteristics of Three Boleo Limestone Samples.”, March 2009.

Minera y Metalurgica Del Boleo S.A. de C.V. (2009),  “Prill & Soda Ash Freight, Operating and Capital Cost Options 6-5-09 Draft 2”, June 2009.

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Minera y Metalurgica Del Boleo S.A. de C.V. (2009),  “Trade-off Study - Ore Haulage Systems for El Boleo Project – Overland Conveyors and Trucking”, June 2009”

Peatfield, G.R., & Smee, B. 1997, (1997).  “Assay Quality Control Report for the Boleo Copper-Cobalt Project, Baja California Sur, Mexico.” Unpublished Internal Company Report, January 1997.

Peatfield, G.R. (1997),  “Update to the Assay Quality Control Report for the Boleo Copper-Cobalt Project, Baja California Sur, Mexico.”  Unpublished Internal Company Report, February 1997

Peatfield, G.R. (1998).  “Analytical Quality Control at the Boleo Copper-Cobalt-Zinc Project, Baja California Sur, Mexico.” Pathways 98, Cordilleran Roundup and Exploration Methods 98, Pathways to Discovery. Short Course #4 Analysis and Quality Control in Mineral Exploration, January 25-26 1998.

Sawlan, M.G. and Smith, J.G., (1984).  “Petrologic Characteristics, Age and Tectonic Setting of Neogene Volcanic Rocks in Northern Baja California Sur, Mexico”, In Frizzel, V.A., ed., Geology of the Baja California Peninsula:Society of Economic Paleontologists and Mineralogists, Pacific Section (Los Angeles), San Diego, California, Symposium Proceedings, p. 237-251, April 18-21, 1984.

Schmidt, E.K., (1975).  “Plate Tectonics, Volcanic Petrology and Ore Formation of the Santa Rosalía  Area, Baja California, Mexico.” Unpublished M.Sc. thesis, Tucson, University of Arizona, 191 p., 1975

Stubens, T.C., Wardrup Engineering Inc. (2009) “Mineral Resource Update on the El Boleo Property, Baja California Sur, Mexico.”, November 2009.

Verret, F-O., Molnar, R., SGS Lakefield (2008).  “An Investigation into the Washability Characteristics of Samples from the Boleo Deposit.”, June 2008

Roggekamp, M., Todd, I., SGS Lakefield, (2008).  “Seven Samples Program on El Boleo Ore Samples Report”, July 2008

Thompson, M. and Howarth, R.J., (1978).  A New Approach to the Estimation of Analytical Precision.  Elsevier Scientific Publishing Company, Journal of Geochemical Exploration, Vol. 9, pp 23-30, 1978

Wilson, I.F., and Rocha, V.S., (1955).  “Geology and Mineral Deposits of the Boleo Copper District. Baja California, Mexico.”  U.S. Geological Survey Professional Paper 273, 134p. 1955.

Wright, F.D. (1997).  “Pre-feasibility Study Final Report: Volume 1, Final Report; Volume 2, Cost Report; Boleo Project.” Unpublished report prepared for Minera Curator, S.A. de C.V. (September 1997)

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Xie, F., Dreisinger, D., University of British Columbia, (2009).  “DSX® Report #1 - The Optimization of the Boleo DSX® Solvent Extraction System”, August 2009

Xie, F., Dreisinger, D., University of British Columbia, (2009).  “DSX® Report #2 - The Optimization of the Boleo DSX® Solvent Extraction System. A Study of the Variation of the LIX 63 Concentration with Constant Versatic 10 Concentration” September 2009

Yeo, W.J.A. Wyche, J., Holmes, M., Norton, E., Hunter, D., Ross, T., Britton, S., Bosworth, G.,  (2007),  “El Boleo Project, Feasibility Study Summary Report”, Filed on Sedar, July 2007

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23 DATE AND SIGNATURE PAGE

This report titled El Boleo (Boleo) Technical Report Update is effective as at March 2, 2010 and was prepared by the following authors:

Dated this 12th day of March, 2010

(Signed) “Michael F. Shaw”

_______________________________

Michael F. Shaw, P.E.

(Signed) “David Dreisinger”

_______________________________

David Dreisinger, PhD, P.Eng

(Signed) “Thomas Gluck”

_______________________________

Thomas Gluck, PhD, FSAIMM, Pr Eng.

(Signed) “Scott Britton”

_______________________________

Scott Britton, P.E.

(Signed) “Terry Hodson”

_______________________________

Terry Hodson, P.Geo

(Signed and sealed) “Timothy Ross”

_______________________________

Timothy Ross, P.E.

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