Exhibit 99.1

Technical Report and Preliminary Economic Assessment for the

Integrated Mother Lode and North Bullfrog Projects,

BULLFROG MINING DISTRICT,

NYE COUNTY, NEVADA

Dated: November 1, 2018

AMENDED: NOVEMBER 8, 2018

Effective Date: September 18, 2018

PREPARED FOR:

CORVUS GOLD INC.

by

QUALIFIED PERSONS:

Scott E. Wilson, C.P.G., SME Resource Development Associates, Inc. 9137 S. Ridgeline Blvd., Ste. 140 Highlands Ranch, CO 80129 720-348-1646	Michael Cole, Mining Eng., SME 381 Washington Ave. SW Roanoke, VA 24016 720-302-3140
Christopher L. Easton, Consulting Metallurgist Easton Process Consulting, Inc. 9532 S. Desert Willow Way Highlands Ranch, CO 80129 303-250-6456	Richard Delong, Environmental Geologist, SME EM Strategies 1650 Meadow Wood Lane Reno, NV 89502 775-826-8822

Date and Signature PagE

Corvus Gold Inc.: Technical Report and Preliminary Economic Assessment for the Integrated Mother Lode and North Bullfrog Projects, Bullfrog Mining District, Nye County, Nevada.

Technical Report Effective Date: September 18, 2018

Dated this 1^st day of November 2018

(signed/sealed) Scott E. Wilson

Scott E. Wilson, SME-RM, CPG

Geologist

(signed) Michael Cole

Michael Cole, SME-RM

Mining Engineer

(signed) Christopher L. Easton

Christopher Easton, MMSA-QP

Consulting Metallurgist

signed) Richard Delong

Richard Delong, MMSA-QP

Environmental Geologist

AUTHOR’S CERTIFICATE

Scott E. Wilson

I, Scott E. Wilson, CPG, SME-RM, of Highlands Ranch, Colorado, as the lead author of the technical report entitled “Technical Report and Preliminary Economic Assessment for the Integrated Mother Lode and North Bullfrog Projects, Bullfrog Mining District, Nye County, Nevada” (the “Technical Report”) with an effective date of September 18, 2018, prepared for Corvus Gold, Inc. (the “Issuer”), do hereby certify:

1. I am currently employed as President by Resource Development Associates, Inc., 10262 Willowbridge Way, Highlands Ranch, Colorado USA 80126.
2. I graduated with a Bachelor of Arts degree in Geology from the California State University, Sacramento in 1989.
3. I am a Certified Professional Geologist and member of the American Institute of Professional Geologists (CPG #10965) and a Registered Member (#4025107) of the Society for Mining, Metallurgy and Exploration, Inc.
4. I have been employed as both a geologist and a mining engineer continuously for a total of 29 years. My experience included resource estimation, mine planning, geological modeling, geostatistical evaluations, project development, and authorship of numerous technical reports and preliminary economic assessments of various projects throughout North America, South America and Europe. I have employed and mentored mining engineers and geologists continuously since 2003.
5. I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.
6. I made personal inspections of the North Bullfrog Project on January 30 and 31, 2012, March 24, 2014, November 2 and 3, 2015, on June 6-8, 2017 and most recently on January 16, 2018.
7. I am responsible for Sections 1 through 12, Section 14, Sections 18 and 19, and Sections 21 through 27 of the Technical Report.
8. I am independent of the Issuer as independence is described in Section 1.5 of NI 43-101.
9. Prior to being retained by the Issuer, I have not had prior involvement with the property that is the subject of the Technical Report.
10. I have read NI 43-101 and Form 43-101F1, and this Technical Report was prepared in compliance with NI 43-101.
11. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the portions of the Technical Report for which I am responsible not misleading.

Dated: November 8, 2018

(signed/sealed) Scott Wilson

Scott E. Wilson, CPG, SME-RM

AUTHOR’S CERTIFICATE

Michael R. Cole

I, Michael R. Cole, Mining Engineer, SME Registered Member of Roanoke, Virginia, as an author of the technical report entitled “Technical Report and Preliminary Economic Assessment for the Integrated Mother Lode and North Bullfrog Projects, Bullfrog Mining District, Nye County, Nevada” (the “Technical Report”) with an effective date of September 18, 2018 prepared for Corvus Gold, Inc. (the “Issuer”), do hereby certify:

1. I am currently employed and an independent consultant and reside at 381 Washington Ave SW, Roanoke, VA 24016.
2. I graduated with a Bachelor of Science degree in Mining and Minerals Engineering from the Virginia Polytechnic Institute and State University in 2005.
3. I am a Registered Member (#4130807) of the Society for Mining, Metallurgy and Exploration, Inc.
4. I have been employed as mining engineer continuously for a total of 13 years. My experiences include resource estimation, mine planning, cash-flow analysis and pit optimizations for numerous technical reports and preliminary economic assessments of various projects throughout North America and South America. I have been involved with mine supervision in several large pits, mining equipment purchases and mine construction projects. I have prepared operational mine plans and budgets that have performed as designed. I have held positions of Senior Mining Engineer, Chief Mining Engineer, Mine Planning Manager and Mine Manager.
5. I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.
6. I made a personal inspection of the North Bullfrog and Mother Lode Project sites on September 11, 2017.
7. I am responsible for sections 15 and 16 of the Technical Report.
8. I am independent of the Issuer as independence is described in Section 1.5 of NI 43-101.
9. Prior to being retained by the Issuer, I have not had prior involvement with the property that is the subject of the Technical Report.
10. I have read NI 43-101, and the portions of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101.
11. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contains all scientific and technical information that is required to be disclosed to make the portions of the Technical Report for which I am responsible not misleading.

Dated: November 8, 2018

(signed) Michael Cole

Michael R. Cole, SME-RM

AUTHOR’S CERTIFICATE

Christopher Easton

I, Christopher L. Easton, Consulting Metallurgist, MMSA-QP, of Highlands Ranch, Colorado, as an author of the technical report entitled “Technical Report and Preliminary Economic Assessment for the Integrated Mother Lode and North Bullfrog Projects, Bullfrog Mining District, Nye County, Nevada” (the “Technical Report”) with an effective date of September prepared for Corvus Gold, Inc. (the “Issuer”), do hereby certify:

1. I am a Consulting Metallurgist and President at Easton Process Consulting, Inc., and reside at 9532 S. Desert Willow Way, Highlands Ranch, CO 80129.

2. I am a graduate of the University of Wyoming with a degree in Chemical Engineering.

3. I am Qualified Person in good standing, of the Mining and Metallurgical Society of America (MMSA).

4. I have worked in the Mineral Processing Industry for a total of 30 years after attending the University of Wyoming. During this time, I have held positions as Plant Metallurgist, Lead Process Engineer, Plant Superintendent, Relieving Metallurgical Manager and Consulting Metallurgist. I have been a practicing consulting engineer since 2002.

5. I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of education, past relevant work experience and affiliation with a professional association (as defined in NI 43-101), I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

6. I made a personal inspection of the North Bullfrog Project on April 22 and 23, 2017 and of the Mother Lode Project on February 17, 2018.

7. I am responsible for the preparation of Sections 13 and 17 of the Technical Report.

8. I am independent of the Issuer as independence is described in Section 1.5 of NI 43-101.

9. Prior to being retained by the Issuer, I have not had prior involvement with the property that is the subject of the Technical Report.

10. I have read NI 43-101, and the portions of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101.

11. As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the portions of the Technical Report for which I am responsible not misleading.

Dated November 8, 2018

(signed) Christopher L. Easton

Christopher L. Easton, MMSA-QP

AUTHOR’S CERTIFICATE

Richard Delong

 
I, Richard Delong, Environmental Geologist, RM-SME, of Reno, Nevada as an author of the technical report entitled “Technical Report and Preliminary Economic Assessment for the Integrated Mother Lode and North Bullfrog Projects, Bullfrog Mining District, Nye County, Nevada” (the “Technical Report”) with an effective date of September 18, 2018 prepared for Corvus Gold, Inc. (the “Issuer”), do hereby certify:

1. I am an Environmental Geologist and President at EM Strategies, located at1650 Meadow Wood Lane, Reno, Nevada 89502.

2. I am a graduate of the University of Idaho with a master degree in Geology and Resource Management.

3. I am Qualified Person (MMSA-01471QP), in good standing, of the Mining and Metallurgical Society of America (MMSA).

4. I have worked in the Mineral Industry for a total of 32 years after attending the University.

5. I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of education, past relevant work experience and affiliation with a professional association (as defined in NI 43-101), I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

6. I made a personal inspection of the North Bullfrog Project and Mother Lode Project on April 3, 2017.

7. I am responsible for the preparation of Section 20 of the Technical Report.

8. I am independent of the Issuer as independence is described in Section 1.5 of NI 43-101.

9. Prior to being retained by the Issuer, I have not had prior involvement with the property that is the subject of the Technical Report.

10. I have read NI 43-101, and the portions of the Technical Report for which I am responsible have

been prepared in compliance with NI 43-101.

11. As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the portions of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the portions of the Technical Report for which I am responsible not misleading.

Dated November 8, 2018

(signed) Richard Delong

Richard Delong, QP-MMSA

Table of Contents

1	SUMMARY	1
2	INTRODUCTION	8
3	RELIANCE ON OTHER EXPERTS	11
4	PROPERTY DESCRIPTION AND LOCATION	12
5	NB-MLP ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	18
6	HISTORY	19
7	GEOLOGICAL SETTING AND MINERALIZATION	23
8	DEPOSIT TYPES	59
9	EXPLORATION	60
10	DRILLING	70
11	SAMPLE PREPARATION, ANALYSIS AND SECURITY	90
12	DATA VERIFICATION	109
13	MINERAL PROCESSING AND METALLURGICAL TESTING	117
14	MINERAL RESOURCE ESTIMATES	161
15	MINERAL RESERVES	218
16	MINING METHODS	219
17	RECOVERY METHODS	229
18	INFRASTRUCTURE	235
19	MARKET STUDIES AND CONTRACTS	240
20	ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT	241
21	CAPITAL AND OPERATION COSTS	244
22	ECONOMIC ANALYSIS	248
23	ADJACENT PROPERTIES	256
24	OTHER RELEVANT DATA AND INFORMATION	257
25	INTERPRETATIONS AND CONCLUSIONS	258
26	RECOMMENDATIONS	260
27	REFERENCES	261

Tables

Table 1-1 Mother Lode Mineral Resource – Pit Constrained Measured, Indicated and Inferred Estimation at a Gold Price of $1,250 Per Ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)	1
Table 1-2 North Bullfrog Project Pit Constrained Mineral Resource Estimate at a Gold Price of $1,250 Per Ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)	2
Table 1-3 NBP Sierra Blanca/YellowJacket Phase I Mineral Resources Pit Constrained at a Gold Price of $1,250/Oz Au (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)	2
Table 1-4 NBP Mayflower Phase I Resources Pit Constrained at a Gold Price of $1,250 Per Ounce (Cut-off >=0.10 g/t Au)	2
Table 1-5 NBP Sierra Blanca Phase II Resources Pit Constrained at a Gold Price of $1,250 Per Ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)	3
Table 1-6 NBP Jolly Jane Pit Constrained at a gold price of $1,250 per ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)	3
Table 1-7 Mother Lode and North Bullfrog, Pit-constrained, Measured, Indicated and Inferred Mineral Resources at Gold Selling Price of USD $1,250 per Ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)	3
Table 1-8 NB-MLP Production Years 1 through 4 Highlights	4
Table 1-9  NB-MLP Life-of-Mine Project Economic Highlights	4
Table 4-1 Summary of Federal Mining Claims at MLP	12
Table 4-2 Summary of Lease Obligations at NBP (All Funds USD)	14
Table 4-3 Summary of the Historic Patented Claims in the Nine NBP Private Land Lease Agreements	16
Table 4-4 Summary of the Original Terms for the Mayflower/Greenspun Group Lease	17
Table 4-5 Summary of the NBP Unpatented Lode Mining Claims	17
Table 6-1 Summary of Companies That Explored NBP	21
Table 7-1  Summary of the Stratigraphy of the Mother Lode Property	27
Table 7-2 Overview of the Stratigraphy of the North Bullfrog Hills	34
Table 7-3 Summary of Zircon Dates from North Bullfrog Hills Volcanic Complex	40
Table 7-4 Adularia and Alunite Ar-Ar Age Determinations	41
Table 10-1 Drill Intercepts for September 2017 to July 2018 Drilling at Mother Lode Core and RC Holes	71
Table 10-2 Drill Intercepts for March-April 2017 NBP Drilling	86
Table 12-1 North Bullfrog Data Verification Samples (Scott Wilson - 2017)	109
Table 12-2 Mother Lode Data Verification Samples (Scott Wilson - 2018)	110
Table 12-3 Comparison of Statistical Indices for Combined, Historic and Corvus Drilling Data	112
Table 12-4 Comparison of Montonic Decay Patterns in Mother Lode Grade Data of Corvus RC and Core Drilling	115
Table 12-5 Distribution of Local Maxima with Greater Than Process Au Cut-off Grade	116
Table 13-1 2017- MLP Pit Sample – Flotation Test Results.	118
Table 13-2 Mother Lode – 2017 Pit Sample – Head Analysis.	120
Table 13-3 2018 Composite Head Analysis	123

Table 13-4 Diagnostic Leach Results	125
Table 13-5 Comminution Parameters	127
Table 13-6 Flotation Tests	128
Table 13-7 Flotation Concentrate Regrind and Cyanidation	129
Table 13-8 Flotation Concentrate POX/Cyanidation Tests	130
Table 13-9 Rougher Tailings Acid Neutralization Potential	130
Table 13-10 2008-2013 Laboratory References	132
Table 13-11 Jolly Jane Bottle Roll and Column Tests 2008-2013	133
Table 13-12 Mayflower Bottle Roll and Column Tests 2008-2013	133
Table 13-13 Savage Valley Bottle Roll and Column Tests 2008-2013	134
Table 13-14 Sierra Blanca Bottle Roll and Column Tests 2008-2013	135
Table 13-15 YellowJacket - Initial Sulphide Bottle Rolls	135
Table 13-16 Summary Metallurgical Results, Bottle Roll Tests, YellowJacket Drill Core Composites, 80%-75 µm Feed Size	136
Table 13-17 Summary of Bottle Roll Tests, YellowJacket YJ PQ Drill Core Composites.	137
Table 13-18 Summary Metallurgical Results, Gold Recovery from Column Percolation Leach Tests, YellowJacket YJ PQ Drill Core Composites (80% -6.3mm and 80% -19mm)	138
Table 13-19 Summary Metallurgical Results, Silver Recovery from Column Percolation Leach Tests, YellowJacket YJ PQ Drill Core Composites (80% -6.3mm and 80% -19mm)	138
Table 13-20 E-GRG Test Results for Gold Recovery from the YJ PQ composites	139
Table 13-21 E-GRG Test Results for Silver Recovery from the YJ PQ composites	139
Table 13-22 Gold Recovered from Gravity Concentrate and CN Leach of Gravity Tail, YJ PQ Composites	140
Table 13-23 Silver Recovered from Gravity Concentrate and CN Leach of Gravity Tail, YJ PQ Composites	140
Table 13-24 Bottle Roll Tests from YJ JV Composites, Gold Recovery at Various Feed Sizes	141
Table 13-25 Bottle Roll Tests from YJ JV Composites, Silver Recovery at Various Feed Sizes	141
Table 13-26 E-GRG Test Results for the YJ JV Composites, Gold and Silver Recovery	141
Table 13-27 Gold and Silver Recoveries from YJ JV Composites with Intense CN leaching of Gravity Concentrate and CN Leach of Tail	142
Table 13-28 Summary of Comminution Test Work on Sierra Blanca, Jolly Jane, and Mayflower Bulk Materials and on YellowJacket PQ3 Core Materials	143
Table 13-29 Vat Leach Test Measurements of Gold Recovery on Large Particles Sizes from Mayflower and Sierra Blanca Dump Materials	144
Table 13-30 Vat Leach Test Measurements of Silver Recovery on Large Particles Sizes from Mayflower and Sierra Blanca Dump Materials	145
Table 13-31 Drill hole Composite Classification	146
Table 13-32 Drill hole Composite Analysis for the North Bullfrog Samples 1 of 2	147
Table 13-33 Drill hole Composite Analysis for the North Bullfrog Samples 2 of 2	148
Table 13-34 Composites for Lithology Flotation	149
Table 13-35 Sierra Blanca Tuff – Flotation Reagents	150

Table 13-36 Sierra Blanca Tuff – Flotation Results	150
Table 13-37 Dacite – Flotation Reagents	150
Table 13-38 Upper and Savage Valley Dacite Flotation Results	151
Table 13-39 Pioneer and Rhyolite – Flotation Reagents	151
Table 13-40 Pioneer and Rhyolite – Flotation Test Results	151
Table 13-41 Flotation Uranium Deportment Test Results	152
Table 13-42 Phase I – AAO and Cyanidation Test Results	154
Table 13-43 Phase II – AAO and Cyanidation Test Results	155
Table 13-44 Phase III – AAO and Cyanidation Test Results	156
Table 13-45 North Bullfrog Assumed Oxide Mill Process Parameters	157
Table 13-46 Proportions of Metal and Estimated Gold and Silver Recovery	157
Table 13-47 Assumed Mother Lode and NBP Sulphide Mill Process Parameters	158
Table 13-48 AAO Metal and Estimated Gold and Silver Recoveries	158
Table 13-49 Assumed Heap Leach Metal Recoveries and Reagent Consumptions	159
Table 14-1 Mother Lode and North Bullfrog, Pit-constrained, Measured, Indicated and Inferred Mineral Resources at Gold Selling Price of USD $1,250 per Ounce (Effective date September 18, 2018)	161
Table 14-2 Gold Assay Statistics at Various Cutoff Grades	166
Table 14-3 Gold Assay Statistics at Various Cutoff Grades	167
Table 14-4 MLP Composite Statistics for Gold	168
Table 14-5 Mother Lode Block Model Framework NAD27 / (UTM Zone 11 North)	170
Table 14-6 High-Grade Estimation Parameters for MLP	171
Table 14-7 Gold Variogram Model Parameters	172
Table 14-8 Gold Ordinary Kriging Estimation Parameters in Vulcan® Format	173
Table 14-9 Oxide Indicator Variogram Parameters	178
Table 14-10 Gold Assay Statistics Sorted By Zone	182
Table 14-11 Silver Assay Statistics Sorted by Zone	182
Table 14-12 Capped Assays by Zone	186
Table 14-13 Capped Assay Statistics for Gold Sorted By Zone	186
Table 14-14 Capped Assay Statistics for Silver Sorted by Zone	187
Table 14-15 Composite Statistics for Gold Sorted By Zone	187
Table 14-16 Composite Statistics for Silver Sorted by Zone	187
Table 14-17 Sierra Blanca Block Model Framework	188
Table 14-18 Lithology Types and Corresponding Specific Gravity Values	188
Table 14-19 Specific Gravity by Stratigraphy	188
Table 14-20 YellowJacket Vein Estimation Parameters for Au and Ag	192
Table 14-21 Liberator Au and Ag Estimation Parameters	194
Table 14-22 Gold Variogram Model Parameters	195
Table 14-23 Gold Ordinary Kriging Estimation Parameters in Vulcan® Format	195
Table 14-24 Silver Variogram Model Parameters	196

Table 14-25 Silver Ordinary Kriging Estimation Parameters in Vulcan® Format	196
Table 14-26 Sulphur Variogram Model Parameters	197
Table 14-27 Sulphur Ordinary Kriging Estimation Parameters in Vulcan® Format	197
Table 14-28 Summary of Assay Statistics for Jolly Jane Mineralized Solid	199
Table 14-29 Summary of 5 M Composite Statistics for Mineralization Solid Jolly Jane	200
Table 14-30 Summary of Jolly Jane Gold and Silver Semi-variogram Parameters	200
Table 14-31 Specific Gravity Determinations for Tuff Units-Jolly Jane	201
Table 14-32 Summary of Kriging Search Parameters for Jolly Jane	203
Table 14-33 Summary of Assay Statistics for Mineralization Solid and Waste- Mayflower.	204
Table 14-34 Summary of Capped Assay Statistics for Mineralization Solid and Waste-Mayflower	205
Table 14-35 Summary of 5 M Composite Statistics for Mineralization Solid-Mayflower	206
Table 14-36 Summary of Semi-variogram Parameters-Mayflower	206
Table 14-37 Specific Gravities Sorted by Lithology - Mayflower	206
Table 14-38 Specific Gravities Sorted by Gold Grade - Mayflower	207
Table 14-39 Summary of Kriging Search Parameters - Mayflower	208
Table 14-40 Pit Constraining Parameters Used for the Corvus Gold September 18, 2018 Mineral Resource Update	211
Table 14-41 North Bullfrog Project Pit Constrained Mineral Resource Estimate. (QP: RDA, Scott Wilson; Effective September 18, 2018)	212
Table 14-42 Mother Lode Mineral Resources (QP: RDA, Scott Wilson; Effective September 18, 2018)	212
Table 14-43 Sierra Blanca Mineral Resources Phase I (QP: RDA, Scott Wilson; Effective September 18, 2018)	213
Table 14-44 Sierra Blanca Mineral Resources Phase II (QP: RDA, Scott Wilson; Effective September 18, 2018)	213
Table 14-45 Jolly Jane Mineral Resources-Phase II (QP: RDA, Scott Wilson; Effective September 18, 2018)	213
Table 14-46 Mayflower Resources Phase 1 (QP: RDA, Scott Wilson; Effective September 18, 2018)	214
Table 16-1 Parameters used in WhittleTM Analysis	220
Table 16-2 Mining production schedule for NB-MLP deposits developed with stockpile scheduler	222
Table 16-3 Process feed production schedule for individual NB-MLP deposits	223
Table 16-4 Total process feed production for NB-MLP mill and heap leach	224
Table 16-5 Classification of Mineral Resources in Individual NB-MLP Open Pit Production	224
Table 16-6 Production Equipment Fleet for the NB-MLP	227
Table 16-7 NB-MLP Mining Personnel Estimates	228
Table 17-1 Process Manpower	234
Table 21-1 NB-MLP – Initial Capital Cost Estimate	245
Table 21-2 NB-MLP Project – Initial Direct Capital Cost Estimate (including Owner’s Cost, EPCM &	245
Table 21-3 NB-MLP Project - Initial Indirect Capital Cost Estimates	245
Table 21-4 NBP – Sustaining Capital Cost Estimates and Remaining Contingency	246
Table 21-5 NB-MLP Project - Average Unit Operating Cost Assumptions	246
Table 22-1 Projected Key Performance Parameters from the NB-MLP Preliminary Economic Assessment (Constant $US, No Escalation, Constant $1,250 per Ounce Gold Price, after-Royalty and after-Tax)	249
Table 22-2 Summary of Physical Data from the NB-MLP PEA Production Schedule	250

Table 22-3 Projected LOM Unit Operating Cost and Capital Cost per Process Tonne and per Produced Au Ounce for the Project (Constant 2018 $US, No Escalation).	251
Table 22-4 Projected Annual Production and Cash Flow (after-Royalty and after-Tax) for the Integrated Project – Base Case (Gold Price $1,250; Silver Price $15.08)	252
Table 22-5 Projected Sensitivity of Net Present Value and Internal Rate of Return to Variation in Gold Price (after-Royalty and after-Tax)	253
Table 26-1 Proposed Work Program to Advance NB-MLP	260

Figures

Figure 2-1 Map Showing the Location of North Bullfrog and Mother Lode along with Local Gold Deposits Near Beatty	9
Figure 4-1 Property Map of the NB-MLP (UTM NAD 27 Zone 11)	12
Figure 5-1 Mayflower Ridge Looking to the Northwest	18
Figure 6-1 Bullfrog District Location Map	20
Figure 7-1 Regional Geology of the Greater Bullfrog District	24
Figure 7-2 Simplified Geologic Map of the Mother Lode Area	26
Figure 7-3 Cross Section 4084410N through the Mother Lode Deposit	31
Figure 7-4 Geologic Map of NBP Showing Target Areas, Resource Outlines and Property Outline	36
Figure 7-5 Map Units Legend for Figure 7-4	37
Figure 7-6 Geologic Cross Section from Savage Valley to Connection Illustrating the Overall Structural Style of the NBP	43
Figure 7-7 Cross-cutting Relationships Between Older Mineralization Types at Sierra Blanca	44
Figure 7-8 Geologic Map of the Jolly Jane Target Area	47
Figure 7-9 Cross Sections through Jolly Jane target area. See Figure 7-8 for section locations.	49
Figure 7-10 Geologic Map of the Greater Sierra Blanca-YellowJacket Area	51
Figure 7-11 Geologic Cross Section Looking North Through Savage Valley Illustrating the Style of Faulting	52
Figure 7-12 Simplified Geologic Cross Section through north Sierra Blanca-YellowJacket	53
Figure 7-13 Geologic Map of the YellowJacket Area Showing Major Faults and Drill Holes Related to Discovery of the High-grade Vein System	55
Figure 7-14 Geologic Map of the Mayflower Area Showing Underground Workings, Drill Holes and Cross Section Locations	57
Figure 7-15 Cross Sections Looking Northwest Through the Mayflower Deposit	58
Figure 9-1 MLP Target Location Map	61
Figure 11-1 Gold Assays of Preparation Duplicates for MLP	91
Figure 11-2 Silver Assays of Preparation Duplicates for MLP	91
Figure 11-3 Comparison of Field Duplicate Gold Assays for MLP	92
Figure 11-4 Comparison of Field Duplicate Silver Assays for MLP	92

Figure 14-1 Corvus Phase1 and Phase2 Drilling (red). Historic holes (black)	163
Figure 14-2 Cross Section 4084410N through Mother Lode Geological Interpretation	164
Figure 14-3 Cross Section 4084410N through Mother Lode 3D Leapfrog Geological Interpretation	165
Figure 14-4  Cross Section 4084410N through Mother Lode 3D Leapfrog Domain model	166
Figure 14-5 Lognormal Probability Plot Assays Capped at 8 g/t Au.  Low Grade and High-Grade Population Identified by Break in line at 0.9 g/t Au.	167
Figure 14-6 Mother Lode Cell Declustering Evaluation	168
Figure 14-7 Oxidation Model, Section 4084410, Looking North	169
Figure 14-8 Section 4084410 Showing High-Grade Indicator Blocks for the Project	170
Figure 14-9 Search Ellipsoids used for Grade Interpolation	172
Figure 14-10 Mother Lode Gold Variogram Model	173
Figure 14-11 Swath Plot Graphical Analysis by Northing	174
Figure 14-12 Swath Plot Graphical Analysis by Easting	174
Figure 14-13 Plan View of Mother Lode Resource Estimate at 970m elevation	175
Figure 14-14 Long-section Through Mother Lode Resource Model at 531,055 East UTM NAD27 Zone 11 North	175
Figure 14-15 Stratigraphy Blocks Defined within Sierra Blanca and YellowJacket (metres).	177
Figure 14-16 Stratigraphic Domains Modeled in Sierra Blanca (N 4,098,050 Looking North, metres).	178
Figure 14-17 Comparison of estimated oxidation to the 2014 oxide/sulphide surface (looking north).	179
Figure 14-18 Isometric View of Modeled YellowJacket Mineralization Grade Shell (Blue)-metres.	180
Figure 14-19 Isometric View of YellowJacket (Blue) and Liberator Veins (Green)-metres	181
Figure 14-20 Isometric View of YellowJacket (Blue) and Disseminated Swale (Yellow)-metres.	182
Figure 14-21 Au Lognormal Graph (within YellowJacket)	183
Figure 14-22 Ag Lognormal Graph (within YellowJacket)	184
Figure 14-23 Au Lognormal Graph (Disseminated)	185
Figure 14-24 Ag Lognormal Graph (Disseminated)	186
Figure 14-25 YellowJacket Contact Profile (Au)	189
Figure 14-26 YellowJacket Contact Profile (Ag)	189
Figure 14-27 Savage Formation vs Pioneer Formation Contact Profile (Au)	190
Figure 14-28 Pioneer Formation vs Sierra Blanca Formation Contact Profile (Au)	190
Figure 14-29 YellowJacket Declustering Au	191
Figure 14-30 YellowJacket Declustering Ag	191
Figure 14-31 YellowJacket Vein Estimation Bearing Changes (Elevation 1,150 m)-metres	193
Figure 14-32 YellowJacket Vein Estimation Plunge Changes (Northing 4098525)-metres	194
Figure 14-33 Gold Variogram Model-metres	195
Figure 14-34 Silver Variogram Model-metres	196
Figure 14-35 Sulphur Variogram Model-metres	197
Figure 14-36 North Bullfrog Project Swath Plot Section Line-metres	198
Figure 14-37 Swath Plot Graphical Analysis-metres	198

Figure 14-38 Isometric View of Jolly Jane Looking Northeasterly Showing Mineralization Solid in Red, Drill holes in Green and Surface Topography in Grey.	199
Figure 14-39 Mayflower Model Looking NW with the Mineralization Solid in Red and Topography in Grey.	204
Figure 14-40 Isometric View Looking W of the Mayflower Geologic Solid in Red with Surface Topography Shown in Grey.	204
Figure 14-41 Lognormal Cumulative Frequency Plot for Au within the Mineralization Solid-Mayflower.	205
Figure 14-42 Isometric View Looking NNW Showing Block Model in White and Drill holes in Magenta-Mayflower.	207
Figure 14-43 Mineral Classification at Mother Lode Section 4084410N	209
Figure 14-44 Kriging Variance vs. Distance, Sierra Blanca Classification-metres	210
Figure 14-45 Long Section through Sierra Blanca/YellowJacket Mineral Resource Model-metres	215
Figure 14-46 Cross Section across Sierra Blanca/YellowJacket Looking North-metres.	216
Figure 14-47 Cross Section through Mayflower Deposit-metres.	216
Figure 14-48 Long Section through Mayflower-metres.	217
Figure 14-49 Cross Section through Jolly Jane-metres.	217
Figure 16-1 Conceptual layout of the Mother Lode site showing open pit, TSF, heap leach pad, process facilities and waste storage area	226
Figure 16-2 Conceptual layout of the North Bullfrog site showing open pits, heap leach pad and waste storage	226
Figure 17-1 Block Flow Diagram for Integrated Processing of NB-MLP Mineralized Materials.	233
Figure 18-1 Map showing the extent of NB-MLP Property Boundary and locations of mining resources, and infrastructure	235
Figure 22-1 Estimated Annual Gold and Silver Production from NB-MLP (81% Measured and Indicated Resources, 19% Inferred Resources)	250
Figure 22-2 Sensitivity of Estimated NPV @ 5% (after-Royalty and after-Tax) for Changes in Cost, Gold Recovery or Gold Price as a Percent of the Base Case at a Gold Price of $1,250 per Ounce, Gold:Silver Price Ratio of 82.9, 78% Gold Recovery and Cost as Defined in Table 22-3	253
Figure 22-3 Sensitivity of Estimated IRR (after-Royalty and after-Tax) for Changes in Cost, Gold Recovery or Gold Price as a Percent of the Base Case at a Gold Price of $1,250 per Ounce, Gold:Silver Price Ratio 0f 82.9, 78% Gold Recovery and Cost as Defined in Table 22-3.	254

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 1

1 Summary

This technical report entitled “Technical Report and Preliminary Economic Assessment for the Integrated Mother Lode and North Bullfrog Projects, Bullfrog Mining District, Nye County, Nevada” (the “Technical Report”) summarizes technical information gathered on the Mother Lode Property (“ML” or “MLP”) and the North Bullfrog Property (“NB” or “NBP”) subsequent to the Corvus Gold Inc. (“CGI” or “Corvus”) Technical Report dated December 15, 2017 (the “2017 Report”).

Prior to the 2017 Report, Corvus acquired the MLP from Goldcorp USA, Inc. on Jun 6, 2017. At the time of the MLP acquisition, there were no MLP Mineral Resources as defined by NI 43-101. Since September 2017, Corvus has been drilling mineralization at the MLP to support the estimate of Mineral Resources subsequent to the drilling data analysis, geological modeling and metallurgical testing that are presented in this Report. The MLP is located 10 kilometres southeast of NBP and the two projects would potentially have shared infrastructure and locally available development resources. Corvus has evaluated the combination of project infrastructure based upon the new Mineral Resources and technical information that have been developed at the MLP. Resource Development Associates, Inc. (“RDA”) have developed a preliminary economic assessment (“PEA”) at this time to reflect the integration of the two properties into a single mining operation (the “Project”). This Report reflects combined economic performance of NBP and MLP now that sufficient technical work on the Project has been completed.

Highlights of the Technical Report, including the PEA, are listed in Table 1-1 through Table 1-9. Table 1-1 and Table 1-2 present a summary of the categorized Mineral Resources for the MLP and NBP. Mineral Resources are reported according to the CIM Definition Standards of May 10, 2014 (“CIM”). The guidance and definitions of CIM are incorporated by reference in Canadian National Instrument 43-101, Standards of Disclosure for Mineral Projects within Canada (“NI 43-101”.) Mineral Resources are pit constrained in order to estimate the portion of MLP and NBP that demonstrates reasonable prospects of eventual economic extraction. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the Mineral Resources will be converted into Mineral Reserves.

1.1 Resource Estimates

The basis for the Mineral Resource estimates are geologic models interpreted by Corvus geologists and constructed in Leapfrog® Software. Geostatistics and estimates of mineralization were prepared by RDA. The current mineralization update focused on developing the MLP stratigraphic and mineralization interpretations around the Mother Lode deposit. No changes were made to the NBP geologic interpretation and no new drill data was available to add to the existing NBP database.

Industry accepted grade estimation techniques were used to develop global mineralization block models. The MLP and NBP Mineral Resource estimates consider three conceptual processing methods; 1) mill processing of oxide and gravity-separable gold and silver, (2) mill processing of sulphide gold and silver mineralization by gravity and flotation concentration followed by Pressure Oxidation of the concentrate, and (3) heap leach processing of oxide gold and silver mineralization. The process plant would include a CIL circuit and a refinery to produce a doré on-site for sale at spot market prices.

Table 1-1 lists the Measured, Indicated and Inferred Mineral Resources at MLP for the various cut-off grades associated with the different processing techniques assumed. The effective date of the Report is September 18, 2018. The MLP Mineral Resource estimation is based on 267 drill holes with 8,296 gold composites.

Table 1-1 Mother Lode Mineral Resource – Pit Constrained Measured, Indicated and Inferred Estimation at a Gold Price of $1,250 Per Ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)

Resource Category	Mill-Sulphide @ 0.63 g/t Cut-off Grade			ROM Heap Leach @ 0.06 g/t Cut-off Grade			Total
	Kt	Au g/t	Kozs	Kt	Au g/t	Kozs	Kt	Au g/t	Kozs
Measured	3,292	1.41	149	20,035	0.29	185	23,327	0.45	334
Indicated	9,934	1.83	583	20,123	0.37	242	30,057	0.85	825
Total M & I	13,226	1.72	733	40,158	0.33	427	53,383	0.68	1,159
Inferred	2,168	1.60	112	14,073	0.29	129	16,241	0.46	241

There are three mineral deposits which make up the NBP mineralization: 1) Sierra Blanca/YellowJacket; 2) Jolly Jane; and 3) Mayflower. The total Mineral Resource at NBP is listed in Table 1-2 without regard to processing requirements. The applicable cut off grades are indicated by the individual items immediately following Table 1-2.

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Preliminary Economic Assessment – NB-MLP	Page 2

Table 1-2 North Bullfrog Project Pit Constrained Mineral Resource Estimate at a Gold Price of $1,250 Per Ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)

	Classification	Tonnes (k)	Au (g/t)	Ag (g/t)	Contained Au (000’s)	Contained Ag (000’s)
Phase I	Measured	10,415	1.08	7.59	362	2,540
	Indicated	24,557	0.69	3.70	542	5,459
	Inferred	5,908	0.31	0.74	59	140
Phase II	Measured	10,129	0.26	1.04	84	338
	Indicated	113,009	0.21	0.61	771	2,227
	Inferred	58,887	0.19	0.48	367	902
Totals	Measured	20,544	0.68	4.36	446	2,878
	Indicated	137,566	0.30	1.16	1,314	5,146
	Inferred	64,785	0.20	0.50	426	1,042

Mineralization within the YellowJacket vein zone and surrounding stockwork is situated within a well-defined zone that holds together at higher cutoff grades within the resource constraining pit. The Sierra Blanca/YellowJacket deposit contains the highest grades. The higher grades at the Mayflower deposit would be economically extracted prior to extracting lower grade heap-leachable mineralization. This portion of the mineral deposit is referred to as Phase I (Tables 1-3, and 1-4). The limits of the constraining pit were determined by assuming only mill mineralization and higher grade disseminated heap leach mineralization would be processed. Mineralization is reported at higher cutoff grades for this Phase I portion of the deposit and listed as the following items:

· Phase I Mill Resources Sierra Blanca (Oxide, Vein and Stockwork Mineralization), >= 0.35 grams per tonne (“g/t”) gold

· Phase I Mill Resources Sierra Blanca (Sulphide Disseminated Mineralization), >= 0.71 g/t Au

· Phase I Heap Leach Resources Sierra Blanca (Oxide Vein Mineralization), (>= 0.15 and <0.35) g/t Au

· Phase I Heap Leach Resources Sierra Blanca (Oxide Disseminated Mineralization), >= 0.15 g/t Au

· Phase I Heap Leach Resources Mayflower (Oxide Mineralization), >= 0.10 g/t Au

Table 1-3 NBP Sierra Blanca/YellowJacket Phase I Mineral Resources Pit Constrained at a Gold Price of $1,250/Oz Au (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)

	Milling			Heap Leach			Sulphides
	>=0.35 g/t Au cut off			>=0.15 g/t Au cutoff			>=0.71 g/t cutoff
Classification	Tonnes (Kt)	Au g/t	Ag g/t	Tonnes (Kt)	Au g/t	Ag g/t	Tonnes (Kt)	Au g/t	Ag g/t	Au Ounces (x1,000)	Ag Ounces (x1,000)
Measured	4,465	1.87	13.97	5,194	0.37	2.42	756	1.32	5.35	362	2,540
Indicated	4,445	1.80	13.40	13,736	0.34	1.63	1,137	1.56	5.83	464	2,850
Inferred	34	1.83	19.21	5,831	0.30	0.59	15	2.07	16.59	59	140

Table 1-4 NBP Mayflower Phase I Resources Pit Constrained at a Gold Price of $1,250 Per Ounce (Cut-off >=0.10 g/t Au)

(Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)

Classification	Tonnes (kt)	Au g/t	Ag g/t	Contained Au (x1,000)	Contained Ag (x1,000)
Indicated	5,239	0.46	0.41	78	69
Inferred	28	0.21	0.24	0	0

Mineral Resources which meet the reasonable prospects of economic extraction in a subsequent phase (Phase II) of resources are listed in Tables 1-5 and 1-6 for Sierra Blanca and Jolly Jane, respectively. Mineral Resources are reported at the breakeven cutoff grades for the mineralization and processing methods described:

· Phase II Mill Resources Sierra Blanca (Oxide, Vein and Stockwork Mineralization), >= 0.35 g/t Au

· Phase II Mill Resources Sierra Blanca (Sulphide Disseminated Mineralization), >= 0.71 g/t Au

· Phase II Heap Leach Resources Sierra Blanca (Oxide Vein Mineralization), (>= 0.10 and <0.35) g/t Au

· Phase II Heap Leach Resources Sierra Blanca and Jolly Jane (Oxide Disseminated Mineralization), >= 0.10 g/t Au

· Phase II NBP Mineral Resources are pit constrained and are reported exclusive of Phase I Mineral Resources in Tables 1-5 and 1-6.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 3

The NB estimate is based on 767 holes comprising 141,286 metres (“metres” or “m”) of length, 89,170 Au samples and 79,236 silver (“Ag” or “silver”) samples.

Table 1-5 NBP Sierra Blanca Phase II Resources Pit Constrained at a Gold Price of $1,250 Per Ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)

	Milling			Heap Leach			Sulphides
	>=0.35 g/t Au cut off			>=0.15 g/t Au cutoff			>=0.71 g/t cutoff
Classification	Tonnes (Kt)	Au g/t	Ag g/t	Tonnes (Kt)	Au g/t	Ag g/t	Tonnes (Kt)	Au g/t	Ag g/t	Au Ounces (x1,000)	Ag Ounces (x1,000)
Measured	397	0.78	4.07	9,331	0.19	0.85	401	1.24	2.48	84	338
Indicated	1,331	0.89	4.16	91,525	0.18	0.58	1,402	1.18	1.82	625	1,973
Inferred	6	0	5.18	50,939	0.19	0.46	61	1.53	2.04	314	760

Table 1-6 NBP Jolly Jane Pit Constrained at a gold price of $1,250 per ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)

Classification	Tonnes (kt)	Au g/t	Ag g/t	Contained Au (x1,000)	Contained Ag (x1,000)
Indicated	18,751	0.24	0.42	146	254
Inferred	7,871	0.21	0.56	53	142

Mineral Resources are not mineral reserves and do not demonstrate economic viability. There is no certainty that all or any part of the Mineral Resource will be converted to mineral reserves. Quantity and grade are estimates and are rounded to reflect the fact that the resource estimate is an approximation.

The combined Mineral Resource Estimate for ML-NBP is summarized in Table 1-7. Mineral Resources in Table 1-7 are summarized based on the above reported Mineral Resources.

Table 1-7 Mother Lode and North Bullfrog, Pit-constrained, Measured, Indicated and Inferred Mineral Resources at Gold Selling Price of USD $1,250 per Ounce (Qualified Person: RDA, Scott Wilson, CPG; Effective September 18, 2018)

	Gold Milling Cutoff Grade 0.63-0.76 Au g/t			Gold Heap Leach Cutoff Grade 0.06-0.15 Au g/t			Gold Total (Milling & Heap Leach)
Resource Classification	Tonnes (x1,000)	Au g/t	Au Ounces (x1,000)	Tonnes (x1,000)	Au g/t	Au Ounces (x1,000)	Tonnes (x1,000)	Au g/t	Au Ounces (x1,000)
Measured	9,311	1.59	475	34,560	0.27	305	43,871	0.55	780
Indicated	18,249	1.68	988	149,374	0.24	1,150	167,623	0.40	2,138
Total M & I	27,560	1.65	1,463	183,934	0.25	1,455	211,494	0.43	2,918
Inferred	2,284	1.61	118	78,742	0.26	549	81,026	0.26	667
	Silver Milling Cutoff Grade 0.63-0.76 Au g/t			Silver Heap Leach Cutoff Grade 0.06-0.15 Au g/t			Silver Total (Milling & Heap Leach)
Resource Classification	Tonnes (x1,000)	Ag g/t	Ag Ounces (x1,000)	Tonnes (x1,000)	Ag g/t	Ag Ounces (x1,000)	Tonnes (x1,000)	Ag g/t	Ag Ounces (x1,000)
Measured	6,019	11.47	2,220	14,525	1.41	658	20,544	4.36	2,878
Indicated	8,315	8.93	2,388	129,251	0.66	2,758	137,566	1.16	5,146
Total M & I	14,334	9.99	4,608	143,776	0.74	3,416	158,110	1.58	8,024
Inferred	116	9.12	34	64,669	0.48	1,008	64,785	0.50	1,042

Mineral Resources are not mineral reserves and do not demonstrate economic viability. There is no certainty that all or any part of the Mineral Resource will be converted to mineral reserves. Quantity and grade are estimates and are rounded to reflect the fact that the resource estimate is an approximation.

1.2 Preliminary Economic Assessment

A summary of the current projected financial performance of the integrated MLP and NBP is listed in Table 1-8 and 1-9.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 4

By definition, a PEA is preliminary in nature, and there is no certainty that the reported results will be realized. The Mineral Resource estimate used for a PEA includes Inferred Mineral Resources which are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves, and there is no certainty that the PEA will be realized. The basis for the PEA is to demonstrate the economic viability of the integrated MLP and NBP. The PEA results are only intended as an initial, first-pass review of the Project economics based on preliminary information.

Table 1-8 NB-MLP Production Years 1 through 4 Highlights

Production Year 1-4
Physicals	Heap Leach	Mill	Total
Process Life (years)	4	4	4
Process Feed (tonnes / day (nominal))	77,200	8,200	85,400
Gold Recovery (%)	68%	84%	76%
Silver Recovery (%)	5%	74%	68%
Total koz. Contained Au(1)	909	914	1,823
Total koz. Contained Ag(1)	2,052	3,337	5,390
Total koz. Produced (Au)(1)	625	764	1,388
Total koz. Produced (Ag)(1)	111	2,480	2,591
Total koz. Produced (Au eq.)(1)(2)	627	794	1,420
Annual koz. Produced Gold(1)	156	191	347

Table 1-9 NB-MLP Life-of-Mine Project Economic Highlights

Production Years 1-9
Physicals	Heap Leach	Mill
Mine Life (years)	9	9
Process Feed (tonnes / day (nominal))	73,500	8,200
Contained Gold (koz.)	1,760	1,497
Contained Silver (koz)	3,592	4,415
Gold Recovery (%)	74%	83%
Silver Recovery (%)	6%	74%
Total koz. Produced Au (1)	1,300	1,242
Total koz. Produced Ag(1)	208	3,280
Total koz. Produced Au eq.(1)(2)	1,303	1,281
Project Financials
Estimated Initial Capital(3) ( M$ US)	$424
Estimated Sustaining Capital(4) ( M$ US)	$60
Total Capital(3),(4) (M$ US)	$484
Estimated Total Cash Cost / Produced Au Eq Oz. (US $)	$680
Estimated Capital Cost / Produced Au eq oz (US $)	$187
Estimated Total Cost/ Produced Au eq oz (US $)(4)	$867
Pre-tax Undiscounted Cash Flow(5)(M$ US)(4)	970
After-tax Undiscounted Cash Flow(6) (M$ US)(3)	$797
After-tax Net Present Value(6) (NPV 5%) (M$ US)	$586
Payback Period(6) (years)	2.1

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 5

1.3 Property Descriptions and Ownership

The MLP is located north of the Bare Mountains of northwestern Nye County, Nevada (Figure 4-1) and is connected to the southeast corner of the NBP property. The property consists of about 3,570 hectares (8,820 acres) of federal unpatented lode mining claims located in Sections 10, 11, 14, 15, 22, 23, 26, 27, 34, 35 and 36 of T11S, R47E; Sections 1, 2, 3, 9, 10, 11, 12, and 13 of T12S, R47E; and Sections 6, 7, 8, 9, 16, 17 and 18 of T12S, R48E, Mount Diablo Base and Meridian (“MDBM”).

The NBP is located in the Bullfrog Hills of northwestern Nye County, Nevada (Figure 4-1).

Corvus’ NBP property covers approximately 7,223 hectares (17,250 acres) of patented and unpatented lode mining claims in Sections 20, 21, 25, 26, 27, 28, 29, 32, 33, 34, 35, and 36 of T10S, R46E; sections 1, 2, 11, 12, 13, and 14 of T11S, R46E; section 31 of T10S, R47E; and section 6, 9, 15, 16 and 17 of T11S, R47E, MDBM.

The Project is accessible as a two and one half hour (260 kilometres) drive north of Las Vegas, Nevada along US Highway 95. US Highway 95 is the major transportation route between Las Vegas, Reno, and Boise. Las Vegas is serviced by a major international airport. MLP lies approximately 10 kilometres east of Highway 95 at the south end of Beatty, NV. NBP lies immediately to the west of the Highway 95 and approximately 10 kilometres north of Beatty, NV. Beatty is the closest town to the Project with a population of about 1,100 and contains most basic services. Access around the two Project areas is by a series of reasonably good gravel roads that extend from Highway 95 to the exploration areas.

Corvus controls the NBP Project through a number of private land leases and unpatented federal load claims listed in Table 4-2 through Table 4-5 of Section 4. Corvus owns and leases several patented lode mining claims and maintains a large contiguous block of federal unpatented lode mining claims. In 2014, Corvus purchased 162 hectares of surface lands in Sarcobatus Flats, approximately 26 kilometres north of the NBP, which includes 1,600 acre feet per year of water rights. In 2018, the Nevada Division of Water Resources granted temporary change of the water right usage to mining application for use at NBP.

1.4 Geology and Mineralization

The Project lays within the Walker Lane mineral belt and the Southwestern Nevada Volcanic Field (SWNVF). The regional stratigraphy includes a basement of Late Proterozoic to Late Paleozoic metamorphic and sedimentary rocks. Basement rocks are overlain by a thick pile of Miocene volcanic and lesser sedimentary rocks of the SWNVF, ranging in age from ~15-7.5 Ma (Figure 7-1). The pre-Tertiary rocks exhibit large-scale folding and thrust faulting, having been subjected to compressional deformation associated with multiple pre-Tertiary orogenic events. The stratigraphy of the SWNVF is dominated by ash flow tuff sheets erupted from a cluster of nested calderas known as the Timber Mountain Caldera Complex. The southwestern edge of the caldera complex lies approximately eight kilometres northeast of the MLP, and ten kilometres east of the NBP (Figure 7-1). The stratigraphy of the SWNVF includes voluminous ash flow tuff sheets, smaller volume lava flows, shallow intrusive bodies, and lesser sedimentary rocks. Many of the volcanic units exposed around the Project include ash flow tuffs that originated from the caldera complex. Other volcanic units are locally sourced outside of the caldera complex, particularly at the NBP.

Mother Lode is characterized as a sediment, intrusive, and locally volcanic-hosted disseminated gold deposit. Mineralization most closely resembles Carlin-type sediment-hosted gold deposits of north-central Nevada. Weiss (1996) was the first to recognize and document the similarity to Carlin-type deposits. Weiss (1996) also cites evidence for a large buried porphyry-type magmatic system associated with the rhyolite dike swarm at eastern Bare Mountain. The Mother Lode deposit formed at ~12.7 Ma, which is much younger than the typical ~40 Ma age in north-central Nevada. The nature of mineralization is rather passive and with very low introduction of secondary silica, suggesting it may have formed at a shallower depth and at lower temperature than typical Carlin-type deposits. Mineralization exhibits geochemical associations between Au and As-Sb-Hg-Tl-Te-Bi-F, with very low Ag and base metals.

Gold mineralization in the NBP is primarily hosted in the middle Miocene Sierra Blanca Tuff. Gold mineralization is also hosted to a lesser extent in monolithic and heterolithic debris-flow breccias, as well as in felsic dikes and plugs. Two district-scale north striking normal faults are the dominant structural features in the Project area, but several smaller-scale faults between them are important controls for distribution of hydrothermal alteration and gold mineralization.

Two styles of precious metal epithermal mineralization are present at the NBP: 1) high-grade, structurally controlled fissure veins and associated stockwork zones; and 2) low-grade disseminated or replacement deposits within altered volcanic rocks. Historic drilling on the NBP, which was not NI 43-101 compliant, outlined areas of important mineralization. Drilling by International Tower Hill Mines Ltd. (“ITH”), a predecessor-in-interest of Corvus, was used to develop initial resource estimates, to better understand precious metal mineralization at Air Track Hill and as initial tests at the Sierra Blanca, Pioneer, Savage and YellowJacket targets.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 6

1.5 Metallurgical Testing

An initial metallurgical testing program was scoped in 2018 as geologic information and sample materials became available from the on-going Mother Lode drilling program. Some historical metallurgical data and production records were available from the heap leach mining projects which operated at the Mother Lode deposit in 1988-1989, and from the Daisy Project which operated in 1990-1996, and were used to guide the testing program design. CN bottle rolls, preg-rob tests and CN shake leach tests were performed on Mother Lode drill samples of oxide mineralization to confirm similarity to historical data which had been developed for design of the heap leach process operations. Those tests indicated high CN solubility (+90% at RC cuttings size) of the mineralization forming the major components (dolomite and rhyolite dikes) of the oxidized mineralization.

The majority of the mineralization at Mother Lode is contained in sulphide minerals (predominantly pyrite) and requires preparation of a sulphide concentrate followed by oxidation to produce a filter cake from which the gold can be CN leached. Two composite samples were prepared from RC drill cuttings consisting of the two primary host rocks. These include the volcaniclastic sediments of the Sedimentary Rocks of Joshua Hollow (Tjvs) and quartz-porphyry rhyolite dike material (Tip). Metallurgical characterization of the samples included CN bottle rolls, detailed mineralogical analyses, grind size optimization, gravity concentration, and flotation concentration. Limited core materials were available and were used to generate some comminution data for Mother Lode mineralization. Bulk samples of concentrate from the two primary mineralization units were prepared and used as feed for oxidation tests which included the currently practiced process methods of Pressure Oxidation, Roasting, Ambient Atmospheric Oxidation and BioOxidation. Filter cakes were then leached to recover the gold using a CIL process. Results, available for Pressure Oxidation, are the basis for the sulphide processing assumed in the PEA. Final results for the other oxidation processes are pending at the time of this report. Maximum gold recoveries to concentrate were 82% for the Tjvs sample and 86% for the Tip sample. The pressure oxidation tests indicated that 96% of the gold could be recovered from the oxidized concentrate. Net gold recoveries were therefore assumed to be 80% for Mother Lode sulphide ores based on weight averaging of the two different gold mineralogies.

During 2012-2013, metallurgical testing was performed using composite samples developed from PQ core materials produced at Mayflower, Sierra Blanca, Savage Valley and Jolly Jane. Column leach testing on up to P80 -19 millimetres (“mm”) indicated relatively high gold recoveries in the range of 80% and confirmed the suitability of heap leach processing on disseminated mineralization. In 2014-2015, composite samples of PQ core materials were developed from YellowJacket vein and stockwork mineralization. Those tests indicate high solubility of contained gold in cyanide leach testing at P80 -150 microns, but reduced gold recoveries at heap leach sized particles. These tests indicated that mill processing would be required on YellowJacket mineralization. Further testing of composite samples, using gravity concentration, intense cyanide leaching of gravity concentrate and cyanide leaching of the gravity tails, indicate gold recoveries in the range of 90% and silver recoveries in the range of 70%. During 2016 and 2017, flotation concentration and AAO oxidation of sulphide mineralization was successfully demonstrated which allowed gold recoveries of +90% by cyanide leaching of filter cake products created from NBP sulphides.

The metallurgical testing reported in the Technical Report has been conducted by McClelland Laboratories of Reno, NV, by Hazen Research of Golden, CO, by Resource Development Inc. of Wheatridge, CO and Blue Coast Research of British Columbia, Canada.

1.6 Current Exploration and Development

Corvus has focused its exploration program and technical program work at the recently acquired MLP. Currently, there are minimum expenditures for development programs in progress at NBP.

1.7 Conclusions

Corvus has invested considerable effort and investment in the advancement of the Project through drilling, permitting, technical and metallurgical evaluations, internally and with the assistance of reputable consulting firms. This PEA evaluation indicates good economic performance through a combination heap leaching and oxide/sulphide milling facility at the Project at the current metal price environment. The Project performance is most sensitive to gold price and gold recovery. Metallurgical data to this point indicates that conventional processing of gold mineralization, as currently practiced throughout Nevada, would be viable for economic extraction of metals at NB-MLP.

The PEA suggests that this is a project that may be put into production for a total capital investment estimate of approximately US $505 million and with the capital being paid back within approximately 2.2 years of startup. Good potential exists for the discovery of additional mill and heap leach Mineral Resources at exploration target areas identified within both the NBP and ML Project claim blocks.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 7

RDA is of the opinion that the current Mineral Resources at Mother Lode and at North Bullfrog are sufficient to warrant continued planning and effort to explore, permit and develop the Project, and that it supports the conclusions and PEA detailed herein.

RDA believes there is sufficient data to support continued exploration, geologic modeling and continuing development of the Project.

1.8 Recommendations

The PEA results for the Project, with a mill facility at MLP and heap leach operations at both NBP and MLP, demonstrate the substantial financial impact on project performance of the higher grade sulphide and vein/vein stockwork types of mineralization that occurs at each site. Therefore, it is recommended that future exploration should focus on the identification and development of other deposits and sources of higher grade mineralization. In addition, RDA recommends that Corvus define and continue to execute programs that enhance and develop the supporting technical information, particularly as related to metallurgical processing of sulphide resources at both locations. Specifically, promising approaches to heap leach processing of sulphide mineralization have recently been tested at different locations and have the potential for substantial and favorable impacts on the project economic performance. These data would be used to refine the preliminary economic assessment of the Project, at the time of the next resource update. These recommended activities are:

· Exploration drilling at identified targets at both MLP and NBP;

· Resource definition drilling at MLP;

· Metallurgical testing of gold bearing mineralization, particularly novel sulphide heap leaching technology;

· Continued environmental baseline characterization;

The projected costs for the next phase of this program are outlined in Table 1-10.

Table 1-10 Proposed Budget to Support Recommended Program at NB-MLP

Activity	Amount
Exploration Drilling Resource Definition Drilling and Data Management	US$ 0.80 M
Baseline Data Collection	US$ 0.05 M
Metallurgical Testing	US$ 0.35 M
Resource Model, Geologic Model and PEA Update	US$ 0.15M
Total	US$ 1.35M

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 8

2 Introduction

2.1 General Statement

NB-MLP (the “Project”) is an advanced stage surface exploration project consisting of both the North Bullfrog Project (“NBP”) and the recently acquired Mother Lode Project (“Mother Lode” or “MLP”). Both properties are 100% controlled by Corvus Gold Inc. (TSX:KOR, OTC:CORVF, “Corvus” or the “Company”). In 2017, Corvus acquired the 13 Mother Lode claims from Goldcorp USA, Inc., which contained the Mother Lode open pit mine which had been operated by Nevada Gold Search Inc. in 1988-1989. Corvus launched a focused exploration drilling program to both verify existing drilling data already in Corvus possession and to expand the extent of the mineralization system that had been indicated by exploration associated with the historical mining operation. Goldcorp USA, Inc. received 1 million shares of Corvus stock as payment for the Mother Lode claims. Corvus has subsequently increased the MLP land position by staking 401 additional unpatented claims on adjacent land where surface geology suggests similar gold mineralization targets. MLP now consist of unpatented mining claims on public lands covering approximately 38 square kilometres in the Mother Lode, MN and ME claim groups (Table 4-1). NBP consists of a mix of patented and unpatented mining claims covering approximately 75 square kilometres. The MLP has been extended to the north and now shares a common border with the southeast corner of the NBP (Figure 2-1).

The report has been amended November 8, 2018 to correct the following editorial issues:

· The columns in the “silver” portion of Tables 1-7 and 14-1 were mislabeled as Au and have been changed to Ag. Overall the columns had been labeled as “Silver” before the amendment;

· The general report header was corrected to read “Preliminary Economic Assessment-NB-MLP” from page 156 onwards;

· On page 225, “Error!Reference source not found” was replaced by reference to Table  16-2;

· On page 227, “As can be seen in the Table 16-3” has been corrected to reference “Table 16-6”;

· The Table 22-4 caption was corrected to read “Base Case( Gold Price $1,250; Silver Price $15.08)”;

· A yellow highlight was removed from the Table 1-8 title in the List of Tables. “

MLP is located approximately 10 kilometres due east of the town of Beatty, Nevada and extends to approximately 15 km north of Beatty to NBP. Mineralization at the Project is related to a large, low-sulphidation epithermal gold system hosted in volcanic and sedimentary rock units. Gold and silver were discovered in the Bullfrog district in 1904. Production records indicate that more than 110,000 ounces of gold and more than 800,000 ounces of silver were produced through 1921. Modern exploration at North Bullfrog began in 1974. In 2009, Corvus was formed as a spin out from International Tower Hill Mines Ltd. (“ITH”), where 100% percent of the Project was transferred to Corvus. Corvus has been actively exploring at NBP since 2009 and extended exploration to MLP in fall of 2017.

This Technical Report incorporates recent drilling, geological interpretation and metallurgical data, developed in 2017 and 2018. The Report includes a maiden Mineral Resource estimate for MLP. The Report delineates a scoping study that describes a conceptual mining operation for the Project.

The independent authors of this Technical Report, Scott Wilson of Resource Development Associates, Inc. (“RDA”), Michael Cole (an independent mining consultant), Christopher Easton of Easton Process Consulting, Inc. (“EPC”) and Richard Delong of EM Strategies (EMS) used various forms of digital data in the Technical Report including geologic models based on surface mapping and drilling, assay data, and metallurgical testing data developed by Corvus, RDA, and EPC.

The NB-MLP is located in northwestern Nye County, Nevada, in the Bare Mountains and the Northern Bullfrog Hills located about 10 kilometres East and 15 kilometres North of Beatty (Figure 2-1). The Project lies within the Walker Lane structural terrain East and North of the historic Bullfrog Mine where Barrick Gold Corp. (and predecessor companies) produced about 2.3 million ounces of gold and 3.0 million ounces of silver from 1989 through 1999 (NBMG MI-2000, page 34). The NB-MLP contains numerous epithermal low-sulphidation volcanic and sediment rock-hosted gold showings that have had limited historic production, as well as the recent and major mining operations at the Bullfrog, Mother Lode and Daisy mining projects.

Corvus controls the Project through a number of private land leases with various land owners (Table 4-2 and Figure 4-1) and through numerous federal unpatented lode mining claim holdings. In 2006, Redstar Gold Corp (“RGC”) initially assembled 213 unpatented lode mining claims and 33 patented lode mining claims through six option agreements, which were the property subject to the original joint venture agreement between RGC and ITH called the North Bullfrog Project Joint Venture (the “NBPJV”). ITH leased an additional 11 patented lode mining claims in the Mayflower area, which became subject to the NBPJV agreement. ITH earned into the NBPJV when in August 2009 they negotiated an agreement to purchase RGC’s interest in the joint venture property. ITH spun out Corvus on August 26, 2010 as the controlling entity of the North Bullfrog property. Corvus completed one additional option agreement on patented lode mining claims in the Jolly Jane area in March 2011, for a total of eight option agreements on private lands. Corvus also expanded the North Bullfrog property in early 2012 by staking a total of 511 federal unpatented lode mining claims. In late 2012, Corvus staked an additional 297 unpatented lode mining claims to the north and east (Figure 4-1), bringing the total unpatented lode mining claim holdings to 808. All of these claims are in good standing with the Bureau of Land Management (the “BLM”) and Nye County.

On March 23, 2013, Corvus announced the purchase of surface rights on five patented lode mining claims from Mr. and Mrs. Gordon Millman to facilitate shorter overburden haul distances for development of the Mayflower deposit. The terms of the purchase as outlined in the Corvus press release (February 21, 2013) were, “…USD 160,000, payable at closing. The terms also included payment by Corvus Nevada of a fee of USD 0.02 per ton of overburden to be stored on the property, subject to payment for a minimum of 12 million short tons. The minimum-tonnage fee (USD 240,000) had an interest rate of 4.2% per annum from closing, as evidenced by a promissory note due on the sooner of the beginning of production or December 31, 2015”. The promissory note was paid off in December 2015.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 9

Figure 2-1 Map Showing the Location of North Bullfrog and Mother Lode along with Local Gold Deposits Near Beatty

 

2.2 Terms of Reference

Corvus requested that this Technical Report be prepared to support a Mineral Resource estimate to include historical drilling results from previous mining operators at the Mother Lode and Daisy Projects plus the 78 new drill holes performed by Corvus in 2017 and 2018. MLP ownership and land positions have been added during the period 2017 and 2018 and are reflected in this Technical Report, however, no material changes in the NBP ownership or land position have occurred at North Bullfrog since publication of the 2017 Report. This Technical Report includes metallurgical data produced on both oxide and sulphide mineralization from MLP drill sample material produced in 2018. This Technical Report also outlines the geology, exploration history, and new drilling data produced by Corvus during 2017 and 2018, as well as metallurgical testing data produced on samples from the 2017 Corvus MLP drilling program. A conceptual Project based on possible exploitation of both NBP and MLP mineralized areas and a common mill processing plant has been created for an economic evaluation of a preliminary nature. Mr. Scott E. Wilson (RDA), Mr. Michael Cole (Mining Consultant), Christopher Easton (EPC) and Richard Delong (EMS) were commissioned by Corvus to prepare this Technical Report.

Mr. Scott Wilson, (CPG #10965, SME 4025107RM), an independent Qualified Person, was the principal author responsible for the overall preparation of this Technical Report, and specifically for Sections 1 through 12, Section 14, Section 18 and 19, and Sections 21 through 27. Mr. Wilson visited the NBP site on January 30 and 31, 2012, on March 24, 2014, on November 2-3, 2015, and NB-MLP on June 6-8, 2017 and on January 16, 2018. Mr. Wilson is independent of Corvus applying all of the tests in Section 1.5 of NI 43-101.

Mr. Michael Cole, (SME), and independent Qualified Person, was responsible for the preparation of Sections 15 and 16 as well as the relevant parts of Sections 1, 21 and 25. Mr. Cole visited the site on September 11, 2017. Mr. Cole is independent of Corvus applying all of the tests in Section 1.5 of NI 43-101.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 10

Mr. Christopher Easton, Consulting Metallurgist and a member of the Mining and Metallurgical Society of America (“MMSA”), as an independent Qualified Person, was responsible for Section 13 and Section 17 of the Technical Report. Mr. Easton visited the NBP site on April 22 and 23, 2017 and the MLP site on February 17, 2018. Mr. Easton is independent of Corvus applying all of the tests in Section 1.5 of NI 43-101.

Mr. Richard Delong, (SME), as an independent Qualified Person, was responsible for the preparation of Sections 20 of the Technical Report. Mr. Delong visited the site on April 3, 2017. Mr. Delong is independent of Corvus applying all of the tests in Section 1.5 of NI 43-101.

All dollar amounts in this document are United States dollars unless otherwise noted.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 11

3 Reliance on Other Experts

No other experts were relied upon in the preparation of this report

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 12

4 Property Description and Location

In 2017 and 2018, Corvus significantly expanded its land position in the Bullfrog district by acquiring the historic Mother Lode property from Goldcorp USA, Inc., staking additional claims in the Mother Lode vicinity, and staking claims between the MLP and the NBP. Corvus’ land position in the Bullfrog district is now collectively referred to as NB-MLP. NB-MLP land positions are shown on Figure 4-1.

Figure 4-1 Property Map of the NB-MLP (UTM NAD 27 Zone 11)

4.1 Mother Lode Property

4.1.1 MLP Area and Location

Mother Lode is located in the northern Bare Mountain area of northwestern Nye County, Nevada. Figure 4-1 shows the boundary defined by public land mining claims in purple. The Project covers approximately 3,570 hectares (8,820 acres) of federal unpatented lode mining claims located in Sections 10, 11, 14, 15, 22, 23, 26, 27, 34, 35 and 36 of T11S, R47E; Sections 1, 2, 3, 9, 10, 11, 12, and 13 of T12S, R47E; and Sections 6, 7, 8, 9, 16, 17 and 18 of T12S, R48E, Mount Diablo Base and Meridian (“MDBM”). A list of the mining claims included in the MLP is provided in Table 4-1. Corvus owns, through its wholly owned subsidiary, Mother Lode Mining Company LLC, the historic Mother Lode property which consists of 13 mining claims (~105 hectares; ~260 acres). The MN and ME claims were staked by Corvus in 2017 and 2018 and are also 100% owned by Corvus.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 13

Table 4-1 Summary of Federal Mining Claims at MLP

Land Holder	Claim Name	BLM Serial Numbers
Mother Lode Mining Company LLC	BVC #5696	262789
Mother Lode Mining Company LLC	BVC #5697	262790
Mother Lode Mining Company LLC	BVC #5698	262791
Mother Lode Mining Company LLC	BVC #2	352232
Mother Lode Mining Company LLC	Mother Lode #15	264628
Mother Lode Mining Company LLC	Mother Lode #20	264633
Mother Lode Mining Company LLC	MF 15	699017
Mother Lode Mining Company LLC	MF 17	699019
Mother Lode Mining Company LLC	MF 19	699021
Mother Lode Mining Company LLC	TWE 41	685492
Mother Lode Mining Company LLC	TWE 42	685493
Mother Lode Mining Company LLC	TWE 43	685494
Mother Lode Mining Company LLC	TWE 44	685495
Corvus Gold Nevada Inc.	ME-01 to ME-22	1143493-1143514
Corvus Gold Nevada Inc.	MN-01 to MN-30	1143463-1143492
Corvus Gold Nevada Inc.	MN-31 to MN-105	1163266-1163340
Corvus Gold Nevada Inc.	MN-106 to MN-159	1166983-1167036
Corvus Gold Nevada Inc.	MN-160 to MN-414	1177223 -11774222

4.1.2 Goldcorp Property Purchase

In May of 2017, Corvus purchased 100% of the historic Mother Lode property by acquiring the Goldcorp Daisy LLC from Goldcorp USA, Inc. Goldcorp Daisy LLC was subsequently renamed the Mother Lode Mining Company LLC, which is a wholly owned subsidiary of Corvus Gold Nevada, Inc. The property consisted of 13 mining claims comprising ~105 hectares (~260 acres) around and adjacent to the historic Mother Lode open pit. Corvus issued Goldcorp USA, Inc., 1,000,000 Corvus Gold Inc. common shares and granted a 1% Net Smelter Royalty (NSR) on any future production from the 13 claims for a gold price up to $1,300 per ounce, increasing to a 2% NSR for a gold price equal to or greater than $1,400 USD per ounce.

4.1.3 Other MLP Property Considerations

All of the public land mining claims at the MLP are administered by the Bureau of Land Management (“BLM”). The mining claims require payment of yearly maintenance fees to the BLM and Nye County (recording fees) of an aggregate of $137,478 (estimated for 2019). These claims give Corvus the right to explore for and mine minerals, including gold and silver, subject to the necessary permits which are described in Section 20.

The current Notice-level exploration permit from the BLM allow Corvus to access, maintain the roads, perform drilling and sampling, and create up to five acres of surface disturbance under permit N-095622. A second Notice-level permit (N-96894) allows construction of up to 17 drill sites with up to 4.39 acres of surface disturbance at the Willys target area (east of Mother Lode) has been granted by the BLM. A third Notice-level permit (N-96991) allows construction of up to 8 drill sites for up to 0.7 acres of disturbance at the Sawtooth target area (west of Mother Lode) and has been acknowledged by the BLM. A Plan of Operations (NVN-096238) to construct up to 550 drill sites for up to 140 acres of disturbance is pending approval from the BLM and the Nevada Department of Environmental Protection (“NDEP”).

Reclamation bonding related to environmental liabilities at MLP is in place to cover proposed disturbance on the property. Such reclamation liabilities are covered by surety bonds issued to the BLM by Lexon Insurance Company in the amounts of $33,216 for the Mother Lode Notice and $9,515 for the Willys NOI. A further surety bond for $5,367 for the pending Sawtooth Notice is anticipated. Additional bonding of an estimated $440,067 will be required when the Mother Lode Plan of Operations is approved. Additional permits and bonding may be required for the expanded exploration program outlined in the Recommendation Section of this Technical Report.

Corvus has submitted an application to the Nevada Division of Water Resources for 275 acre-feet of water rights used historically in the Daisy Project mining and processing operation.

4.2 North Bullfrog Property

4.2.1 NBP Area and Location

The NBP is located in the Bullfrog Hills of northwestern Nye County, Nevada (Figure 4-1). The Project covers approximately 7,223 hectares of patented and unpatented lode mining claims in Sections 19, 20, 21, 23, 24, 25, 26, 27, 28, 29, 30, 31, 32, 33, 34, 35, and 36 of T10S, R46E; Sections 1, 2, 3, 4, 5, 6, 10, 11, 12, 13, 14 and 15 of T11S, R46E; Section 19, 30, 31, 32 of T10S, R47E; and Sections 4, 5, 6, 7, 8, 9, 15, 16, 17 and 18 of T11S, R47E, Mount Diablo Base and Meridian (“MDBM”). Corvus has a total of nine option/lease agreements in place that give it control of 303 hectares (748 acres) of private land derived from 51 historic patented lode mining claims which are listed in Table 4-2. Corvus has also purchased surface rights to an additional 37 hectares (91 acres) of private land derived from 5 historical patented claims in the Mayflower area.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 14

The mining claims and lease agreements give Corvus the right to explore the property subject to required regulatory permits which are described in Section 4.14. Corvus currently has permits for exploration on the public and private land from the BLM and Nevada Department of Environmental Protection (“NDEP”) Bureau of Mining Regulation and Reclamation (“BMRR”). The permits allow non-exclusive access by Corvus and its contractors for surface disturbance associated with exploration drilling, engineering characterization, and baseline environmental data collection as defined in the permits. The claims and lease agreements also give Corvus the rights to conduct mining operations to extract the Mineral Resources subject to future permits described in Section 20.

4.2.2 Redstar Option/Joint Venture/ITH Purchase of Land

Redstar Gold Corporation (RGC) originally staked 213 public land mining claims and optioned private land related to 21 historic patented lode mining claims from six private parties in 2005-2006 at the NBP. International Tower Hill Mines (ITH) optioned the original NBP land package from RGC in 2006, creating the North Bullfrog Property Joint Venture (NBPJV). In 2007, ITH added the Mayflower Property related to 11 historic patented lode mining claims to the NBPJV under the Greenspun lease agreement. In 2008, RGC added further private land related to 12 historic patented lode mining claims (Connection and adjacent properties) to the NBPJV under the first lease agreement with Lunar Landing LLC. In August 2009, ITH purchased a 100% interest in the NBPJV from RGC by paying RGC CAD$250,000 and issuing 200,000 ITH common shares. The land holdings were then transferred to Corvus during the spin out from ITH. In March 2011, Corvus completed the Sussman option agreement on private land related to two historic patented lode mining claims in the Jolly Jane area. In May 2014, Corvus amended its existing lease agreement with Kolo Corp. to add the historic Yellowrose and Yellowrose No. 1 patented claims. In March 2015, Corvus added a second option agreement with Lunar Landing LLC, to lease the Sunflower, Sunflower No. 1 and Sunflower No. 2 historic patented claims. Table 4-2 summarizes the obligations of the nine private land lease agreements which are part of Corvus’ responsibilities on the Project. Table 4-3 lists the historic patented claim names and Patent Numbers associated with the nine lease obligations. The principal author has verified that all lease obligations have been met and are paid in full as of the date of this Technical Report.

Table 4-2 Summary of Lease Obligations at NBP (All Funds USD)

Party	Area	Claims/Acres	Next Payment	Property Taxes	NSR	Signing Date	Term (yrs)	Term Extension	Option to Purchase Property	Option to Purchase Royalty	NSR Option Term
Gregory	North Pioneer	1/8.2	$3,600	na	2%	6/16/2006	10	yes	no	$1 M/%	na
Wylie1	Savage	3/45.7	$8,600	na	2%	5/22/2006	5	yes	no	$1 M/%	na
Kolo Corp	Jolly Jane & Yellowrose	4/81.7	$6,000	$258	3%	5/8/2006	10	yes	no	$0.85/%	na
Milliken	Pioneer	3/24.5	$5,400	na	2%	5/8/2006	10	yes	no	$1M/%	na
Pritchard	Pioneer	12/203.0	$20,000	na	4%	5/16/2006	10	yes	no	$1M/%	na
Lunar Landing LLC	Connection	12/195.0	$16,200	$207	4%	10/27/2008 Amended 5/28/2014	10	yes	$1 M	$1M/%	35 yrs
Lunar Landing LLC	Sunflower	3/59.2	$5,000	$180	4%	3/30/2015	4	7 yrs	$0.3 M	$0.5/%	35 yrs
Greenspun 2	Mayflower	11/183.05	$10,000	$214	4%	08/25/2017	10	3 yrs	$7.5 M	No	3 yrs
Sussman	Jolly Jane	2/37.4	$30,000	$113	2%	3/14/2011	10	10 yrs	Inclusive in Royalty Purchase	$1M/%	na
Total	-	51/748.7	$104,800	$972	-	-	-	-		-	-

¹ Original title transferred from Hall to Wylie due to death in the family

² Plus 50,000 ITH shares and 25,000 Corvus shares

4.2.3 Mayflower Area

ITH, through its subsidiary Talon Gold Nevada Inc. (now Corvus Gold Nevada, Inc.), entered into a mining lease with the Greenspun Group with an option to purchase 74 hectares (183 acres) of patented lode mining claims that cover much of the Mayflower prospect. The Mayflower lease requires Corvus to make payments and complete work programs as outlined in Table 4-4. During the term of the lease, any production from the Mayflower property is subject to a sliding scale royalty, also outlined in Table 4-4. Corvus has the right to purchase a 100% interest in the Mayflower property for $7.5 million plus a 0.5% NSR (if gold is less than $500) or 1.0% (if gold is above $500) at any time during the term of the lease (subject to escalation for inflation if the option is exercised after the 10th year of the lease). The annual property taxes to be paid by Corvus for the Mayflower property are $214. On February 11, 2015, the Mayflower mining lease with option to purchase was amended with the addition of an anti-dilution clause applying to the ITH shares and with an increase in the annual payment to include 25,000 Corvus shares. On November 22, 2017, the Mayflower lease was extended until 2027 with the original terms and an annual payment of $10,000, and 50,000 ITH shares and 25,000 Corvus shares.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 15

On February 21, 2013, Corvus signed a purchase agreement for the surface rights to 37 hectares (91 acres) of private land related to five historic patented lode mining claims owned by Mr. and Mrs. Gordon Millman. The claims are located immediately east of the Mayflower deposit. The purchase was subsequently closed on March 27, 2013. Corvus purchased the surface rights for $160,000. This ground could be used for overburden storage at the Mayflower deposit as well as improving access to the Mayflower deposit, in general. Additionally, Corvus agreed to pay Mr. and Mrs. Millman a fee of $0.02 per ton of any potential overburden storage subject up to a minimum storage of 12 million short tons of material. The minimum storage fee of $240,000 was financed by a promissory note at an interest rate of 4.2% per annum from the closing date and due on December 31, 2015. The promissory note was paid off in December 2015.

4.2.4 Other NBP Property Considerations

Corvus staked 652 additional public land mining claims between 2012 and 2015, including 595 claims staked in 2012, and 57 claims staked in 2015. Table 4-5 lists the total of 865 mining claims on federal land at the NBP. All of these properties are held through Corvus Gold Nevada, Inc. (“CGN”), which is a wholly owned subsidiary of Corvus Gold, Inc. The Corvus companies were created as a spin-out from ITH on August 26, 2010.

All of the mining claims on public land are administered by the BLM. The mining claims require payment of yearly maintenance fees to the BLM and Nye County (recording fees) of an aggregate of $144,465 (estimated for 2019). Annual property taxes to be paid by Corvus for some of the properties subject to the original six RGC leases and subsequent leases are tabulated in Table 4-2. These claims give Corvus the right to explore for and mine minerals, including gold and silver, subject to the necessary permits described in Section 20.

The current exploration permits from BLM and NDEP allow Corvus surface access, maintenance of roads, and exploration drilling with a defined amount of accompanying surface disturbance. Current exploration activities are covered by a Plan of Operations (NVN-083002) with the BLM. Two Plans of Operations are in place with the Nevada Department of Environmental Protection (“NDEP”) (NDEP #0280 and #0290) that fulfill the State of Nevada permitting obligations on the private and public lands, respectively. Reclamation bonds, related to environmental liabilities to which the NBP is subject, are in place to cover activities on the property. Corvus’ reclamation liabilities are covered by surety bonds issued by Lexon Insurance Company in the amount of $284,386 for 103 acres of disturbance on public land with the BLM and $209,070 for 20.3 acres of disturbance on private land with NDEP.

Two Notice-level operating permits have been obtained for disturbance in areas outside the Plan of Operations area. These include the North Bullfrog Baseline Collection Project and the East Bullfrog Exploration Project. Corvus’ reclamation liabilities for these permits are covered by surety bonds issued by Lexon Insurance Company in the amounts of $47,068 and $31,143, respectively. Additional permits and bonding may be required for the expanded exploration program outlined in the Recommendation Section of this Technical Report.

In December 2013, the Company completed the purchase of 170 hectares (420 acres) parcel of private land 16 kilometres north of the NBP, which includes 1,600 acre feet of irrigation water rights within the Sarcobatus Flat water basin. The cost of the land was USD$1,000,000. The Company has registered the purchase of water rights with the Nevada State Engineer (“NSE”) and has received Permit 87214 which allows a consumptive use of 1,277 acre-feet of water per year for the purpose of mining. The Company has also received Permit 87745T which allows the transfer of the extraction point of a portion of the water right to NBP and allows the consumption of 67 acre-feet of water per year to support its future drilling programs. The water right requires annual renewal and has currently been extended through June 11, 2019.

None of the authors know of any significant factors and risks that may affect access or title to the NBP, or the right or ability to perform work on the Project. To the extent known, the authors know of no other royalties, back-in rights, payments or other agreements and encumbrances to which the property is subject.

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 16

Table 4-3 Summary of the Historic Patented Claims in the Nine NBP Private Land Lease Agreements

Owner/Lessor	Claim Name	Patent  Number
Gregory	Jim Dandy	448055
Wylie	Gold Basin	330227
Wylie	Savage	330227
Wylie	Savage 2	330227
KoloCorp.	Black Jack	163170
Kolo Corp	ZuZu	261838
Kolo Corp	Yellowrose	369130
Kolo Corp	Yellowrose No. 1	369130
Milliken	Indiana 1	245488
Milliken	Indiana 2	245488
Milliken	Indiana 3	245488
Pritchard	Banker's Life	493623
Pritchard	Bimettalic 1	46204
Pritchard	Bimettalic 2	46204
Pritchard	Bimettalic 3	46205
Pritchard	Bluff	493623
Pritchard	Conservative	611953
Pritchard	KK1	504301
Pritchard	Mutual	493623
Pritchard	Penn Mutual	493623
Pritchard	Prudential	493623
Pritchard	Sunrise 1	114544
Pritchard	Sunrise 2	114544
Lunar Landing LLC	Dewey Bailey	269019
Lunar Landing LLC	Four Aces	269019
Lunar Landing LLC	Parson Haskins	269019
Lunar Landing LLC	Bull Con	269019
Lunar Landing LLC	Ugly	296019
Lunar Landing LLC	Hardtack	341527
Lunar Landing LLC	Connection Mine	342533
Lunar Landing LLC	Equity	342533
Lunar Landing LLC	Geraldine 3	342533
Lunar Landing LLC	Grey Eagle 2	342533
Lunar Landing LLC	Grey Eagle 4	342533
Lunar Landing LLC	Vinegarroan	342533
Lunar Landing LLC	Sunflower	369130
Lunar Landing LLC	Sunflower No.1	369130
Lunar Landing LLC	Sunflower No. 2	369130
Greenspun	Mayflower	2548
Greenspun	Mayflower No. 1	2548
Greenspun	Mayflower No. 2	2548
Greenspun	Mayflower No. 3	2548
Greenspun	Moonlight	2640
Greenspun	Moonlight No. 1	2640
Greenspun	Moonlight No. 2	2640
Greenspun	Starlight No. 4	2640
Greenspun	Starlight No. 5	2640
Greenspun	Starlight No. 6	2640
Greenspun	Starlight No. 7	2640
Sussman	Jolly Jane	402672
Sussman	Valley View	402672

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 17

Table 4-4 Summary of the Original Terms for the Mayflower/Greenspun Group Lease

Term: 5 Years Beginning December 1, 2007

Five additional years to December 7, 2017, plus an additional 3 year period or so long thereafter as commercial production continues

Ten Additional years to December 7, 2027, plus an additional 3 year period or so long thereafter as commercial production continues

Lease Payments: Due on Each Anniversary Date of the Lease

On regulatory acceptance - US$5,000 and 25,000 ITH shares

Each of first – fourth anniversaries, US$5,000 and 20,000 ITH shares

Each of fifth – ninth anniversaries, US$10,000, 50,000 ITH shares and 25,000 Corvus Common Shares

Each tenth to 20^th anniversaries, US$10,000, 50,000 ITH shares and 25,000 Corvus Common Shares

Work Commitments: Excess Expenditures in Any Year Can Be Carried Forward, or if under Spent the Unspent Portion Paid to Greenspun Group

Years 1-3 US$100,000 each year the lease is in effect

Years 4-6 US$200,000 each year the lease is in effect

Years 7-10 US$300,000 each year the lease is in effect

Years 11-20 US$300,000 each year the lease is in effect

Retained Royalty: Production Sliding Scale Net Smelter Return Based on Price of Gold Each Quarter

2% if gold is less than US$300 per ounce

3% if gold is between US$300 and US$500 per ounce

4% if gold is more than US$500 per ounce

Advance Minimum Royalty Payments (if not in commercial production by the twentieth anniversary, in order to extend lease for an additional three years)

Years 21-23 US$100,000 each year the lease is in effect and commercial production has not been achieved

Purchase Option:

During first 10 years property can be purchased for US$ 7.5 million plus an 0.5% NSR (if gold is less than US$ 500) or 1.0% (if gold is above US$ 500)

After the tenth anniversary the US$ 7.5 million purchase price escalates by the Consumer Price Index, using the CPI immediately prior to the tenth anniversary as a base

Table 4-5 Summary of the NBP Unpatented Lode Mining Claims

Land Holder	Claim Name	US Bureau of Land Management Serial Number
Corvus Gold Nevada, Inc.	NB 1 – NB 149	922928 – 923076
Corvus Gold Nevada, Inc.	NB 150	943108
Corvus Gold Nevada, Inc.	NB-151A	1078379
Corvus Gold Nevada, Inc.	NB 152 – NB 154	943110 – 943112
Corvus Gold Nevada, Inc.	NB-155A	1078381
Corvus Gold Nevada, Inc.	NB 156 – NB 161	943114 – 943119
Corvus Gold Nevada, Inc.	NB 162 – NB 213	989863 – 989914
Corvus Gold Nevada, Inc.	NB 214 – NB 510	1069332 – 1069628
Corvus Gold Nevada, Inc.	NB 511	1078379
Corvus Gold Nevada, Inc.	NB 512 – NB 808	1085130 – 1085426
Corvus Gold Nevada, Inc.	NB 809 – NB 865	1109343 – 1109399

4.3 Environmental Liabilities

Corvus currently has permits to conduct exploration activities at NBP with both the Nevada Division of Environmental Protection (NDEP)-Bureau of Mining Regulation and Reclamation (BMRR) and the Bureau of Land Management (BLM). Those permits allow 20 acres and 120 acres of surface disturbance on the private and public land, respectively. The permits for activities on the public lands are based on an Environmental Assessment that contains environmental baseline data on wildlife, climate and local physical characteristics.

To the extent applicable the Author knows of no significant factors and risks that may affect access or title to NBP or the right or ability to perform work on the Project. The Author Knows of no other royalties, back-in rights, payments or other agreements and encumbrances, environmental liabilities, permits or any other significant factors and risks that may affect access, title, or the right or ability to perform work on the property.

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5 NB-MLP Accessibility, Climate, Local Resources, Infrastructure and Physiography

NB-MLP is accessible from Beatty, Nevada, approximately 2.5 hours’ drive (193 kilometres, 120 miles) north of Las Vegas, Nevada via US Highway 95. US Highway 95 is the major transportation route between Las Vegas, Nevada, Reno, Nevada and Boise, Idaho. Las Vegas is serviced by a major international airport. Beatty is the closest town to the Project having a population of about 1,100 and providing most basic services.

NBP lies 10 kilometres due north of Beatty and is accessed via several dirt roads to the west from US Highway 95 between mile markers 66 and 72 (6-12 road miles north of Beatty on US 95). A network of typically well-maintained dirt roads provides access to most of the important exploration areas. MLP lies 8 kilometres due east of Beatty. MLP is accessed from US Highway 95 just one mile south of Beatty by taking the Fluorspar Canyon road east. Access around the Project is also by a network of reasonably good dirt roads.

NB-MLP is in Western Nevada’s high desert which receives about 15 centimetres (“cm”) of precipitation per year, mostly as modest snowfall in the winter and thunderstorms in the summer. The average daily temperature varies from a low of 5°C (40.8°F) in January to a high of 27 °C (80.8 °F) in July. Due to the mild climate at NBP, the operating season is year-round, though occasional thunderstorms may prohibit drilling for short periods due to safety concerns about lightning strikes.

The hills at NB-MLP are covered with sparse low brush including creosote, four-wing saltbush, rabbit brush and ephedra. The Project is in the Basin and Range province. Topographic relief is several hundred feet. Topography varies from low hills and desert plains to locally very steep, rocky and rugged hills. The elevation of the Project ranges from 1,100m (3,600 feet) to 1,500m (4,800 feet). Most of the Project is characterized by low hills separated by modest width valleys (Figure 5-1).

As described in Section 4, Corvus maintains sufficient surface rights to support mining operations; including waste disposal areas, tailings storage areas, heap leach pads and mill sites, subject to necessary permits which are defined in Section 20. Claim blocks are contiguous, power is readily accessible and Corvus has secured water rights. The towns of Beatty, Pahrump and Tonopah support an ample population for mining personnel.

Figure 5-1 Mayflower Ridge Looking to the Northwest

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6 History

The Bullfrog district is informally divided into three subdistricts: 1) Main Bullfrog; 2) North Bullfrog; and 3) Bare Mountain. Figure 6-1 shows the North Bullfrog (blue boundary) and Mother Lode (green boundary) claim blocks, the historic (green asterisks) and modern (red) producing mines and the general areas of the subdistricts. Gold was first discovered in the main Bullfrog district by Frank “Shorty” Harris and Ernest Cross on August 9, 1904 at the location of the Original Bullfrog mine (Elliot, 1966). The discovery sparked a rush of prospectors and within a few weeks the district was staked “for nine miles in all directions from the sagebrush flats, including part of the desert to the tip of every summit in sight” (Elliot, 1966). Lincoln (1923) reported that 111,805 ounces of gold and 868,749 ounces of silver were produced in the district between 1905 and 1921. According to Lincoln (1923), the Montgomery-Shoshone mine was the most important mine in the district, operating between 1907 and 1910. A number of other small mines in the Main Bullfrog and North Bullfrog subdistricts contributed to the total reported production.

The Bare Mountain subdistrict is also known historically as the Fluorine district. Gold was first discovered on the east side of Bare Mountain in 1905 (Lincoln, 1923), as prospectors flocked to the Beatty area after the discovery at Original Bullfrog mine. Gold was first discovered in Fluorspar Canyon in 1906 near the Daisy shaft (Papke, 1979). Unreported amounts of gold were produced from a number of small operations around Bare Mountain between 1907 and 1929 (Kral, 1951). Gold prospecting led to the discovery of a number of fluorite vein deposits associated with gold mineralization including the Daisy (or Crowell), Goldspar and Mary mines (Papke, 1979). Daisy was by far the largest fluorspar mine producing 204,508 tons between 1919 and 1986. Fluorspar Canyon was named because of a number of fluorite occurrences in the Daisy area.

Mercury was discovered in 1908 at the Harvey (also known as the Telluride) mine on the northeast flank of Bare Mountain just one kilometer south-southwest of the Mother Lode deposit. The Harvey mine operated intermittently between 1912 and 1943 producing 72 flasks of mercury (Bailey and Phoenix, 1944). Another mercury occurrence known as the Tip Top mine lies along the Flatiron jasperoid just 800 metres southwest of the Mother Lode deposit. The Tip Top mine has no documented production, but may have produced up to 100 flasks of mercury. The Harvey and Tip Top mines are known to have gold associated with cinnabar, but no reported gold production. Mercury was later discovered at the Thompson mine in 1929. The Thompson mine is located 5 kilometres northeast of Mother Lode, and is reported to have had only minor mercury production (Bailey and Phoenix, 1944).

Ceramic grade high-purity silica was produced at the Silicon mine between 1919 and 1929, located 1.5 kilometres northwest of the Thompson mine (Kral, 1951).

Modern exploration for precious metals in the main Bullfrog subdistrict began as early as 1982, when geologists from St. Joe Minerals Corporation became interested in the Montgomery-Shoshone area. St. Joe Minerals conducted extensive exploration in the area of the Montgomery-Shoshone and Senator Stewart mines, resulting in the discovery of the Bullfrog vein deposit in 1987 (Jorgenson et al, 1989). Several company acquisitions resulted in Barrick Gold Corporation being the final owner and operator of the mine. The Bullfrog mine produced gold and silver from three deposits including: 1) Bullfrog (open pit and underground); 2) Montgomery-Shoshone (open pit); and 3) Bonanza Mountain (open pit). Between 1989 and 1999, the Bullfrog mine produced 2.31 million ounces (“Moz”) of gold and 3.0 Moz of silver (NBMG MI-2000, page 34). The Gold Bar open pit mine, located 4 kilometres northeast of the Original Bullfrog mine, produced a small (unreported) amount of gold in the late 1980’s.

Modern exploration for precious metals in the North Bullfrog subdistrict began as early as 1974 (see Section 6.1, NBP History). There has been no reported modern production from the North Bullfrog subdistrict.

Modern exploration for precious metals in the Bare Mountain subdistrict began in the 1973 when Cordilleran Exploration Company (Cordex) staked claims in the Sterling mine area. Between 1973 and 1977, Cordex discovered and delineated bulk-tonnage sediment-hosted gold mineralization at the Sterling mine (Ennis et al, 2017). In 1978, Cordex leased the Sterling property to Saga Exploration Company (Saga). Saga continued to explore the property, and in 1980 formed the Sterling Mine Joint Venture (SMJV). The SMJV began producing gold in April 1980 (Ennis et al, 2017).

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Cordex explored the Paleozoic sedimentary rocks in Fluorspar Canyon starting in the mid-1970’s, focusing on gold mineralization associated with the fluorite occurrences. In 1979, U.S. Borax (Pacific Coast Mines) staked the original BVC claims in the Fluorspar Canyon area. U.S. Borax was primarily interested in uranium, but recognized the potential for bulk-tonnage volcanic-hosted disseminated gold mineralization, and discovered the Secret Pass deposit in 1981. In 1985, Cordex entered into a joint venture agreement with U.S. Borax, and by 1989 had earned 100% interest in the U.S. Borax holdings, which would become the core of the Daisy mine property (Roscoe Postle Associates Inc., 1995). In 1990-1991, Inter-Rock Gold, Inc., through a series of transactions, acquired an option to earn 100% interest in the Daisy mine property, subject to a 25% back-in right to Cordex. Between 1985 and 1994, Cordex was involved in the discovery and delineation of two new sediment-hosted gold deposits at Daisy South and Daisy West, as well as expanding the Secret Pass deposit. In 1994, Inter-Rock had earned 100% interest in the Daisy mine property through the funding of ongoing exploration, and Cordex (now Rayrock Mines) elected to exercise their back-in rights to hold 25% interest in the property. In 1995, the Mother Lode property was purchased by Rayrock Mines from USNGSJV (Roscoe Postle Associates, 1998) and incorporated into the Daisy mine property. The Daisy mine partnership subsequently consisted of Rayrock Mines as the operator (35%) and Inter-Rock Gold, Inc. as the majority owner (65%). The Daisy mine partnership began mining the Secret Pass and Daisy West deposits in late 1996 and ceased mining operations in late 1999. Inter-Rock sold all interest in the Daisy mine property back to Rayrock Mines in 1998. Gold production from the Daisy leach pad continued into 2001.

Reported modern production of 351,744 ounces from Bare Mountain subdistrict has come from the following deposits: 1) 212,744 ounces produced at the Sterling mine (open pit and underground) between 1980 and 2015 (Ennis et al, 2017); 2) 104,000 ounces produced from the combined Secret Pass and Daisy West open pit mines between 1996 and 2001 (NBMG MI-2000); and 3) 35,000 ounces produced from the Mother Lode open pit mine between 1989-1991 (Weiss, 1996). The discovery and production history of the MLP is presented in Section 6.2 below.

Figure 6-1 Bullfrog District Location Map

 

6.1 MLP History

In 1983, shortly after U.S. Borax had discovered the Secret Pass deposit, Galli Exploration (GEXA) prospected and staked claims in the Mother Lode-Joshua Hollow areas east of Secret Pass. In 1985, GEXA drilled 54 shallow holes in the Flatiron jasperoid area along the Fluorspar Canyon Fault. Mineralized intercepts were generally narrow, low grade, and confined to the fault zone. GEXA geologists recommended drilling the pediment area where the Fluorspar Canyon Fault was projected to intersect a northwest-trending rangefront fault. The Mother Lode deposit was discovered in 1987 with drill hole ML-59, which intersected 47m (155 ft) of 1.7 g/t (0.049 ounces per ton) gold (Mapa, 1990). GEXA also discovered the SNA deposit under pediment to the south of Mother Lode in 1987.

The Mother Lode deposit was delineated by drilling in 1987 and 1988. In 1988, GEXA formed U.S. Nevada Gold Search Joint Venture (USNGSJV) with U.S. Precious Metals and contract miner N.A. Degerstrom. Mining of the Mother Lode deposit began in 1989. Between 1989 and 1991, the USNGSJV produced ~35,000 ounces of gold at Mother Lode at an average grade of 1.8 g/t Au. The bulk of the gold production came in 1990. Mining ceased in 1991 as oxide reserves were mined-out.

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Rayrock Mines Inc. purchased the Mother Lode property in 1995 from the USNGSJV; incorporating the Mother Lode and SNA deposit resources into the Daisy mine property. Rayrock continued to explore the Mother Lode-SNA area through 1997. Rayrock also intended to mine sulfide resources at Mother Lode as a part of their Daisy mine plan but a declining gold price in 1998 prohibited this development. Some mining development was done in the Mother Lode pit between 1998 and 1999, but no separate gold production was reported.

Glamis Marigold Mining Company acquired all of the assets of Rayrock in March 1999, including 100% interest in the Daisy mine property. Glamis continued limited gold recovery from the Daisy leach pad into early 2001. The majority of the Daisy mine property was dropped by Glamis in 2001 and 2002, and the Daisy Project began reclamation work. Glamis retained a small portion of the Daisy Project property including the 13 mining claims covering the Mother Lode mine area. Goldcorp (US) Inc. acquired all of the assets of Glamis Marigold Mining Company in 2006, including the 13 claims at Mother Lode. The Mother Lode claims were transferred into Goldcorp Daisy LLC (a wholly owned subsidiary of Goldcorp (US) Inc.) in 2014. In May 2017, Corvus acquired Goldcorp Daisy LLC from Goldcorp (US) Inc. In 2017-2018, Corvus has staked 436 additional mining claims which comprise the current MLP land position shown in Figure 6-1.

In May 2016, Marigold Mining Company was notified by the Bureau of Land Management (BLM) that the BLM was closing the Plan of Operations and authorizing a reclamation cost estimate reduction of 100%. In June 2016, Goldcorp Daisy LLC received notification from the BMRR that all regulatory requirements for permanent closure of the Daisy mine had been met, and that the Water Pollution Control Permit was terminated. Goldcorp Daisy LLC was relieved of all obligations under Nevada Revised Statutes 445 A and the Nevada Administrative Code 445A. BMRR also concurred with the BLM’s decision and terminated the Reclamation Permit 0031 noting that all reclamation and surety liability associated with the permit had been successfully completed.

Sterling Gold Mining Corp (Imperial Metals) began acquiring land and staking open ground in the Daisy-Secret Pass-Mother Lode area in 2005, and by 2007 had assembled a substantial land position at northern Bare Mountain covering all the former Daisy mine property and surrounding Glamis’ small Mother Lode claim block.

6.2 NBP History

Also known as the Pioneer district, the early history of the NBP property is comingled with the main Bullfrog district. A rush of prospectors came to the Bullfrog Hills soon after the Original Bullfrog discovery, which resulted in discoveries at the Mayflower mine in 1906 and the Pioneer mine in 1907. The Pioneer and Mayflower were the principal mines at the NBP. The Mayflower mine was intermittently active between 1906 and 1940 (Kral, 1951). The Pioneer mine was intermittently active between 1908 and 1920 with 1909-1910 being the most productive years (Kral, 1951). There are no accurate production figures, but limited records suggest that head grades were between 0.5 to 1 ounce of gold per ton at both Mayflower and Pioneer. Underground development at Sierra Blanca, Jolly Jane, Savage Valley and YellowJacket between 1910 and 1914 had no reported production.

Modern exploration for precious metals at NBP started with Cordex at the Connection prospect in 1974. Table 6-1 outlines a number of companies that worked in various target areas up to 1996. These programs consisted of a variety of activities including surface mapping and sampling, underground mapping and sampling, and drilling. Through the 1996 Barrick program, approximately 249 rotary and reverse-circulation drill holes were drilled on the Project by eight different operators (see Section 10 for detailed description of these programs). Barrick had dropped all interest in the NBP area by 1998.

Table 6-1 Summary of Companies That Explored NBP

Company	Years of Activity	Principal Target
Cordex Exploration Co.	1974-1982	Connection, Pioneer
US Borax Incorporated	1982	Mayflower
GEXA/Galli Exploration	1984-1991	Pioneer, Connection
CR Exploration	1984-1985	Mayflower
Western States	1987	West Mayflower
Bond Gold/Sunshine JV	1988-1994	Sierra Blanca, YellowJacket
Pathfinder Minerals	1991, 1992	Pioneer
Barrick Gold Corporation	1995-1996	Jolly Jane, Sierra Blanca, Mayflower, etc

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With the downturn in gold price around the start of the 21^st Century, interest in the NBP area was essentially nonexistent. RGC became attracted to the North Bullfrog area in late 2005, and began staking unpatented claims and acquiring leases on patented claims in 2006. In March 2007, RGC granted ITH the right to earn an interest in the NBP and thereafter formed the NBPJV. In December 2007, ITH completed a lease of the Mayflower property, which was included in the NBPJV. Following the execution of the NBPJV option/joint venture agreement, ITH commenced active exploration on the NBP. In October 2008, RGC completed a lease of the Connection property, which was also included in the NBPJV. On August 4, 2009, ITH purchased RGC’s interests in the property and continued the exploration program as sole owner/lessor. On August 26, 2010, ITH spun out Corvus as a separate public company in a transaction that resulted in Corvus acquiring all of the interest and responsibilities in the NBP.

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7 Geological Setting and Mineralization

7.1 Regional Geological Setting

NB-MLP lies within the Walker Lane mineral belt and the Southwestern Nevada Volcanic Field (SWNVF). The regional stratigraphy includes a basement of Late Proterozoic to Late Paleozoic metamorphic and sedimentary rocks. Basement rocks are overlain by a thick pile of Miocene volcanic and lesser sedimentary rocks of the SWNVF, ranging in age from ~15-7.5 Ma (Figure 7-1). The pre-Tertiary rocks exhibit large-scale folding and thrust faulting, having been subjected to compressional deformation associated with multiple pre-Tertiary orogenic events. The stratigraphy of the SWNVF is dominated by ash flow tuff sheets erupted from a cluster of nested calderas known as the Timber Mountain Caldera Complex. The southwestern edge of the caldera complex lies approximately eight kilometres northeast of the MLP, and ten kilometres east of the NBP (Figure 7-1). The stratigraphy of the SWNVF includes voluminous ash flow tuff sheets, smaller volume lava flows, shallow intrusive bodies, and lesser sedimentary rocks. Many of the volcanic units exposed around NB-MLP include ash flow tuffs that originated from the caldera complex. Other volcanic units are locally sourced outside of the caldera complex, particularly at the NBP.

The Bullfrog and Fluorspar Hills comprise a somewhat isolated structural domain within the Walker Lane, where both pre-Tertiary and Miocene rocks have been subjected to large-scale, W- to WNW-directed, syn-volcanic extension (i.e. down-to-the-west normal faulting and east-tilting of stratigraphy). Extensional faulting was coincident with magmatism and volcanic activity between ~15-9.4 Ma. Hydrothermal alteration and gold mineralization were also episodic through this time period. If present, through-going right-lateral faults of the Walker Lane are poorly exposed in the SWNVF. One possible example of a NW-trending, through-going Walker Lane structure cuts through the historic Silicon and Thompson mine areas to the northeast of the MLP (Figure 1). Despite the dominance of caldera volcanism in the region, little or no mineralization is associated with any caldera ring fracture system. Rather, mineralization is typically associated with extensional faults outside of the caldera complex.

Extension is accommodated by the Bullfrog Hills Fault System (BHFS); a complex group of kinematically linked faults that facilitate WNW-directed extension and east-tilted block rotation. The primary structure of the BHFS is the Southern Bullfrog Hills Fault (SBHF). The SBHF is an east-west-trending, north-dipping, district-scale, low-angle detachment fault (Eng et al. 1996). West of Beatty, in the main Bullfrog district, the SBHF separates Proterozoic metamorphic rocks in the footwall from weakly metamorphosed Paleozoic sedimentary and Tertiary volcanic rocks in the hanging wall. East of Beatty, in the Bare Mountain sub-district, the same fault is called the Fluorspar Canyon Fault (FCF, Figure 7-1). The FCF cuts up-section from west to east such that in Fluorspar Canyon it separates Paleozoic sedimentary rocks (footwall) from brittle Tertiary volcanic rocks (hanging-wall). The FCF continues to cut up-section to the east until it eventually separates brittle Tertiary rocks (footwall) from brittle Tertiary rocks (hanging-wall) at Mother Lode. The magnitude of displacement along the SBHF-FCF appears to decrease from west to east across the greater Bullfrog district. The northward dip of the SBHF-FCF generally increases from west to east including: ~20^o north at the Bullfrog mine, 25-30^o north in lower Fluorspar Canyon, 45^o north at the Secret Pass mine, and 55-65^o northwest at Mother Lode.

Secondary structures of the BHFS include large-displacement, NNW- to NNE-trending, moderately to steeply west-dipping, down-to-the-west, normal faults. These faults accommodate the east-tilting of the Tertiary units throughout the Bullfrog and Fluorspar Hills. Hydrothermal alteration and gold mineralization are often spatially associated with these large-displacement faults. Such faults are expected to have listric shapes at depth. The MP fault, which hosts the Bullfrog vein at the Bullfrog mine, is an example of a listric fault. Another example is the Contact Fault, which truncates the north side of the Montgomery-Shoshone deposit. The Contact fault also hosts low-grade mineralization under Rhyolite Valley. The Road Fault at the NBP is the northern continuation of the Contact Fault. Such faults are interpreted to sole into the SBHF-FCF detachment fault at depth.

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Figure 7-1 Regional Geology of the Greater Bullfrog District

 

The long period of syn-volcanic extension in the western SWNVF between ~15-9.4 Ma (depending on location) includes at least two periods of accelerated extension documented in the Bullfrog and Fluorspar Hills. Connors, et al. (1998) have identified a major period of extension between ~12.7 Ma and 11.6 Ma culminating with the eruption of the Timber Mountain Group ash-flows between 11.6 and 11.45 Ma. Evidence for this period of extension lies west of Mother Lode where the 12.7 Ma Tiva Canyon Tuff is tilted up to 45^o east, but the nearby 11.6 Ma Rainier Mesa Tuff is essentially horizontal. Block rotation of up to 45^o is documented between 12.7 and 11.6 Ma in the Fluorspar Hills. Mineralization at Mother Lode is coincident with the onset of this period of accelerated extension. Connors et al. (1998) postulate a second period of accelerated extension between 11.4 Ma and 10.5 Ma, which resulted in major block rotation, rapid erosion and the deposition of the Rainbow Mountain Debris Flow Sequence. In the Mayflower area, the base of the Rainbow Mountain Debris Flow Sequence is tilted as much as 55^o east, whereas the top of the sequence is tilted only 25^o east. Block rotation of up to 30^o is documented in this area while the Mayflower basin was filling with debris. In the southern Bullfrog Hills east of Rainbow Mountain, the 11.45 Ma Ammonia Tanks tuff is tilted as much as 70^o east, whereas the base of the overlying Rainbow Mountain sequence is tilted only 35^o east. The block rotation of 35^o is documented in this area between 11.4 and 10.5 Ma. Extensional faulting continued through ~9.5 Ma, with an additional 25^o of eastward rotation of the Rainbow Mountain Sequence. Extensional faulting ceased in the Bullfrog Hills prior to the eruption of the relatively flat-lying 9.4 Ma Pahute Mesa Tuff (Connors, et al., 1998). The 10 Ma age of Bullfrog and Mayflower mineralization and the 9.5-10.2 Ma age of the Eastern Steam-heated Zone at NBP coincide with the culmination of extensional tectonism in the Bullfrog Hills.

Extensional faulting in the Bullfrog and Fluorspar Hills created fault-bounded sedimentary basins which filled with basement- and volcanic-derived sediments (e.g. the Sedimentary Rocks of Joshua Hollow, the Jolly Jane Formation, and the Rainbow Mountain Debris Flow Sequence). During younger periods of extension, older normal faults in the hanging-walls of large-displacement listric faults may have experienced significant reactivation and subsequent eastward rotation, such that they may exhibit relative reverse displacement.

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7.2 MLP Geology

7.2.1 Stratigraphy

The stratigraphy of the Fluorspar Hills has been described in pamphlets that accompany published geologic maps by Monsen et al. (1992) and later by Fridrich et al. (2007). The local stratigraphy at MLP has been refined by Corvus based on pit mapping and exploration drilling. Table 7-1 outlines the MLP stratigraphy using map symbols consistent with those used by Corvus at the NBP. Where possible, the map symbols of Monsen et al. (1992) and Fridrich et al. (2007) have been preserved. Brief descriptions of units present in the MLP area are given below. Figure 7-2 is a simplified geologic map of the Mother Lode deposit area.

7.2.1.1 Paleozoic Sedimentary Rocks – Psd, Psq, Pss

Due to the position of the main claim block, surface exposures of Paleozoic sedimentary rocks are limited at the MLP (Figure 7-2). The Paleozoic rocks encountered in drilling are differentiated only by lithofacies (Table 7-1). The lithofacies-based units are informally correlated to major Paleozoic units, and interpreted as separated by thrust faults. The most abundant lithology is a thick sequence of light to medium grey massive dolomite (Psd), which correlates with the Silurian Lone Mountain Dolomite (Monsen et.al., 1992). Overlying the dolomite in drill holes and in outcrop south of the Mother Lode pit is a heterogeneous siliciclastic sequence (Psq) dominated by quartzite with lesser dolomitic quartz sandstone, sandy dolomite, and sandy limestone. The Psq sequence is interpreted to be a highly attenuated, dismembered thrust slice including Eureka Quartzite, Antelope Valley Limestone and possibly other Ordovician through Devonian lithologies. The Psq unit structurally overlies Psd along a thrust fault identified herein as the SNA thrust fault (Figure 7-2). Overlying the Psq sequence is another dismembered sequence dominated by black siliceous graphitic argillite with lesser dark grey limestone and dolomite (Pss). Lithologies of the Pss sequence resemble Mississippian and Devonian lithologies exposed in the Joshua Hollow area south of Mother Lode. The Pss sequence is interpreted to be a dismembered thrust slice overlying Psq along a second thrust fault identified herein as the Joshua thrust.

Despite variations in thickness, the Paleozoic stratigraphy described above is recognized consistently in reverse circulation drill holes at Mother Lode. The SNA thrust places older Ordovician(?) rocks over younger Silurian rocks. The Joshua thrust places younger Devonian and Mississippian(?) rocks over older Ordovician(?) rocks. Both thrust faults appear to dip moderately to gently east. Paleozoic sedimentary rocks, primarily Psd, host 18% of the samples >0.2 g/t gold in the Mother Lode assay database.

The age of the platy, silty carbonate rocks in the southeast corner of the Mother Lode pit has been the subject of much debate by previous workers. The well-indurated and variably folded nature of the platy rocks suggests they are older pre-Tertiary rocks. Based on Corvus drilling, the platy rocks overlie Tertiary volcaniclastic and conglomeratic sediments. The current interpretation is that these platy rocks are indeed Tertiary in age and are assigned to the Sedimentary Rocks of Joshua Hollow (SRJH, see Section 7.2.1.1.2 below). If the platy rocks can be determined through micro-fossil work to be pre-Tertiary, then this exposure would be interpreted as a gravity slide block of older Paleozoic rocks shed into the SRJH basin.

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Figure 7-2 Simplified Geologic Map of the Mother Lode Area

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Table 7-1 Summary of the Stratigraphy of the Mother Lode Property

Au min		Major Unit Name	Symbol	Formation	Description
		Quaternary Cover	Qc		Unconsolidated alluvial deposits, talus
		Gravels of Sober-up Gulch	Tgs		Older unconsolidated terrace gravels
		Timber Mountain Group	Tma	Ammonia Tanks Tuff	Moderate to densely welded crystal-rich rhyolite tuff
			Tmr	Rainier Mesa Tuff	Moderate to densely welded crystal-rich rhyolite tuff
		Pre-Rainier Mesa Tuffs and Lavas	Tprr			Rhyolite lava flows
			Tprt			Bedded crystal-lithic tuff
	Mother Lode	Pre-Rainier Mesa Debris Flow Breccia	Tox			Debris flow bx between Tp and Tprt
		Paintbrush Group	Tp	Tpc -Tiva Canyon Tuff	Aphanitic phenocryst-poor welded tuff
				Tpt - Topopah Spring Tuff	Aphanitic phenocryst-poor welded tuff
Secret Pass		Crater Flat Group	Tc	Tcb - Bullfrog Tuff	Variably welded rhyolitic crystal tuff
				Tct - Tram Tuff	Variably welded rhyolitic crystal tuff
	Mother Lode	Lithic Ridge Tuff	Tlr	Lithic Ridge Tuff		Rhyolite porphyry lava flows?
						Tys - sediments at top of Lithic Ridge Tuff
						Tlr -biotite-chlorite-rich lithic-crystal tuff
		Sedimentary Rocks of Joshua Hollow	Tjs	Sedimentary Rocks of Joshua Hollow		Tjs - fine grained carbonaceous sediments
			Tip	Rhyolite quartz-porphyry dikes
			Tjs Tjvs Tjc	Sedimentary Rocks of Joshua Hollow		Tjs - fine grained carbonaceous sediments
						Tjvs - volcaniclastic sandstone
						Tjt - grey welded tuff
						Tjc - gritty pebble-cobble conglomerate
SNA			Tjs1			Tjs1 - older tuffaceous sediments
		Paleozoic Basement	Ps/Pss	Dev. & Miss. argillite, siltstone, limestone?		Upper Plate of Joshua Thrust
			Psq	Ord. Eureka Quartzite and Antelope Valley Limestone?		Upper Plate of SNA Thrust
			Psd	Silurian Lone Mtn Dolomite		Lower Plate of SNA Thrust

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7.2.1.2 Sedimentary Rocks of Joshua Hollow

The Sedimentary Rocks of Joshua Hollow (SRJH) comprise a heterogeneous sedimentary sequence subdivided into five lithofacies units including: 1) older tuffaceous sediments (Tjs1); 2) conglomerate (Tjc); 3) volcaniclastic sandstone (Tjvs); 4) grey welded tuff (Tjt); and 5) fine-grained carbonaceous sediments (Tjs). The stratigraphic order is generally ascending, but the subunits typically inter-finger with each other laterally and are locally repeated vertically. There is substantial lithologic variability in the SRJH between drill holes. The SRJH are collectively up to 150 metres thick in the Mother Lode area. Tjvs and Tjs comprise the bulk of the SRJH sequence, and are the only sub-units exposed in the Mother Lode pit. Tjc, Tjvs and Tjs comprise a generally fining-upward sequence deposited in a local Tertiary basin marginal to the ancestral Bare Mountain structural high. The SRJH were originally named by Monsen et al. (1992) for sparse recessive exposures south of Mother Lode in lower Joshua Hollow. The age of the SRJH is unknown, but pre-dates the 14 Ma quartz-porphyry rhyolite dike swarm of eastern Bare Mountain. The SRJH are considered time-stratigraphic equivalents of the Tr1 sediments of Eng et al. (1996) in the southern Bullfrog Hills (>14 Ma). It is possible that the lower SRJH includes sediments as old as Oligocene. All subunits of the SRJH are known to be mineralized at Mother Lode. The SRJH hosts 40% of the samples >0.2 g/t gold in the Mother Lode assay database.

7.2.1.2.1 Basal Tuffaceous Sediments – Tjs1

The basal sedimentary sequence includes variegated pink, light green and light brown tuffaceous siltstone, fine tuffaceous sandstone, and tuffaceous-conglomeratic sandstone. When present, Tjs1 lies unconformably on Paleozoic rocks along the basal Tertiary unconformity. Thickness varies from 0-30 metres in drill holes, suggesting the unit was deposited on an erosional unconformity of significant relief or may have been subjected to erosion. Tjs1 is not known to crop out anywhere in the Mother Lode area. Tjs1 is distinctive from the overlying Tjvs and Tjs sub-units, suggesting that it was derived from a different provenance area. Tjs1 may be, in part, a time-stratigraphic equivalent of the Oligocene Titus Canyon Formation recognized elsewhere in the Bullfrog district.

7.2.1.2.2 Conglomerate – Tjc

Overlying and interbedded with Tjs1 is a distinctive siliceous to locally calcareous conglomerate unit. The unit includes fine to coarse angular lithic gritstone, gritty pebble conglomerate and coarse pebble to cobble conglomerate. Tjc contains clasts of a variety of Paleozoic sedimentary lithologies including chert, quartzite, dolomite, limestone, sandstone and siltstone. Tjc also contains clasts of Mesozoic (?) granitoid rocks, Tjs1 lithologies, and minor Tertiary volcanic rocks. Where Tjs1 is absent, Tjc conglomerate lies directly on Paleozoic rocks along the basal Tertiary unconformity. Tjc often contains conspicuous rounded, broken, deformed and healed Paleozoic clasts that resemble those found in conglomerates mapped as Titus Canyon Formation elsewhere in the Bullfrog district. Tjc varies in thickness from 0-30 metres, apparently being deposited in channels along an unconformity of significant relief. Because of their inter-fingering nature, Tjc and Tjs1 have been lumped together as a single unit for the purpose of geologic modeling. The combined Tjc-Tjs1 units host 5% of the samples >0.2 g/t gold in the Mother Lode assay database.

7.2.1.2.3 Volcaniclastic Sandstone - Tjvs

Overlying and locally interbedded with Tjc are intervals of fine to coarse, bedded to non-bedded, calcareous and siliceous volcaniclastic sandstone. Tjvs sandstone typically exhibits an equigranular salt and pepper texture as a result of detrital grains of white feldspar, black biotite, minor grey quartz, and small black carbonaceous lithic fragments. Highly irregular rip-up clasts of carbonaceous Tjs up to 20 centimetres are common within the volcaniclastic sandstones. The drilled thickness of individual sandstone intervals varies from <1 to >30 metres. Volcaniclastic sandstone is interpreted as channel-filling debris derived predominantly from Tertiary volcanic rocks. Tjvs is well-mineralized where it exhibits the light grey color of strong illite-pyrite alteration. Tjvs hosts 18% of the samples >0.2 g/t gold in the Mother Lode assay database.

7.2.1.2.4 Welded Tuff - Tjt

A distinctive light grey welded tuff subunit has been recognized in a few of the northeastern-most drill holes. The tuff is similar in color and composition to the volcaniclastic sandstone but exhibits a welding fabric defined by white flattened pumice clasts. The tuff subunit varies in drilled thickness from 10 to 30 metres. It generally occurs in the lower portion of the SRJH and is expected to thicken to the northeast in future drilling. Tjt is mineralized in every hole that it has been recognized. Tjt may prove to be a significant mineralized marker unit in future drilling. The grey welded tuff currently hosts <1% of the samples >0.2 g/t gold in the Mother Lode assay database.

7.2.1.2.5 Carbonaceous Sediments - Tjs

Fine grained carbonaceous sediments comprise the bulk of the upper portion of the SRJH. Lithofacies include variably carbonaceous, mostly calcareous to locally siliceous, laminated mudstone and siltstone, massive or flaser-bedded water-lain tuff, and variably argillaceous laminated lacustrine limestone (i.e. marl). Carbonaceous sediments are commonly interbedded with lenses of Tjvs volcaniclastic sandstone. The drilled thickness of individual intervals of Tjs varies from 10-50 metres within the SRJH. Some of the highest grades in the Mother Lode deposit are hosted in Tjs carbonaceous sediments. Tjs hosts 17% of the samples >0.2 g/t gold in the Mother Lode assay database.

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7.2.1.3 Lithic Ridge Tuff - Tlr

The Lithic Ridge Tuff is recognized by Corvus geologists as a distinct stratigraphic unit in drill holes and on the north high-wall of the Mother Lode pit. The tuff had been erroneously logged by previous operators as a xenolith-rich porphyritic dacite intrusive phase. The Lithic Ridge Tuff consists of poorly to moderately welded, generally lithic-rich, quartz-biotite-feldspar-bearing ash-flow tuff. Lithic content varies from <5 to 20+%, including common green altered porphyritic dacite clasts and a variety of Paleozoic clasts. The Lithic Ridge Tuff overlies the upper beds of the SRJH along an irregular mixed contact zone, suggesting a significant volume of SRJH soft sediments were ripped-up into the base of Tlr as it was being deposited. The Lithic Ridge Tuff is a well-known regional ash-flow unit of the SWNVF dated at 14 Ma (Sawyer et al., 1994). Tlr is ubiquitously hydrothermally altered, sulphidized and locally mineralized in the Mother Lode deposit area. Tlr hosts 4% of the samples >0.2 g/t gold in the Mother Lode assay database.

7.2.1.4 Rhyolite Porphyry Dikes – Tip

Rhyolite porphyry dikes in the Mother Lode area are part of a NNE-trending, west-dipping swarm of hydrothermally altered dikes that intrude pre-Tertiary rocks for over 14 kilometres along the eastern flank of Bare Mountain. The dikes at Mother Lode comprise the northernmost exposures of the dike swarm, and the only exposures known to intrude Tertiary rocks. The dikes form somewhat tabular to highly irregular, discontinuous bodies in both the hanging wall and footwall of the Fluorspar Canyon Fault. Where relatively unaltered, the rhyolite has an aphanitic to granophyric groundmass with phenocrysts of quartz, plagioclase, sanidine, biotite and minor hornblende. Quartz, biotite and plagioclase phenocrysts >5 mm are common. Where intensely altered and mineralized, the groundmass is replaced by a mixture of illite-smectite. Mafic minerals are progressively altered to chlorite and replaced by pyrite in mineralized rhyolite. Feldspars are altered to calcite, illite, or adularia. Rhyolite porphyry dikes have yielded age dates ranging from 14.4 to 13.8 Ma (Weiss, 1996), with an accepted average age of ~14 Ma. An interpretation from new drilling at Mother Lode is that rhyolite dikes intrude the Lithic Ridge Tuff and may have formed a local lava flow overlying the Lithic Ridge Tuff. Dike emplacement and the eruption of the Lithic Ridge Tuff are closely related to magmatic activity at ~14 Ma. Tip is the dominant host lithology at Mother Lode, hosting 33% of the samples >0.2 g/t gold in the assay database.

7.2.1.5 Younger Sediments – Tys

A second sedimentary sequence has been recognized overlying the Lithic Ridge Tuff. The unit includes limonitic to carbonaceous, siliceous to calcareous, fine volcaniclastic sandstone and siltstone similar to Tjs and Tjvs. North of the Mother Lode pit, the Tys unit varies in thickness from 0 to 10 metres and increases to a maximum known thickness of 70 metres in ML 18-072 west of the pit. Tys is lumped in with Tlr in the geologic model. Tys is not known to be mineralized at Mother Lode.

7.2.1.6 Crater Flat Group – Tct and Tcb

The regionally extensive Crater Flat Group has been described in detail by Carr et al. (1986). The Tram Tuff (Tct) and Bullfrog Tuff (Tcb) members crop out to the northeast of Mother Lode and in Fluorspar Canyon (Fridrich et al., 2007). Neither tuff has been identified in-situ in the Mother Lode drilling area. The tuffs are similar in texture and composition, both being moderately crystal-rich welded tuffs that exhibit ubiquitous hydrothermal alteration. The younger Bullfrog Tuff has been dated at 13.25 Ma (Sawyer et al., 1994). The Bullfrog Tuff is the primary host unit at the Secret Pass Deposit.

7.2.1.7 Paintbrush Group – Tpt and Tpc

In the Fluorspar Hills, the Paintbrush Group is comprised of two members: the 12.8 Ma Topopah Springs Tuff (Tpt), and 12.7 Ma Tiva Canyon Tuff (Tpc) (Fridrich et al., 2007 and Sawyer et al., 1994). The Paintbrush tuffs are typically reddish brown, aphanitic, phenocryst-poor and densely welded. Paintbrush tuffs are distinctly shard-rich and phenocryst-poor compared to other major ash flow tuffs of the SWNVF. The Paintbrush Group has an exposed thickness of nearly 600 metres north of the Secret Pass Deposit but is not recognized in the Mother Lode area. The contact between the underlying Bullfrog Tuff and basal vitrophyre of the Topopah Springs Tuff is well-exposed on the north high-wall of the Secret Pass Pit.

7.2.1.8 Pre-Rainier Mesa Debris Flow Breccia – Toxh and Toxm

A substantial thickness of heterolithic debris flow breccia (Toxh) overlies the Tiva Canyon Tuff ~1.5 kilometres west of the Mother Lode deposit. Such breccias in the Mother Lode area are mapped as Tox (i.e. older breccia) by Fridrich et al. (2007). The Tox unit is analogous to younger post-Rainier Mesa debris flow breccia deposits of the Rainbow Mountain Sequence at NBP (see Section 7.3.1.8.1). Tox is predominantly heterolithic but may locally contain monolithic breccias (Toxm) of rhyolite porphyry, SRJH, Lithic Ridge Tuff, or Crater Flat Group tuffs. Heterolithic debris flow breccias are generally non-bedded, very poorly sorted, and contain sand- to large boulder-size clasts of all older units. Tox exists only in the hanging wall of the FCF, where it unconformably overlies Tp, Tip, Tlr, Tys and SRJH. Tox is interpreted to have been deposited on an angular erosional unconformity of significant relief. Up to 150 metres of Toxh has been penetrated in the hanging-wall of the Fluorspar Canyon Fault west of the Mother Lode pit. Tox is interpreted as mass-wasted scarp breccias and alluvial fan deposits that have been shed off fault scarps, including the FCF, during the period of accelerated extension between 12.7-11.6 Ma. Based on stratigraphic relations, Tox is considered a post-mineral unit at Mother Lode (post-12.7 Ma), but it is also known to contain gold. The distribution of oxide gold mineralization in Tox suggests that the mineralization is detrital in nature, having been mass-wasted off the mineralized footwall of the FCF. Tox hosts 2% of the samples >0.2 g/t gold in the assay database.

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7.2.1.9 Pre-Rainier Mesa Tuffs And Lavas – Tprt and Trr

The lower part of the Timber Mountain Group includes rhyolitic tuffs and lava flows that are known as the Pre-Timber Mountain tuffs (Tprt) and lavas (Tprr) respectively. These units are undated but occupy the stratigraphic position between the Paintbrush Group (or Tox, if present) and the large volume Timber Mountain tuffs (see Section 7.2.1.10). Tprt consists of light-colored, non-welded to bedded, lapilli rich, pumaceous lithic-rich crystal tuff. Tprt is up to 150 metres thick in the Mother Lode drilling area. Tprt appears to blanket paleo-topography cut on the Toxh debris flow sequence. Tprt was deposited during or just after the period of accelerated extensional tectonism associated with the deposition of Tox. Tprt is not known to be mineralized at Mother Lode but exhibits ubiquitous steam-heated alteration north and west of the deposit.

Tprt represents pyroclastic deposits associated with Tprr rhyolite volcanism. Tprr consists of aphanitic to fine porphyritic, massive to flow-banded, rhyolite lava flows (i.e. flow-dome complexes) at the Twisted Canyon and Sawtooth areas of MLP and at the Spicerite target areas at NBP (see Figure 9-2). The rhyolite varies in appearance from relatively fresh, grey to black, glassy perlite or obsidian, to tan to pale green strongly devitrified zeolitic rhyolite. Tprr is not present in the Mother Lode area.

7.2.1.10 Timber Mountain Group – Tmr and Tma

Regionally there are two large-volume ash flow sheets that comprise bulk of the Timber Mountain Group. These include the 11.6 Ma Rainier Mesa Tuff (Tmr) and 11.45 Ma Ammonia Tanks Tuff (Tma, Sawyer et al., 1994). The Timber Mountain tuffs are typically densely welded, crystal-rich ashflow tuffs, with 15-20+% phenocrysts of quartz, feldspar and biotite. The units can be distinguished from each other by the presence of sphene and bluish chatoyant sanidine phenocrysts in the Ammonia Tanks Tuff. In the Mother Lode area, the Rainier Mesa Tuff directly overlies steam-heated Tprt. The Rainier Mesa Tuff may overlie Tprr or Tp elsewhere in the district. The Rainier Mesa Tuff is sub-horizontal, relatively unaltered, and entirely post-mineral in the Mother Lode area. The Ammonia Tanks Tuff is recognized only at the Sawtooth target area (see Figure 9-1).

7.2.1.11 Gravels of Sober-Up Gulch - Tgs

Older unconsolidated gravel deposits of late-Miocene to Pliocene age occur at the Willys target area along the eastern rangefront area of Bare Mountain, and at the Twisted Canyon and Baileys Gap target areas on the east side of Oasis Valley (see Figure 9-1). The gravels form sub-horizontal terraces that are commonly, deeply incised. Underlying older rocks are often poorly exposed in the bottoms of incised ravines below the terrace gravels. The older gravels at the MLP are correlated to the Gravels of Sober-up Gulch on the west side of Oasis Valley at NBP. The thickness of Tgs is highly variable, but not expected to exceed 100 metres. Monsen et. al. (1992) reports an 8.2 Ma age for a tuff unit within the Gravels of Sober-up Gulch east of Bare Mountain.

7.2.1.12 Quaternary Cover - Qc

The Mother Lode deposit is largely covered with unconsolidated Quaternary gravels. Quaternary gravels range in thickness from 1-20 metres in the Mother Lode drilling area. The Quaternary gravels are variably caliche-cemented and form stable vertical walls in the Mother Lode pit.

7.2.2 Structure

The most notable structural feature at Mother Lode is the Fluorspar Canyon Fault (FCF). The NE-trending, NW-dipping FCF is exposed at the Flatiron target where it hosts an auriferous, cinnabar-bearing, silicified fault breccia (Figure 7-2). The fault breccia is juxtaposed against a massive jasperoid hosted in the footwall dolomite. Drilling at Flatiron indicates a dip of 65^o northwest from the outcrop to 300 metres down-dip. The FCF continues to the northeast from Flatiron for about 700 metres where it intersects a north-trending rhyolite dike-filled structural zone (Figure 7-2). Just before it enters the Mother Lode pit, the FCF is deflected to the northerly strike of the dike-filled structural zone. Displacement along the FCF appears to be transferred to the dike orientation at this structural intersection. This intersection is a vital structural feature facilitating the formation of the Mother Lode deposit. The earliest movement on the FCF clearly post-dates dike emplacement.

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The rhyolite dike swarm along eastern Bare Mountain is hosted by structures that exhibit little or no discernable displacement. Displacement on the dike-filled structure just south of Mother Lode appears to be nil as the dike does not offset the SNA thrust (Figure 7-2). However, cross section analysis of the Mother Lode deposit indicates notable displacement on many dike margins north of the intersection with the FCF. Fault gouge on dike margins is a common occurrence in the Mother Lode deposit. The evidence suggests that large-scale displacement associated with the FCF has been transferred to the dike swarm orientation and has reactivated older sub-parallel dike-filled structures north of the structural intersection. Reactivation of dike-filled structures has enhanced porosity and permeability for fluids to ascend around and through the structurally modified dikes. The dike zone widens rapidly to the north of the intersection, and branches into three or more mineralized dikes (Figure 7-3). The subparallel FCF and dike swarm strike northerly and dip 50-70^o west for ~500 metres to the northern limit of drilling. Stratigraphic relationships suggest the main splay of the FCF has 400 metres or more of down-to-the-west normal displacement within the Mother Lode deposit.

Mineralized rhyolite occurs in both the hanging-wall and footwall of the FCF. There is evidence that dikes cut the unconformity, but no conclusive evidence that dikes cut the FCF. Rather, cross section analysis suggests that both footwall dikes and hanging-wall dikes are truncated by the main splay of the FCF (Figure 7-3). The FCF bisects and offsets the large rhyolite body in the Mother Lode pit, where mineralization is only present in the footwall portion of the rhyolite. Elsewhere in the deposit, mineralized rhyolite exists only in the hanging-wall of the FCF (Figure 7-3). The implication is that additional sub-parallel mineralized dikes may be found in the hanging-wall of the FCF west of the known deposit. The FCF apparently dismembers individual dikes as it propagates through the dike swarm in a subparallel orientation. Reactivated older dike-filled structures may be the primary structural conduits for ascending fluids. However, the main splay of the FCF is mineralized at both Mother Lode and Flatiron. Deeper drilling has extended the deposit to the west, where it appears that the FCF plays a more significant role as a high-grade structural feeder. Future drilling will continue to test the FCF as a primary feeder structure at depth to the west.

Figure 7-3 Cross Section 4084410N Looking North through the Mother Lode Deposit

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The timing of the latest movement along the FCF in the Mother Lode area is bracketed by the 12.7 Ma Tiva Canyon Tuff and 11.6 Ma Rainier Mesa Tuff. The Tiva Canyon Tuff is tilted as much as 45^o east in the hanging-wall of the FCF just 1.5 kilometres west Mother Lode. Just northwest of Mother Lode, the Rainier Mesa Tuff is sub-horizontal, exhibiting little or no east-tilting in the hanging-wall of the FCF. The evidence suggests that an accelerated period of extensional faulting and eastward block rotation, accommodated by the FCF, occurred between 12.7 and 11.6 Ma in the Mother Lode area. The onset of this period of extension is coincident with the age of Mother Lode mineralization. Extension continued along the FCF after mineralization at Mother Lode, resulting in un-mineralized hanging-wall rocks being juxtaposed against mineralized footwall rocks. The amount of post-mineral displacement on the FCF is unknown. Structural and stratigraphic relationships suggest that a faulted-off portion of the Mother Lode deposit may exist in the hanging-wall of the FCF at depth to the west.

A major unconformity between Paleozoic and Tertiary rocks appears to have played a role in dike emplacement and the propagation of ascending hydrothermal fluids. The porosity and permeability contrast between the rigid, non-porous Paleozoic rocks and the less rigid, highly porous Tertiary rocks apparently allowed the mineralizing system to open and broaden above the unconformity. Rhyolite dikes appear to expand in width just below and above the unconformity (Figure 7-3). The grade and thickness of mineralization generally increases with proximity to, and above, the unconformity. Gold grade in all mineralized units generally increases with proximity to mineralized dikes. The grade and thickness of mineralization is significantly better adjacent to dikes in the Tertiary sediments above the unconformity.

A large-scale, low-amplitude, north-plunging antiform involving Sedimentary Rocks of Joshua Hollow (SRJH) and the Lithic Ridge Tuff can be observed on the north wall of the Mother Lode pit. On cross sections, the antiform is mimicked by the basal Tertiary unconformity (Figure 7-3). The feature appears to be a large-scale drag-fold in the footwall of the FCF. Bedding in the west limb of the fold appears to steepen to the west and merge with the FCF.

There is evidence of compressional deformation (i.e. folding) within the SRJH in the Mother Lode pit and in core holes. Small-scale, low amplitude folding is interpreted as gravity-related compressional deformation of soft sediments in a rapidly subsiding and filling Tertiary basin during active extensional tectonism.

7.2.3 Mineralization

Mother Lode is characterized as a sediment, intrusive, and locally volcanic-hosted disseminated gold deposit. Mineralization most closely resembles Carlin-type sediment-hosted gold deposits of north-central Nevada. Weiss (1996) was the first to recognize and document the similarity to Carlin-type deposits. Weiss (1996) also cites evidence for a large buried porphyry-type magmatic system associated with the rhyolite dike swarm at eastern Bare Mountain. The Mother Lode deposit formed at ~12.7 Ma, which is much younger than the typical ~40 Ma age in north-central Nevada. The nature of mineralization is rather passive and with very low introduction of secondary silica, suggesting it may have formed at a shallower depth and at lower temperature than typical Carlin-type deposits. Mineralization exhibits geochemical associations between Au and As-Sb-Hg-Tl-Te-Bi-F, with very low Ag and base metals.

The Mother Lode deposit model consists of structurally and stratigraphically-controlled disseminated gold mineralization hosted primarily in rhyolite porphyry dikes (Tip) and the Sedimentary Rocks of Joshua Hollow (SRJH). Lesser volume hosts include Paleozoic sedimentary rocks (Psd and Psq), Tertiary volcanic rocks (Tlr), and debris flow breccias (Tox). Mineralization in Tip, SRJH and Tlr is mostly sulphide, but may be oxidized depending on depth. Mineralization in Psd, Psq and Tox is mostly oxide. The current interpretation is that oxide mineralization in Tox is detrital, with mineralized rock having been mass-wasted as scarp breccias into the hanging-wall of the FCF. Paleozoic rocks are most commonly mineralized in proximity to dike margins but are also mineralized along subtle non-dike-filled structures and as pseudo-stratiform jasperoid bodies.

The primary structural control feeding mineralization at Mother Lode is a series of north-trending, 50-70^o west-dipping rhyolite dike-filled structures. Mineralization is both semi-tabular and highly irregular as fluids ascended along dike-filled structures in the underlying Paleozoic rocks, through the Tertiary unconformity, and expanded upward into the Tertiary section. Mineralizing fluids appear to have bled out laterally away from mineralized dikes into favorable permeable lithologies and secondary structures, including the FCF. Deeper drilling has extended the deposit to the west, where it appears that the FCF plays a significant role as a high-grade structural feeder (Figure 7-3). The potential for additional mineralized dikes in the vicinity of the FCF at depth to the west below current drilling is a highly attractive target.

Rhyolite dikes along the entire Bare Mountain swarm were subjected to high-temperature, sanidine- and biotite-stable alteration shortly after emplacement (Weiss, 1996). At Mother Lode, biotite-stable alteration was followed by a lower temperature pervasive illite-smectite-pyrite event that affected both the dikes and surrounding wall rocks. The illite-smectite-pyrite event penetrates well up into the Lithic Ridge Tuff, forming a large sulphidation halo around the gold mineralization. Illite-smectite-pyrite alteration typically exhibits strongly elevated As, Sb and Tl, and may be well-mineralized or only anomalous in gold. The evidence suggests an early relatively barren illite-pyrite event is followed by an overprinting main stage Au-Te sulphidation event.

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Trace element geochemistry associated with gold mineralization in drill hole samples includes: As (max 9262 ppm), Sb (max 1429 ppm), Te (max 16 ppm), Tl (max 17 ppm), Hg (max 30 ppm) and Bi (max 25 ppm). The elements with the strongest correlation to gold are tellurium and arsenic. Base metals are very low at Mother Lode. Elevated base metal values are largely associated with basement rocks. Weiss (1996) reports the presence of fluorite veining and a strong association between gold and fluorine at Mother Lode. Fluorine is not included in the multi-element analysis used by Corvus, but fluorite has been identified in the Mother Lode pit. Petrographic studies by Weiss (1996) on sulphide concentrates from drill holes at Mother Lode identified growth zones in pyrite grains exhibiting concentrations of arsenic. Weiss (1996) concluded that the Mother Lode deposit formed at a depth of 500-1000 metres or less, from gold-bearing fluids with temperatures between 200-240^oC. High-grade gold appears to be associated with remobilized carbon, particularly in Tjs and Tjvs in the upper portion of the deposit. Remobilized carbon appears to have accumulated in a blanket-like zone above the upper reaches of the rhyolite dikes.

Jasperoid in Paleozoic rocks appears to have both structural and stratigraphic control. Jasperoid intervals in drill holes commonly contain significant Ag values up to 100 ppm. The average Ag/Au ratio from jasperoid samples is 15:1. The average Ag/Au ratio in non-jasperoid Tertiary rock-hosted samples is <0.5:1. Jasperoid mineralization in Paleozoic rocks may be a separate mineralizing event that predates the main Mother Lode gold system. Jasperoid is typically oxidized and yields high cyanide shake leach recoveries.

7.2.4 Hydrothermal Alteration

Rhyolite dikes along the entire Bare Mountain swarm were subjected to a sanidine- and biotite-stable (potassic), high-temperature alteration shortly after emplacement, as evidenced by secondary hypersaline fluids inclusions in quartz phenocrysts (Weiss, 1996). At Mother Lode, rhyolite dikes and surrounding wall rocks were later subjected to lower temperature, pervasive, illite-smectite alteration and associated disseminated pyrite sulphidation as an early event associated with gold mineralization. This alteration in Mother Lode dikes is progressive in intensity, probably due to proximity to open feeder structures. Biotite goes to chlorite and eventually chlorite gets replaced by pyrite. The groundmass and feldspar phenocrysts get replaced by illite-smectite, calcite and adularia(?).

Published age dates on alteration minerals from fluorine-related gold deposits in the Bare Mountain area suggest the deposits formed between 12.9-12.2 Ma, more than a million years after dike emplacement. Weiss (1996) reports K-Ar ages of 13.1 and 12.2 Ma on mixed-layer illite-smectite from a mineralized dike at Mother Lode. The average of these dates is 12.7 Ma, which is consistent with an adularia date of 12.9 Ma from a similar mineralized dike at the historic Goldspar mine (Noble, et al., 1991). Stratigraphic relations suggest that the Secret Pass deposit, hosted in the 13.25 Ma Bullfrog Tuff, formed soon after deposition of the 12.8 Ma Topopah Spring Tuff and possibly before the deposition of the 12.7 Ma Tiva Canyon Tuff. Alunite dates of 12.2 and 11.2 Ma from silicified fault breccia at the Flatiron target suggest continued or renewed episodic hydrothermal activity in the Mother Lode area (Weiss, 1996). Alunite dates of 12.9 and 11.6 Ma from the Thompson and Silicon mines, respectively (McKee and Bergquist, 1993), reflect hydrothermal activity similar in age to Mother Lode and YellowJacket (NBP).

Steam-heated alteration is ubiquitously present in the post-mineral pre-Rainier Mesa tuff unit (Tprt), which overlies the Mother Lode deposit. The alteration zone is characterized by thick intervals of pervasive kaolinite-alunite alteration and local low temperature opaline silica replacement of the tuff unit. The Tprt unit and alteration zone are up to 150 metres thick. Steam-heated alteration is also known to affect the upper portion of the underlying Tox debris flow sequence. No significant gold has been detected in rocks affected by steam-heated alteration at Mother Lode, but gold is known to occur with cinnabar and alunite at the Flatiron target and the nearby Telluride mine. The steam-heated alteration may be related to the waning stages of the Mother Lode gold system, but likely post-dates gold mineralization. The age dates suggest episodic hydrothermal activity in the Fluorspar Hills between ~13-11.2 Ma.

7.3 NBP Geology

7.3.1 Stratigraphy

The stratigraphy of the northern Bullfrog Hills was most recently described in published mapping by Connors et al. (1998). Where possible, the terminology of Connors et al. (1998) has been preserved in the unit assignments at NBP. New geochronology has shown that some units were incorrectly correlated and required new names. The most significant examples are the Sierra Blanca Tuff and the Pioneer Formation, which were previously included in the Crater Flat Group. Based on drilling and geochronology studies conducted by Corvus (see Section 7.3.2.1), the local stratigraphy has been significantly refined to warrant the identification of the North Bullfrog Hills Volcanic Complex (NBHVC, Table 7-2). Brief descriptions of the stratigraphy at the NBP area are given below. Figure 7-4 is a compiled geologic map of the NBP and surrounding areas. An explanation of the map units on Figure 7-4 is presented in Figure 7-5.

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7.3.1.1 Paleozoic Basement

7.3.1.1.1 Wood Canyon Formation - Pzw

The Wood Canyon Formation is Lower Cambrian in age and made up of variably calcareous shale, siltstone, and sandstone, with occasional beds of massive to finely laminated limestone. Lenses of massive siliceous quartzite are also present. The minimum thickness is 350 metres (Connors, et al., 1998).

7.3.1.1.2 Zabriskie Quartzite - Pzz

The Zabriskie Quartzite is Lower Cambrian in age and generally consists of massive, fine- to medium-grained ortho-quartzite with poorly preserved bedding. The minimum thickness is around 370 metres (Connors, et al., 1998).

7.3.1.1.3 Carrara Formation - Pzc

The Carrara Formation is Middle to Lower Cambrian in age and consists largely of thin- to medium-bedded limestone. The lower parts of the Carrara Formation contain cherty, argillaceous and silty limestone interbeds. The minimum thickness is around 280 metres (Connors, et al., 1998)

7.3.1.2 Jolly Jane Formation - Tjj

The Jolly Jane Formation consists of a basal Tertiary conglomerate overlain by heterogeneous sediments including hematitic mudstone, siltstone, sandstone and conglomeratic sandstone. The conglomerate characteristically contains abundant clasts of pre-Tertiary basement rocks. The type locality is in drill holes at Jolly Jane. It consists of up to 50 metres of heterogeneous sediments that appear to have accumulated in isolated structural basins prior to and during the onset of volcanism. The sediments are typically siliceous and hematitic but may include minor calcareous or carbonaceous intervals. The thickness is highly variable, having been deposited on a Tertiary erosional unconformity of significant relief. The Jolly Jane Formation may be time transgressive along the basal Tertiary unconformity. It is considered the litho-stratigraphic equivalent of the Titus Canyon Formation of Connors et al. (1998).

7.3.1.3 North Bullfrog Hills Volcanic Complex

The North Bullfrog Hills Volcanic Complex (NBHVC) is a name that has been given to a sequence of largely locally-sourced lavas and pyroclastic rocks exposed in the Western Resource Area of the NBP (Figure 7-4). Some of these rocks were incorrectly correlated with the Crater Flat Group by Connors, et al. (1998). The subdivisions identified below are based on age-dating and geologic logging, which have clarified both the stratigraphic order and aerial distribution of the units. With ages ranging from 15-14 Ma, the NBHVC correlates with some of the oldest volcanic rocks in the SWNVF. The NBHVC occupies the Tr1 time-stratigraphic position of Eng et al. (1996).

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Table 7-2 Overview of the Stratigraphy of the North Bullfrog Hills

Au min		Major Unit Name	Symbol	Formation	Lithodeme	Major Unit Description
		Quaternary Cover	Qc			Quaternary alluvium, colluvium, talus, and mine dumps
		Gravels of Sober-up Gulch	Tgs	Gravels of Sober-up Gulch		Semi-consolidated cobble and boulder gravels
		Pumiceous Sediments	Tps			Light colored tuffaceous sandstone, conglomerate with pumice clasts
		Rainbow Mountain Sequence	Trl	Donovan Mountain Latite		Latite and quartz latite lava flows and flow breccias
			Trt	Tuffs and Lavas of Rainbow Mountain		Non-welded crystal-lithic rhyolite ash-flow tuff and aphanitic flow-banded rhyolite flows and domes with minor sedimentary interbeds
Connection	Mayflower
			Tdf	Rainbow Mountain Debris Flow Sequence		A sequence of intercalated heterolithic and monolithic debris flow breccias derived from local stratigraphy. Heterolithic sequences are poorly sorted, consisting of sand- to boulder-size clasts of volcanic and Pz sedimentary rocks. Monolithic breccias are interpreted as landslide megabreccia deposits shed off local fault scarps
		Timber Mountain Group	Tma	Ammonia Tank Tuff		Moderate to densely welded crystal-rich rhyolite ash-flow tuff
			Tmr	Rainier Mesa Tuff		Moderate to densely welded crystal-rich rhyolite ash-flow tuff
			Tprr			Variably flow-banded rhyolite and rhyolite flow breccia
			Tprt			Light-colored, non-welded, locally bedded, crystal-lithic ash-flow tuff
		Paintbrush Group	Tp	Paintbrush Tuff		Aphanitic phenocryst-poor welded rhyolite ash-flow tuff
		Crater Flat Group	Tcb	Bullfrog Tuff		Variably welded crystal-lithic rhyolite ash-flow tuffs. Probable equivalent of Bullfrog Member of the Crater Flat Group
	Sierra Blanca, Jolly Jane and YellowJacket	Lithic Ridge Tuff	Tlr	Lithic Ridge Tuff		Variably welded, lithic-rich "dacitic" tuff
		North Bullfrog Hills Volcanic  Complex	Td	Savage Valley Dacite	North Bullfrog Suite: rhyolitic to dacitic dikes, sills and domes of ambiguous origin	Upper Member consists of intercalated lava flows, breccias and pyroclastics of dacitic composition. Probable stratigraphic correlation to Tr1g quartz latite unit in Southern Bullfrog Hills
						Lower Member consists of complex mixture of rhyolite lava flows, tuffs, breccias and sedimentary rocks covering post-Sierra Blanca erosional surface
			Tsb	Sierra Blanca Tuff		Large compound cooling unit of variably welded crystal-lithic rhyolite ash-flow tuff.  Lies stratigraphically within the NBHVC, but likely sourced from an unknown older caldera related to the Timber Mountain caldera complex. May be correlative to Upper Tuff of Sawtooth Mountain
			Tpf	Pioneer Formation		Upper Epiclastic Member: mixed bedded to non-bedded heterogeneous epiclastics: poorly sorted silty, sandy, pebbly and cobbly sediments
						"Green tuff" of Sierra Blanca; Heterogeneous  non-welded to semi-welded, lithic-poor to lithic-rich crystal ash-flow tuffs with scattered intervals of bedded tuff and epiclastics.   May be correlative to Lower Tuff of Sawtooth Mountain
			Tnb
			Tsf	Savage Formation		Sequence of intercalated lava flows, intrusives and epiclastic debris of dacitic to rhyolitic compostion
		Jolly Jane Formation	Tjj	Jolly Jane Formation		Heterogeneous sedimentary sequence consisting of mudstone, siltstone, sandstone and conglomerate accumulated in localized structural  basins
	SB-JJ
		Paleozoic Basement	PzC	Carrara Limestone		Micritic and argillaceous carbonaceous limestone
			PzZ	Zabriskie Quartzite		Massive quartzite

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7.3.1.3.1 Savage Formation - Tsf

The Savage Formation consists of locally-sourced lava domes, flows, pyroclastics and associated intrusive rocks of generally dacitic composition. The Savage Formation is recognized in drill holes under south Savage Valley, Air Track Hill and Jolly Jane, where it overlies and inter-fingers with sediments of the Jolly Jane Formation. The Savage Formation may also include variably carbonaceous epiclastic intervals of re-worked dacite. The thickness of the Savage Formation varies greatly from 0-100 metres, possibly reflecting both fault-bounded basins and the areal distribution of individual domes, flows and associated pyroclastic or epiclastic aprons. The Savage Formation is correlative to the lower portion of the Trl unit as described in the Southern Bullfrog Hills by Eng et.al. (1996). The Savage Formation represents the onset of SWNVF volcanism in the northern Bullfrog Hills. It is locally mineralized in the Sierra Blanca and Jolly Jane areas.

Figure 7-4 Geologic Map of NBP Showing Target Areas, Resource Outlines and Property Outline

See Figure 7-5 below for the explanation of map units.

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Figure 7-5 Map Units Legend for Figure 7-4

7.3.1.3.2 Pioneer Formation - Tpf

The Pioneer Formation consists of a rather monotonous sequence of light green, variably welded, lithic-lapilli ash-flow and air-fall tuffs. The tuffs show marked variations in clast size ranging from coarse tuffaceous sedimentary breccias to fine lithic-lapilli tuffs. Bedded epiclastic intervals have also been observed throughout the Pioneer Formation. The type locality for the Pioneer Formation is west of the Pioneer Mine. In the subsurface north of Sierra Blanca, the Pioneer Formation is interbedded with rhyolite bodies assigned to the North Bullfrog Intrusive Suite. The rhyolite bodies are interpreted as lava flows and shallow intrusive bodies that are genetically related to, and comingled with, the pyroclastic deposits of the Pioneer Formation. The lithic content increases significantly near the top of the unit, grading into a coarse heterogeneous tuffaceous epiclastic sequence known as the Upper Epiclastic Member. The Upper Epiclastic Member forms a semi-continuous marker horizon at the top of the unit. The thickness of the Pioneer Formation varies from just a few metres at south Savage Valley and Jolly Jane, to several hundred metres north of Sierra Blanca. Divergence of compaction foliation directions between the Pioneer Formation and the Sierra Blanca Tuff indicate that some tilting and erosion of the Pioneer Formation took place prior to the eruption of the Sierra Blanca Tuff. The Pioneer Formation is widely mineralized but is generally lower grade than the overlying Sierra Blanca Tuff.

7.3.1.3.3 North Bullfrog Intrusive Suite - Tnb

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The North Bullfrog Intrusive Suite consists of plugs, domes, dikes, sills and flows of generally rhyolite composition that are recognized over much of the NBP. Rocks assigned to the intrusive suite intrude Paleozoic sediments and nearly all volcanic units of the North Bullfrog Hills Volcanic Complex. A lithodeme classification has been created in order to deal with the ambiguity of the emplacement mechanism (intrusive or extrusive) and relative age of many of these bodies (Table 7-2). Plug or dome-like bodies of rhyolite are exposed at south Savage Valley, the north flank of Sober-Up Peak, and at south Jolly Jane. Field relationships suggest that these bodies are intrusive into the Pioneer and Savage Formations. There is no significant age difference between the Pioneer Formation and Tnb rhyolites, and for the most part they are geochemically indistinguishable.

There is one suite of rhyolite bodies that have no compositionally similar pyroclastic rocks. These rhyolites are relatively depleted in light rare earths such as cerium and lanthanum compared to the other rhyolites. They occur as intrusive plugs and flow-domes both above and below the Sierra Blanca Tuff in the Sierra Blanca area. These rhyolites are the host rocks for significant higher-grade disseminated and vein mineralization in the northern part of the Sierra Blanca resource area. Their genetic association with the mineralization events at Sierra Blanca has yet to be determined.

7.3.1.3.4 Sierra Blanca Tuff - Tsb

The Sierra Blanca Tuff is a large cooling unit of rhyolitic ash flow tuff that underlies much of the Western Resource Area of the NBP (Figure 7-4). It is named for the exposures at Sierra Blanca and North Sierra Blanca ridges. The unit varies in thickness from 70 metres at Jolly Jane to 170 metres at North Sierra Blanca. The caldera source of the Sierra Blanca Tuff is unknown but is likely located outside of the Bullfrog Hills. Based on lithology and age, it is correlated to the Upper Tuff of Sawtooth Mountain mapped by Maldonado and Hausback (1990) in the southern Bullfrog Hills.

The Sierra Blanca Tuff has a 5-15-metre-thick distinctive shard-rich interval (Tsb1) at the base in the Sierra Blanca area. Tsb1 commonly overlies the Upper Epiclastic Member of the Pioneer Formation. Above the shard-rich marker, the tuff is a relatively homogeneous densely welded crystal tuff with well-developed compaction foliation exhibited by flattened pumice. Thickness variations are partly due to filling of paleo-topography at the base, and an erosional unconformity at the top. The brittle nature of the densely welded Sierra Blanca tuff facilitated significant brittle fracturing. The increased permeability from fracturing likely played a significant role in focusing hydrothermal fluids through the unit. The Sierra Blanca Tuff is the most important host at NBP for both disseminated- and vein-style mineralization.

7.3.1.3.5 Savage Valley Dacite - Td

The Savage Valley Dacite consists of locally-sourced dacite lava domes, flows and associated pyroclastic and epiclastic deposits overlying the Sierra Blanca Tuff. The Lower Member (Tdl) is a heterogeneous sequence of dacitic pyroclastic and epiclastic deposits of highly variable thickness. The Lower Member locally includes a rhyolite flow breccia, which in some drill holes is of sufficient volume to be differentiated as a rhyolite flow of the North Bullfrog Intrusive Suite. The Lower Member grades upward into more homogenous dacite lava flows of the Upper Member. The Upper Member (Td) consists of fine- to medium-grained porphyritic lava flows of dacitic to andesitic composition. Where it is relatively unaltered, the Savage Valley Dacite is strongly magnetic. The Savage Valley Dacite overlies the Sierra Blanca Tuff along an erosional unconformity of significant relief. Locally the Savage Valley Dacite appears to be deposited on the Pioneer Formation. Some or all of the Sierra Blanca stratigraphy may have been eroded or not deposited in these areas. The Savage Valley Dacite is correlative to the Tr1g unit as described in the southern Bullfrog Hills by Eng et.al. (1996). The Savage Valley Dacite is mineralized in the Jolly Jane and Sierra Blanca resource areas.

7.3.1.4 Lithic Ridge Tuff - Tlr

The Lithic Ridge Tuff is a regional tuff unit that was first recognized at NBP in 2015. Previously thought to be the Tram Tuff of the Crater Flat Group, it occupies the appropriate regional stratigraphic position above the Savage Valley Dacite and below the Bullfrog Tuff. The Lithic Ridge Tuff consists of poorly to moderately welded, generally lithic-rich, crystal-bearing ash-flow tuff. Lithic content varies from <5 to 20%. It is generally biotite-rich and commonly contains green altered clasts of Savage Valley Dacite. It was originally described as “dacite tuff” and assigned to the upper part Savage Valley Dacite. It has been differentiated in several drill holes in the YellowJacket area as overlying and locally inter-fingering with the Savage Valley Dacite.

7.3.1.5 Crater Flat Group - Tcb

The regionally extensive Crater Flat Group has been described in detail by Carr et al. (1986), and three tuff members are recognized in the SWNVF by Sawyer et al. (1994). The largest of these members, the Bullfrog Tuff (Tcb), is exposed along the east side of YellowJacket, Savage Valley, and Jolly Jane. The Bullfrog Tuff is a moderately crystal rich, densely welded, ash-flow tuff. The densely welded middle portion grades upward into a weak to moderately welded upper portion. The contact between the Bullfrog Tuff and underlying Lithic Ridge Tuff is exposed along the east side of Savage Valley, where an interval of bedded tuffaceous epiclastic rocks separates the two units. The Bullfrog Tuff has been dated at 13.25 Ma (Sawyer et al., 1994). The only gold mineralization interpreted to be hosted in the Bullfrog Tuff at NBP is in drill hole NB-11-80 under the Connection area.

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7.3.1.6 Paintbrush Group - TP

The Paintbrush Group in the Bullfrog Hills includes the 12.8 Ma Topopah Springs Tuff and the 12.7 Ma Tiva Canyon Tuff (Sawyer et al., 1994). The Paintbrush tuffs are distinctly shard-rich and phenocryst-poor compared to other major ash flow sheets in the SWNVF. The Paintbrush Group varies from 190 to >240 metres in thickness in the northern Bullfrog Hills (Connors et al., 1998). The two subunits have not been differentiated in published mapping. At the NBP, the Paintbrush Group exists primarily as large slide blocks or monolithic breccias within the Rainbow Mountain Debris Flow Sequence. No significant mineralization has been found to date within the Paintbrush Tuff or monolithic Paintbrush breccias. However, Paintbrush breccias are hydrothermally altered in a number of places at NBP and are considered potential host rocks as they were present at the time of YellowJacket and Mayflower mineralization.

7.3.1.7 Timber Mountain Group – Tmr and Tma

Regionally there are two large-volume ash flow sheets that make up the bulk of the Timber Mountain Group. These include the 11.6 Ma Rainier Mesa Tuff (Tmr) and the 11.45 Ma Ammonia Tanks Tuff (Tma). The Timber Mountain tuffs are rather distinctive because of their large (2-4 mm) and abundant (20%) phenocrysts of quartz and feldspar. The units can be distinguished from each other by the presence of sphene and bluish chatoyant sanidine phenocrysts in the Ammonia Tanks Tuff. The lower part of the Timber Mountain Group includes smaller volume rhyolitic tuffs and lava flows that are known as the pre-Timber Mountain tuffs (Tprt) and lavas (Tprr) respectively, which underlie the Rainier Mesa Tuff.

In-situ bedrock exposures of the Timber Mountain Group are found only in the southern and eastern portions of the NBP area. The Timber Mountain Group exhibits widespread steam-heated alteration at the Spicerite and Baileys target areas at the Eastern Steam-Heated Zone (Figure 7-4). All sub-units of the Timber Mountain Group have potential to host younger disseminated and vein type mineralization at NBP. Over much of the NBP, the Rainier Mesa and Ammonia Tanks tuffs occur only as large slide blocks or monolithic breccia bodies within the Rainbow Mountain Debris Flow Sequence. Clasts of the Rainier Mesa and Ammonia Tanks tuffs are a significant component of heterolithic portions of the Rainbow Mountain Debris Flow Sequence.

7.3.1.8 Rainbow Mountain Sequence

The Rainbow Mountain Sequence consists of a complex group of sedimentary and volcanic deposits that record the onset of a major phase of extensional tectonism and erosion in the Bullfrog Hills between 11.4 and 10.5 Ma (Connors et al., 1998). The sequence includes in ascending order: 1) the Rainbow Mountain Debris Flow Sequence; 2) the Tuffs and Lavas of Rainbow Mountain; and 3) the Donovan Mountain Latite.

7.3.1.8.1 Rainbow Mountain Debris Flow Sequence - Tdf

The Rainbow Mountain Debris Flow Sequence is the most heterogeneous unit at NBP. It consists of thick sequences of non-bedded, poorly sorted, heterolithic and monolithic sedimentary breccias, and interbedded fluvial sediments. Heterolithic debris flow breccias contain sand- to large boulder-size clasts of Miocene volcanic and Paleozoic sedimentary rocks. These deposits are largely the result of the re-working of volcanic and basement rocks via gravity sliding and alluvial fan development around fault-bounded structural highs. Relatively intact blocks of monolithic breccias are interpreted as landslide megabreccia deposits that were shed off local fault scarps. The volcanic debris is derived from many of the SWNVF units including the NBHVC, the Crater Flat Group, Paintbrush Group and Timber Mountain Group (Table 7.2). The Debris Flow Sequence lies unconformably on an erosional surface (i.e. angular unconformity) cut on nearly all pre-Rainbow Mountain Sequence map units. The thickness of Debris Flow Sequence exceeds 300 metres in the Mayflower area. Gold mineralization at Mayflower, Connection and Cat Hill is hosted in the Rainbow Mountain Debris Flow Sequence.

7.3.1.8.2 Tuffs and Lavas of the Rainbow Mountain Sequence – Trt and Trr

Generally overlying and inter-fingering with the Debris Flow Sequence are light-colored, poorly- to non-welded, pumiceous crystal- and lithic-rich tuffs. The Rainbow Mountain Tuff includes three separate units in ascending order: Trt, Trt₂, and Trt₃. The most volumetrically significant tuff sub-unit is the middle 10.5 Ma Trt₂, which overlies the Debris Flow Sequence over much of the NBP. Trt₂ is correlative to the Tr11 unit of the Rainbow Mountain Sequence in the southern Bullfrog Hills (Eng et al., 1996). Trt₂ is up to 300 metres thick in the Mayflower-Pioneer area. The base of Trt₂ is locally mineralized at Mayflower. Trt₂ has yielded an Ar-Ar date of 10.1 Ma (Connors et al., 1998), and a Zircon date of 10.5 Ma (Valencia date, Table 7-3). The Trt tuffs are genetically related to and inter-finger with locally-sourced, flow-banded rhyolite plugs, domes and flows (Trr).

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7.3.1.8.3 Donovan Mountain Latite – Tl

The uppermost sub-unit of the Rainbow Mountain Sequence is the Donovan Mountain Latite. It consists of numerous lava flows and flow breccias of dark-colored, relatively unaltered, porphyritic latite and quartz latite. The latite occurs primarily in the hanging wall (west) of the Donovan Mountain Fault. Light colored tuffs of Trt₃ commonly occur as pyroclastic intervals between individual lava flows. Latite flows of similar composition cap the Rainbow Mountain Sequence throughout the Bullfrog Hills. Ar-Ar dates of between 10.0 and 10.7 Ma are reported by Connors et al. (1998).

7.3.1.9 Pumiceous Sediments - Tps

An unnamed unit of heterogeneous pumiceous sediments overlies the various units of the Rainbow Mountain Sequence. The sediments include white, light-grey, greenish-grey and light brown, weakly indurated, bedded tuffaceous sandstone and conglomerate. The beds may be moderately- to poorly- sorted, commonly containing abundant small pumice and other volcanic clasts. The unit overlies and inter-fingers with the Rainbow Mountain Debris Flow Sequence at the Eastern Steam-heated Zone, and overlies the Donovan Mountain Latite west of YellowJacket. The thickness varies from ranging from 0 to 40 metres, probably filling paleo-topography.

7.3.1.10 Gravels of Sober Up Gulch - Tgs

Gently dipping, ridge-forming terraces of older alluvial deposits are named for extensive exposures in the Sober-up Gulch area just south of the NBP. The unit consists of semi-consolidated, heterolithic boulder gravels of probable late Miocene to Pliocene age (Maldonado and Hausback, 1990). The gravels unconformably overlie older Miocene units throughout the NBH. Tgs is similar to Tdf, but contains abundant conspicuous boulders of Donovan Mountain Latite. Tgs typically forms a gently east-dipping (<5^o) pediment surface. The pediment surface has been deeply incised by Quaternary erosion, resulting in a series of gently-dipping gravel terraces along the eastern side of the NBP. The gravel terraces overlap and conceal steam-heated alteration at the Eastern Steam-heated Zone. The gravel unit is not known to be mineralized, but contains clasts of altered and mineralized rock.

7.3.1.11 Quaternary Cover - Qc

Quaternary Cover includes unconsolidated deposits of alluvium, colluvium, talus and mine dump material.

7.3.2 Geochronology

Many of the major units at the NBP are separated by significant erosional unconformities, which together with pervasive alteration and structural juxtaposition have confounded local and regional stratigraphic correlations. For this reason a series of samples were submitted by Corvus for geochronological studies. Laser ablation inductively coupled plasma (“ICP”) analysis on zircons was used to determine the eruptive ages of the different volcanic units, and Ar-Ar dating on adularia was used to date the vein mineralization.

Table 7-3 Summary of Zircon Dates from North Bullfrog Hills Volcanic Complex

Map Unit	Age (Ma)	2s	Locality	Description	Lab	Lab ID	Number of Dates
Trt2	10.5	0.1	Mayflower Mine	Rainbow Mountain Tuff	Valencia	115904	27
Trt2	11.4	0.2	Mayflower Mine	Rainbow Mountain Tuff	AtoZ	1335-003	26
Tp	12.7	0.2	Jolly Jane	Paintbrush Tuff	Valencia	115923	20
Tcb	13.3	0.2	Ladd Mountain	Crater Flat Tuff, Bullfrog Member	AtoZ	1335-001	25
Tcb	13.4	0.2	East Savage Valley	Crater Flat Tuff, Bullfrog Member	Valencia	115902	35
Tcb	13.5	0.2	Ladd Mountain	Crater Flat Tuff, Bullfrog Member	Valencia	115906	47
Tcb	13.5	0.2	Ladd Mountain	Crater Flat Tuff, Bullfrog Member	Valencia	115907	29
Tcb	13.5	0.2	Jolly Jane	Crater Flat Tuff, Bullfrog Member	Valencia	115924	34
Tct	13.7	0.2	East Savage Valley	Crater Flat Tuff, Tram Member	Valencia	115901	19
Tct	13.8	0.2	YellowJacket	Crater Flat Tuff, Tram member	Valencia	M610285A	13
Tct	14.3	0.2	YellowJacket	Crater Flat Tuff, Tram Member	AtoZ	1293-06	28
Td2	14.4	0.2	Jolly Jane	Savage Valley Dacite	AtoZ	1293-10	21
Td2	15.6	0.4	YellowJacket	Savage Valley Dacite	AtoZ	1293-01	8
Td1	14.2	0.3	Air Track Hill	Savage Valley  Felsic Facies	Valencia	115908	33
Td1	14.8	0.2	YellowJacket	Savage Valley  Felsic Facies	Valencia	115905	22
Td1	15	0.3	YellowJacket	Savage Valley  Felsic Facies	AtoZ	1293-02	28

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Map Unit	Age (Ma)	2s	Locality	Description	Lab	Lab ID	Number of Dates
Trl	14.9	0.2	Sawtooth Mtn	Sawtooth Mtn Tuff, Upper Unit	AtoZ	1335-005	26
Tsb	14.4	0.2	YellowJacket	Sierra Blanca Tuff	Valencia	P346369	30
Tsb	14.5	0.2	East Jolly Jane	Sierra Blanca Tuff	AtoZ	1293-09	27
Tsb	14.7	0.3	Jolly Jane	Sierra Blanca Tuff	Valencia	115925	34
Tsb	15	0.2	YellowJacket	Sierra Blanca Tuff	AtoZ	1293-07	28
Tsb	15.1	0.2	Pioneer	Sierra Blanca Tuff	AtoZ	1335-002	29
Tpf	14.5	0.2	YellowJacket	Pioneer Formation	AtoZ	1293-05	27
Tpf	14.6	0.2	YellowJacket	Pioneer Formation	AtoZ	1293-08	28
Tnb	14.3	0.3	YellowJacket	Rhyolite - spherulitic	Valencia	NB171608	32
Tnb	14.5	0.2	YellowJacket	Rhyolite - flow banded	Valencia	P346202	20
Tnb	14.7	0.2	Radio Tower Hill	Rhyolite - flow banded	Valencia	115903	26
Tnb	14.7	0.2	YellowJacket	Rhyolite - spherulitic	Valencia	P346250	26
Tnb	14.8	0.2	YellowJacket	Rhyolite - flow banded	AtoZ	1293-04	28
Tnb	15.3	0.3	Gold Pit	Dacite Porphyry in Cambrian Basement	Valencia	P347979	28
Tnb	15.8	0.3	Jolly Jane	Pre-Pioneer Formation Dacite	AtoZ	1293-12	24
Tnb	15.9	0.3	Radio Tower Hill	Rhyolite - flow banded	AtoZ	1335-004	34
Tnb	16.1	0.3	Savage Valley	Pre-Pioneer Formation Dacite	AtoZ	1293-11	26
Td	1599.5	20.7	YellowJacket	Basement Xenocrysts in Dacite	AtoZ	1293-03	3

7.3.2.1 Zircon Dating

Samples were submitted for analysis to two different laboratories: Apatite to Zircon, Inc. (A to Z) and Victor Valencia. Duplicate samples were included to confirm the analytical precision of the dates (Table 7-3). In general the match between the labs is reasonable, however, in many instances the A to Z dates are substantially older with differences far exceeding the analytical precision (Table 7-3). The Valencia dates match the published Ar-Ar ages for the Rainbow Mountain Tuff (Trt₂), the Paintbrush Tuff and the Bullfrog Tuff. However, the A to Z dates of the Rainbow Mountain (Trt₂) and the Lithic Ridge tuffs are almost 1 Ma older than the published ages. Similarly, the A to Z age on the duplicate rhyolite from Sober-up Peak is also approximately 1 Ma older than the Valencia date (Table 7-3). For this reason it appears that some of the dates from A to Z are too old.

7.3.2.2 Ar-Ar Dating

Only four of the eight samples submitted to the University of Alaska, Fairbanks for Ar-Ar dating of vein adularia returned statistically valid ages (Benowitz and Layer, 2013). The valid age samples came from four different YellowJacket drill holes. These age dates constrain mineralization at YellowJacket between 11.7-11.2 Ma (Table 7-4). The new age dates for YellowJacket vein mineralization confirm an earlier 11.3 Ma date published by Connors et al. (1998). Connors et al. (1998) also published adularia dates of 11.0 Ma at the East Savage Vein (Figure 7-4) and ages of 10.0 Ma and 9.9 Ma from the Mayflower Mine. The Mayflower deposit is the same age as the Bullfrog vein deposit in the southern Bullfrog Hills.

One of two alunite samples submitted to the University of Nevada, Las Vegas for Ar-Ar dating returned a valid age-date. Coarse vein alunite collected from the Alunite Hill area of the Eastern Steam-heated Zone returned an age of 9.5 Ma (Table 7-4). This age is similar to the 10.2 Ma age of alunite obtained from the Bailey’s Hot Springs area (Weiss, et al., 1994), and similar to the published adularia age dates at Mayflower and Bullfrog. The alunite age dates highlight the potential for the discovery of a new Bullfrog-age, high-grade vein system under the extensive 14 square kilometre Eastern Steam-heated Zone.

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Table 7-4 Adularia and Alunite Ar-Ar Age Determinations

Hole ID	Sample	Mineral	Integrated Age  (Ma)	Plateau Age (Ma)	Plateau Information	Isochron Age (Ma)	% Atmospheric 40Ar	Lab
NB-12-127	M610395	Adularia	11.2 ± 0.1	11.2 ± 0.1	5 of 7 fractions	__	15.8	University of Alaska Fairbanks
					98.4% 39Ar release
					MSWD = 1.09
NB-12-139	M612038	Adularia	11.6 ± 0.1	11.6 ± 0.2	4 of 7 fractions	__	13.5	University of Alaska Fairbanks
					99.1% 39Ar release
					MSWD = 2.43
NB-12-126b	M610140	Adularia	15.5 ± 2.1	11.7 ± 0.4	4 of 7 fractions	__	93	University of Alaska Fairbanks
					51.3% 39Ar release
					MSWD = 1.21
NB-12-138	M611584	Adularia	11.7 ± 0.4	11.7 ± 0.4	6 of 7 fractions	11.4 ± 0.4	42.7	University of Alaska Fairbanks
					97.6% 39Ar release
					MSWD = 0.51
Alunite Hill	Alun SW107	Alunite	9.73 ± 0.5	9.52± 0.5	10 of 14 fractions	_	_	University of Nevada Las Vegas
					96% 39Ar release

Adularia samples are all from the YellowJacket Zone. The Alunite sample is from the Eastern Steam-Heated Zone north of Alunite Hill.

7.3.3 Regional Correlation

The geochronological studies have significantly refined the stratigraphy of the NBP. Prior to obtaining the zircon age dates, the Sierra Blanca Tuff was correlated with the Bullfrog Tuff of the Crater Flat Group. The zircon dates confirm the age difference between rocks of the NBHVC and the Bullfrog Tuff, and indicate that volcanism of the NBHVC extended from around ~15 Ma to 14 Ma (Table 7-3). With age range of 15-14 Ma, the NBHVC correlates with some of the oldest volcanic rocks in the SWNVF, as well as the Tr1 time-stratigraphic position of Eng et.al. (1996). Zircon dating also appears to confirm the proposed correlation between the Sierra Blanca Tuff and the Upper Tuff of Sawtooth Mountain (Table 7-3).

The NBHVC rocks have experienced several significant periods of tectonic reorganization in the Bullfrog Hills extensional domain. Connors, et al. (1998) have identified a major period of extension between 12.7 and 11.6 Ma culminating with the eruption of the Timber Mountain Group ash-flows between 11.6 and 11.45 Ma. These events are coincident with the age of YellowJacket vein mineralization. Connors, et al. (1998) postulate a second major period of extension occurred between 11.4 Ma and 10.5 Ma, which resulted in major block rotation, rapid erosion and the deposition of the Rainbow Mountain Debris Flow Sequence. The 10 Ma Bullfrog vein, the 9.9 Ma Mayflower deposit, and the 9.5 Ma Eastern Steam-heated Zone all coincide with the waning stages of extensional tectonism in the Bullfrog Hills. Extension appears to have ceased by 9.4 Ma (Connors, et al., 1998).

7.3.4 Structure

Both pre-Tertiary and Miocene rocks have been subjected to large-scale, W- to WNW-directed, syn-volcanic and syn-mineral extension between ~15-9.4 Ma. Such extension is largely exhibited by down-to-the-west normal faulting and east-tilting of stratigraphy. The extensional tectonism was apparently comprised of two periods of intense faulting and tilting (i.e. accelerated extension), within a protracted period of less intense but episodic extension. Extensional faulting events were accompanied by episodic hydrothermal alteration and gold mineralization throughout the Bullfrog district.

Extension at NBP was accommodated by the Bullfrog Hills Fault System (BHFS). The BHFS at NBP consist of two basic fault types: 1) large-displacement NNW- to NNE-trending, moderately to steeply west-dipping, down-to-the-west normal faults; and 2) generally smaller displacement, steeply east-dipping, down-to-the-east, antithetic faults (Figure 7-4 and 7-6). The down-to-the-west faults are interpreted to have listric shapes at depth, similar to the MP fault at the Bullfrog mine. These faults are interpreted to sole into the Southern Bullfrog Hills Fault (detachment) at unknown depth under the NBP. Steeply east-dipping, down-to-the-east faults commonly occur in the hanging walls of the large displacement down-to-the-west faults, and are interpreted to be truncated at depth by the down-to-the-west faults (Figure 7-6).

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Figure 7-6 Geologic Cross Section from Savage Valley to Connection Illustrating the Overall Structural Style of the NBP

The section location for Figure 7-6 is shown on Figure 7-4.

Four major splays of the BHFS cross the NBP including: the Donovan Mountain Fault, the West Jolly Jane Fault, the Road Fault and the Spicerite Fault (Figure 7-4). These faults have apparent down-to-the-west, dip-slip displacements of ~600 to >1000 metres (Figure 7-6). The Road Fault is considered the northern continuation of the Contact Fault from the Bullfrog mine area in the southern Bullfrog Hills.

Some of the oldest faults at the NBP are represented by the down-to-the-east Liberator, East Jolly Jane and Quartzite Faults (Figure 7-6 and 7-10). Evidence for an older age on these structures is the fact that the Bullfrog Tuff is only preserved on the downthrown side of these faults. It seems likely that down-to-the-east faults developed before or during the 12.7-11.6 Ma deformation event just after the deposition of the Bullfrog Tuff. At Jolly Jane, the Bullfrog Tuff is strongly quartz-adularia altered in the hanging-wall of the East Jolly Jane Fault. The evidence suggests that the major down-to-the-east faults were present during the older disseminated mineralization event at the NBP (see discussion in Section 7.3.5.1) and played a role in the process of mineralization. The Liberator and East Jolly Jane Faults appear to truncate the eastern sides of the Sierra Blanca-YellowJacket and Jolly Jane deposits respectively (Figure 7-4, 7-6, 7-12), suggesting they were also active after the older mineralization event. There are many faults at the NBP with relatively minor displacements, the ages of which are difficult to constrain.

7.3.5 Mineralizing Events

All of the mineralizing events known to date at the NBP can be classified as low-sulphidation epithermal mineralization. Two general styles of mineralization are present at NBP: 1) pervasive alteration-style disseminated mineralization; and 2) structurally controlled vein and stockwork mineralization. There are at least three distinct periods of mineralization present at the NBP:

· pre-11.7 Ma pervasive alteration-style disseminated (Sierra Blanca and Jolly Jane)

· 11.7-11.2 Ma structurally-controlled alteration-style enrichment and late vein (YellowJacket)

· ~10 Ma structurally-controlled alteration-style disseminated and late vein (Mayflower).

Pre-11.7 Ma pervasive disseminated mineralization is hosted in ~15-14 Ma rocks of the NBHVC. The presence of pervasive quartz-adularia alteration in the relatively barren Bullfrog Tuff east of Jolly Jane suggests the 13.25 Ma Bullfrog Tuff was affected by this early quartz-adularia mineralization, bracketing the earliest event between 13.25 and 11.7 Ma. A barren jasper (quartz-hematite) vein event overprints the pervasive quartz-adularia mineralization at Sierra Blanca-YellowJacket. The jasper vein event was subsequently overprinted by higher grade, structurally-controlled, alteration-style enrichment mineralization; which was closely followed by (or contemporaneous with) the YellowJacket vein and stockwork mineralization. Based on overprinting relationships observed in core (Figure 7-7), it is apparent that multiple events have contributed to the gold endowment that had accumulated by 11.2 Ma. Pre-11.2 Ma mineralization will be discussed below as “Older” mineralization and the 10 Ma mineralization at Mayflower and Connection will be discussed as “Younger” mineralization.

7.3.5.1 Older Mineralization Styles

7.3.5.1.1 Older Pervasive Alteration-Style Mineralization

The most widespread mineralization at the NBP is associated with pervasive quartz-adularia alteration and pyritization of iron minerals in the volcanic host rocks. The grade of this mineralization often reflects the intensity of quartz-adularia or illite-adularia alteration, as well as the original iron content of the host rocks. Gold grade in the Pioneer Formation and Sierra Blanca Tuff, which average 1% iron, is on the order of 200-400 ppb gold. In contrast, grades in the Savage Valley Dacite containing 2-5% iron may reach several thousand ppb gold. Pervasive alteration associated with disseminated pyrite mineralization generally shows a progressive change from illite-smectite, to illite-adularia, and to quartz-adularia+illite as the degree of mineralization increases. The silver to gold ratio of the alteration-style mineralization is approximately 1:1.

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Figure 7-7 Cross-cutting Relationships Between Older Mineralization Types at Sierra Blanca

 

In Figure 7-7 (above) the cross-cutting relationships are as follows: 1) Early pervasive quartz-adularia alteration with gold related to sulphidation of iron; 2) Barren jasper veins filling brittle fractures in adularized tuff; 3) White illite-adularia overprint with an increase in gold; and 4) Grey translucent stockwork veinlets with high-grade gold.

The distribution of pervasive alteration-style mineralization appears to reflect a combination of rock matrix permeability and structural permeability. Although both the Sierra Blanca Tuff and the Pioneer Formation are pervasively altered, it appears that the brittle nature of the densely-welded Sierra Blanca Tuff enhanced the permeability relative to the less welded tuffs of the Pioneer Formation. In contrast, alteration-style mineralization in the underlying Savage Formation and the overlying Savage Valley Dacite appears to be controlled largely by fault-related permeability.

Pervasive alteration-style mineralization has been observed locally in the Paleozoic limestone and calcareous sediments. Jasperoid is developed in the Carrara Formation limestone just below the Tertiary unconformity in many drill holes at both Jolly Jane and Savage Valley. The jasperoid is typically anomalous in gold, and locally yields 200-400 ppb values. The jasperoid target area (Figure 7-4 and see Section 9) consists of a bedding-parallel replacement jasperoid body in calcareous Wood Canyon Formation sediments. There is a surface gold anomaly in rocks and soils around the jasperoid occurrence. Barrick drill hole RDH-767 intercepted gold values up to 1.1 g/t, and has never been followed-up (see Section 9.2.1.8). Paleozoic rocks do not appear to be well-mineralized at NBP, but cannot be ruled-out as potential host rocks for future exploration.

7.3.5.1.2 Older Structurally-Controlled Mineralization

Structurally controlled mineralization consists of two distinct styles which may represent two periods of mineralization. The first is a structurally-controlled alteration-style mineralization and the second is quartz vein-style mineralization. The onset of structurally-controlled mineralization is marked by the formation of a distinctive suite of essentially barren jasper (quartz-hematite) veins. The jasper veins crosscut the older pervasive quartz-adularia disseminated mineralization (Figure 7-7). After the formation of the jasper veins, movement on major fault structures (e.g. Liberator, YellowJacket, NE20, NE30, NW10 Faults) apparently resulted in a second stage of more structurally-controlled sulphidation and gold enrichment. This structurally-controlled mineralization can be distinguished from the older earlier alteration-style event by a higher As/Au ratio, and is generally associated with a white to light brown illite or illite-adularia alteration overprint (Figure 7-7). Structurally-controlled alteration-style mineralization frequently yields grades >1 g/t, and sometimes >10 g/t if the fluids encounter the higher iron contents of dacitic lithologies. Structurally-controlled alteration-style mineralization is clearly crosscut by high-grade, low sulphidation quartz veins of the YellowJacket vein event.

The YellowJacket Vein Zone consists of a massive quartz vein surrounded by hanging wall and footwall quartz stockwork zones. Such quartz vein and stockwork mineralization is found at YellowJacket and along the crest of North Sierra Blanca ridge. Observed textures that are typical of low sulphidation epithermal veins include bladed quartz pseudomorphs after calcite, crustiform banding and milky chalcedonic quartz with distinct but fuzzy banding. Veins with these textures may be relatively barren or have high-grade gold. The most common and best mineralized veins at YellowJacket are grey translucent quartz stockwork veins with little distinctive internal structure (Figure 7-7). Grains of native gold can often be observed in this quartz. There is generally little wall rock alteration associated with grey translucent quartz veins. However, white illite overprint of earlier quartz-adularia is often observed in the general vicinity of grey translucent stockworks. The illite overprint can locally be rather intense creating selvages around jasper veins and destroying all the feldspar in the rock (Figure 7-7).

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The primary minerals associated with the vein-style mineralization are gold, electrum, acanthite (Ag₂S) and pyrite. Petrographic studies have also documented pyrargyrite (Ag₃SbS₃), stromeyerite (AgCuS), proustite (Ag₃AsS₃), chalcopyrite (CuFeS₂) and covellite (CuS). Sphalerite has been observed as a late cavity infill. In general, the silver to gold ratio associated with vein mineralization is >6:1 and locally can be 100:1 or more.

7.3.5.2 Younger Mineralization Styles

A number of deposits in the NBP area formed after the deposition of the 10.5 Ma Rainbow Mountain Tuff (Trt₂). These include disseminated and structurally-controlled deposits at Mayflower, Connection and Cat Hill. Younger mineralization exhibits similar styles to older mineralization, including alteration-style, disseminated quartz-adularia and quartz-calcite vein mineralization. Steam-heated alteration along the Road Fault and the Eastern Steam-heated Zone are also related to this younger period of mineralization.

7.3.5.2.1 Younger Alteration-style Mineralization

At Mayflower, the debris flow sediments and Trt₂ tuff have been affected by quartz-adularia alteration developed around a central zone of faulting. The quartz-adularia alteration is directly associated with disseminated gold mineralization. Although the deposit is completely oxidized at present, it appears that the original mineralizing process was sulphidation during the quartz-adularia event. The process is similar to, but significantly younger than the quartz-adularia-pyrite mineralization at Sierra Blanca and Jolly Jane. The distribution of quartz-adularia alteration appears to reflect a combination of structural and stratigraphic permeability. The brittle nature of the quartz-adularia alteration enabled fracturing and development of open space during later faulting events associated with vein mineralization.

7.3.5.2.2 Younger Structurally-Controlled Mineralization

At Mayflower, quartz veining is only seen in the southeastern part of the deposit on the dumps at the Mayflower Shaft. The quartz can be white, pink and grey translucent, and contains visible gold. It is frequently banded and may have very fine acicular textures indicating replacement of earlier adularia. Bladed pseudomorphs of quartz after calcite have not been observed at Mayflower. Some of the gold at Mayflower is associated with grey manganiferous calcite veining. Observations from both macroscopic and microscopic studies at Mayflower have shown visible gold in the calcite bands rather than in the quartz. Gold grains are often found at the calcite-wallrock contact. In the David Adit, the only cavity infill phase is coarse euhedral manganiferous calcite. Calcite veining is not always mineralized, but historical sampling combined with underground observations shows that the highest grade areas have calcite cavity linings. The gold-calcite association at Mayflower is very different from YellowJacket, where calcite is generally not associated with high-grade mineralization. Another difference is the Ag:Au ratio in the Mayflower vein zones; it is generally <0.5:1, whereas at YellowJacket it is generally >6:1.

7.3.5.2.3 Steam-Heated Alteration

Steam-heated alteration is characterized by low-temperature silica replacement, pervasive kaolinite-alunite alteration and late alunite veining. It is the result of wall rocks reacting with steam being generated by hot acidic waters at depth. In modern geothermal systems, steam-heated alteration develops above a groundwater table, which in turn lies above a boiling zone at depth where gold is being deposited. Steam-heated alteration has been identified over an extensive 14 square kilometre area at the “Eastern Steam-heated Zone” (Figure 7-4). Steam-heated alteration also occurs in several places along the Road Fault, particularly in the vicinity of the Connection and Cat Hill target areas. Anomalous gold has been detected in rocks affected by steam-heated alteration at Alunite Hill, Vinegaroon, Cat Hill and Yellow Rose (Figure 7-4). There is a reasonably high probability that this alteration may be associated with a mineralized Bullfrog-Mayflower age (~10 Ma) vein system at depth. Based on drilling in 2015 at the Spicerite target in the Eastern Steam-heated Zone, the target depth is >300 metres below the current topographic surface.

7.4 Mineral Resource Areas

Corvus and previous operators exploring in the northern Bullfrog Hills have defined targets in areas of historic mines or prospects, as well as targets associated with high level epithermal alteration. Mineral Resources have been identified in several areas discussed in detail in Section 7 including: Jolly Jane, Sierra Blanca, YellowJacket, Air Track Hill, Air Track West, and Mayflower (Figure 7-4). Refer to Section 9 for descriptions of all other target areas at NBP. In 2007-2008, ITH/Redstar (NBPJV) drilled several holes at Air Track Hill and Mayflower, with two holes each at Sierra Blanca, Pioneer and Savage Valley. Between 2010 and 2017, Corvus has drilled numerous holes at Sierra Blanca, YellowJacket, Jolly Jane, Mayflower and Connection, leading to the Mineral Resource estimates presented in this document.

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7.4.1 Jolly Jane

The pseudo-stratabound nature of disseminated mineralization within the Sierra Blanca Tuff at Jolly Jane was recognized by Barrick Gold in 1995, but was not of sufficient grade to be pursued at that time. This style of mineralization was the main focus of Corvus’ drilling program in 2010-11, when twenty-seven reverse circulation (RC) holes totaling 4,128.5 metres (13,545 feet) were drilled at Jolly Jane. In 2012 and 2013, 34 additional holes were drilled at Jolly Jane totaling 4,234 metres (13,891 feet). These included three PQ3 core holes for metallurgical samples, 29 infill RC holes on the ZuZu patented claim, and two step-out RC holes to the north of the Mineral Resource area (Figure 7-8). Eight surface rock chip/channel lines totaling 384 metres (1,260 feet) have been sampled at 5 foot intervals to resemble drill holes. The results of the 2010-13 work, along with data from the Barrick drilling, are the basis for the Mineral Resources presented in this document.

7.4.1.1 Jolly Jane Stratigraphy

The stratigraphy of Jolly Jane includes the following units in ascending stratigraphic order: 1) the Cambrian Carrara Formation; 2) the Jolly Jane Formation; 3) the Savage Formation; 4) the Pioneer Formation; 5) the rhyolite bodies of the North Bullfrog Intrusive Suite; 6) the Sierra Blanca Tuff; 7) the Savage Valley Dacite; 8) the Lithic Ridge Tuff, 9) the Bullfrog Tuff; and 10) monolithic debris flow breccias of Paintbrush Tuff.

The Carrara Formation consists of primarily micritic and argillaceous limestone, with lesser calcareous shale. Jasperoid is common at or near the basal Tertiary unconformity in many drill holes. The jasperoid is considered hydrothermal in nature as it is anomalous in gold (up to 0.372 ppm), and occurs in proximity to gold mineralization in the overlying Sierra Blanca Tuff throughout the deposit area.

The Jolly Jane Formation was deposited unconformably on the Carrara Formation. It includes a heterogeneous sequence of: 1) siliceous hematitic conglomerate, sandstone and siltstone; 2) calcareous and variably carbonaceous lithic-volcanoclastic sediments; and 3) locally intercalated monolithic debris flow breccias of the Carrara Formation limestone. The conglomerate is Paleozoic-clast dominated. It typically occurs directly along the unconformity, grading upward into finer pebbly sandstone and red hematitic siltstone. The thickness of the Jolly Jane unit varies dramatically from 0-50 metres between drill holes. Lithologic variations are also quite dramatic between drill holes.

The Savage Formation consists of aphanitic to porphyritic dacite and associated dacitic pyroclastic and epiclastic rocks. It is interpreted as a dacitic flow-dome complex that pre-dates the Pioneer Formation. The basal flows of the Savage Formation inter-finger with the underlying sedimentary rocks. The Savage Formation varies dramatically in thickness from ~100 metres in the south end of Jolly Jane to zero in the northern portion of the Jolly Jane area.

The Pioneer Formation is only a few tens of metres thick at Jolly Jane. Thickness can vary dramatically across faults, suggesting that there was active tectonism and erosion between the eruptions of the Pioneer and Savage Formations. The preserved portion of the Pioneer Formation coincides with the uppermost intervals found at Sierra Blanca, suggesting that Jolly Jane was a topographic high area during that time. A number of flow-banded aphanitic rhyolite bodies intrude the Savage and Pioneer Formations at southern Jolly Jane. The rhyolites appear to be cross-cutting intrusions and have been assigned to the North Bullfrog Intrusive Suite.

The Sierra Blanca Tuff is the dominant host rock for mineralization at Jolly Jane. The Sierra Blanca Tuff is ubiquitously quartz-adularia-altered. The preserved thickness of the unit at Jolly Jane is approximately 70 metres. At Jolly Jane, there is an interval of Savage Valley Dacite (lava flow or sill?) within the middle of the Sierra Blanca Tuff. While the Sierra Blanca Tuff appears to be a single cooling unit, there may have been time during the eruptive cycle for a simultaneous local dacite eruption to have occurred. The result is a lava flow of Savage Valley Dacite within the Sierra Blanca Tuff.

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Figure 7-8 Geologic Map of the Jolly Jane Target Area

The Savage Valley Dacite overlies the Sierra Blanca Tuff, and consists of a heterogeneous sequence of lava flows, pyroclastics and epiclastics of predominantly dacitic composition. There is an angular unconformity between the top of the Sierra Blanca Tuff and the base of the Savage Valley Dacite at Jolly Jane. The Lithic Ridge and Bullfrog tuffs overlie the Savage Valley Dacite east of Jolly Jane, and are juxtaposed against the Savage Valley Dacite along the East Jolly Jane Fault (Figure 7-8, Lithic Ridge Tuff not shown). The Lithic Ridge Tuff was not recognized at NBP until late 2015 and was previously lumped into the Crater Flat Group. Drilling to date at Jolly Jane has not identified any significant gold mineralization in the Lithic Ridge or Bullfrog Tuffs.

A large semi-tabular mass of Paintbrush Tuff and monolithic debris flow breccia of Paintbrush Tuff caps the ridge just east of Jolly Jane (Figure 7-8). Relatively consistent compaction foliation in the welded tuff dips 20^o to 30^o to the east indicating a significant angular discordance with the underlying Bullfrog Tuff. It is possible that the entire section represents a large, relatively intact, slide block of Paintbrush Tuff overlying the Bullfrog Tuff. The monolithic breccia block has been included with the Rainbow Mountain Sequence Debris Flow Sequence (Tdf_p on Figure 7-8). The Paintbrush breccias are not mineralized at Jolly Jane.

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7.4.1.2 Jolly Jane Structure

The geology of Jolly Jane is complex, mostly due to the existence of faults and erosional unconformities between each of the major stratigraphic units. The Jolly Jane gold deposit is preserved as a horst between the West Jolly Jane and the East Jolly Jane Faults (Figures 7-6, 7-8 and 7-9). The horst block is cut by steeply dipping NE- and NNW-trending faults that appear to control deposition of the Jolly Jane Formation sediments (Figure 7-9, Sections Long 04 and G). At the north end of Jolly Jane, the structure is relatively simple with large coherent zones of pseudo-stratabound disseminated mineralization between faults (Figure 7-9, Sections G and K). In contrast, the structure in the south becomes increasingly complex with larger volumes of Savage Formation and Tnb rhyolite intrusions present beneath the Sierra Blanca Tuff (Figure 7-9, Section B). The Jolly Jane mineralized zone is truncated on the west side by the West Jolly Jane Fault (Figure 7-8 and 7-9). The West Jolly Jane Fault exhibits >600 metres of down-to-the-west normal displacement, which repeats the Jolly Jane stratigraphy and gold mineralization under the Savage Valley area to the west (Figure 7-6).

7.4.1.3 Jolly Jane Mineralization

The Mineral Resources at Jolly Jane consist of the older alteration-style quartz-adularia-pyrite disseminated mineralization. The primary host rock is the Sierra Blanca Tuff, and secondary host rocks include the Savage Valley Dacite, Pioneer Formation, Savage Formation and rhyolite of the North Bullfrog Suite. Mineralization is controlled by a combination of small-displacement, high-angle feeder structures and the highly brecciated Sierra Blanca Tuff. The deposit is largely pseudo-stratabound within the Sierra Blanca Tuff. Minor quartz stockwork veining occurs throughout the Jolly Jane area, but no YellowJacket-style gold-silver enrichment event has been identified to date. All of the current Mineral Resources at Jolly Jane are oxide.

Wide spaced drilling at North Jolly Jane, including five additional holes drilled in 2015-2017, has encountered thick intervals of low-grade quartz-adularia-altered Sierra Blanca Tuff. Both low-grade oxide and sulphide mineralization is present in the Sierra Blanca Tuff and Savage Valley Dacite at North Jolly Jane. Low-grade mineralization has been extended ~750 metres north of the Jolly Jane Mineral Resource, but there is insufficient drill density to necessitate an update of the Jolly Jane Mineral Resource in this document.

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Figure 7-9 Cross Sections through Jolly Jane target area. See Figure 7-8 for section locations.

Drill traces are colored by gold assay values in the figure above.

7.4.2 Sierra Blanca

The greater Sierra Blanca area includes the Savage Valley, Sierra Blanca, North Sierra Blanca, YellowJacket, Air Track Hill and Air Track West areas (Figure 7-10). In 2010-11, Corvus drilled 44 reverse circulation (RC) holes totaling 12,785 metres (41,945 feet) in the Sierra Blanca area. In 2012, Corvus drilled 16 additional holes totaling 3,548 metres (11,640 feet) including: 4 PQ3 holes for metallurgical samples, 6 HQ3 exploration holes and 6 step-out/infill RC holes. In 2013, Corvus drilled 87 additional holes totaling 19,000 metres (62,340 feet) including: 35 HQ3 core holes, two PQ3 core holes, and 50 RC holes. Fifteen channel sample lines totaling 1,070 metres (3,510 feet) were completed along new road cuts in 2013-14. In 2014, Corvus drilled 48 additional core holes totaling 11,000 metres (36,100 feet) including: 36 HQ3 core holes and 12 PQ3 holes. Between 2015 and 2017, Corvus drilled 67 additional holes (4 core and 63 RC) totaling 19,192 metres (62,965 feet) in the greater Sierra Blanca area. The Sierra Blanca Mineral Resource is based on this drilling.

7.4.2.1 Sierra Blanca Stratigraphy

The stratigraphy of the greater Sierra Blanca area includes the following major units in ascending stratigraphic order: 1) Paleozoic basement rocks including the Wood Canyon Formation, Zabriskie Quartzite and Carrara Formation; 2) the Jolly Jane Formation; 3) the Savage Formation; 4) the Pioneer Formation; 5) the Sierra Blanca Tuff; 6) the Savage Valley Dacite; 7) rhyolites of the North Bullfrog Intrusive Suite; 8) the Crater Flat Group; 9) monolithic and heterolithic debris flow breccias of the Rainbow Mountain Sequence; 10) the Trt₂ tuff of the Rainbow Mountain Sequence; and 11) the Sober Up Gulch Gravels (Table 7-2). Figure 7-10 is a geologic map of the Sierra Blanca area updated in 2016.

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The Wood Canyon Formation (PzW) and Zabriskie Quartzite (PzZ) crop out along the southwest side of Savage Valley, and both units were penetrated in drill holes under Savage Valley (Figure 7-10). The Wood Canyon Formation consists of calcareous shale, siltstone and sandstone with occasional beds of massive to finely laminated limestone. The Zabriskie consists of light brown, pink or light grey, non-calcareous to weakly calcareous vitreous quartzite. The Carrara Formation (PzC) overlies the Zabriskie and consists of primarily carbonaceous calcareous shale, argillaceous limestone, micritic limestone and sandy limestone. PzC is the primary bedrock unit penetrated under Savage Valley.

The Jolly Jane Formation at Sierra Blanca is only known from drilling under Savage Valley. It includes a heterogeneous sequence of: 1) Paleozoic clast-dominated conglomerate; 2) calcareous and carbonaceous lithic-volcanoclastic sediments that appear to be largely re-worked dacite; and 3) rare monolithic debris flow breccias of the Carrara Formation. The Jolly Jane Formation varies dramatically in thickness from 0-35 metres between drill holes, and is generally thinner at Sierra Blanca than at Jolly Jane. Overlying the Jolly Jane Formation, the Savage Formation consists of dacitic tuffs, flows and intercalated dacitic sediments. The lava flows thicken to the south where they appear connected with the Savage Plug: a rhyo-dacite porphyry plug-dome that is part of the North Bullfrog Intrusive Suite.

The Pioneer Formation consists of felsic pyroclastic rocks including pale green crystal-lithic tuff and lithic lapilli tuff with common white rhyolite clasts. Near the upper contact, the lithic content increases significantly and grades into a coarse heterogeneous tuffaceous epiclastic sequence known as the Upper Epiclastic Member. The Pioneer Formation ranges in thickness from tens of metres in the south to over 250 metres in the north. The base of the unit has never been drilled in the North Sierra Blanca area. The Pioneer Formation exhibits varying degrees of quartz-adularia alteration and hosts significant gold mineralization.

In the northern and central Sierra Blanca area, rhyolite bodies of the North Bullfrog Intrusive Suite occur stratigraphically below, within and above the Pioneer Formation. The rhyolites exhibit flow-banding, spherulitic textures, and are often highly brecciated. Rhyolite breccias may be intrusive or locally extrusive. Based on zircon age dating and geochemical similarities, Tnb rhyolite volcanism is in part coeval with deposition of the Pioneer Formation. Rhyolite flows are also present within the Savage Valley Dacite stratigraphy above the Sierra Blanca Tuff. The various rhyolite bodies appear to represent episodic intrusive and extrusive phases of rhyolite volcanism contemporaneous with the Pioneer Formation, Sierra Blanca Tuff and Savage Valley Dacite.

The Sierra Blanca Tuff varies in thickness from >160 metres in the north to 30 metres in the south. A 5-15 metre thick distinctive welded shard-rich interval (Tsb1) at the base of the Sierra Blanca Tuff serves as a marker for the bottom of the unit. Above the shard-rich marker, the tuff is a relatively homogeneous, densely welded, crystal tuff with well-developed compaction foliation exhibited by flattened pumice fragments. The tuff is typically fractured and brecciated, and is the primary host rock for gold mineralization.

The Sierra Blanca Tuff is unconformably overlain by the Savage Valley Dacite. The Lower Member is a heterogeneous pyroclastic and epiclastic sequence of dacitic composition. The Lower Member contains a discontinuous rhyolite flow breccia that is assigned to the North Bullfrog Intrusive Suite. The Upper Member is a more uniform sequence of dacitic to andesitic lava flows and breccias. There are dramatic lateral changes in the unit that make it difficult to correlate internal stratigraphy between drill holes. Structurally-controlled alteration-style mineralization is common in the Savage Valley Dacite.

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Figure 7-10 Geologic Map of the Greater Sierra Blanca-YellowJacket Area

 

The Lithic Ridge Tuff overlies and inter-fingers with the Savage Valley Dacite. The Lithic Ridge Tuff is a poorly to moderately welded crystal-lithic ash-flow tuff, which typically exhibits abundant chloritized biotite phenocrysts. Lithic clast content varies from 5-20+% lithic clasts. The most common lithic clasts are from the underlying Savage Valley Dacite. The Lithic Ridge is highly altered, locally mineralized near the Liberator Fault, but not a significant volume host to gold mineralization.

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The Bullfrog Tuff of the Crater Flat Group is locally exposed in the eastern side of the Sierra Blanca resource area where it overlies and is in fault contact with the Lithic Ridge Tuff (Figure 7-10). There is significant alteration in the Bullfrog Tuff along the Liberty Vein structure. No gold mineralization has been encountered to date in the Bullfrog Tuff at Sierra Blanca.

A significant interval of monolithic debris flow breccia of Paintbrush Tuff unconformably overlies the Bullfrog Tuff east of YellowJacket and Savage Valley. The Paintbrush breccia is not known to be mineralized at Sierra Blanca, but locally exhibits strong hydrothermal alteration and high level quartz veining, which may be related to the waning YellowJacket vein event. Heterolithic debris flow breccias overlie and locally inter-finger with the Paintbrush breccias. East of YellowJacket and Savage Valley, monolithic and heterolithic debris flow breccias are unconformably overlain by the crystal-lithic Trt₂ tuff of the Rainbow Mountain Sequence (Figure 7-4 and 7-10). The base of the Trt₂ tuff is marked by bedded epiclastic rocks that locally fill a significant erosional channel cut through the Paintbrush breccia into the Bullfrog Tuff. The 10.5 Ma Trt₂ tuff is relatively unaltered and appears to have been deposited after the last known mineralization event affecting the Sierra Blanca area.

The last unit of importance at Sierra Blanca is the Gravels of Sober-Up Gulch. These younger gravels fill the valley to the west of Air Track Hill, directly overlying the Donovan Mountain Latite. The gravel sequence in this area is only known from reverse circulation drilling. The mineralization at Air Track West appears to be a large slide block of quartz-adularia-altered rhyolite (Tnb?) hosted within the gravel unit.

Figure 7-11 Geologic Cross Section Looking North Through Savage Valley Illustrating the Style of Faulting

 

The location of Figure 7-11 shown in Figure 7-10.

7.4.2.2 Sierra Blanca Structure

The structural setting of the Sierra Blanca area is similar to Jolly Jane, but the area is at least four times larger. The relative timing of the complex structural setting at Sierra Blanca is not completely understood. The Sierra Blanca area is divided in two blocks by the Cairn Fault (Figure 7-10), which separates the Savage Valley block from the North Sierra Blanca block. The Cairn Fault appears to consist of multiple E-W-trending, north-dipping, en echelon splays, comprising a down-to-the-north zone of displacement exhibiting both pre- and post-mineral movement. The Cairn Fault is interpreted to be an accommodation structure similar to the Pioneer Shear.

The structure of the Savage Valley block is relatively simple with a series of down-to-the-east faults on the western side of the valley (Quartzite and Gap Faults) and the Savage Fault on the eastern side (Figures 7-10 and 7-11). The stratigraphy dips steeply to the east between the Savage and West Jolly Jane Faults, reflecting dramatic rotation in the hanging-wall of the West Jolly Jane Fault.

The structure in the North Sierra Blanca block north of the Cairn Fault is more complex, with an array of faults that all dip to the west, but have mixed apparent normal and reverse displacements (Figure 7-12). The dip of units in this block is quite variable but is generally 20-45° to the east-southeast. Significant down-to-the-west faults at Sierra Blanca include the Donovan Mountain, NE20, NE30 and YellowJacket Faults. The largest fault is the Donovan Mountain Fault (Figure 7-10 and 7-12), which exhibits >1 kilometre of down-to-the-west displacement. Much of this displacement appears to be post-mineral, placing unaltered Donovan Mountain Latite against mineralized Pioneer Formation. The Air Track Hill, NE30, NE20 and NW10 all appear to be truncated by post-mineral movement of the Donovan Mountain Fault (Figure 7-10 and 7-12). The Gravels of Sober-up Gulch west of Sierra Blanca overlap and cover the Donovan Mountain Fault.

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Significant down-to-the-east faults include the Air Track Hill, NE40, Liberator, Gap and Quartzite faults. The largest down-to-the-east fault is the Liberator. The Liberator locally hosts both vein and alteration-style mineralization, and also exhibits significant post-mineral movement. Some down-to-the-west faults are apparently truncated on the east by the Liberator Fault, including the YellowJacket fault and vein zone (Figure 7-12). The Liberator and YellowJacket faults are also cut by NE-trending cross faults (NE50, NE60 and other smaller faults). The NE-trending cross faults appear to have the youngest movement in the area. The Liberator Fault is apparently truncated on the north by the NW10, and on the south by the Cairn Fault. The NW10 fault is a WNW-trending, down-to-the-north fault at the north end of the YellowJacket Vein Zone. The NW10 fault is mineralized but also interpreted to have post-mineral movement.

The fault pattern at Sierra Blanca consists of generally smaller-scale faulting that has affected the blocks between the larger faults (Figure 7-10, 7-12 and 7-13). With the onset of accelerated extension during the 11.4-10.5 Ma event, any older faults riding in hanging-walls would be reactivated and rotated to the east. This means that early east-dipping faults have rotated and steepened, and early west-dipping faults would have rotated and flattened. The NE40 fault at Sierra Blanca currently exhibits apparent reverse displacement, probably as a result of eastward rotation of what was a formerly east-dipping normal fault (Figures 7-12 and 7-13).

Figure 7-12 Simplified Geologic Cross Section Looking North through Sierra Blanca-YellowJacket

The location of the section shown in Figure 7-12 is shown on Figure 7-10.

7.4.2.3 Sierra Blanca Mineralization

Mineralization at Sierra Blanca can be classified into the following styles:

· Early, large-volume, pervasive low-grade disseminated gold mineralization associated with quartz-adularia-pyrite alteration.

· Middle, structurally-controlled disseminated gold mineralization associated with overprinting illite-adularia-pyrite alteration and gold enrichment (NW10, NE30, NE50, NE60, Liberator Faults).

· Late, high-grade gold associated with quartz veins and quartz stockwork veining (YellowJacket Vein Zone).

The metallurgical characteristics of these styles of mineralization have resulted in the differentiation of three metallurgical classes for the estimate the Sierra Blanca Mineral Resource: 1) disseminated oxide; 2) disseminated sulphide; and 3) quartz vein and stockwork mineralization.

7.4.2.3.1 Pervasive Disseminated Mineralization

The Sierra Blanca Tuff and Pioneer Formation exhibit alteration-style disseminated mineralization over virtually the entire Sierra Blanca area. The Sierra Blanca Tuff is ubiquitously altered to a fine-grained mixture of quartz and adularia with disseminated pyrite. Quartz-adularia alteration is not as widespread or as well developed in the underlying Pioneer Formation, which is more typically altered to a light green smectite-illite+adularia-chlorite assemblage. The tuffs are progressively altered to smectite-illite, illite-adularia and finally to quartz-adularia as the alteration intensity and gold grade increases. The specific controls on alteration development are difficult to constrain with wide-spaced drilling. However, it appears that alteration is controlled by a combination of high-angle structures feeding fluids into permeable stratigraphic intervals.

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Disseminated mineralization in the un-oxidized portions of the deposit is directly related to the pyrite content, which in turn reflects the original iron content of the rock. Pyrite morphologies in disseminated mineralization include fine to coarse disseminated, biotite-replacement, lithic clast-replacement and veinlet infill; indicating a complex history with multiple generations of pyrite growth. Some pyrite grains show zoned gold and arsenic, but no consistent pattern was observed in the grains studied (AMTEL Report 11/34, 2011). A gold deportment study carried out on the disseminated mineralization revealed that most of the gold is held in the lattice of the disseminated pyrite. When the pyrite is oxidized, the gold is readily recoverable with simple cyanide (AMTEL Report 11/34, 2011, also see Section 13).

7.4.2.3.2 Structurally-controlled Disseminated mineralization

Structurally-controlled alteration-style mineralization is characterized by the development of disseminated pyrite within faults and in proximal wallrock directly adjacent to faults. These zones may be over 10 metres wide locally. Many faults host this style of mineralization including the NE20, NE30, NE40, NE50, NE60, NW10, Air Track Hill, and the Liberator Faults. Structurally-controlled alteration-style mineralization overprints the earlier pervasive quartz-adularia disseminated mineralization, and results in consistently higher grades (1-17 g/t gold). In the Sierra Blanca Tuff, the illite overprint results in a bleached white illite-adularia-pyrite assemblage (Figure 7-7). Where developed in the Savage Valley Dacite, the alteration is a brown illite-pyrite to illite-adularia-pyrite assemblage, which commonly contains higher grades. This is particularly notable along the Liberator Fault, where gold grades of up to 17 g/t have been encountered in the dacite. This mineralization has a consistently higher Ag/Au ratio than earlier disseminated mineralization.

Disseminated mineralization has been modeled as oxide disseminated and sulphide disseminated. Oxide disseminated is amenable to heap leaching. Sulphide disseminated is amenable to flotation concentration and either ambient air oxidation or pressure oxidation recovery of gold from a sulphide concentrate.

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Figure 7-13 Geologic Map of the YellowJacket Area Showing Major Faults and Drill Holes Related to Discovery of the High-grade Vein System

7.4.2.3.3 Quartz Vein Mineralization

The YellowJacket Vein Zone occurs in the northeast portion of the Sierra Blanca resource area (Figure 7-10 and 7-13). The YellowJacket Vein Zone consists of the massive YellowJacket Vein surrounded by hanging wall and footwall stockwork vein zones. The YellowJacket Vein Zone was discovered with drill hole NB-12-138, and was systematically drilled-out with core in 2013-14. The YellowJacket Vein Zone strikes north-northwest and dips between 65-75° west. The zone varies between 15-35 metres wide, and persists over a strike length of ~850 metres. The main YellowJacket Vein is continuous in drill holes along ~700 metres of strike length. The vein zone is entirely blind and not recognizable at the surface. The surface projection of the YellowJacket Vein Zone is shown on Figure 7-13. The continuity of the vein and stockwork zone along strike is remarkably consistent. Quartz vein and stockwork mineralization appears to overprint all earlier alteration-style mineralization.

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High-grade quartz veins often exhibit crustiform banding and bladed quartz pseudomorphs after calcite typical of low-sulphidation epithermal veins. The quartz vein mineralogy is very simple and consists of native gold and electrum with varying amounts of acanthite and accessory silver sulphosalts. The high grades of the quartz stockwork zones are typically carried by <1-2 cm grey translucent quartz veinlets containing visible gold and trace amounts of pyrite. Metallurgical testing has shown that the massive quartz vein and quartz stockwork mineralization is free milling. Other YellowJacket-style veins and stockwork zones have been penetrated by drilling in the North Sierra Blanca area outside of the YellowJacket Vein Zone. These are generally small-volume veins or stockwork zones that locally carry high-grades. Most of these subsidiary vein zones appear to be controlled by NE-trending cross faults (i.e. NE50 and the Rhyolite Vein in the vicinity of the NE20). The NE faults are kinematically linked to the YellowJacket Vein structure and served as vein fluid conduits. Potential exists to expand NE-trending or other subsidiary vein zones with additional drilling. The better grade-thickness intercepts are likely associated with intersections between the YellowJacket Fault and NE-trending cross faults.

7.4.2.3.4 Gold and Tellurium Mineralization

At Air Track Hill, another style of mineralization consists of quartz-free disseminated gold with anomalous tellurium. Unfortunately, this mineralization is completely oxidized, so the original character is not known. This style of mineralization is amenable to heap leaching. Based on the core drilled through this interval in NB-13-364, the mineralization is most likely related to the occurrence of hairline pyrite veins in volcanic rocks.

7.4.3 Air Track West

Air Track West is a zone of mineralization located under alluvial cover to the west of Sierra Blanca (Figure 7-10). Air Track West was originally discovered by Sunshine Mining in 1991 when they found what appears to be detrital boulders of strongly quartz-adularia-altered volcanic rock in the pediment area ~500 metres west of Air Track Hill. The boulders have yielded gold values up to 0.273 g/t. Sunshine Mining drilled the discovery hole GS-45 which yielded 17.8 m (15.2-32 m) grading 1.81 g/t gold. Sunshine Mining did some additional drilling in the area and produced additional low-grade intercepts. Corvus drilled one hole (NB-12-117) in 2012 and confirmed this mineralization with an intercept of 15.2 m (10.7-25.9 m) grading 2.36 g/t gold. The mineralization at Air Track West appears to be a large, dismembered slide block of quartz-adularia-altered rhyolite (Tnb?) hosted within the Gravels of Sober-up Gulch. The mineralized block is interpreted to have slid westward from the Sierra Blanca-Air Track Hill area into the hanging-wall of the Donovan Mountain Fault. An alternative interpretation is that the gravels have been mineralized by an unknown younger gold event. A small Mineral Resource has been defined at Air Track West.

7.4.4 Mayflower

The historic Mayflower mine was developed on en echelon quartz-calcite veins and stockwork zones along the NW-striking, steeply SW-dipping Mayflower Fault Zone (MFZ). The MFZ consists of multiple SW-dipping fault splays with a complex network of fractures linking the main strands together. Based on the displacement of the base of Trt₂ on Section 31 in Figure 7-14, the total apparent vertical displacement across the MFZ is approximately 60 metres (Figure 7-15). There appears to be three main fault splays within the MFZ in the southeastern part of the deposit. Two of these splays merge in the vicinity of the David Adit and only two splays remain in the northwestern end of the deposit (Figures 7-15). High-grade mineralization appears to be best developed in the fracture zones along and adjacent to the main fault splays. Dilation caused by differential movement between the fault splays appears to be the main control on disseminated mineralization. The entire Mayflower deposit is oxidized.

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Figure 7-14 Geologic Map of the Mayflower Area Showing Underground Workings, Drill Holes and Cross Section Locations

 

The mineralized zone is traceable for ~900 metres along strike (Figure 7-4 and 7-14). The bulk of the mineralization occurs in debris flow sediments of the Rainbow Mountain Sequence, but locally extends upward into the overlying Rainbow Mountain Tuff (Trt₂). Alteration-style quartz-adularia disseminated mineralization surrounds a steeply dipping zone of vein-filled breccias. Quartz-adularia alteration extends out into the debris flow sediments for several tens of metres around the main fault zone. Certain horizons appear more permeable than others and therefore are altered and mineralized for greater distances. Alteration pinches to only a few metres wide along faults in Trt₂.

Vein and stockwork mineralization is proximal to individual fault splays. Late stage black manganiferous calcite occurs as veins and breccia fillings along the mineralized structures. This calcite is mineralized and is thought to be a late vein stage of the waning hydrothermal system. Narrow high-grade gold zones (shoots) are known from the historic workings and from Corvus drilling. Overall, the mineralized zone appears to narrow with depth, and has a steeper more planar hanging-wall than footwall (Figure 7-15). There is a clear correlation between higher gold grades and arsenic, both of which are associated with adularization of the host rocks (Myers, 2008).

The Mayflower inclined shaft was initially developed in the early 20th Century on the main Mayflower splay of the MFZ. Historical records indicate the shaft developed four levels, with mining occurring on the 200, 300 and 400 levels. The bulk of the production came from the 300 level (Spencer, 1919). Based on the Spencer (1919) map, it appears that approximately 17,000 tonnes of material had been extracted by that time. The David Adit and Starlight workings were developed on the David Adit splay, which is the footwall to the Mayflower splay. The David Adit was driven to explore the northwest extension of the system.

The Mayflower prospect was the focus of modern exploration and drilling by numerous companies starting in 1982. Drilling results have been collected for most of the drill holes. Original assay certificates are available for the Barrick drilling. In 2008, Corvus (International Tower Hill under the NBPJV) drilled 24 reverse circulation (RC) holes totaling 5,953 metres (19,531 feet) in the Mayflower area. In 2012, Corvus drilled 52 additional holes totaling 7,503 metres (24,615 feet) including: 1) 14 PQ3 core holes totaling 1,922 metres (6,306 feet); 2) 26 in-fill/definition RC holes totaling 3,077 metres (10,095 feet); 3) seven condemnation RC holes totaling 1,218 metres (3,500 feet); 4) four water monitor wells (RC) totaling 981 metres (3,220 feet); and 5) one water pilot RC hole totaling 305 metres (1,000 feet). The PQ3 core holes have been used for additional metallurgical testing including bottle roll and column leach tests. Data for both the Corvus drilling and historic Barrick drilling are used in the Mineral Resource estimate in Section 14.

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Figure 7-15 Cross Sections Looking Northwest Through the Mayflower Deposit

For the cross section locations of Figure 7-15 see Figure 7-14.

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8 Deposit Types

Gold mineralization in the Bare Mountain sub-district includes a variety of types: 1) sediment- and intrusive-hosted disseminated Carlin-like (Sterling, Mother Lode, SNA); 2) sediment-hosted Carlin-like (Daisy West and South); 3) volcanic-hosted low sulfidation disseminated epithermal (Secret Pass); and 4) sediment-hosted quartz vein stockwork (Reward). Types 1, 2, and 3 are associated with late fluorite vein mineralization, high As-Sb-Hg, and very low base metals. Type 4 has no fluorite association, very low As-Sb-Hg, and elevated base metals (Weiss, 1996). Most of the gold deposits in the greater Bullfrog district range in age from ~13 Ma to 9.5 Ma. However, quartz stockwork veins associated with the Reward deposit are cut by unmineralized andesitic dikes dated at 26 Ma (Monsen, 1992). The Reward deposit is older and apparently unrelated to the other deposits in the district (Weiss, 1996).

Mineralization at MLP most closely resembles Carlin-type sediment-hosted gold systems of northern Nevada. The Mother Lode deposit model includes structurally and stratigraphically-controlled disseminated sulfide and oxide mineralization hosted in primarily in Tertiary sedimentary rocks and rhyolite porphyry dikes. Paleozoic sedimentary rocks comprise a smaller volume host in proximity to dikes and the basal Tertiary unconformity. Tertiary volcanic rocks and debris flow breccias are also mineralized at Mother Lode. Alteration types associated with gold mineralization include passive decalcification and illite-pyrite alteration. Unique features of Mother Lode-style mineralization compared to typical Carlin-type deposits include: the association with fluorine, the significant volume of mineralization that is rhyolite porphyry dike-hosted, elevated tellurium and bismuth, and the generally passive, innocuous, very low-silica nature of the mineralization. Carlin-type gold deposits are generally known to form at depths of >2 kilometres, at temperatures between 180-240^oC. Mineralization at Mother lode may have formed at depths of 500-1000 metres or less, and at temperatures <240^oC (Weiss, 1996).

Gold mineralization in the main Bullfrog and North Bullfrog sub-districts is characterized as volcanic-hosted low-sulfidation epithermal type. Two styles of epithermal precious metal mineralization are present: 1) disseminated mineralization associated with sulfidation of iron in the volcanic host rocks; and 2) open-space filling quartz and/or carbonate veins, which are controlled by boiling. Pervasive quartz-adularia alteration of volcanic rocks is intimately associated with the disseminated mineralization and is an important ground preparation process for later vein forming events. Epithermal deposits form at shallow depth, from the surface to generally <2 kilometres. Temperatures of formation range between 150 to 300°C. Mineralization at NBP is typical of other low-sulfidation type gold systems in and around the Walker Lane trend such as: Bullfrog, Round Mountain, Rawhide, Aurora, Bodie and Comstock. These deposits commonly contain higher grade gold in vein and stockwork zones surrounded by zones of lower grade disseminated mineralization. This is the accepted exploration model at NBP.

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9 Exploration

Exploration is ongoing at both the MLP and NBP as discussed separately below. Exploration at the Project has been conducted through drilling and surface mapping and rock sampling.

9.1 MLP Exploration

The exploration potential of the MLP. The focus of 2017 Phase 1 and 2018 Phase 2 drilling programs has been to delineate open pit resources at the Mother Lode deposit area. The deposit remains open at depth to the west, north and east of the current limit of drilling. There is still excellent potential for additional open pit and underground mineable resources. The primary focus of Phase 3 drilling in the fall of 2018 will be extending the mineralization at Mother Lode West. Exploration target areas for MLP are shown on Figure 9-1 include Flatiron, Willys, Sawtooth, Coronet, Twisted Canyon, and Baileys Gap. Each of the exploration targets are described briefly in the follow sections.

9.1.1 Flatiron

The Flatiron target is located 700 metres southwest of the Mother Lode deposit. GEXA was originally attracted to the Mother Lode area in 1983 because of the gold and cinnabar-bearing silicified breccia along the Fluorspar Canyon Fault (FCF) at Flatiron. The first 54 holes at the original Mother Lode property were drilled at Flatiron by GEXA in 1985. Nearly all of these early holes intersected between 3-15 metres of low grade gold mineralization (0.5 to 1 g/t, with values up to 5 g/t). Three holes were drilled by Corvus in early 2018 to test the down-dip extension of the known mineralization along the FCF. All three holes (ML18-056, -057 and -058) cut gold mineralization similar to the historic drilling (see Table 10-1). The style of mineralization at Flatiron is identical to Mother Lode. The Flatiron mineralization is interpreted to be a distal extension of the Mother Lode deposit, and ultimately may connect with Mother Lode mineralized zone at depth and along strike. Follow-up drilling is planned for the fall of 2018.

9.1.2 Willys

The Willys target area is located on the ME claim block, 2 km east of the Mother Lode pit. The Willys target area is underlain by the Sedimentary Rocks of Joshua Hollow (Tjs), the Lithic Ridge Tuff (Tlr), pre-Rainier Mesa Tuff (Tprt), the Rainier Mesa Tuff (Tmr) and older gravels (Tgs). There are also small outcrops of altered quartz-porphyry rhyolite dikes exposed beneath Tgs gravels along the west side of the claim block. The dikes appear to be intruding Tjs and Tlr, in a similar setting to the Mother Lode deposit. The first pass of surface rock sampling yielded no significant gold values. Recessive exposures of altered and iron-stained Tjs and Tlr have yielded arsenic values up to 439 ppm. The alteration and trace element geochemistry exhibited in the Tlr unit at Willys is similar to what is exposed on the north wall of the pit above the Mother Lode orebody. The Willys area has excellent potential for discovery of new blind Mother Lode-style gold mineralization hosted in Tjs and Tip dikes below Tlr. The dikes at Willys lie along strike of a major north-northeast-trending dike swarm that is well-exposed just one kilometre to the south. Four to six RC holes are planned at the Willys target area in the fall of 2018.

9.1.3 Sawtooth

The Sawtooth target is located 7 kilometres west of Mother Lode on the western portion of the MN claim block in lower Fluorspar Canyon. Geologic mapping has identified the Sawtooth fault: a large-displacement, NNE- trending, down-to-the-west normal fault. The Sawtooth fault juxtaposes Ammonia Tanks Tuff (Tma) in the hanging wall against pre-Rainier Mesa Tuff (Tprt) in the footwall. The fault is largely covered by alluvium. The footwall of the fault also includes exposures of lower Bullfrog Tuff (Tcb) below the Tprt unit. The Tram Tuff (Tct), the Lithic Ridge Tuff (Tlr) and possibly Tjs are all expected to lie at depth in the footwall of the fault. The down-to-the-west displacement on the fault could be 500 metres or more. Limited rock sampling has identified anomalous arsenic (up to 160 ppm) and antimony (up to 15 ppm) in strongly altered Tcb in the footwall of the fault. The structural and stratigraphic setting is similar to the MP Fault at the Bullfrog mine. The Sawtooth fault is expected to sole into the Fluorspar Canyon Fault just 400 metres south of the MN claim block. A couple of holes are planned in the fall of 2018 to test for blind vein or disseminated gold mineralization in the footwall rocks of the Sawtooth Fault.

9.1.4 Coronet

The Coronet target area is located 1.5 kilometres west of Mother Lode on the eastern portion of the MN claim block. Initially, the target was identified as having a west-dipping, down-to-the-west fault that projects to the north-northeast from Secret Pass pit area. Subsequent mapping has resulted in some ambiguity about the nature of the structure and stratigraphy in the area. However, first pass rock sampling has yielded Au values up to 104 ppb and As up to 247 ppm. Historic shallow drilling in the Coronet area has not intersected any significant gold values. Additional work is required to refine the Coronet as a drill target.

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9.1.5 Twisted canyon

The Twisted Canyon target area is located 1.5 kilometres north of Coronet, and 2.5 kilometres northwest of Mother Lode (Figure 9-1). The area lies along the eastern margin of what Fridrich (2007) describes as the Twisted Canyon Topographic Wall. First pass mapping and sampling has yielded a sample with 65 ppb Au and 2.37 ppm Hg associated with high level argillic alteration. The alteration is hosted in pre-Rainier Mesa rhyolite and tuff, indicating a younger age of alteration and mineralization than Mother Lode. Additional field work is required to develop a drill target in this area.

Figure 9-1 MLP Target Location Map

 

9.1.6 Baileys Gap

The Baileys Gap target area is located 6-9 kilometres northwest of Mother Lode, just southeast of the Baileys target area in the Eastern Steam-heated Zone of the NBP (Figure 9-1). The entire target area is covered by post-mineral Tertiary tuffs and gravels, including the Spearhead Member of the Stonewall Flat Tuff (7.5 Ma, Sawyer et al, 1994) and the Gravels of Sober-up Gulch (Tgs). The Baileys target area exhibits intense quartz-alunite-kaolinite steam-heated alteration, which appears to extend under cover to the southeast beneath the Baileys Gap area. Corvus believes the Eastern Steam-heated Zone at NBP is contiguous with the large steam-heated alteration zone present at the Silicon mine area to the southeast (see Figure 9-1). Additional work, including geophysical surveys, is required to generate drill targets at Baileys Gap.

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9.2 NBP Exploration

Despite the substantial amount of work that has been done, the exploration potential of the NBP is still significant and the Project remains under-explored. The blind discovery of the Yellowjacket high-grade vein/stockwork deposit in 2012 and the identification of the extensive largely untested Eastern Steam-heated Zone in 2014 indicate significant exploration potential for the discovery of new blind high-grade deposits. Opportunities for expanding the NBP resources include: 1) possible continued expansion of the Yellowjacket vein deposit at depth; 2) new discoveries of blind, high-grade Yellowjacket-style vein systems adjacent to or within current disseminated resources in the Western Resource Area; 3) expanding or identifying new disseminated mineralization at target areas outside of the existing Mineral Resource boundaries; and 4) new discoveries of either high-grade vein or disseminated mineralization under the Bullfrog-age Eastern Steam-heated Zone (Figure 9-2).

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Figure 9-2 NBP Exploration Target Location Map

 

9.2.1 Western Resource area

9.2.1.1 Yellowjacket Vein Zone

The Yellowjacket Vein Zone has largely been closed off by drilling along strike, but remains open at depth on some sections. Potential for the discovery of additional blind Yellowjacket-style high-grade veins exists within and adjacent to the disseminated resource areas. Much of the early resource definition drilling of the disseminated mineralization was vertical and was not effective at defining through-going, steeply dipping vein targets. Infill drilling of the low-grade oxide resources may discover new blind high-grade veins, particularly hosted in the Sierra Blanca Tuff.

9.2.1.2 Swale

The Swale target lies along strike of the north-northwest projection of the Yellowjacket Vein Zone (Figure 9-2). Several east-directed angle holes were drilled in the Swale area looking for the northern extension of the Yellowjacket Vein. Most of the holes drilled at Swale encountered 1+ g/t Au sulfide mineralization, some of which is associated with quartz stockwork veining. Much of this sulfide mineralization appears to be too deep to support open pit mining. The current drilling has not confirmed a northern extension of the Yellowjacket Vein Zone. However, hydrothermal alteration and sulfide mineralization in the Sierra Blanca Tuff, Pioneer Formation and underlying Tnb rhyolite indicate the mineralizing system is still strong in the northernmost drilling at Swale. Additional deep drilling (400-600+ metres below surface) will be required to continue to test the underground vein and disseminated sulfide potential along strike of the Yellowjacket Vein Zone.

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9.2.1.3 Cat Hill

Cat Hill lies in the footwall of the Road Fault, just south of the Connection area (Figure 9-2). As at Connection, Cat Hill is underlain by heterolithic and monolithic debris flow breccias of the Rainbow Mountain Sequence. Alteration at Cat Hill is characterized by overlapping assemblages of both steam-heated alteration and silicified ribs with quartz veining exhibiting boiling textures (quartz pseudomorphs of bladed calcite). Surface rock chip sampling yielded up to 1.4 g/t Au from quartz vein material. Fine vuggy, northeast trending silicified ribs with quartz replacing calcite coexist at the same elevation as an opal-kaolinite-alunite assemblage that locally exhibits alunite veinlets. The current interpretation of the alteration at Cat Hill is that it suggests a fluctuating paleo-groundwater table, which has resulted in the juxtaposition of contrasting styles of hydrothermal alteration.

Five east-directed angle holes were drilled at Cat Hill between 2015 and 2017. Hole NB-15-284 encountered multiple 10-40 metre intercepts at >0.15 g/t Au, including two individual samples at >1 g/t Au. Two other holes also have significant low-grade oxide intercepts. Most of the mineralization encountered in the drilling at Cat Hill is oxidized. The Corvus drilling suggests that a small oxide resource can be developed at Cat Hill with additional drilling.

9.2.1.4 Connection

The historic Connection shaft and prospects were developed at Connection in the early 1900’s. Connection lies just east of and in the footwall of the Road Fault. Between 1974 and 1982, Cordex drilled a number of shallow holes and delineated a small mineralized zone.

Five general lithologic units have been identified in the Connection drill area including: 1) probable rooted Bullfrog Tuff at depth; 2) a monolithic debris flow breccia unit of Paintbrush Tuff (Tdf_p); 3) a monolithic debris flow unit of Rainier Mesa Tuff (Tdf_mr); 4) a massive slide block of mixed Paleozoic lithologies (Tdf_C); and 5) a quartzite-clast-dominated heterolithic debris flow (Tdf_h) which caps Connection Hill. All of these local units lie within the Rainbow Mountain Debris Flow Sequence. The true thickness of each unit is unknown and is expected to be highly variable in such a debris flow environment.

Corvus has drilled a total of four holes in the Connection area and has not been able to expand the mineralized zone sufficiently to define significant Mineral Resource. Since the host rocks are part of the Rainbow Mountain Debris Flow Sequence, this mineralization is most likely part of the younger ~10 Ma hydrothermal activity. Given the style of sulfide mineralization in the Bullfrog Tuff at depth, the area may have also been affected by older mineralizing events. It is possible that the Tdf_C slide block was mineralized elsewhere and then mass-wasted to the present location during later extensional deformation.

9.2.1.5 Liberty Vein

The Liberty Vein is located along two historic workings that lie along strike of the southeast projection of the Yellowjacket Vein Zone (Figure 7-8). The workings contain a NNW-trending, steeply west-dipping quartz vein/replacement zone, hosted in strongly clay-altered Bullfrog Tuff. A NE-trending set of cross fractures is also present. The target concept was that the Liberty Vein is a high-level surface expression of the southeast extension of the Yellowjacket Vein Zone. A fence of two angle holes (NB-15-427, 428) was drilled under the historic workings. The Bullfrog Tuff and underlying Lithic Ridge Tuff were not significantly mineralized, and no significant quartz veining was found in either hole. The Savage Valley Dacite in the bottom of NB-15-428 is weakly mineralized (sulfide, up to 0.19 ppm). The Liberty Vein structure exposed at the surface does not appear to be rooted by any significant gold mineralization. However, these holes failed to target the structure at a deep enough elevation to test for a vein zone in the Sierra Blanca Tuff.

9.2.1.6 Cloud 9

The Cloud 9 target was generated in 2016 from additional mapping along the West Jolly Jane Fault to the north of the North Jolly Jane target area (Figure 9-2). An historic prospect and two small outcrops exhibiting banded manganiferous calcite and quartz veining were discovered along the West Jolly Jane Fault ~1.2 kilometres north of the drilling at North Jolly Jane. The hanging wall unit is Trt2 tuff, and the footwall unit is heterolithic debris flow breccia dominated by Paleozoic clasts. Rock chip sampling has not yielded any significant gold or trace element geochemistry. One east-directed angle hole (NB-16-307) was drilled to test the fault for veining at shallow depth below the vein occurrence. No significant gold was encountered on the fault, but the footwall rocks at depth were found to be relatively unaltered, in situ Wood Canyon Formation basement. The drill hole is significant in that it shows that pre-Tertiary basement stratigraphy has returned to near surface in the footwall block of the West Jolly Jane Fault. This implies that the Sierra Blanca Tuff may also return to near surface between North Jolly Jane and Cloud 9. The evidence suggests the existence of a new shallow oxide target in Sierra Blanca Tuff between North Jolly Jane and Cloud 9. Additional drilling is warranted in this area.

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9.2.1.7 Jim Dandy

The Jim Dandy target is named for the Jim Dandy patented claim, which lies ~300 metres north-northeast of the Pioneer Mine (Figure 9-2). The Jim Dandy fault is NNE-trending, steeply west-dipping and has yielded anomalous Au in surface rock samples. Historic drill hole P92-9 in this area encountered 1+ g/t Au in the top of the hole along the trace of the Jim Dandy fault. Two east-directed angle holes (NB-17-439, 440) were drilled in 2017 to test the Jim Dandy fault. Both holes encountered intensely altered rocks with quartz veins and stockwork zones. However, the holes intersected only narrow zones of anomalous gold (max. 0.42 ppm). No future work is recommended at Jim Dandy.

9.2.1.8 East Savage Vein

The East Savage Vein lies just east of the south end of Savage Valley (Figure 9-2). The East Savage Vein was identified by previous explorers at NBP, but had only one historic drill hole. The vein has yielded an adularia date of 11.0 Ma (Connors et al., 1998), a similar age to the Yellowjacket Vein. The East Savage Vein is NNE-trending, steeply west dipping, and cuts through Sierra Blanca Tuff, Pioneer Formation, and a large Tnb rhyolite body. There is anomalous Au (max. 0.640 ppm) in surface rock samples. The vein zone persists along strike for nearly 500 metres. Two east-directed angle holes (NB-17-441, 442) were drilled to test nearly 250 metres of strike length. Both holes drilled thick intervals of intensely quartz-adularia-pyrite-altered Tnb rhyolite, and eventually penetrated into PzC basement. No significant quartz veins were encountered at depth, and only anomalous gold (max. 0.130 ppm) was found in the intensely altered rhyolite. No future work is recommended at the East Savage Vein.

9.2.1.9 Jasperoid

The Jasperoid target is located south of the Sierra Blanca Mineral Resource (Figure 9-2). The target is an occurrence of mineralized bedding-parallel jasperoid in limey beds of the Wood Canyon Formation, Barrick drill hole RDH-767 intersected 21 metres at 0.35 g/t oxide in this area (including 1.52 metres at 1.1 g/t). This intercept has never been followed up. The Wood Canyon Formation is the host unit at the Reward deposit south of Beatty. Little attention has been given to the Wood Canyon Formation as an ore host at NBP. Additional drilling is recommended in this area.

9.2.1.10 Road Fault

The Road Fault is the northern continuation of the Contact Fault from the southern Bullfrog Hills. It is one of the largest displacement faults at NBP. It has gold mineralization in the immediate footwall at Cat Hill and Connection, and steam-heated alteration widely distributed in the hanging wall. The surface characteristics of the fault and its linear extent suggest the Road Fault was a significant structural conduit for hydrothermal fluids. It remains largely untested at depth along the entire strike length across the NBP.

9.2.1.11 West Connection

The West Connection vein zone lies ~500 metres west of the Connection area (Figure 7-14). The vein consists of a zone of high-level chalcedonic quartz and quartz-flooded breccia up to 50 feet wide, hosted within silicified monolithic debris flow breccias of the Paintbrush and Rainier Mesa Tuffs. The vein zone strikes N5E, dips 70-80° to the east, and persists along strike for ~250 metres. The vein zone has formed along one or more hanging wall splays of the West Connection Fault, which is antithetic to the Road Fault. The fault and vein geometry suggests that the vein fluids may have ascended from the Road Fault at depth.

In 1992, Pathfinder Exploration drilled a shallow hole (P92-3), which encountered 6.1 metres of 0.243 g/t Au from 48.7 to 54.9 metres at the south end of the vein zone. This intercept has better gold grades than any of the surface rock sampling, indicating that the gold tenor may be increasing with depth. The vein zone has yielded surface trace element values up to 503 ppm As, 33 ppm Sb, and 2.74 ppm Hg. Corvus drilled one hole at West Connection in 2011 (NB-11-77). The hole did not encounter significant quartz veining, but it intersected a zone of anomalous gold (>0.1 ppm) gold between 94 to 107 metres, and several zones with anomalous arsenic and antimony. Well crystallized hydrothermal kaolinite was found in a number of intervals in this hole, possibly linking this structure to the opalite alteration along the Road Fault. There has been no follow-up drilling in this area since 2011.

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9.2.1.12 Pioneer

The historic Pioneer workings are located immediately north of the Mayflower Mine (Figure 9-2). A series of underground workings were developed at Pioneer in the early 1900’s. Little is known about the production or the nature of the mineralization extracted. Based on maps of historical underground workings, mineralization appears to occur along intersecting northeast and northwest striking faults. Alteration styles from the waste dumps include silicification, adularization, argillization, and minor quartz veining. Fault zones in the Pioneer area also host argillized dacite dikes, which are compositionally similar to the Savage Valley Dacite.

Much of the historic drilling as well as surface and underground sampling demonstrates that the bulk of the unmined mineralization at Pioneer is low-grade (<1 g/t Au). Most of the higher grade gold samples came from the upper levels of the Pioneer mine with grades over a few metres of 1-14 g/t Au. During 2007, the NBPJV drilled two holes to investigate the Pioneer mineralization. The first hole targeted the down dip extension of the mineralization. Low grade mineralization was intersected across 130 metres, with a maximum value of 0.26 g/t Au in quartz-adularia-altered Sierra Blanca Tuff. A second hole was designed to drill across known higher grade mineralization. This hole encountered a total of eight metres of 2 g/t Au, including 17.6 g/t over 0.4 m, on either side of a 3.5 metre wide stope. The high-grade interval in the core hole is in a clay altered fault zone without visible quartz veining. Bladed pseudomorphs of quartz after calcite are found in outcrop at Pioneer, but such veining lacks significant grade. No new drilling has been undertaken at Pioneer since 2007, and no Mineral Resource has been established.

9.2.2 Eastern Steam-heated Zone

Geologic mapping, rock and soil sampling, age dating and a gravity survey was conducted by Corvus over the Eastern Steam-heated Zone (“ESHZ”) in 2014 and 2015. In early 2015, a gravity survey was completed in order to define possible structures beneath the extensive alteration zone which covers approximately 142 kilometres (Figure 9-3). Also in 2015, an extensive soil grid consisting of 3,672 samples was completed (Figure 9-4). The ESHZ is a broad area characterized by resistive low-temperature opal-chalcedonic silica accumulations (ribs and mounds) surrounded by recessive kaolinite-alunite alteration. Low-temperature residual silica accumulation is interpreted to define a paleo-groundwater table at the ESHZ. Residual silica forms erosional remnants of a flat tabular zone at similar elevations across much of the ESHZ. The steam heated alteration and residual silica are largely barren of metals, as would be expected at this level of erosion over a productive vein system. The target concept at the ESHZ is to test for high-grade veins in a hypothetical boiling zone below the steam-heated alteration and the paleo-groundwater table. The work that has developed this concept has defined a series of exploration targets within and around the ESHZ (Figure 9-2).

9.2.2.1 Spicerite

The Spicerite area is located at the southeastern corner of the NBP (Figure 9-2). The Spicerite area is underlain by (in ascending order): the Rainier Mesa Tuff, Pre-Rainier Mesa rhyolite flows, and heterolithic to monolithic debris flow breccias. The stratigraphy is cut by at least three through-going NNW-trending, moderate to steeply west-dipping, down-to-the-west normal faults. The most notable of these faults is the Spicerite No. 1. A fourth NNW-trending fault is down-to-the-east, and forms a graben in the hanging wall of the Spicerite No. 1 Fault. The graben has been filled with largely non-steam-heated volcanic debris flow breccia, but also contains cobbles and boulders of steam-heated rocks. It is hypothesized that the graben fill may be concealing a primary vein target at depth. The area has yielded an age date of 10.2 Ma from alunite (Weiss, et al., 1994). The evidence observed in the Spicerite area suggests a very dynamic period of faulting, hydrothermal alteration and erosion between 9.5-10.2 Ma, which is equivalent in age to the Bullfrog deposit.

Corvus drilled one angled core hole (NB-15-429) and three angled RC holes (NB-15-263, 264 and 265) on an E-W fence across the Spicerite target area. No significant gold or other metal values were encountered at the elevations reached by these holes (~300 metres below surface). All holes bottomed in low-temperature opal-kaolinite-alunite alteration. The stratigraphy encountered in the drill holes demonstrates that the Spicerite No. 1 Fault has ~700 metres of down-to-the-west displacement. This fault is likely the deep structural conduit feeding the hydrothermal fluids at Spicerite. The Spicerite No. 1 Fault is an analog of the MP fault that hosts the Bullfrog vein deposit, and should be the primary target for future drilling. The Spicerite No. 1 Fault requires deep drilling down-dip below existing holes. This represents one of the best targets for a new Bullfrog-age vein system at depth under the ESHZ.

9.2.2.2 Alunite Hill

Alunite Hill is located on the western side of the Eastern Steam-heated Zone, in an area of transition from steam-heated alteration to illite and adularia alteration (Figure 9-2). Alunite Hill is named for the abundant hypogene alunite veining in strongly silicified Paintbrush Tuff. The primary structural feature is the NW-trending, moderate to steeply SW-dipping Alunite Hill Fault. The Alunite Hill Fault juxtaposes the Paintbrush and Bullfrog Tuffs in the hanging wall against a dacite porphyry intrusive body. The wall rocks of the Alunite Hill Fault exhibit both steam-heated and non-steam-heated alteration. The fault hosts a discontinuous Au-Ag-bearing quartz vein that locally exhibits spectacular bladed quartz pseudomorphs after calcite. Rock sampling has yielded up to 0.746 ppm Au and 13 ppm Ag. The Alunite Hill and Spicerite No. 1 Faults may be the same structure propagating under the cover of the altered debris flow breccia sequence.

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Corvus drilled three holes (NB-15-260, 261 and 262) on the Alunite Hill Fault. Holes 260 and 261 comprise a fence of two angle holes under the best developed portion of the vein at the surface. Both holes intercepted quartz stockwork veining on the fault, each having with 4-6 metres intervals of low-grade Au-Ag mineralization. The drilling indicates a much flatter SW-dip of ~35^o in contrasts to the 50-75^o dips measured at surface. The initial test of the Alunite Hill Fault was not successful in identifying high-grade gold. The fault still has potential for high-grade mineralization at depth and along strike. Additional drilling is recommended at depth and along strike to fully test the Alunite Hill Fault.

Figure 9-3 Location of 2015 Gravity Stations on Complete Bouger Anomaly Data with 3rd Order Trend Removed

 

9.2.2.3 Vinegaroon

The Vinegaroon area is located just north of Alunite Hill in the hanging wall of the Vinegaroon Fault (Figure 9-2). The Vinegaroon Fault is major E-W-trending, moderate to steeply north-dipping, basement-bounding fault that juxtaposes Tertiary volcanic rocks against Paleozoic basement rocks. The Vinegaroon Fault projects eastward into the Eastern Steam-heated Zone from the Road Fault, and is truncated by the Road Fault on the west. Much of the Vinegaroon target area is in an alteration transition zone, exhibiting both quartz-adularia and steam-heated alteration assemblages. Numerous NNE- to NNW-trending high angle faults cut the debris flow breccias. Silicified ribs with anomalous gold occur along several of the faults. The interpretation is that hydrothermal fluids have ascended into these hanging wall faults from the Vinegaroon Fault at depth. Hypogene alunite, similar to that of Alunite Hill, is present in this area and has been dated by Corvus at 9.5 Ma (Table 7-4). Anomalous gold is also present associated with apparently stratabound quartz-adularia-pyrite alteration.

Corvus drilled seven holes (NB-15-286 through 292) testing a number of high-angle silicified structures, a stratabound quartz-adularia zone, and the Vinnegaroon Fault itself. The holes intersected several scattered narrow low-grade gold zones including 14 metres at 0.20 g/t Au and 1.45 g/t Ag (max. value 0.35 g/t Au), but no quartz veins. This initial wide-spaced drilling of the Vinegaroon area was unsuccessful in finding mineralization of sufficient continuity for resource definition. The area hosts significant alteration and gold mineralization, and warrants additional work in the future.

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9.2.2.4 Haul Road

The Haul Road target is located along the D & H Mining haul road to the Gold Pit. Haul Road was newly identified in 2016 (Figure 9-2). An irregular quartz vein exposed in a partially reclaimed prospect has yielded gold values up to 0.104 g/t. The vein is NE-trending, steeply NW-dipping, and hosted in monolithic debris flow breccia of Wood Canyon Formation lithologies. The target remains untested by drilling.

9.2.2.5 Burro

The Burro target area was defined primarily from a gold anomaly in soil sampling (Figures 9-2 and 9-4). There was reason to question the validity of this anomaly given that it lies in the middle of steam-heated alteration, which is generally barren of gold. Clasts of altered Paleozoic lithologies were noted weathering out of generally recessive kaolinite-alunite-altered rocks in the anomalous area. Gold values up to 1.13 ppm were obtained from the Paleozoic clasts. The conclusion is that false anomalies may be present in the Rainbow Mountain Debris Flow Sequence. Regardless of this conclusion, the Burro area exhibits strong alteration in the middle of the ESHZ, and has never been drilled. This is an ideal area for a fence of deep holes across the ESHZ to test for gold mineralization at depth below the steam-heated alteration.

9.2.2.6 Sinter

The Sinter target area is located east of Vinegaroon along the eastward projection of the Vinegaroon Fault (Figure 9-2). The Sinter target area was named for outcrops of opalized sediments, which were originally speculated to be sinter. The area hosts intense opaline and chalcedonic silicification in a debris flow sequence with no apparent structural control. The silicification at Sinter is interpreted to represent residual silica at a paleo-groundwater table. Similar to the Burro target, this is an ideal area for a fence of deep holes to test for gold mineralization at depth below the residual paleo-groundwater water table silica.

9.2.2.7 Yellow Rose

The Yellow Rose target area lies partially on patented claims near the northernmost exposure of the ESHZ (Figure 9-2). Shallow drilling in the surrounding area by Galli Exploration failed to encounter significant gold mineralization. The area is underlain by steam-heated alteration hosted in heterolithic and monolithic debris flow breccias. There is a series of N-S to NNW-trending high angle structures that cut the host rocks. There is anomalous gold in rock samples at the Yellow Rose adit in steam-heated alteration. Gold values up to 0.230 ppm area associated with illite-adularia-alteration ~200 metres west of the adit. The Yellow Rose area exhibits both steam-heated and non-steam heated alteration assemblages at similar elevations, suggesting a fluctuating paleo-groundwater table and the potential for buried gold mineralization. A fence of deep drill holes has been proposed at Yellow Rose, but has yet to be drilled.

9.2.2.8 Baileys

Limited mapping and rock sampling has been done at the Baileys target area east of US Highway 95 (Figure 9-2). Steam-heated alteration hosted in the Ammonia Tanks Tuff includes resistive residual silica ribs and mounds surrounded by recessive kaolinite-alunite alteration. As expected, there is no significant surface geochemistry found in the current rock sampling. Additional mapping is necessary before targeting deep drilling at Baileys. Additional claims were staked in 2018 which connect the NBP with the MLP, extending the Baileys target area into the Baileys Gap area to the southeast (see Figure 9-1).

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Figure 9-4 Image of Gridded Gold Values in Soils Over the Eastern Steam Heated Zone Target Areas

 

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10 Drilling

After completing the drill program at NBP in April 2017, an updated Mineral Resource estimate and Technical Report was developed (Wilson, et al, 2017). The Mother Lode Property was acquired in May 2017, and Corvus reprioritized exploration for the higher grade mineralization indicated to exist at MLP. Corvus initiated the MLP drilling program in 2017 with fall Phase 1 drilling and Phase 2 drilling in the spring of 2018.

10.1 MLP Drilling

The current Mother Lode claim group at MLP was once a part of a much larger claim block. The purchase of the Mother Lode group included the data for 520 drill holes totaling 66,932 metres (219,592 feet). The data included exploration holes and water wells from within and surrounding the current claim block. The first known drilling at the MLP was conducted by Galli Exploration (GEXA). In 1985, GEXA drilled 54 shallow holes totaling 4,371 metres (14,340 ft) along the Flatiron target southwest of Mother Lode. Between 1987 and 1991, GEXA and the USNGSJV drilled 318 holes totaling 40,842 metres (133,964 ft) in conjunction with discovery, delineation and development of the Mother Lode and SNA deposits. The GEXA-USNGSJV drilling included four PQ core holes at Mother Lode and four PQ core holes at SNA. The GEXA-USNGSJV drilling also included six water wells drilled in 1988. Between 1987 and 1997, Cordex/Rayrock drilled 132 holes totaling 19,502 metres (63,984 ft) on the ground surrounding the original Mother Lode property. A significant portion of this drilling was at Mother Lode proper in 1997 after the property was acquired by Rayrock in 1995. The 1997 Rayrock drilling included two PQ core holes at Mother Lode. Rayrock also drilled two shallow water wells (monitor wells?) near the Daisy Leach Pad. The existing drilling data from the earlier exploration and production operations has been compared to the results of the Corvus 2017-2018 drilling as part of the data verification program, which is discussed in Section 12. No exploration drilling had occurred in the Mother Lode area for a period of nearly 20 years between 1998 and the fall of 2017.

Starting in September 2017 through July 2018, Corvus has drilled 78 holes totaling 25,823 metres (84,721 ft), including three core holes and 75 reverse circulation holes. All three core holes (ML17-001, 002 and 003) were drilled HQ diameter from the surface. ML17-003 was reduced to NQ lower in the hole when the HQ rods were stuck. The three core holes were drilled early in the program and have been essential to recognizing and understanding the stratigraphy at Mother Lode. Most aspects of drilling and sampling procedures at the MLP are done with the same methodology and quality control developed and used at the NBP since 2010.

Reverse circulation drilling uses 5½ inch downhole hammer bits. Most holes were completed with a hammer bit to depths of up to 440 metres (1445 ft). Tricone rotary drill bits are used to complete those holes that had significant groundwater inflow. Perched water was been encountered in a number of holes. Corvus measures static water level daily as a hole is drilled. Historic drilling at Mother Lode indicated localized, perched water zones, and the Corvus exploration drilling has produced similar information. The data have not demonstrated a predictable static water table. A few of the deeper holes past 400 metres have encountered significant fracture-controlled groundwater up to 30 gallons per minute. When significant groundwater is encountered, it is measured by a timed 5 gallon bucket test taken from the discharge at the splitter.

Reverse circulation samples are collected at continuous 5 foot (1.52 metre) intervals starting from the top of each hole. Two duplicate samples for each interval are captured in large sample bags placed in 5 gallon buckets. The custom-made, heavy duty, sample bags have a white barcode tag for the samples going to the assay lab, and a red barcode tag for the duplicate samples being kept for other purposes (field duplicates, metallurgical testing, etc). The sample hose and rotary splitter are cleaned thoroughly with a high pressure water sprayer prior to drilling of each 20 foot rod (6.1 metre). In order to minimize contamination between 5 foot intervals, the splitter is also quickly sprayed out after each interval is drilled, but before the sample bags are pulled, without stopping drill penetration. Individual samples bags are tied-off without pouring off the contained water, and placed in orderly rows at the drill site for natural decanting of the excess water. The sampling associated with reverse circulation drilling is supervised by an on-site Corvus rig geologist.

Within 3-5 days the samples are sufficiently dry to allow transport. The samples are loaded into super sacks (bulk bags) and transported to a staging area at Corvus’ core shack/field office. Pre-selected blanks and reference standards are placed inside the super sack, and it is sealed with a large, numbered plastic zip-tie. The super sacks are stored in a secure area until they are loaded onto the assay lab truck. The assay lab truck comes to the project for sample pick-ups on an as-needed basis. Chain of custody is transferred to the assay lab personnel at pick-up time.

The drill chips are cursorily logged at the drill site, and later logged in greater detail in an office setting using a binocular microscope. Magnetic susceptibility and an HCL acid “fizz-test” are also measured on the chips for each five foot interval. The geologic characteristics that are determined routinely on drill chips include: color, lithology and stratigraphic unit assignment, alteration style and intensity, vein type and percentage, sulfide type and percentage, and oxide type and relative intensity. The following five oxide classes are used to quantify the oxidation state of each sample:

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Class 1: Unoxidized; total sulfide, no oxide present

Class 2: Mostly unoxidized; sulfide with minor oxide present

Class 3: Mixed oxide/sulfide; generally both oxide and sulfide in nearly equal proportions

Class 4: Mostly oxidized; oxidized with minor fresh sulfide present

Class 5: Completely oxidized; total oxide; no sulfide present

Oxide classes 5, 4 and 3 have consistently yielded favorable gold recoveries in cyanide solubility tests (see Section 13). Resource model blocks assigned to classes 5, 4 and 3 comprise the oxide mineralization category. Oxide classes 2 and 1 have consistently yielded un-favorable gold recoveries in cyanide solubility tests. Model blocks assigned to oxide classes 2 and 1 comprise the sulfide mineralization category.

Significant intercepts from the MLP drilling between September 2017 and July 2018 are listed in Table 10-1. Drilled intercepts are not true widths. No known drilling, sampling or recovery factors have been identified that materially impact the accuracy and reliability of the results. Note, in Table 10-1 reported intercepts are not true widths as there is currently insufficient data to calculate true orientation in space. Mineralized intervals are calculated using a 0.3 g/t cutoff unless otherwise indicated below.

Table 10-1 Drill Intercepts for September 2017 to July 2018 Drilling at Mother Lode Core and RC Holes

Core Holes

Drill Hole #	From (m)	To (m)	Interval (m)	Gold (g/t)	Silver (g/t)	Comment
ML17-001 AZ 90 dip -60	94.70	154.53	59.83	1.95	1.32	Upper Zone
inc	96.28	121.01	24.73	3.67	3.67	1 g/t cut
inc	96.28	107.29	11.01	5.78	3.09	ML structural zone
	178.99	204.59	25.6	2.17	2.21	Lower Zone
ML17-002 AZ 090 dip -50	75.94	134.88	58.94	2.77	1.30	Upper Zone
inc	79.07	81.90	2.83	10.25	3.64	ML structural zone
inc	104.47	112.85	8.38	6.48	2.18	ML structural zone
	162.39	171.84	9.45	0.98	2.69	Lower Zone
	248.72	262.83	14.11	1.06	33.55	Lost hole @ 262.83
ML17-003 AZ 090 dip -70	249.02	252.07	3.05	0.32	0.63
	275.55	341.83	66.28	1.24	2.66	Upper & Lower Zones, Lost hole in 1.64 g/t Au & 30.4 g/t Ag

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RC Holes

ML17-004 AZ 90 dip -65	121.92	164.59	42.67	1.97	1.40	Upper Zone
inc	121.92	150.88	28.96	2.57	1.83	1 g/t cut
inc	121.92	129.54	7.62	4.86	3.10	ML structural zone
	205.74	220.98	15.24	1.18	1.53	Lower Zone
inc	208.79	216.41	7.62	1.87	1.89	1 g/t cut
	254.51	262.13	7.62	1.12	2.11	FCF
inc	256.03	260.60	4.57	1.55	3.13	1 g/t cut
	274.32	275.84	1.52	1.21	1.73	Bottom of lost hole

ML17-005 AZ 090 dip -50	150.88	213.36	62.48	2.10	1.07	Upper & Lower Zone Merge
inc	152.40	163.07	10.67	6.27	3.05	ML structural zone
	230.12	240.79	10.67	0.46	0.82	Low-grade halo for
	281.94	289.56	7.62	0.38	2.05	Possible deep target?
ML17-006 AZ 090 dip -45	97.54	143.26	45.72	1.35	1.35	Upper Zone
inc	97.54	121.92	24.38	1.75	1.66	1 g/t cut
	172.21	175.26	3.05	0.49	0.49	Upper Zone
	179.83	199.64	19.81	1.35	2.27	Upper Zone
inc	205.74	239.27	33.53	2.54	2.03	Lower Zone
inc	211.84	231.65	19.81	3.54	2.15	ML structural zone
	243.84	252.98	9.14	1.77	5.27	Lower Zone
inc	243.84	248.41	4.57	2.97	4.96	1 g/t cut
	257.56	269.75	12.19	0.76	10.49	New Deep Zone?
ML17-007 AZ 090 dip -50	137.16	179.83	42.67	1.82	0.93	Upper Zone
	210.31	265.18	54.87	2.70	2.48	Lower Zone

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inc	246.89	256.03	9.14	6.96	4.48	ML structural zone
	281.94	284.99	3.05	0.51	3.98	Possible deep target?
ML17-008 AZ 090 dip -48	91.44	149.35	57.91	0.79	0.95	South low-grade Oxide zone
inc	118.87	132.59	13.72	1.28	1.03	South low-grade Oxide zone
	153.92	182.88	28.96	0.43	0.72	South low-grade Oxide zone
	192.02	202.69	10.67	0.61	4.52	South low-grade Oxide zone
	225.55	233.17	7.62	0.83	0.47	South low-grade Oxide zone
	237.74	246.89	9.15	1.14	3.77	South low-grade Oxide zone
ML17-009 AZ 090 dip -65	114.30	131.06	16.76	0.39	0.28	New CF Zone
	138.68	225.55	86.87	1.44	1.07	Upper Zone
inc	143.26	146.30	3.05	6.17	1.31	ML structural zone
	269.75	284.99	15.24	0.97	6.30	Lower Zone (lost hole)
ML17-010 AZ 090 dip -65	167.64	175.26	7.62	0.37	0.32	CF oxide low-grade zone
	198.12	201.17	3.05	0.34	0.14	CF oxide low-grade zone
	236.22	248.41	12.19	2.88	2.37	Upper Zone
	300.23	303.28	3.05	1.65	5.29	Lost hole in Lower Zone
ML17-011 AZ 090 dip -70	97.54	135.64	38.10	2.67	1.28	Upper Zone
	102.11	115.82	13.71	5.71	2.52	Mother Lode Structural Zone
	181.36	195.07	13.71	1.48	1.89	Lower Zone
	204.22	234.70	30.48	0.83	3.74	Lower Zone expansion
inc	204.22	211.84	7.62	2.05	2.10	1 g/t cut
ML17-012 AZ 090 dip -87	149.35	152.40	3.05	0.37	0.42	CF oxide low-grade zone

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	187.45	196.60	9.15	0.61	0.48	CF oxide low-grade zone
	205.74	234.70	28.96	1.53	1.82	Upper Zone
	240.79	249.94	9.15	0.96	2.60	Possible Lower Zone
ML17-013 AZ 090 dip -70	210.31	213.36	3.05	0.64	0.08	CF oxide low-grade zone
	233.17	243.84	10.67	0.67	1.24	CF oxide low-grade zone
	256.03	350.52	94.49	1.20	1.86	Upper & Lower Zone
inc	275.84	288.04	12.20	2.00	1.97	1 g/t cut
inc	292.61	300.23	7.62	1.89	2.00	1 g/t cut
inc	310.90	341.38	30.48	1.62	2.78	1 g/t cut
ML17-014 AZ 090 dip -50	112.78	172.21	59.43	1.68	1.67	Upper Zone
	182.88	195.07	12.19	0.43	1.56
	236.22	272.80	36.58	1.49	3.75	Lower Zone
inc	257.56	272.80	15.24	2.28	6.95	1 g/t Au cut
ML17-015 AZ 090 dip -75	303.28	327.66	24.38	1.69	2.37	Upper Zone
	336.80	339.85	3.05	0.58	0.84
ML17-016 AZ 090 dip-45	140.21	143.26	3.05	1.04	0.26	South end of deposit
	158.50	167.64	9.14	1.11	0.20
inc	161.54	166.12	4.58	1.62	0.22
	175.26	184.40	9.14	0.98	1.96	New Deep In-pit Zone
inc	179.83	182.88	3.05	2.36	3.31	New Deep In-pit Zone
	216.41	228.60	12.19	0.96	1.61	New Deep In-pit Zone
inc	217.93	219.46	1.52	5.52	4.09	New Deep In-pit Zone
ML17-017 AZ 090 dip-70	268.22	272.80	4.58	0.68	0.59	Northwest Extension
	291.08	326.14	35.06	1.97	2.09	Northwest Extension

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 75

 

inc	292.61	324.61	32.00	2.10	2.17	Northwest Extension
	347.47	352.04	4.57	0.55	0.33	Northwest Extension
ML17-018 AZ 090 dip-65	105.16	187.45	82.29	1.11	n/a	Drilled west of ML17-006
inc	105.16	132.59	27.43	1.57	1.32
inc	156.97	172.21	15.24	1.72	2.75
	202.69	208.79	6.10	0.60	n/a
	220.98	227.08	6.10	0.62	0.51
	236.22	245.36	9.14	1.09	1.65	New Deep Main Zone
inc	240.79	245.36	4.57	1.41	2.61	New Deep Main Zone
ML17-019 AZ 090 dip-87	92.96	109.73	16.77	0.47	0.57	Drilled west of ML17-004
	146.30	213.36	67.06	1.32	1.92
inc	147.83	166.12	18.29	1.49	3.08
inc	170.69	176.78	6.09	1.13	0.73
inc	182.88	211.84	28.96	1.58	1.99
	291.08	294.13	3.05	0.96	0.47
	336.80	345.95	9.15	0.86	1.36	New Deep Main Zone
inc	336.80	339.85	3.05	1.95	3.19	New Deep Main Zone
ML17-020 AZ 090 dip-65	120.40	216.41	96.01	1.35	1.24	Using 0.1 g/t cutoff
	121.92	143.26	21.34	0.49	0.60	Drilled west of ML17-007
	147.83	167.64	19.81	1.82	1.23
	172.21	210.31	38.10	2.12	1.98
inc	173.74	196.60	22.86	3.08	2.21
	259.08	265.18	6.10	0.49	2.22	New Deep Main Zone
	271.27	298.70	27.43	1.31	3.90	New Deep Main Zone
inc	275.84	283.46	7.62	3.46	9.98	New Deep Main Zone

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 76

 

ML17-021 AZ 090 dip-45	100.58	126.49	25.91	0.44	0.49	Drilled west of ML17-005
	147.83	178.31	30.48	2.54	1.92
	227.08	262.13	35.05	1.95	2.48
inc	227.08	246.89	19.81	2.28	2.05
inc	251.46	262.13	10.67	2.07	4.20
ML17-022 AZ 090 dip-85	102.11	124.97	22.86	0.41	0.37	Drilled west of ML17-009
AZ 090 dip-85	129.54	153.92	24.38	0.74	0.80
inc	131.06	135.64	4.58	1.34	1.71
	170.69	216.41	45.72	1.33	1.25
inc	172.21	207.26	35.05	1.57	1.29
	329.18	338.33	9.15	0.40	0.30	New Deep Main Zone
ML17-023 AZ 360 dip-90	64.01	67.06	3.05	0.81	1.96
AZ 360 dip-90	79.25	99.06	19.81	0.73	1.23
inc	89.92	94.49	4.57	1.21	1.93
	103.63	111.25	7.62	2.69	2.60	New Deep In-pit Zone
ML17-024 AZ 360 dip-90	0.00	15.24	15.24	2.76	1.42
	19.81	41.15	21.34	0.58	0.30
	79.25	112.78	33.53	0.94	1.37
inc	82.30	100.58	18.28	1.40	1.79
	117.35	129.54	12.19	0.56	0.38	New Deep In-pit Zone
	140.21	149.35	9.14	0.55	2.72	New Deep In-pit Zone
ML17-025 AZ 360 dip-90	0.00	50.29	50.29	1.74	1.17
inc	0.00	19.81	19.81	2.26	1.60
inc	24.38	42.67	18.29	1.99	1.13

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 77

 

	86.87	108.20	21.33	1.59	2.11
inc	86.87	103.63	16.76	1.90	2.53
	138.68	141.73	3.05	2.90	0.85	New Deep In-pit Zone
	150.88	164.59	13.71	1.41	4.39	New Deep In-pit Zone
inc	152.40	158.50	6.10	2.09	6.61	New Deep In-pit Zone
ML17-026 AZ 090 dip-65	126.49	172.21	45.72	1.33	1.73	West of hole ML17-014
inc	126.49	129.54	3.05	1.47	2.01	>1g/t cut
inc	144.78	169.16	24.38	1.72	2.03	>1g/t cut
	178.31	185.93	7.62	0.57	2.44
	242.32	266.70	24.38	0.93	2.28	New deep in-pit zone
inc	249.94	256.03	6.09	1.75	2.86	>1g/t cut
ML17-027 AZ 090 dip-65	124.97	156.97	32.00	0.49	1.02	West of hole ML17-008
inc	150.88	153.92	3.05	1.03	1.62	>1g/t cut
	172.21	176.78	4.57	0.78	0.94
	181.36	213.36	32.00	0.63	1.28
inc	188.98	192.02	3.05	1.33	3.52	>1g/t cut
	233.17	251.46	18.29	0.44	0.88	New deep in-pit zone
ML17-028 AZ 090 dip-65	73.15	124.97	51.82	1.86	0.88	East of hole ML17-016
inc	83.82	103.63	19.81	3.43	1.28	>1g/t cut
inc	115.82	121.92	6.10	2.37	2.60	>1g/t cut
	167.64	172.21	4.57	0.74	2.10	New deep in-pit zone
ML17-029 AZ 090 dip-85	103.63	112.78	9.15	0.43	0.19	West of hole ML17-010
	117.35	123.44	6.09	0.73	1.59
	161.54	179.83	18.29	1.84	1.46
	193.55	196.60	3.05	0.46	1.56

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 78

 

	204.22	240.79	36.57	1.85	3.15
inc	211.84	237.74	25.90	2.37	3.75	>1g/t cut
	277.37	286.51	9.14	1.17	2.41
inc	278.89	284.99	6.10	1.56	2.39	>1g/t cut
	320.04	338.33	18.29	0.90	3.47
inc	321.56	326.14	4.57	2.55	6.59	>1g/t cut

ML17-030 AZ 090 dip-65	59.44	62.48	3.04	0.93	0.58	West of hole ML17-023
	100.58	105.16	4.57	0.67	0.12
	112.78	117.35	4.57	0.89	0.06
	144.78	155.45	10.67	0.97	0.99
	161.54	164.59	3.05	0.35	1.55
	169.16	172.21	3.05	0.37	0.43
	176.78	179.83	3.05	0.35	0.53
	185.93	198.12	12.19	0.80	0.55	New deep in-pit zone
inc	190.50	193.55	3.05	1.58	1.01	>1g/t cut
ML17-031 AZ 090 dip-70	313.94	347.47	33.53	1.60	3.13	NW of hole ML17-003
inc	313.94	332.23	18.29	2.47	4.12	>1g/t cut
ML17-032 AZ 360 dip-90	137.16	172.21	35.05	0.93	1.22	West of hole ML17-016
inc	161.54	172.21	10.67	1.78	1.24	>1g/t cut
	181.36	195.07	13.71	1.24	1.36
inc	187.45	193.55	6.10	2.02	2.58	>1g/t cut
	199.64	213.36	13.72	0.86	2.45	New deep in-pit zone
	231.65	243.84	12.19	0.81	0.63	New deep in-pit zone
ML17-033 AZ 090 dip-70	294.13	306.32	12.19	2.31	2.05	North of hole ML17-003

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 79

 

inc	294.13	301.75	7.62	3.39	2.58	>1g/t cut
ML17-034 AZ 360 dip-90	100.58	123.44	22.86	1.18	3.27	Drilled SW of Pit New Upper Ox Zone
	134.11	150.88	16.77	0.50	7.28
	254.51	257.56	3.05	0.39	0.44
ML17-035 AZ 180 dip-85	80.77	85.34	4.57	1.55	n/a	Drilled SW of Pit

ML17-036 AZ 090 dip-85	143.26	152.40	9.14	0.76	n/a	West of ML17-027 New Upper Ox Zone
inc	147.83	150.88	3.05	1.45	2.31	>1 g/t
	158.50	164.59	6.09	0.97	1.69	New Upper Ox Zone
	230.12	236.22	6.10	0.60	2.50
	257.56	265.18	7.62	0.71	0.48
	280.42	324.61	44.19	1.33	1.40	New Deep In-Pit Zone
inc	280.42	284.99	4.57	2.16	1.10	>1 g/t
inc	295.66	313.94	18.28	2.10	2.50	>1 g/t
ML17-037 AZ 090 dip-70	188.98	211.84	22.86	0.77	0.52	East of ML17-13
inc	196.60	199.64	3.05	1.03	0.51	>1g/t cut
inc	205.74	208.79	3.05	1.21	0.56	>1g/t cut
	246.89	259.08	12.19	2.69	2.64	New East Zone
inc	248.41	257.56	9.15	3.31	3.03	>1g/t cut
ML17-038 AZ 090 dip-85	132.59	185.93	53.34	1.57	1.96	West of ML17-018
inc	132.59	144.78	12.19	2.29	2.29	>1g/t cut
inc	149.35	179.83	30.48	1.61	1.85	>1g/t cut
AZ 090 dip-85	281.94	304.80	22.86	0.49	0.37
	309.37	321.56	12.19	0.87	3.29	New Deep In-Pit Zone
inc	315.47	320.04	4.57	1.22	5.66	>1g/t cut

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 80

 

ML17-039 AZ 090 dip-85	100.58	105.16	4.57	0.46	0.15	West of ML17-021 New Upper Ox Zone
	112.78	124.97	12.19	0.56	0.40	New Upper Ox Zone
	182.88	252.98	70.10	1.45	1.98	West Ext - Main Zone
inc	192.02	246.89	54.87	1.72	2.18	>1g/t cut

ML17-040 AZ 090 dip-85	106.68	109.73	3.05	0.36	n/a	West of ML17-020 New Upper Ox Zone
	124.97	128.02	3.05	0.36	n/a	New Upper Ox Zone
	140.21	152.40	12.19	0.49	n/a	New Upper Ox Zone
	185.93	216.41	30.48	1.80	n/a	West Ext – Main Zone
inc	185.93	213.36	27.43	1.91	n/a	>1g/t cut
ML17-041 AZ 090 dip-65	89.92	115.82	25.90	0.77	0.38	East of ML17-001
inc	96.01	99.06	3.05	2.65	1.06	>1g/t cut
	150.88	166.12	15.24	0.89	1.99	New East Zone
inc	150.88	160.02	9.14	1.28	2.63	>1g/t cut
ML17-042 AZ 090 dip-70	24.38	32.00	7.62	0.31	0.33	East of ML17-011
	150.88	179.83	28.95	2.02	1.86	New East Zone
inc	150.88	172.21	21.33	2.62	2.43	>1g/t cut
	239.27	245.36	6.09	0.43	5.46	Lower Ox Zone
ML18-043 AZ 085 dip-70	124.97	128.02	3.05	0.63	n/a	West of ML17-027 New Upper Ox Zone
	137.16	143.26	6.10	0.55	n/a	New Upper Ox Zone
	220.98	248.41	27.43	1.81	n/a	West ext. – Main Zone
inc	224.03	246.89	22.86	2.05	n/a	>1 g/t
	333.76	341.38	7.62	0.44	n/a
ML18-044 AZ 085 dip-85	207.26	222.50	15.24	0.37	n/a	West of ML17-022 New Upper Ox Zone

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 81

 

	230.12	251.46	21.34	0.92	n/a	New Upper Ox Zone
inc	234.70	239.27	4.57	1.22	n/a	>1 g/t
inc	243.84	249.94	6.10	1.14	n/a	>1 g/t
	256.03	281.94	25.91	2.32	n/a	West ext. – Main Zone
inc	259.08	280.42	21.34	2.69	n/a	>1 g/t
ML18-045 AZ 085 dip-80	169.16	176.78	7.62	0.52	n/a	West of ML17-039 New Upper Ox Zone
	184.40	187.45	3.05	0.46	n/a	New Upper Ox Zone
	256.03	295.66	39.63	2.45	n/a	West ext. – Main Zone
inc	256.03	289.56	33.53	2.84	n/a	>1 g/t
ML18-046 AZ 083 dip-55	21.34	51.82	30.48	2.05	n/a	West of ML17-027 Upper East Zone
inc	35.05	47.24	12.19	4.45	n/a	>1 g/t
	85.34	91.44	6.10	0.77	n/a
	216.41	225.55	9.14	0.95	n/a	Lower East Zone
inc	217.93	220.98	3.05	2.05	n/a	>1 g/t
ML18-047 AZ 085 dip-60	193.55	205.37	11.82	1.77	n/a	West of ML17-022 Lower East Zone
inc	193.55	201.17	7.62	2.57	n/a	>1 g/t
	216.41	220.98	4.57	0.32	n/a
	256.03	266.70	10.67	0.42	n/a
ML18-048 AZ 085 dip-65	51.82	54.86	3.05	0.41	n/a	East of ML17-037 Upper East Zone
	210.31	214.88	4.57	3.61	n/a	East H-G Zone
	320.04	341.38	21.34	0.77	n/a	New Far-East Zone
inc	332.23	339.85	7.62	1.20	n/a	>1 g/t
ML18-049 AZ 085 dip-70	230.12	237.74	7.62	21.77	n/a	East of ML17-010 East H-G Zone
	248.41	252.98	4.57	0.27	n/a

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 82

 

	277.37	280.42	3.05	1.20	n/a
ML18-050 Az 085 dip-65	345.95	367.28	21.33	2.00	n/a	West of ML17-013 Main ML Zone
ML18-051 AZ 085 dip-70	356.62	385.57	28.95	1.02	n/a	West of ML 17-017 East H-G Zone
inc	356.62	371.86	15.24	1.56	n/a	1 g/t cut
	400.81	416.05	15.24	0.45	n/a	Lower OX Zone to the EOH
ML18-052 AZ 085 dip-75	132.59	155.45	22.86	0.67	n/a	West of ML 17-021 Upper LG Zone
inc	149.35	153.92	4.57	1.30	n/a	1 g/t cut
	230.12	283.46	53.34	1.90	n/a	Main ML Zone
inc	230.12	277.37	47.25	2.09	n/a	1 g/t cut
	385.57	426.72	41.15	1.52	n/a	Lower OX Zone to the EOH
inc	385.57	399.29	13.72	2.09	n/a	1g/t cut
ML18-053 AZ 085 dip-85	79.25	94.49	15.24	0.52	n/a	West side of ML Pit Upper L-G OX Zone
	114.30	126.49	12.19	0.38	n/a	Upper L-G OX Zone
	134.11	141.73	7.62	0.40	n/a	Upper L-G OX Zone
	146.30	149.35	3.05	0.67	n/a	Upper L-G OX Zone
	155.45	163.07	7.62	0.44	n/a	Upper L-G OX Zone
	173.74	176.78	3.05	1.17	n/a	Upper L-G OX Zone
	214.88	248.41	33.53	2.20	n/a	Upper & Lower Zone
inc	216.41	220.98	4.57	1.14	n/a	1 g/t cut
inc	227.08	246.89	19.81	3.29	n/a	1 g/t cut
ML18-054 AZ 085 dip-75	131.06	161.54	30.48	1.13	n/a	West of ML18-053 Upper L-G OX Zone
inc	135.64	141.73	6.09	1.50	n/a	1 g/t cut
inc	149.35	155.45	6.09	2.73	n/a	1 g/t cut

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 83

 

	166.12	172.21	6.09	0.72	n/a	Upper L-G OX Zone
	233.17	243.84	10.67	0.40	n/a	Upper L-G OX Zone
	262.13	288.04	25.91	1.50	n/a	Upper & Lower Zone
inc	262.13	271.27	9.14	2.65	n/a	1 g/t cut
inc	283.46	288.04	4.58	1.49	n/a	1 g/t cut
ML18-055 AZ 085 dip-75	188.98	214.88	25.90	0.94	n/a	West of ML18-054 Upper LG OX Zone
inc	195.07	201.17	6.10	1.30	n/a	1 g/t cut
inc	207.26	210.31	3.05	1.87	n/a	1 g/t cut
	303.28	306.32	3.05	0.41	n/a	Upper L-G OX Zone
	310.90	315.47	4.57	0.75	n/a	Upper L-G OX Zone
	326.14	338.33	12.19	1.27	n/a	Upper Main Zone (OX)
inc	327.66	335.28	7.62	1.72	n/a	1 g/t cut
	342.90	353.57	10.67	0.34	n/a	L-G OX Zone
	361.19	367.28	6.09	1.18	n/a	Lower Main Zone

ML18-056 AZ 145 dip-75	184.40	204.22	19.82	0.43	n/a	Drilled at structure below hole ML18-057
inc	184.40	185.93	1.52	3.73	n/a	1 g/t cut
	213.36	230.12	16.76	0.24	n/a
ML18-057 AZ 145 dip-50	213.36	230.12	16.76	0.24	n/a	Shallow angle, drilled at base of Jasperoid
inc	143.26	144.78	1.52	1.83	n/a	1 g/t cut
ML18-058 AZ 140 Dip-70	297.18	304.80	7.62	0.74	0.27	Northern Scout Hole OX on structure
inc	297.18	300.23	3.05	1.51	0.43	1 g/t cut
ML18-059 AZ 085 dip-70	181.36	185.93	4.57	0.50	n/a	West side of ML Pit Upper L-G OX Zone
	246.89	277.37	30.48	0.70	n/a	Upper Zone
inc	256.03	259.08	3.05	1.16	n/a	1 g/t cut

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 84

 

	281.94	304.80	22.86	1.48	n/a	Lower OX Zone
inc	281.94	300.23	18.29	1.73	n/a	1 g/t cut
	324.61	330.71	6.10	0.81	n/a	Lower OX to end of hole
ML18-060 AZ 085 dip-80	294.13	312.42	18.29	0.95	n/a	West of ML18-053 Upper Zone
	318.52	321.56	3.05	0.48	n/a
	345.95	370.33	24.38	3.10	n/a	Lower Zone
inc	347.47	368.81	21.34	3.47	n/a	1 g/t cut
ML18-061 AZ 085 dip-70	283.46	286.51	3.05	0.32	n/a	West of ML18-054 Upper LG OX Zone
	291.08	303.28	12.20	0.54	n/a	Upper OX
	309.37	321.56	12.19	1.13	n/a	Lower OX to the end of the hole
inc	309.37	315.47	6.10	1.83	n/a	1 g/t cut

ML18-062 AZ 070 dip-65	62.48	67.06	4.57	0.14	n/a	North of ML18-049 Upper LG OX Zone
	239.27	242.32	3.05	0.30	n/a
	291.08	323.09	32.01	0.16	n/a
	327.66	339.85	12.19	0.37	n/a	Lower OX Zone
ML18-063 AZ 085 dip-70	257.56	262.13	4.57	0.40	n/a	North of ML18-062 Upper LG OX Zone
	271.27	289.56	18.29	0.29	n/a
	300.23	303.28	3.05	0.30	n/a
	338.33	347.47	9.14	0.26	n/a	Lower OX Zone

ML18-064 AZ 085 dip-65	294.13	315.47	21.34	0.97	n/a	East of ML17-033 Main Zone
inc	294.13	304.80	10.67	1.42	n/a	1 g/t cut
	320.04	323.09	3.05	0.43	n/a	Lower OX Zone
	367.28	370.33	3.05	0.74	n/a	Lower OX Zone
	373.38	377.95	4.57	0.30	n/a	Lower OX Zone
ML18-065 AZ 085 dip-50	33.53	36.58	3.05	0.59	n/a	East of ML17-042 Upper OX Zone

Corvus Gold Inc.
Preliminary Economic Assessment – NB-MLP	Page 85

 

ML18-066 AZ 085 dip-50	0.00	15.24	15.24	1.56	n/a	South of ML18-046 Upper OX Zone
inc	0.00	9.14	9.14	2.24	n/a	1 g/t cut
	68.58	79.25	10.67	0.92	n/a	Main Zone
inc	76.20	79.25	3.05	1.30	n/a	1 g/t cut
	105.16	108.20	3.05	0.73	n/a	Lower OX
	149.35	155.45	6.10	0.61	n/a	Lower OX
ML18-067 AZ 085 dip-70	201.17	227.08	25.91	0.68	n/a	West of ML18-045 Upper OX Zone
inc	222.50	225.55	3.05	1.18	n/a	1 g/t cut
	233.17	263.65	30.48	0.67	n/a	New Zone
inc	257.56	262.13	4.57	1.15	n/a	1 g/t cut
	297.18	316.99	19.81	3.76	n/a	Main Zone
inc	298.70	316.99	18.29	4.02	n/a	1 g/t cut
ML18-068 AZ 085 dip-70	233.17	342.90	109.73	1.40	n/a	East of ML17-033 Main Zone
inc	233.17	288.04	54.87	1.86	n/a	1 g/t cut
inc	298.70	320.04	21.34	1.26	n/a	1 g/t cut
inc	329.18	338.33	9.15	1.33	n/a	1 g/t cut
ML18-069  AZ 085 dip-65	292.61	306.32	13.71	0.94	n/a	North of ML17-061 Upper OX Zone
inc	294.13	303.28	9.15	1.22	n/a
inc	336.80	339.85	3.05	1.94	n/a
ML18-070 AZ 085 dip-60	217.93	220.98	3.05	0.41	0.46
	339.895	342.90	3.05	0.95	0.33
	349.00	350.52	1.52	0.61	0.38

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353.57 359.66 6.09 2.54 1.20

ML18-071  AZ 085 dip-50	240.79	243.84	3.05	0.52	0.69
ML18-072 AZ 080 dip-70	362.71	368.81	6.10	0.36	0.63
	402.34	416.05	13.781	1.80	4.33
ML18-073  AZ 080 dip-60	321.56	370.33	48.77	1.19	1.22
ML18-074 AZ 080 dip-87	298.70	320.04	21.34	1.44	0.87
inc	300.23	315.47	15.24	1.74	0.95	1 g/t cut
	352.04	391.67	39.63	1.67	6.28
inc	364.24	390.14	25.90	2.34	9.37	1 g/t cut
ML18-075  AZ 080 dip-85	463.30	469.39	6.09	1.00	0.57
ML18-076  AZ 075 dip-80	423.67	432.82	9.15	1.39	1.32
ML18-077  AZ 075 dip-85	342.90	402.34	59.44	2.36	2.36
ML18-078  AZ 075 dip-85	371.8	390.14	18.28	1.49	1.82
inc	373.38	385.57	12.19	2.97	2.34	1 g/t cut

10.2 NBP Drilling

Significant intercepts from the NBP drilling between March and April 2017 are listed in Table 10-1. These intercepts comprise the last of the holes drilled at the NBP in 2017. Drilled intercepts are not true widths. No known drilling, sampling or recovery factors have been identified that materially impact the accuracy and reliability of the results. Table 10-2 reported drill intercepts are not true widths and mineralized thickness is calculated at 0.10 g/t cutoff with internal intervals calculated at 0.5 g/t cutoff.

Table 10-2 Drill Intercepts for March-April 2017 NBP Drilling

Drill Hole #	from (m)	to (m)	Interval (m)	Gold (g/t)	Silver (g/t)	Comment
NB-17-335 AZ 090 dip-55	64.01	92.96	28.95	0.49	1.97
inc	67.06	79.25	12.19	0.78	2.34

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	99.06	160.02	60.96	0.31	6.28
inc	132.59	135.64	3.05	1.04	7.00
inc	140.21	141.73	1.52	1.01	18.00
inc	147.83	149.35	1.52	0.76	59.00
NB-17-336 AZ 085 dip-45	42.67	48.77	6.10	0.26	0.37
	57.91	65.53	7.62	1.04	0.69
inc	59.44	62.48	3.05	2.18	0.94
	109.73	114.30	4.57	0.38	0.74
	137.16	140.21	3.05	0.31	0.80
	146.30	179.83	33.53	0.34	0.41
inc	150.88	152.40	1.52	0.95	0.41
inc	160.02	164.59	4.57	0.61	0.66
NB-17-337 AZ 090 dip-50	30.48	41.15	10.67	0.36	0.46
inc	38.10	39.62	1.52	1.21	1.34
NB-17-338 AZ 090 dip-50	45.72	70.10	24.38	0.71	1.33
inc	47.24	62.48	15.24	0.98	1.65
	76.20	91.44	15.24	0.26	0.53
NB-17-339 AZ 090 dip-55	132.59	134.11	1.52	0.34	0.49
NB-17-340 AZ 085 dip-60	65.53	68.58	3.05	0.58	0.45
inc	67.06	68.58	1.52	1.03	0.60
	73.15	150.88	77.73	0.43	2.67
inc	76.20	79.25	3.05	0.93	2.38
inc	105.16	117.35	12.19	0.88	11.22
inc	129.54	137.16	7.62	0.94	0.65
NB-17-430 AZ 085 dip-60	129.54	135.64	6.10	0.14	0.58
	164.59	182.88	18.29	0.17	0.35
NB-17-431 AZ 085 dip-50	44.20	265.18	220.98	0.21	0.94

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	272.80	304.80	32.00	0.23	0.39	Hole ended in mineralization
NB-17-432 AZ 085 dip-55	53.34	65.53	12.19	0.75	2.20
	80.77	82.30	1.52	1.32	2.15
	146.30	156.97	10.67	0.36	0.46
inc	152.40	155.45	3.05	0.73	0.67
	166.12	198.12	32.00	0.56	1.41
inc	167.64	179.83	12.19	0.76	1.50
inc	184.40	188.98	4.58	0.77	1.69
	233.17	286.51	53.34	0.27	0.71	Hole ended in mineralization
NB-17-433 AZ 085 dip-55	131.06	234.70	103.64	0.33	1.10
inc	185.93	199.64	13.71	0.56	1.26
inc	213.36	217.93	4.57	0.58	1.23
inc	222.50	225.55	3.05	0.53	1.36
	251.46	286.51	35.05	0.35	0.63
inc	275.84	280.42	4.58	0.66	0.64	Hole ended in mineralization
NB-17-434 AZ 085 dip-45	62.48	65.53	3.05	0.39	1.00
	70.10	89.92	19.82	0.38	0.97
inc	73.15	79.25	6.10	0.59	1.18
	193.55	220.98	27.43	0.14	0.42
NB-17-435 AZ 085 dip-45	254.51	262.13	7.62	0.11	0.22
NB-17-436 AZ 090 dip-47	77.72	83.82	6.10	0.16	0.05
NB-17-437 AZ 085 dip-50	36.58	42.67	6.09	0.13	0.04
	103.63	105.16	1.52	0.27	0.06
NB-17-438 AZ 085 dip-50	27.43	32.00	4.57	0.14	0.05
	68.58	100.58	32.00	0.28	0.25

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	117.35	141.73	24.38	0.22	0.13
	158.50	166.12	7.62	0.41	0.41
	233.17	243.84	10.67	0.24	1.33
NB-17-439 AZ 085 dip-55	12.19	22.86	10.67	0.12	0.19
	38.10	51.82	13.72	0.13	0.19
	150.88	155.45	4.57	0.23	0.16
NB-17-440 AZ 105 dip-45	266.70	269.75	3.05	0.12	0.25
NB-17-441 AZ 095 dip-45	No significant results
NB-17-442 AZ 100 dip-45	135.64	137.16	1.52	0.13	1.24
NB-17-443 AZ 100 dip-70	118.87	128.02	9.15	0.12	0.20
	161.54	172.21	10.67	0.24	0.19
	196.60	320.04	123.44	0.14	1.27
NB-17-444 AZ 090 dip-50	108.20	129.54	21.34	0.26	0.88
	135.64	185.93	50.29	0.35	1.08
inc	149.35	164.59	15.24	0.69	1.18
	211.84	245.36	33.52	0.21	0.77
	251.46	262.13	10.67	0.12	0.37
NB-17-445 AZ 090 dip-50	167.64	173.74	6.10	0.24	0.26
	181.36	210.31	28.95	0.40	0.53
inc	192.02	196.60	4.58	1.29	0.68
	216.41	281.94	65.53	0.24	1.24
inc	219.46	222.50	3.05	0.58	1.21
	329.18	390.14	60.96	0.22	2.12

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11 Sample Preparation, Analysis and Security

11.1 MLP QA/QC Program

This section summarizes Quality Assurance and Quality Control (“QA/QC”) data related to drill hole sample assaying carried out between September 2017 and June 2018 by Corvus.

11.1.1 RC Drilling

Corvus drilling between September 2017 and June 2018 included 3 core holes and 67 reverse circulation (RC) holes totaling 22,551 metres. Standard procedures put in place during the 2010-2011 North Bullfrog RC sampling program were followed. All samples were sent to American Assay Laboratories (“AAL”) in Sparks, Nevada, which is an independent lab (discussed later in 11.1.3).

11.1.2 Core Drilling

Three (3) core holes were drilled at Mother Lode in 2017, totaling 911.53 metres. HQ3 core was drilled and extracted using triple-tube tooling to ensure the best recovery through highly fractured intervals. Triple tube tooling minimizes core separation and rotation within the extraction tube. Selected lengths of each hole were sampled with continuous intervals based on careful logging of geological characteristics. In conjunction with the logging, sample intervals were marked in the core box and assigned unique sample numbers in a sequence that included pre-selected QA/QC samples every tenth sample. Each hole starts with a blank QA/QC sample, and alternates between blanks and reference standards. Once a hole is logged and tagged for sampling, each box is photographed within a fabricated lighting and reference frame. The reference frame allows rectification of the image so that in future applications true lengths can be measured on the core using the photos. Once a hole, or a group of boxes in a hole, are photographed, the photos are reviewed for adequacy and the photo files renamed using hole number and box number. All sample intervals of core are cut in half with a core saw, with one half of the core being placed in a sample bag to be assayed, and the other half returned to the original core box for archive.

11.1.3 Accredited Laboratories

Assaying for the MLP holes has been performed by AAL in Sparks, Nevada. Corvus has no business relationship with AAL beyond being a customer for analytical services. The Sparks laboratory is Standards Council of Canada, Ottawa, Ontario Accredited Laboratory No. 536 and conforms with requirements of CAN-P-1579, CAN-P-4E (ISO/IEC 17025:2005).

Check assaying has been performed by Bureau Veritas (Metals & Mining) North America (“BVM”, formerly Inspectorate America Corporation), in Sparks, Nevada and Vancouver, Canada, and ALS Minerals Laboratories (“ALS Minerals”), in Sparks, Nevada. Corvus has no business relationship with BVM or ALS Minerals beyond being a customer for analytical services. The BVM laboratory is Accredited Laboratory No. 720 and conforms to requirements of CAN-P-1579, CAN-P-4E (ISO 9001:2008) and ALS is Accredited Laboratory No. 660 and conforms to requirements of CAN-P-1579, CAN-P-4E (ISO/IEC 17025:2005).

11.1.4 Transport and Security

Prior to shipment, all core samples were photographed, cut, weighed and then placed in bulk bags that are sealed with a security tag. RC samples are not weighed until they reach the lab, but are also shipped in bulk bags which are sealed with a security tag prior to shipment. Samples from each drill hole were sent to AAL in Sparks, Nevada as a separate shipment with a chain of custody document to certify that the seals were intact when the shipment was received.

11.1.5 Duplicates

Duplicate samples were used to monitor the precision of the assays that would be incorporated into the Mineral Resource estimate. Duplicates monitor three sources of variation: sampling method, preparation and assaying. Preparation duplicates are used to monitor the sample preparation process, field duplicates are used to document the precision associated with sampling at the drill site, and pulp duplicates are used to monitor the assaying process. Corvus uses all three types of duplicates to monitor the precision of the gold and silver analyses. However, field duplicates are only collected for RC holes.

11.1.5.1 Preparation Duplicates

Sample preparation duplicates (Prep Duplicates) were created by crushing the sample at the lab and then splitting it in half. The two halves were then processed as separate samples. Five prep duplicates were created for each drill hole. The selection of preparation duplicates was made by the geologist logging the hole, based on their interpretation of lithologies and degree of mineralization. Comparison of the Prep Duplicates is shown in Figures 11-1 and 11-2, for gold and silver, respectively. The graphs plot the original assay versus the duplicate assay and indicate the results falls between the ± 10% precision lines.

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Figure 11-1 Gold Assays of Preparation Duplicates for MLP

 

Figure 11-2 Silver Assays of Preparation Duplicates for MLP

 

Figure 11-1 and Figure 11-2 show that, in general, the preparation duplicates reproduce very well for both gold (coefficient of variation 22%) and silver (coefficient of variation 28%). Not all sample intervals were assayed for silver, and Figure 11-2 only displays available silver data for holes 1 to 42 and 56-58.

11.1.5.2 Field Duplicates

Field duplicates were selected by the Project Manager after the gold assay results for an RC hole have been received. At the drill site, two samples were taken for every 1.52 m (5 foot) interval, with a primary sample being labeled with a white tag. The secondary sample was labeled with a red tag with an “M” suffix added to the original sample number. The secondary (red tag bags) were used for the field duplicates. The selected field duplicates would have a gold value of 0.1 ppm or higher, based on the results of the primary sample. Field duplicates have undergone the same transport and security procedure as all other RC and rock samples. The field duplicates were used to check the accuracy and precision of the sample splitting at the drill site. Figure 11-3 and 11-4 show the comparsion of the primary sample (original) and the secondary (or field duplicate) assay results, and indicated that the splitting of duplicate samples at the drill site was both accurate and precise for gold (coefficient of variation 18%) and silver (coefficient of variation 23%).

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Figure 11-3 Comparison of Field Duplicate Gold Assays for MLP

 

Figure 11-4 Comparison of Field Duplicate Silver Assays for MLP

 

s

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11.1.5.3 Pulp Duplicates

Pulp Duplicates were used to indicate the homogeneity of the pulp material that is subjected to the fire assay and variations generally reflect the nugget effect in gold samples. In this instance, AAL routinely ran pulp duplicates as part of their internal QA/QC program and these assays are analyzed as part of assay QA/QC. The resulting variation in assay results for the Pulp Duplicates are compared in Figures 11-5 and 11-6, where the assay value reported (Reported Assay) in the Mother Lode drill hole database is plotted versus the AAL Pulp Duplicate assay, for Gold (Au) and Silver (Ag), respectively. Figure 11-5 and Figure 11-6 show that the AAL internally selected pulp duplicates compared well for both gold (coefficient of variation 15%) and silver (coefficient of variation 27%).

Figure 11-5 AAL Gold Assays of Pulp Duplicates for MLP

 

Figure 11-6 AAL Silver Assays of Pulp Duplicates for MLP

 

11.1.6 Check Assays

Four hundred and two (402) check samples were analyzed for gold and multi-elements in 3 separate check assay jobs at BVM and ALS to check the results of AAL. The samples submitted included both drill produced materials from MLP drilling and standard samples supplied by Rocklabs Ltd. of Auckland, New Zealand and Geostats Pty Ltd. of O’Connor, Western Australia. The standard samples include both Certified Reference Materials (CRM) and Blank Material (Blank). The first job, ML180215CA, had 135 samples, the second, ML180410CA, had 128 samples, and the third, ML180524CA, had 139 samples. Of the standards submitted for the three jobs, 33 were standard material and 6 were blank material. ML180215CA contained 12 with standard material, ML180410CA contained 7 with standard material and 6 with blank material, and ML180524CA contained 14 with standard material.

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Overall, the comparison of the test results indicates good agreement between the different labs. The lab comparisons are shown graphically in Figures 11-7, 11-8, and 11-9 for gold assays, and 11-10, 11-11, and 11-12 for silver assays. The graphs indicate that there is a slight low bias in the gold assays at AAL when compared to BVM as shown in Figure 11-7 for the first two check assay jobs (green and blue dots). The third check assay job checked by BVM (red dots, Figure 11-7) shows an even distribution for gold assays. ALS plotted against AAL also shows a slight low gold bias for AAL for the first two check assay jobs (green and blue dots) and an even distribution for the third job (red dots) as seen in Figure 11-8. The comparison between the three labs for silver assays is a bit more scattered as shown in Figure 11-10 and Figure 11-11, especially for the second check assay job (blue dots). Interestingly, BVM and ALS compare best to each other for both gold and silver assays, shown in Figure 11-9 (gold assays) and Figure 11-12 (silver assays).

Figure 11-7 BVM vs AAL Gold Assays for MLP

 

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Figure 11-8 ALS vs AAL Gold Assays for MLP

 

Figure 11-9 BVM vs ALS Gold Assays for MLP

 

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Figure 11-10 BVM vs AAL Silver Assays for MLP

 

Figure 11-11 ALS vs AAL Silver Assays for MLP

 

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Figure 11-12 BVM vs ALS Silver Assays

 

11.1.7 Blanks

Blank samples were inserted into the sample sequence at a ratio of 1:20 to monitor for carryover contamination and to ensure that there was not a high bias in the assay. Carryover is a process where a small portion of the previous sample contaminates the next sample. AAL allows a total of 1% carryover from preparation and analytical processes combined. Each blank that assays higher than three times the detection limit was evaluated to see if the value reflected carryover or some other problem. For example, if a blank assayed 0.009ppm Au for the FA-PB30-ICP method and the previous sample ran 1ppm Au then the blank was not investigated because acceptable carryover could explain up to 0.01ppm. However, if the blank had assayed 0.015ppm Au it would be more than can be explained by carryover from a 1ppm previous sample and an investigation would be initiated. The investigation included a rerun of the blank and surrounding samples as well as any documentation that was associated with the work order at AAL. There were cases where the investigation does not resolve the reason for the higher than expected value. Figure 11-13 shows the gold assay values for blank samples submitted in the MLP Quality Control Program from September 2017 to June 2018.

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Figure 11-13 MLP Blank Sample Assays Over Time

 

11.1.8 Certified Reference Materials

Certified Reference Materials (“CRMs” or “standards”) were used to monitor the accuracy of the assay results reported by AAL. CRMs were inserted into the sample sequence at a ratio of 1:20 and served to monitor both accuracy and sample sequence errors. A number of different CRMs covering a range of grades and mineral compositions were used for Mother Lode assay quality control. The reported assay value was controlled by the 10% analytical precision specified by AAL, and therefore, in the following discussion the performance of the CRM’s was discussed relative to the specified AAL precision.

CRM’s used in the 2017-2018 drilling campaign were analyzed using the FA-PB30-ICP analytical method. All the CRM values fall within the theoretical analytical precision quoted by AAL and shown in Figure 11-14. The specific CRM label was also shown on the graph, and as shown in Figure 11-15, illustrates that a range of concentrations were being monitored with the suite of CRMs used for MLP assay quality control.

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Figure 11-14 CRM Gold Assays

 

Figure 11-15 CRM Performance Over Time

 

Some of the CRMs used by Corvus have certified silver values while others that are certified for gold only have “reported” silver values. Nevertheless, these values were all used to monitor the accuracy and precision of the silver assays as shown in Figure 11-16. It is clear that most of the CRM silver values do report within the analytical precision of the ICP-61-UT method. There are clearly problems with precision in the CRMs with values less than 0.1ppm silver. Figure 11-17 illustrates that a range of concentrations are being monitored with the CRM suite used at the MLP.

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Figure 11-16 CRM Silver Assays

 

Figure 11.17 Silver Performance of CRMs Over Time

 

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11.2 NBP QA/QC Program

11.2.1 RC Drilling

Corvus drilling between 2015 and 2017 included 97 reverse circulation (RC) holes with a total length of 26,345 metres. Standard procedures, put in place during the 2010-2011 RC sampling program, were followed. All samples were sent to ALS Minerals Laboratories (“ALS Minerals”).

11.2.2 Accredited Laboratories

Assaying for the NBP was performed by ALS Minerals primarily in Reno, Nevada, with some work performed in Vancouver, British Columbia. Corvus has no business relationship with ALS Minerals beyond being a customer for analytical services. The Reno laboratory is Standards Council of Canada, Ottawa, Ontario Accredited Laboratory No. 660 and conforms with requirements of CAN-P-1579, CAN-P-4E (ISO/IEC 17025:2005). The North Vancouver, British Columbia laboratory is Standards Council of Canada, Accredited Laboratory No. 579 and conforms with requirements of CAN-P-1579, CAN-P-4E (ISO/IEC 17025:2005).

Check assaying was performed by Inspectorate America Corporation, Sparks, Nevada. Corvus has no business relationship with Inspectorate America Corporation beyond being a customer for analytical services. The Laboratory is Accredited Laboratory No. 720 and conforms to requirements of CAN-P-1579, CAN-P-4E (ISO/IEC 17025:2005).

11.2.3 Transport and Security

Individual RC samples were not weighed and were grouped by hole in bulk bags which were sealed with a security tag prior to shipment. Each drill hole was sent to ALS Minerals in Reno, Nevada as a separate shipment with a chain of custody document to certify that the seals were intact when the shipment was received.

11.2.4 Duplicates

Duplicates were used to monitor the precision of the assays that were incorporated into the Mineral Resource estimate. Duplicates monitored the three sources of variation: sampling method, preparation and assaying. Preparation duplicates (Prep Duplicates) were used to monitor the sample preparation process, field duplicates were used to document the precision associated with sampling at the drill site, and pulp duplicates were used to monitor the assaying process.

Sample preparation duplicates were created by crushing the sample and then splitting it in half. The two halves were then processed as separate samples. Five Prep Duplicates were created for each drill hole. The selection of preparation duplicates was made by geologists logging the hole, based on their interpretation of lithologies and degree of mineralization.

Figure 11-18 and Figure 11-19 are graphs of the original sample assay versus the duplicate sample assay, for gold and silver, respectively. The ±10% precision lines are shown in the graph to illustrate the trend of the data pairs. The results show that the preparation duplicates reproduced the original assay very well for both gold (coefficient of variation 27%) and silver (coefficient of variation 18%).

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Figure 11-18 Preparation Duplicate Gold Assays from Sierra Blanca

 

Figure 11-19 Preparation Duplicate Silver Assays from Sierra Blanca

 

11.2.4.1 Field Duplicates

Field duplicates are selected by the Project Manager after the gold assay results for an RC hole have been received. At the drill site, two samples are taken for every five foot interval, with the primary sample labeled with a white tag. The secondary sample is labeled with a red tag with an “M” suffix added to the original sample number. The red tag bags are used for the field duplicates. The selected field duplicates would have a gold value of 0.1ppm or higher. Field duplicates underwent the same transport and security procedure as all other RC and rock samples. The field duplicates were used to check the accuracy and precision of the sample splitting at the drill site. Figure 11-20 and 11-21 are graphs of the original assay versus the duplicate assay, and show that the splitting of duplicate samples at the drill site was both accurate and precise for gold (coefficient of variation 26%) and silver (coefficient of variation 22%) content, respectively.

Corvus Gold Inc.
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Figure 11-20 Field Duplicate Gold Assays from Sierra Blanca

 

Figure 11-21 Field Duplicate Silver Assays from Sierra Blanca

 

11.2.4.2 Pulp Duplicates

Pulp duplicates reflect the homogeneity of the pulp material that is subjected to the fire assay and variations generally reflect the nugget effect in gold samples. In this instance ALS Minerals routinely run pulp duplicates as part of their internal QA/QC program and these assays were reported as part of the assay package.

Figure 11-22 and Figure 11-23 show that the gold and silver assay values for ALS internal pulp duplicates reproduced accurately for gold (coefficient of variation 22%) and silver (coefficient of variation 14%), respectively.

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Figure 11-22 ALS Pulp Duplicate Gold Assays for NBP

 

Figure 11-23 ALS Pulp Duplicate Silver Assays for NBP

 

11.2.5 Check Assays

One hundred and ninetyfour (194) samples were sent to Inspectorate America Corporation to check results from ALS. Figure 11-24 plotted the ALS assay (ICP21/22) and the Inspectorate assay (FA330) for gold, and Figure 11-25 plots the ALS assay (ME-MS61) versus the Inspectorate assay (MA200) for silver, and show that there was very good agreement between the ALS and Inspectorate values, for gold and silver, respectively.

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Figure 11-24 Comparison of ALS (ICP21/22) and Inspectorate (FA330) Gold Analyses in Duplicate NBP Samples

 

Figure 11-25 Comparison of ALS (ME-MS61) and Inspectorate (MA200) Silver Analyses in Duplicate NBP Samples

 

11.2.6 Blanks

Blank samples were inserted into the sample sequence at a ratio of 1:20 to monitor for carryover contamination and to ensure that there is not a high bias in the assay values. Carryover is a process where a small portion of the previous sample contaminates the next sample. ALS Minerals allows a total of 1% carryover from preparation and analytical processes combined. Each blank that assays higher than three times the detection limit was evaluated to see if the value reflected carryover or some other problem. For example, if a blank assayed 0.006ppm Au for the Au-ICP22 method and the previous sample ran 1ppm Au then the blank was not investigated because acceptable carryover could explain up to 0.01ppm. However, if the blank had assayed 0.015ppm Au which was more than could be explained by carryover from a 1ppm previous sample then an investigation was initiated. The investigation included a rerun of the blank and surrounding samples, as well as any documentation that was associated with the work order at ALS Minerals. There were cases where the investigation did not resolve the reason for the higher than expected value. Figure 11-26 shows the performance of blank samples submitted for the NBP Quality Control progam from 2015 to 2017.

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Figure 11-26 Sierra Blanca Blanks for Au-Icp22/ICP21 Methods

 

11.2.7 Certified Reference Materials

Certified Reference Materials (“CRMs” or “standards”) were used to monitor the accuracy of the assay results reported by ALS Minerals. CRMs were inserted into the sample sequence at a ratio of 1:20 and served to monitor both accuracy and sample sequence errors. A number of different CRMs covering a range of grades and mineral compositions were used at the NBP. Each CRM comes with a certified concentration with a stated uncertainty. However, the precision on the assay is ultimately controlled by the 10% analytical precision reported by ALS Minerals. Therefore, in the following discussion the performance of the CRMs was discussed relative to the specfied ALS Minerals precision.

CRMs used in the 2015-2017 drilling campaign were analyzed using the Au-ICP22 and Au-ICP21 analytical methods. All the CRM values fell within the theoretical analytical precision quoted by ALS Minerals as shown in Figure 11-27. Figure 11-28 illustrated that a range of concentrations are being monitored with the CRM suite used at the NBP was greater than the generally expected assay values.

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Figure 11-27 Au-ICP22 Certified Reference Material Gold Assays

 

Figure 11-28 Performance of CRMs Over Time

 

Some of the CRMs used by Corvus have certified silver values while others that are certified for gold only have “reported” silver values. Nevertheless, these values were used to monitor the accuracy and precision of the silver assays as shown in Figure 11-29. It was clear that most of the CRM silver values reported within the analytical precision of the ME-MS61 method. There were clearly some problems with precision in the CRMs with values less than 0.1ppm silver. Figure 11-30 illustrates that the range of concentrations that were monitored by the CRM suite used at the NBP covered the range of expected assay values.

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Figure 11-29 Silver Assays for CRMs

 

Figure 11-30 Silver Performance of CRMs Over Time

 

11.3 Data Adequacy

In the opinion of the author, sample preparation, security and analytical procedures as described in this Section 11 are adequate and can be relied upon in the estimation of Mineral Resources and for the PEA, each as described herein.

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12 Data Verification

The principal author, Mr. Wilson, has verified the data used in this Technical Report by:

· Visiting the Project and confirming the geology and mineralization;

· Visiting the core and RC storage areas and inspecting the core cutting facility;

· Reviewing drill core;

· Verifying the location of drill holes in the field;

· Reviewing the QA/QC protocols;

· And, reviewing the quality analysis of RC drilling data.

The principal author, Mr. Wilson, concludes that:

· Exploration drilling, drill hole surveys, sampling, sample preparation, assaying, and density measurements have been carried out in accordance with CIM Best Practice Guidelines and are suitable to support the Mineral Resource estimates and PEA contained herein

· Exploration and drilling programs are well planned and executed and supply sufficient information for Mineral Resource estimates and Mineral Resource classification and the PEA contained herein

· Sampling and assaying includes sufficient quality assurance procedures

· Exploration databases are professionally constructed and are sufficiently error free to support Mineral Resource estimates and the PEA contained herein

Therefore, in the opinion of the principal author, such data is adequate and can be relied upon to estimate Mineral Resources for the Project and for the purposes of the PEA as described in this Technical Report.

12.1 Database Error Checks

The drill database was reviewed by Mr. Wilson by selecting 10% of the gold sample records in the database. The certified assay certificates were crosschecked with the data entry in the database. The data entry procedures have been verified by Mr. Wilson and are accurate as compared to the certificates.

12.2 Data Verification Samples

Scott Wilson independently collected seven field duplicates during his visit to the North Bullfrog property on June 6-8, 2017 and collected 10 field duplicates during his visit to the Mother Lode Property on January 16, 2018. The two groups of verification samples were submitted to Inspectorate America Corporation in Sparks, Nevada and ALS Global’s laboratory in Reno, Nevada, respectively. The purpose of these data verification samples was to independently verify the existence of the mineralization and to review the reproducibility of the original Corvus assays. No limitations were placed on the author’s ability to review data or to independently verify the data used in the Mineral Resource estimate and the PEA. Samples were marked by Scott Wilson with information regarding the selected sample (Date, Sample#, Hole ID, From, To, Original Assay). The results show that check samples grades range within acceptable limits compared to the original individual sample grades. Table 12-1 and 12-2 compare the results of the data verification testing for the North Bullfrog and Mother Lode sites, respectively.

Table 12-1 North Bullfrog Data Verification Samples (Scott Wilson - 2017)

Scott Wilson Data Verification Samples – June 6-8, 2017
Sample #	Hole ID	From (m)	To (m)	Original (ALS) Value Au (g/t)	Verification (IAC) Value Au (g/t)	Lithology
NB187273	Blank	n/a	n/a	0.00	0.00	Blank*
NB187274	NB-17-329	290	295	0.722	0.767	Tsb
NB191471	NB-16-314	915	920	0.931	0.976	Tpf
NB191592	NB-16-315	860	865	0.700	0.724	Tnb
NB191667	NB-16-316	560	565	0.610	0.554	Tnb
NB191773	NB-16-318	620	625	1.070	0.718	Tpf
NB191794	NB-16-319	515	520	0.525	0.531	Tnb
NB192224	NB-16-325	595	600	0.736	0.733	Tsb
G913-1	NB-36	n/a	n/a	0.820	0.783	CRM**
*-Blank standard material ** - Certified Reference Material

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Table 12-2 Mother Lode Data Verification Samples (Scott Wilson - 2018)

Scott Wilson Data Verification Samples – January 16, 2018
Sample #	Hole ID	From	To	Original Value Au (g/t)	Verification Value Au (g/t)	Lithology
P352689	ML17-001	110.34	111.8	1.65	1.555	Tjvs
P352929	ML17-002	96.15	97.7	5.36	5.34	Tjs
NB196744M	ML17-010	167.64	169.16	0.383	0.307	Tjvs
NB196800M	ML17-010	245.36	246.89	1.31	1.245	Tjc
NB197922M	ML17-014	114.3	115.82	1.68	1.505	Tjvs
NB197935M	ML17-014	132.59	134.11	1.03	0.851	Tip
NB200108M	ML17-017	300.23	301.75	1.7	1.545	Tip
NB200122M	ML17-017	318.52	320.04	1.95	2.03	Tip
NB204922M	ML17-026	152.4	153.92	2.36	2.53	Tip
NB204926M	ML17-026	158.5	160.02	1.11	1.14	Tip

12.3 Verification of historical mother Lode drilling data

With the purchase of Goldcorp Daisy LLC, Corvus received historic drill data from the immediate area of Mother Lode which had been developed in the previous exploration and mining operations conducted by various companies between 1985 and 1997. This included a total of 520 drill holes and 66,932 m of drill and log data. A group of 164 of these drill holes were in close proximity to Mother Lode and had been used in historical estimates of the mining resource. In 2017 and 2018, Corvus drilled an additional 52 holes, which provided a basis to evaluate the quality of the historic information. Table 12-1 lists statistical indices which provide a basis to compare the 2 population of drilling data.

Figure 12-1 presents a map showing the perimeter of the historic Mother Lode open pit, and the locations of drill collars from the 2 drilling data groups. Figure 12-1 shows the drill hole collars for the historic drilling data in green and the modern drilling data performed by Corvus in blue, see below.

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Figure 12-1 Map Showing Historical Surface Topography and the Location of the Historic Mother Lode Open Pit

 

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Table 12-3 Comparison of Statistical Indices for Combined, Historic and Corvus Drilling Data

	Raw Data Not Capped			Data Capped at 13 g/t
Data Group	Combined	Historic	Corvus	Combined	Historic	Corvus
No. of Samples	8170	5163	3007	8159	5152	3007
Minimum	0.100	0.103	0.100	0.100	0.103	0.100
Maximum	25.749	25.749	12.998	12.998	12.926	12.998
Mean	1.075	1.134	0.974	1.053	1.099	0.974
Standard Deviation	1.457	1.575	1.222	1.321	1.374	1.222
Variance	2.122	2.480	1.492	1.745	1.888	1.493
Coeff. Of Variance	1.355	1.389	1.255	1.255	1.251	1.255

The highest grade of the assays in the Historic data was 25.749 g/t while the highest grade for Corvus drilling is 12.998 g/t. In total there are 11 assays in the Historic data higher than 13 g/t. The 11 assay values are spread such that they have very little anomalous effect on the data, an average grade of 1.134 vs. 1.053, with very close standard deviations. Ignoring the Historic grades higher than 13 g/t as anomalous, the comparison in Table 12-1 indicates that there is very little difference in the two assay databases, suggesting that the Historic data collection and analyses was carried out with a standard care suitable for the professional analysis of the mineral deposit.

The distribution of drill assay data in the 3 groups of data (Combined, Historic and Corvus) is illustrated by the histograms and cumulative frequency plots in Figure 12-2, 12-3 and 12-4, respectively. There were no apparent statistical trends which would give concerns in the use of the Historic assays in estimating a NI 43-101 compliant resource, including classification of the resources. Variances of the data suggest assay reproducibility was not an issue. Coefficients of variance suggest the Corvus drilling data lie within the same population of assays indicated by the Historic data. Comprehensive statistical analysis has not identified any anomalies in the Goldcorp data.

Scott Wilson (2018) sees no reason for concern with the inclusion of the Historic data in estimating Mineral Resources for Motherlode, and recommends that it is acceptable to use the Historic assays for the estimation of Mineral Resources.

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Figure 12-2 Histograms and Cumulative Frequency of Occurrence for the Combined Grade Data

 

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Figure 12-3 Histograms and Cumulative Frequency Occcurrence for the Historic Grade Data

 

12.4 Potential for Downhole contamination in RC drilling

The Corvus Reverse Circulation drilling data has been evaluated to assure that downhole contamination by mineralized material falling downhole has been minimal during the drilling operations. Evidence of downhole contamination in Corvus data was evaluated by examining the data for:

· Patterns of “decay” in the grade values of adjacent sample intervals, where there is a monotonic decrease in grade below zones of mineralization;

· Patterns of spikes in the grade (“cyclicity”) of adjacent sample intervals that occur when new rod connections are made in the drilling process;

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· Material of a different lithology than predominat for an interval observed to be mixed into the drill chips logged by the geologists.

Analysis of North Bullfrog RC drill data had been reported in a previous technical report (Steininger et. al., 2013) based on decay analysis and cyclicity analysis, plus comparison of core and RC twin holes. Similar analyses for the Corvus Mother Lode RC drilling have been performed and are reported here. The comparison included the 75 RC holes and 3 core holes drilled during 2017-2018.

Corvus collects data during the drilling and logging operations to support the detection of contamination problems in the data. The depth of each rod connection is noted in the data record, so that patterns of cyclicity can be detected. Spreadsheets have been developed to evaluate the grade data to assure minimal contamination.

The evaluation of contamination of the drilling data indicated 21 samples out of 3563 rod connections (0.5 %) may have been contaminated and should be removed from the database.

12.4.1 Decay Analysis

Spreadsheets have been developed to identify monotonic decreases of three and four samples below a mineralized interval. Analysis of the patterns in the Corvus RC data has been compared to an identical analysis of the patterns in the 3 Corvus core holes drilled at Mother Lode, in Table 12-4. The analysis indicates similar patterns in the occurrence of intervals of monotonic increase and decrease in the core versus RC data sets. Since downhole contamination cannot occur in the core data set, the comparison indicates that downhole contamination was not a common occurrence in the Corvus RC drilling.

Table 12-4 Comparison of Montonic Decay Patterns in Mother Lode Grade Data of Corvus RC and Core Drilling

Drill Type	Sampled Intervals	No. of intervals with 4 sample monotonic decrease	No. of intervals with 3 sample monotonic decrease	No. of intervals with 4 sample monotonic increase	No. of intervals with 3 sample monotonic increase
RC Drilling -samples	14,187	431	1101	99	277
RC Drilling - % of total	-	3.0	7.8	0.7	2.0
Core Drilling - samples	499	7	31	11	29
Core Drilling - % of total	-	1.4	6.2	2.2	5.8

12.4.2 Cyclicity Analysis

The locations of drill rod connections in the drill depth record have been recorded for all RC drilling done by Corvus at Mother Lode. A test for cyclicity was based on testing for local grade maximums that occured in sample intervals immediately after a rod connection. The standard rod length was 20 feet (6.098 m) and the standard sample interval was 5 feet (1.524 m), so samples were assigned a “position” index (1, 2, 3, and 4) indicating the relative location of a sample below a rod connection. The first sample in each hole was assigned a position zero (“0”), so that it was not treated as a rod connection.

The test for a local maximum was done by comparing the Au grade of the sample in position 1 with the three samples above and below (an interval of 35 feet [10.67 m]). If the Au grade of a sample in position 1 had the maximum grade of the 7 samples in the interval, its grade was compared to the average grade of the 7 samples in the interval. If the maximum Au grade was more that 10% above the average grade, the location was flagged as potentially anomalous.

The number of local grade maxima in individual holes was analyzed to establish the variability. The data in the 75 RC drill holes had 3563 rod changes, and the total number of rod changes with local grade maxima was 567. The average percentage of rod changes with local grade maxima in individual holes was 15.7% with a standard deviation of 6.7%. A total of 5 holes were flagged for further review because the number of anomalous rod changes exceeded 25.7% (mean + 1.5 standard deviations).

A further test for evidence of contamination was to search the downhole grade data for clusters of local grade maxima that might indicate contamination problems during a specific period of drilling. The drilling rig was operated on a single shift per day basis, so there was always good continuity of operator personnel. If local grade maxima occurred in 3 successive rod connections, the grade data were examined. The number of intervals in the grade data with 3 or more successive local grade maxima was 31, and of those intervals, 24 intervals had grades that were less than 2 times the assay detection limit of 0.003 ppm Au. The 7 holes where successive grade maxima were greater than 2 and the grades in the interval were greater than 2 times the detection limit were flagged for further review.

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The 567 rod connections with local grade maxima were sorted to identify their characteristics. The number of rod connections where the local maxima was less thatn 2 times the detection limit was 202 or 36%. The remaining rod connections were evaluated with respect to mineralization type (oxide or sulfide) and with respect to the projected cut-off grade of the process alternative (0.15 ppm for oxide, 0.5 ppm for sulfide). The local maxima grade at 109 of the rod connections (19%) was greater than the projected cut-off grade for the indicated mineralization type. The grade of samples around each of the rod connections with local maxima greater than the indicated process cut-off grade was examined using the following criteria:

· The local maxima is not a spike – the grades of samples surrounding the local maxima have similar values;

· The flagged local maxima is higher than the surrounding samples, but only marginally, therefore possibly contaminated;

· The flagged local maxima are distinctly higher and unlike the grade of surrounding samples, therefore it is probably contaminated.

Table 12-5 lists the results of the evaluation of the 109 local maxima with grades greater than the process cut-off and indicates that 21 samples had characteristics that suggested possible or probable contamination. The 21 samples where the numerical review for cyclicity indicated possible or probable contamination were reviewed in more detail by inspecting the chip trays and by comparing the geology and geochemistry of the sample. This review indicated that the majority of grade maxima was real and could be attributed to localized geology and geochemistry. Only 3 of the identified 21 local grade maxima at the rod connection could not be explained by other data.

Table 12-5 Distribution of Local Maxima with Greater Than Process Au Cut-off Grade

Criteria	No Spike	Possible contamination	Probable contamination
No. of Local Maxima	88	10	11

12.4.3 Visible analysis of chips and geochemistry

Visible contaminated intervals in the chip trays were logged routinely for all 2017-18 RC holes. Once assay results were received, a review of the chip trays and geochemistry was conducted for each hole to verify where mineralized intervals may be influenced by contamination coming downhole from mineralized intervals above. A total of 16 of the 75 RC holes have been identified as having at least one contaminated sample. A total of 134 samples from these 16 holes were identified for potential removal from the data used for the Mother Lode Mineral Resource and were evaluated by the Qualified Person responsible for the Mineral Resource estimate.

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13 Mineral Processing and Metallurgical Testing

The samples tested as part of the MLP and NBP metallurgical programs and presented here are representative of the various types and styles of mineralization found at the Mother Lode and NBP. The testing data are representative and the author knows of no known processing factors or deleterious elements that could have a significant effect on potential economic extraction.

MLP metallurgical test work has been performed on samples from core and RC drill cuttings generated during the 2017-2018 exploration programs at MLP and on a “grab” sample from the MLP pit in 2017. The data consist of the results of cyanidation bottle roll tests, gravity concentration with cyanidation leaching of tails, flotation, comminution, diagnostic leaching, mineralogy, concentrate pressure oxidation and concentrate roasting followed by cyanidation, and gravity/flotation/cyanidation tests. The metallurgical test work addresses both the oxidized portion of the mineralization, suitable for heap leach processing and sulphide portion, which is amenable to cyanidation after concentration by flotation and pressure oxidation of sulphides.

NBP metallurgical test work has been performed on samples from core and RC drill cuttings generated during the exploration programs at NBP and on bulk sample materials collected from surface outcrops and from dumps resulting from previous underground mining at the NBP. The data consist of the results of cyanidation bottle roll tests at different particle sizes, cyanidation column leach tests at 80% passing (P80) -6.3mm, -12.5 mm (½ inch), -19 mm (3/4 inch) and -51 mm (2 inch) and gravity concentration with cyanidation leaching of tails tests. The metallurgical test work addresses both the disseminated, oxidized portion of the mineralization, suitable for heap leach processing and the higher-grade, vein and stockwork mineralization from the YellowJacket zone which will require milling with gravity concentration and cyanide leaching of tails. These results demonstrate high cyanide solubility of the contained gold and silver in YellowJacket mineralization at 100-200 mesh. The data indicate that simple process systems based on cyanide leaching are suitable for both the disseminated, oxidized low-grade mineralization and for the higher-grade vein and vein stockwork mineralization at YellowJacket. The data indicate North Bullfrog Sierra Blanca tuff, Pioneer Formation tuff, Dacite, and Rhyolite refractory sulphide materials are cyanide amenable after sulphide and gold concentration by flotation and alkaline sulphide oxidation.

Metallurgical testing related this Technical Report was performed by McClelland Analytical Services Laboratories Inc. of Sparks Nevada (“McClelland”), McClelland is an ISO 17025 accredited facility that supplies quantitative chemical analysis in support of metallurgical, exploration and environmental testing using classic methods and modern analytical instrumentation. McClelland has met the requirements of the IAS Accreditations Criteria for Testing Laboratories (AC89), has demonstrated compliance with ANS/ISO/IEC Standard 17025:2005, General requirements for the competence of testing and calibration laboratories, and has been accredited, since November 12, 2012. Hazen Research Inc. (“Hazen”), an independent laboratory, has performed flotation, AAO testing and cyanide leach testing on samples of sulphide mineralization from the YellowJacket zone and Swale area of Sierra Blanca. Hazen performed a roasting test on Mother Lode flotation concentrate. Hazen holds analytical certificates from state regulatory agencies and the US Environmental Protection Agency (the “EPA”). Hazen participates in performance evaluation studies to demonstrate competence and maintains a large stock of standard reference materials from the National Institute of Standards and Technology (NIST), the Canadian Centre for Mineral and Energy Technology (CANMET), the EPA and other sources. Hazen’s QA program has been developed for conformance to the applicable requirements and standards referenced in 10 CFR 830.120 subpart A quality assurance requirements, January 1, 2002.

Other scoping tests have been performed by lab organizations and are presented for completeness, including:

· Hazen – roasting tests of MLP material and analysis of cyanide bottle roll tests performed at RDi. NBP bottle roll testing of six samples collected from drilling in the Mayflower area, and refractory material treatment by sulphide and gold flotation, flotation concentrate sulphide oxidation by AAO and cyanidation of oxidized material.

· Kappes, Cassidy and Associates, Reno, Nevada (“KCA”) – bottle roll testing on materials from the Connection area, core material from the Sierra Blanca area and on RC sample materials from the Sierra Blanca and Jolly Jane areas;

· Advanced Mineral Technology Laboratory Limited (“AMTEL”), London, Ontario, Canada - gold deportment studies of two samples of leached tail fractions of un-oxidized samples from the Savage Valley area.

· Bureau Veritas Commodities Canada Ltd. – Quemscan analysis of gravity concentrate samples YellowJacket vein and stockwork materials and MLP sulphide composites.

· Resource Development Inc. (“RDi”) completed pressure oxidation and cyanidation test on MLP sulphide concentrates.

· Blue Coast Group Research, Parksville, B.C., conducted scoping test work on a grab sample from the MLP pit that included standard cyanidation, flotation, minerology, and whole ore atmospheric alkaline oxidation.

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13.1 Metallurgical Development at MLP

During 2017, before the drill program began, a single bulk sample of sulphide material was obtained from the bottom of the Mother Lode pit and shipped to the Blue Coast Group, in Parksville, B.C. Initial concerns regarding sulfide material test work on material that had been exposed to the atmosphere and weathered for twenty-one years were validated during the test program. The program was prematurely halted in anticipation of fresh material from the 2017-2018 drilling program. Blue Coast, 2017 metallurgical test results are presented for completeness, but are not included in the current metallurgical recovery estimate.

13.1.1 2017 - Metallurgical Test Program

A surface sample of sulfide material from the bottom of the MLP pit was shipped to Blue Coast Research for metallurgical testing. The material had been exposed to the atmosphere and weathered for twenty-one years, after initial test results were reviewed, the test program was halted in anticipation of fresh material from the 2017-2018 drilling program.

13.1.1.1 Head Analysis

Head analysis indicated the sample contained 1.7 g/t gold and less than 10 g/t silver. Total sulfur and sulfide sulfur analyses indicated 0.78 wt% and 0.71 wt%, respectively. Total and organic carbon analyses indicated 0.973 wt% and 0.063 wt%, respectively.

Metallic screen analysis at 75 µm indicated the plus 75 µm fraction contained 11.6% of the weight, 18.4% of the gold, and assayed 2.52 g/t Au. The minus 75 µm fraction contained 88.4% of the weight, 81.5% of the gold, and assayed 1.58 g/t Au.

Preg-robbing test indicated marginal preg-robbing at 9.6% of the spiked sample. Table 13-2 presents analysis of the Mother Lode Pit Material.

13.1.1.2 Cyanidation

A standard bottle roll and CIL bottle roll tests were completed at a grind P₈₀ of 100 µm. The tests were completed in 40% solids slurry with 1 g/l sodium cyanide for 48 hours. The standard and CIL bottle roll tests indicated 4.9% and 8.4% gold dissolution, respectively. Gold in the material was not amenable to conventional cyanidation.

13.1.1.3 Whole Sample – Atmospheric Oxidation

A single atmospheric oxidation test was completed on a whole sample. The sample was ground to produce P₈₀ 40 µm and leached in 25 wt% solids at 75^o C for 72 hours. Soda ash was added to maintain the pH. The leach residues at 24, 48, and 72 hours were analyzed for sulfur. Sulfide oxidation at 24, 48 and 72 hours was 28%, 48%, 56%, and 44% on neutralized solids.

The oxidized residue was leached in sodium cyanide. Gold dissolution after 48 hours with 1 g/l NaCN was 41%. Cyanide consumption was 1.7 kg/t concentrate. The material was leached for an additional 24 hours and showed a 7% decrease in gold extraction. Gold accountability was poor at 85.2%.

13.1.1.4 Flotation

Ten rougher flotation tests were completed. The material particle size ranged from P₈₀ 74 to 142 µm. Rougher mass pull averaged 22.7% ranging from 16% to 30.9%. Gold recovery averaged 66.2% and ranged from 56.9% to 73.1%. Rougher flotation gold grades averaged 5.0 g/t and ranged from 3.8 to 5.9 g/t. Sulfide recovery averaged 82.7% and ranged from 77.4 to 86.5%. Concentrate sulfide grade averaged 2.8% and ranged from 2.0 to 3.6%. Flotation results are summarized in Table 13-1.

Table 13-1 2017- MLP Pit Sample – Flotation Test Results.

 

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The highest gold recovery was obtained in Test F-8. Test F-8 was ground to a P₈₀ 100 µm. Flotation reagents were 67 g/t of W22C, 240 g/t PAX, and 100 g/t copper sulfate, and the flotation time was 30 minutes. Rougher concentrate mass recovery was 23.8%. Gold recovery and grade were 73.1% and 4.88 g/t, respectively. Sulfide recovery and grade were 84.5% and 2.79%, respectively.

13.1.1.5 Mineralogy-Pit Sample

A sample of Mother Lode master composite was ground to a P₈₀ 94 µm and screened into the following size fractions:

• +75µm – 28.1% retained

• -75µm / +38µm-18.8% retained

• -38µm (dry sieve) 53.0% passing

• -38µm (wet sieve)

The -38 µm fraction from both the wet and dry fractions were kept separate to allow for a more detailed analysis of the clay component. Material from each of the size fractions was submitted for preparation into polished sections and automated mineral analysis using a TESCAN-TIMA system. The head analysis of the material is presented in Table 13-2.

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Table 13-2 Mother Lode – 2017 Pit Sample – Head Analysis.

Parameter	Units	Result
Au	g/t	1.657
Ag	g/t	<= 10
Ctot	wt %	0.973
Corg	wt %	0.063
Total S	wt %	0.783
Sulphide (S-2)	wt %	0.707
Preg Rob	%	9.6
Al	wt %	0.48
As	wt %	0.1
Ba	wt %	0
Bi	wt %	<= 0.001
Ca	wt %	2.18
Cd	wt %	<= 0.001
Co	wt %	<= 0.001
Cr	wt %	<= 0.001
Cu	wt %	0
Fe	wt %	1.01
Hg	mg/kg	<= 10
K	wt %	0.32
Mg	wt %	0.57
Mn	wt %	0.06
Mo	wt %	<= 0.001
Na	wt %	0.11
Ni	wt %	<= 0.001
P	wt %	0.02
Pb	wt %	0
S	wt %	0.84
Sb	wt %	0
Se	wt %	<= 0.001
Ti	wt %	0
Zn	wt %	0.01

13.1.1.5.1 Modal Minerology

Mother Lode material is dominated by quartz (45%) and orthoclase (44%). On a size-by-size basis the quartz is more prevalent in the coarser fractions, while the orthoclase is concentrated slightly in the finer (-38µm) fraction, suggesting that these materials may have different breakage characteristics in the mill, with the orthoclase more likely to slime and the quartz more resistant to breakage.

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Pyrite was the predominate sulphide mineral present (2.9%) with minor amounts of arsenopyrite noted (0.2%). The presence of montmorillonite (a swelling clay) was noted.

13.1.1.5.2 Sulfide Liberation

Liberation of pyrite and arsenopyrite was evaluated. In the following analysis the following descriptions apply:

• Liberated material: >80% of the surface area of the target mineral is exposed

• Middling: >30%, <80% of surface area of target mineral is exposed

• Locked material: <30% of the surface area of the target mineral is exposed

Pyrite liberation was less than 50% across all size fractions and points to some finer textures which may be inhibiting effective flotation recovery. The release curve for pyrite flattens out at finer size fractions and suggests that a portion of the pyrite is present as very fine mineral textures which may make extremely high recoveries difficult to achieve.

An analysis of the pyrite grain sizes indicates that most of the pyrite is reasonably coarse. However, the presence of some fine-grained pyrite is noted specifically in the slimes fraction (-38µm wet).

The locked pyrite is predominately associated with quartz and orthoclase. The fines fraction (-38µm) showed a greater proportion of locking with orthoclase, while the coarser fractions showed slightly greater amounts of locking with quartz.

This assessment highlights that while most of the pyrite is reasonably coarse, some fine-grained textures are present. This fine-grained pyrite is predominately locked with quartz and orthoclase. Finer grinding may assist with recovery of this pyrite.

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13.1.2 2018 - Mother Lode Test Program

During 2018, fresh materials from core assaying and field duplicates from the RC drilling were obtained and composites developed for oxide material, two sulphide materials, and comminution tests. The sulfide materials are predominantly associated with two different geologic units; Tip1 – a porphyritic rhyolite intrusive, and Tjvs – sediments of Joshua Hollow.

The 2018 test program took place at McClelland Laboratories, Inc. (MLI, 2018), Hazen Research, Inc. (Hazen, 2018), and Resource Development, Inc. (RDi, 2018).

Three metallurgical composites were prepared from drill hole reject samples to represent oxide material, sulfide material, type Tip1, and sulfide material type Tjvs. The samples comprising these composites were RC drill duplicate samples and were nominally minus 6.3 mm (1/4" in.) size.

Two additional composites were prepared for comminution testing. The samples were broken core, nominally 25.4 mm (1") diameter in size.

A separate set of 52 oxide samples were investigated for preg-rob assaying, as well as a series of bottle roll tests.

McClelland Laboratories completed:

· Head Analyses: conducted on the three metallurgical composites included fire assay, cyanide shake analysis, carbon and sulfur speciation analyses, and diagnostic leach tests.

· Comminution Testing: a crusher work index, abrasion index and ball mill work index test was conducted on each of the two comminution composites.

· Gravity Concentration Testing: a whole sample gravity concentration test was conducted on each of the two sulfide composites to determine response to whole sample gravity concentration at an P₈₀-75µm feed size.

· Flotation Testing: Whole sample rougher flotation testing was conducted on the two sulfide composites. Parameters evaluated included feed size, reagent schemes, solids density and desliming before flotation. Bulk, 14 kg, flotation tests were also conducted to generate concentrate for ultra-fine grinding/cyanidation testing and concentrate oxidation tests.

· Flotation Product Cyanidation Testing: Flotation concentrate from bulk whole flotation tests conducted on the sulfide composites. Concentrate samples were sent to ALS (Kamloops) for ultra-fine grinding (UFG) to generate feeds for cyanidation leach tests. Duplicate cyanidation tests, both with and without activated carbon added, were conducted on the reground concentrate from each composite. Bottle roll cyanidation tests were also conducted on flotation tailings from three selected tests.

· Bulk Gravity Concentration/Flotation Testing: A bulk (49 kg) gravity concentration test was conducted on each of the two sulfide composites at a P₈₀ -75µm feed size, using a Knelson concentrator to generate concentrate and feed for bulk flotation testing. The resulting gravity concentrate was cleaned by panning. The resulting cleaner tails and rougher tails were recombined and used as feed for a bulk flotation test.

· Rougher flotation was conducted on each of the two gravity tailings to generate concentrate for testing. The gravity cleaner concentrates and flotation rougher concentrates were combined for down-stream process testing.

· Acid base accounting (ABA) testing on the corresponding flotation tailings were also completed.

Additionally,

· Concentrate roasting tests and cyanidation of roasted residues was completed at Hazen Research Inc.

· Concentrate pressure oxidation (POX) and cyanidation of residues was completed at Research Development Incorporated, RDi.

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13.1.2.1 Composite Characterization

13.1.2.1.1 Head Analysis

Composite Head analysis is presented in Table 13-3.

Table 13-3 2018 Composite Head Analysis

 

13.1.2.1.2 Cyanide Shake Tests

Cyanide shake tests on the Oxide, Tip1, and Tjvs composites indicated gold dissolutions of 49.4%, 4.0%, and 2.4%, respectively.

13.1.2.1.3 Preg-Robbing

Preg-robbing tests on the oxide, Tip1, and Tjvs composite indicate preg-robbing (PR) values of 6.6, 3.2, and 16.7, respectively, all very slightly preg-robbing. A value of 0.0, indicates no preg robbing, a positive value indicates preg robbing, a marginal negative value may indicate analytical errors in the procedure.

Preg-robbing was evaluated on an additional 52 oxide samples. The average preg-robbing factor (PR) was minus 0.7 with a range from minus 8.5 to 4.2. The oxide samples tested did not indicate significant preg-robbing.

13.1.2.1.4 Sulfide Composite Desliming Evaluation

De-sliming the sulfide composites was evaluated. The Tip1 composite indicated the sands fractions contained 79.6% of the mass and 83.8% of the gold at a grade of 2.2 g/t. The slime fraction contained 20.4% of the mass and 16.2% of the gold at a grade of 1.68 g/t.

Tjvs composite indicated the sands fractions contained 74% of the mass and 76.5% of the gold at a grade of 2.37 g/t. The slime fraction contained 26.0% of the mass and 23.5% of the gold at a grade of 2.07 g/t. The results suggest upgrading by de-sliming would not be advantageous.

13.1.2.1.5 Standard and CIL Cyanide Leach Tests

13.1.2.1.5.1 Oxide Material Bottle Roll Leach

Oxide material bottle roll leach tests were completed on material ground to a P₈₀ of 74 µ. The tests were conducted for 72 hours with the initial cyanide concentration of 1 g/l NaCN and lime added to maintain the pH between 10.5-11.0. Average pH in four leach tests was 10.9 and averaged dissolved oxygen 6.3 mg/l.

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Leach results indicated the oxide composite sample tested was preg robbing. Non-CIL bottle roll gold and silver recoveries averaged 53% and 22.6%, respectively. CIL bottle roll recoveries for gold and silver averaged 81.1% and 20.0 %, respectively. Sodium cyanide and lime consumptions averaged 0.2 and 2.1 kg/t, respectively.

An additional 52 oxide samples were obtained and tested for preg-robbing which indicted no preg robbing. Additional bottle roll leach tests on Tip1 oxide material P₈₀ 74 µm to 3900 µm indicated gold extractions from 85.5% to 97.1% and averaged 90.5%. Sodium cyanide and lime consumptions averaged 0.1 and 2.35 kg/t.

Additional oxide bottle roll samples at sizes from 74 µm to 3900 µm indicated gold extractions from 87.2% to 93.4% and averaged 91.3%. Sodium cyanide and lime consumptions averaged 0.1 and 1.4 kg/t.

Average gold dissolution for all oxide samples tested was 84.9% with average sodium cyanide and lime consumptions of 0.13 kg/t and 1.95 kg/t, respectively.

13.1.2.1.5.2 sulphide material Bottle Roll Leach

Sulphide material bottle roll leach tests were completed on material ground to a P₈₀ of 106 µm. The tests were conducted for 72 hours with the initial cyanide concentration of 2 g/l NaCN and lime added to maintain the pH between 10.5-11.0.

Leach results indicated the Tip1 composite gold recovery was 32.6%. Sodium cyanide and lime consumptions were 23.0 and 28.2 kg/t, respectively.

Leach results indicated the Tjvs composite gold recovery was 44.0%. Sodium cyanide and lime consumptions were 23.7 and 24.0 kg/t, respectively.

13.1.2.1.6 Diagnostic Leach Tests

Diagnostic leaching tests on the three composites were completed. The results are discussed below as cumulative dissolution and presented in Table 13-4 below.

The Oxide composite indicated 81.9% gold dissolution in sodium cyanide, 85.3% gold dissolution after HCl/NaCN treatment, 92.2% gold dissolution after HCL-HNO₃/NaCN treatment, and 96.5% gold dissolution after roasting/NaCN treatment. The Oxide composite indicated 27.5% silver dissolution in sodium cyanide, 33.73% silver dissolution after HCl/NaCN treatment, 35.1% silver dissolution after HCL-HNO₃/NaCN treatment, and 35.1% silver dissolution after roasting/NaCN treatment.

The Tip1 composite indicated 32.6% gold dissolution in sodium cyanide, 44.1% gold dissolution after HCl/NaCN treatment, 93.1% gold dissolution after HCL-HNO₃/NaCN treatment, and 94.2% gold dissolution after roasting/NaCN treatment. The Tip1 composite indicated 7.1% silver dissolution in sodium cyanide, 17.1 % silver dissolution after HCl/NaCN treatment, 85.1% silver dissolution after HCL-HNO₃/NaCN treatment, and 94.3% silver dissolution after roasting/NaCN treatment.

The Tjvs composite indicated 44.0% gold dissolution in sodium cyanide, 62.9% gold dissolution after HCl/NaCN treatment, 85.1% gold dissolution after HCL-HNO₃/NaCN treatment, and 94.2% gold dissolution after roasting/NaCN treatment. The Tjvs composite indicated 12.0% silver dissolution in sodium cyanide, 22.9% silver dissolution after HCl/NaCN treatment, 73.2% silver dissolution after HCL-HNO₃/NaCN treatment, and 95.6% silver dissolution after roasting/NaCN treatment.

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Table 13-4 Diagnostic Leach Results

 

13.1.2.2 Minerology

BV Minerals - Metallurgical Division, Bureau Veritas Commodities Canada Ltd., characterized the gold and silver deportment mineralogy. The analysis identified gold and silver minerals, gold and silver deportments by the bearing minerals, as well as the liberation and associations of gold and silver with sulphide and non-sulphide gangue minerals. The fragmentation characteristics of the dominant sulphide and non-sulphide minerals were also addressed.

13.1.2.2.1 Gold Deportment

The two composite samples, labeled as Tip1 and Tjvs, assayed about 2.0 grams per tonne gold. QEMSCAN Trace Mineral Search (TMS) on the unsorted samples as-received identified only a few gold grains. A Knelson gravity concentrations was made from the two composite samples to observe more gold occurrences.

QEMSCAN Trace Mineral Search on Knelson concentrated samples indicated that 80 to 85 percent of the gold in the two composites were present as either native gold or gold electrum. The remainder of the composite gold were mainly contained in fischesserite (Ag₃AuSe₂) and petzite (Ag₃AuTe₂) and some acanthite/argentite (Ag₂S) in the two composites carried trace amounts of gold.

QEMSCAN TMS gold search, a total of 21 gold particles were observed in the two composites. The majority of the gold grains were sized finer than 2.0 microns, and over half were sized finer than 1.0 microns in circular diameter. The averaged circular diameters of the two composites were measured at 2.3 microns and 0.8 microns, respectively.

The relatively coarse-grained gold, sized coarser than 20 microns, existing in the composites was also confirmed in this study. The coarsely grained gold (20 µm in circular diameter) carried about a quarter of the total gold in the composite Tip1 sample. In the Tjvs composite no gold grains greater than 2.0 microns were observed.

At the primary grind size of P₈₀ 75 µm, about a quarter of the gold in composite Tip1 was liberated when estimated in two dimensions. Liberated gold was not observed in the Tjvs composite.

The unliberated gold in the two composites was dominantly associated with non-sulphide gangue, pyrite and arsenopyrite in binary or multiphase forms. The gold-gangue binaries mostly contained less than 10 percent gold by area; these binaries probably will be the sources that cause the gold losses during a sulphide flotation.

The gold locking characteristics data indicates that 42 to 80 percent of the unliberated gold in the two composites presented exposed surfaces, either liberated or in the form of adhesions attaching to other minerals.

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13.1.2.2.2 Silver Deportment

Both composites contained about 2.0 grams per tonne silver. The silver distributions by silver bearing mineral species and by silver grain size data indicate 5 to 15 percent of the silver in the two composites was contained in gold or gold bearing minerals (fischesserite and petzite). Recovering the gold will accordingly recover part of the silver in the ore. The remainder of the composite silver was mainly carried by chlorargyrite (AgCl), stephanite (Ag₅SbS₄), acanthite/argentite (Ag₂S), and hessite (Ag₂Te). The silver in sulphide form or gold likely will be recovered during a sulphide flotation.

QEMSCAN TMS for silver, 25 silver particles (including the gold carrying silver) were detected in the two composites. All the silver grains were sized finer than 5 microns in circular diameter and 64 to 86 percent of the silver occurrences were sized finer than 1.0 micron.

The silver liberation and locking characteristics indicate the particle size of about P₈₀ 75 µm, nearly half of the composite silver presented exposed surfaces. The locked silver was dominantly associated with non-sulphide gangue and pyrite in binary form.

Difficulties in achieving a good silver recovery can be anticipated, especially when sulphide flotation or cyanidation leach will be used to process these ores.

13.1.2.2.3 Chemical and Mineral Composition

The two composites contained about 3.0 to 4.2 percent by weight sulphide minerals, dominated by pyrite and arsenopyrite and over 98 percent of the composite sulphur was contained in these iron sulphide minerals. Other observed sulphide minerals were chalcopyrite, sphalerite and galena, but all were in trace amounts. Arsenopyrite was the main arsenic-bearing mineral in the two composites. No other arsenic minerals were identified.

The sulphide minerals were hosted in silicon rich non-sulphide gangue, which mostly occurred as different types of silicates: quartz, muscovite/illite, K-feldspar and plagioclase feldspar. Both composites contained about 4.5 percent by weight carbonates and 0.3 percent by weight organic carbon.

13.1.2.2.4 Mineral Fragmentation

QEMSCAN Particle Mineral Analysis (PMA) protocols indicate at the defined particles size, one third to half of the iron sulphides (including pyrite and arsenopyrite) were liberated when estimated in two dimensions. The unliberated iron sulphides were mainly interlocked with non-sulphide gangue in binary form. The quality of iron sulphides binary particles reveals that 85 to 90 percent of the iron sulphide bearing particles contained greater than 20 percent by weight iron sulphides.

13.1.2.2.5 Mineraology Conclusions and Recommendations

The two composites, Tip1 and Tjvs, contained 3.0 to 4.2 percent by weight sulphide minerals and just above 2.0 grams per tonne gold. The dominant sulphide minerals in the two composites were pyrite and arsenopyrite. About 80 to 85 percent of the composite gold was present as native gold and electrum. The remainder of the gold was distributed between fischesserite and petzite. It is of importance to note that majority of the detected gold occurrences were sized finer than 2 microns in circular diameter.

At the particle size of P₈₀ 75 µm, the two-dimensional liberation of gold in the composite Tip1 was measured at 26 percent. No liberated gold was detected in the composite Tjvs. The gold locking characteristics indicate that about 42 to 80 percent of the gold in these two composites presented exposed surfaces, occurring as either liberated gold grains or gold adhesions. The locked gold was mostly fine-grained and associated with iron sulphides (pyrite and arsenopyrite) and non-sulphide gangue.

The mineral fragmentation characteristics data suggest that a sulphide concentrate, assaying 10 to 15 percent sulphur at an over 85 percent sulphur recovery, can be theoretically achieved at the current primary grind size. The gold locked with non-sulphide gangue probably will be the sources that cause the gold losses during sulphide flotation.

13.1.2.3 Mother Lode Comminution Tests

Comminution parameters were obtained from ½ core composites for the sulphide materials.

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Table 13-5 Comminution Parameters

 

^* Samples for the crusher work index, CWi testing (1/2 HQ Core) do not meet specifications for the standard Bond test, but may provide a general estimate of CWi.

13.1.2.4 Gravity Concentration Tests

Gravity concentrates were obtained by processing material through a Knelson concentrator and upgrading the concentrate on a table.

Tip1 material ground to P₈₀ 75 µm produced a Knelson rougher concentrate containing 8.4% of the gold and 1.9% of the silver in 0.93% of the mass. Gold and silver grades were 19.4 g/t and 3.0 g/t, respectively. Cleaning the rougher concentrate produced a cleaner table concentrate containing 0.33% the mass and recovering 4.2% of the gold and 0.5% of the silver. Gold and silver grades in the cleaner concentrate were 27.1 g/t and <3 g/t, respectively. Cleaner table tails recovered 4.2 of the gold and 1.4% of the silver in 0.6% of the mass. Gold and silver grades were 15.1 g/t and <3.0 g/t, respectively.

Tjvs material ground to P₈₀ 75 µm produced a Knelson rougher concentrate containing 9.1% of the gold and 1.3% of the silver in 0.85% of the mass. Gold and silver grades were 25.7 g/t and 3.0 g/t, respectively. Cleaning the rougher concentrate produced a cleaner table concentrate containing 0.28% the mass and recovering 4.2% of the gold and 0.4% of the silver. Gold and silver grades in the cleaner concentrate were 36.1 g/t and <3 g/t, respectively. Cleaner table tails recovered 4.9 of the gold and 0.9% of the silver in 0.85% of the mass. Gold and silver grades were 20.6 g/t and <3.0 g/t, respectively.

13.1.2.5 Flotation tests

Whole sample rougher flotation testing was conducted on the two sulfide composites. Parameters evaluated included feed size, reagent schemes, solids density and desliming before flotation. Bulk, 14 kg, flotation tests were also conducted to generate concentrate for ultra-fine grinding/cyanidation testing and concentrate oxidation tests. Flotation results are presented in Table 13-6.

Tip1 composite at a P₈₀ 75 µm grind recovered 86.9% gold in 26.4% of the weight at a gold grade of 7.74 g/t. Sulphur recovery was 92.6% at a grade of 5.9 wt%.

Tjvs composite at a P₈₀ 75 µm grind recovered 82.3% gold in 26.4% of the weight at a gold grade of 8.12 g/t. Sulphur recovery was 85.4% at a grade of 2.9 wt%.

The reagent utilized were 0.25-0.3 kg/t sodium metasilicate, 0.025 kg/t PAX, 0.05 kg/t Aero 208, and 0.005 kg/t Aerofroth 65.

Flotation tests with inert gas are underway, results are pending.

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Table 13-6 Flotation Tests

 

 

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13.1.2.6 Rougher Concentrate regrind and cyanidation

Gold in rougher floatation concentrates was not amenable to concentrated cyanide leach methods. Rougher flotation concentrates were reground to P₈₀ 6.2-7.9 µm and leached in 2 g/l or 10 g/l sodium cyanide solution for 96-hours. Concentrated gold recovery ranged from 7.7-17.4% and averaged 10.8%. Concentrate silver recovery ranged from 3.3-42.4% and averaged 19.8%. Sodium cyanide consumption ranged from 3.7-15.7 kg/ton concentrate and averaged 8.7 kg/ton concentrate. Lime consumption ranged from 4.9 to 7.7 kg/ton concentrate and averaged 5.8 kg/ton concentrate. Table 13-7 presents flotation concentrate cyanide leach tests results.

Table 13-7 Flotation Concentrate Regrind and Cyanidation

 

13.1.2.7 Gravity/flotaion/cyanidation tests -

Tip1 gravity tests recovered 4.2% gold in 0.33% of the mass at a grade of 27.1 g/t. Cleaner tails contained 0.6% of the mass and 4.9% of the gold at a gold grade of 15.1 g/t. Rougher flotation test results on recombined gravity cleaner and rougher tails indicated a mass pull of 14.7% would recover 68.6% of the gold at a grade of 3.37 g/t. Rougher tails contained 31.4% of the gold in 85.3% of the mass at a grade of 0.5 g/t. Cyanidation of rougher flotation tailings indicated an average of 10.4% gold dissolution in three different flotation tailings tests. POX cyanidation results indicate 96.2% gold dissolution after pressure oxidation. The combined gold recovery for a gravity/concentrate – flotation – POX/cyanidation - rougher tail/cyanidation circuit, was calculated to be 70.9%.

Tjvs gravity tests recovered 4.2% gold in 0.28% of the mass at a grade of 36.1 g/t. Cleaner tails contained 0.57% of the mass and 4.9% of the gold at a gold grade of 20.6 g/t. Rougher flotation test results on recombined gravity cleaner and rougher tails indicated a mass pull of 15.5% would recover 73.2% of the gold at a grade of 5.4 g/t. Rougher tails contained 26.8% of the gold in 84.45% of the mass at a grade of 0.76 g/t. Cyanidation of rougher flotation tailings indicated an average of 1.0% gold dissolution in three different flotation tailings tests. POX cyanidation results indicate 95.1% gold dissolution after pressure oxidation. The combined gold recovery for a gravity/concentrate – flotation – POX/cyanidation - rougher tail/cyanidation circuit, was calculated to be 72.4%.

The tests suggest gravity concentration before flotation, and cyanidation of rougher flotation tails may not be beneficial.

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13.1.2.8 Concentrate Roastng and Cyanidation Tests

Concentrate roasting test were conducted at Hazen on both concentrates at temperature ranges from 500^o C to 700^o C in single stage roasting. Cyanidation after roasting indicated gold dissolution decreased from 81-84% to 55-62% as the roasting temperature increased from 500^o C to 700^o C for both concentrates. In two-state roasting tests, conducted at 1^st stage/2^nd stage temperatures of 550/650^o C and 600/650^o C, gold dissolution decreased from 84% to 78% as the temperature increased, for both concentrates.

13.1.2.9 Pressure Oxidation Cyanidation Tests

Tip1 flotation concentrates, “as-received” and reground, were oxidized in an autoclave at 225^o C with 690 kPa Oxygen overpressure (100 psi) for 1 hour. The oxidized slurries were filtered, washed, and filter solids leached in sodium cyanide with activated carbon, reference Table 13-8.

Tip1 concentrate was pre-acidified to pH 2.0 with sulfuric acid. Sulphuric acid consumption ranged from 71 kg H₂SO₄/ton POX feed to 111 kg H₂SO₄/ton POX feed for “as-received” and reground concentrates, respectively. Pressure oxidation residue was cyanide leached with activated carbon in 30% solids at pH 11 with 1 g/l sodium cyanide for 48 hours. Gold extraction on unground concentrate was mixed in two tests, 30.6% and 96.3%. Gold recovery on reground concentrate averaged 96.2%. Average sodium cyanide and lime consumptions were 0.5 to 0.7 kg/ton grinding mill feed and 2 to 30 kg/ton grinding mill feed. The high lime consumption suggests a hot cure step will be required in a commercial process.

Tjvs flotation concentrate was pre-acidified to pH 2.0 with sulfuric acid. Sulphuric acid consumption ranged from 79 kg H₂SO₄/ton POX feed to 87 kg H₂SO₄/ton POX feed for “as-received” and reground concentrates, respectively.

Gold extraction on unground concentrate was 88.6 and 93.4%. Gold recovery on reground concentrate was mixed in two tests 77.6% and 95.1%. Average sodium cyanide and lime consumptions were 0.5 to 1.1 kg/ton grinding mill feed and 2 to 30 kg/ton grinding mill feed. The high lime consumption suggests a hot cure step will be required in a commercial process.

Table 13-8 Flotation Concentrate POX/Cyanidation Tests

 

13.1.2.10 Rougher Flotation Tailings - Acid Base Accounting

Acid base accounting tests indicate the net neutralization potential of Tip1 flotation tailings is 58.8 tons of CaCO₃/1000 tons of tailings and Tjvs is 45.4 tons of CaCO₃/ 1000 tons of rougher tailings, reference Table 13-9. Flotation tailings will be utilized to neutralize pressure oxidation wash solution instead of limestone or site dolomite.

Table 13-9 Rougher Tailings Acid Neutralization Potential

 

13.1.2.11 Mother Load Gold Dissolution and Reagent Consumption

13.1.2.11.1 Mother lode - oxide gold dissoltuion

Mother Lode oxide samples tested show high gold dissolutions and low reagent consumptions. In the absence of percolation column test work, the ROM extractions are based on historical oxide recovery information from the Daisy mine materials that indicated actual ROM gold recovery in the high 70’s and dissolution curve analysis of production data that suggest gold dissolution in the low 80’s. Mother Lode ROM recovery for this study is assumed to be 74%. Field adjusted sodium cyanide and lime consumptions are estimated to be 0.05 kg/t and 2 kg/t, respectively. It is recommended column test work be completed at the ROM size to firmly establish gold dissolution.

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13.1.2.11.2 Mother lode Flotation/POX gold dissoltuion

Tip1 flotation gold recovery of 86.9% and POX oxidized concentrate cyanidation gold recovery of 96.3% give a weighted gold dissolution of 83.7%. Field adjusted sodium cyanide and lime consumptions are estimated to be 0.2 kg/t and 25 kg/t.

Tjvs flotation gold recovery of 82.3% and POX oxidized concentrate cyanidation gold recovery of 92.4% give a weighted gold dissolution of 76.1%. Field adjusted sodium cyanide and lime consumptions are estimated to be 0.2 kg/t and 25 kg/t.

Weighted gold dissolution for Tip1 and Tjvs material concentrated by flotation, pressure oxidized, and treated in a conventional CIL circuit is estimated to be 80%. Sodium cyanide and lime consumptions are estimated to be 0.2 kg/t and 25 kg/t.

13.2 North Bullfrog Metallurgical Testing

Mining activities at NBP took place in the period 1909 – 1926, however, there are no metallurgical data available for that period. Data from subsequent exploration drilling at NBP between 1972 and 1996 by the various organizations listed in Table 6-1 do not contain any records of metallurgical test work. All of the known metallurgical test work on NBP mineralization have been produced by Corvus beginning in 2010.

This section presents a summary the metallurgical testing performed from 2008-2017 and available for inclusion in this Technical Report.

The majority of the metallurgical program has been completed by McClelland. McClelland has performed bottle roll testing on RC cuttings and core sample materials from the Sierra Blanca, the Jolly Jane, the Savage Valley and the Mayflower areas and column leach tests at P₈₀ of -12.5 mm (1/2 inch), -19 mm (3/4 inch) and -51 mm (2 inch) for the Sierra Blanca, Savage Valley, Jolly Jane and Mayflower areas, bottle roll testing on these column leach composites at nominal P₈₀ -75 µm (-200 mesh) and 1.7 mm (-10 mesh), 6.3 mm (-1/4 inch) and 19 mm (-3/4 inch), bottle roll test results for vein and stockwork materials representative of the YellowJacket mineralization at various feed sizes and gravity concentration with cyanide leaching of tail materials on YellowJacket vein and vein stockwork mineralization.

The first section summarizes bottle roll and column tests by deposit for the Connection, Jolly Jane (JJ), Mayflower (MF), Savage Valley, (SV), Sierra Blanca (SB), and preliminary bottle roll tests on Yellow Jacket The test programs were completed 2008-2013 at the laboratories referenced in Table 13-10. The second section details metallurgical development of the the higher-grade Yellow Jacket oxide material. The third section details development of the Yellow Jacket sulphide material, and the final section summarizes the NBP metallurgy.

The gold mineralization at the NBP contains various amounts of silver. Silver ratios in the Sierra Blanca mineralization average 3.5 silver to 1.0 gold, the Jolly Jane mineralization averages 1.6 silver to 1.0 gold and the Mayflower mineralization averages 0.64 silver to 1.0 gold. YellowJacket mineralization averages approximately 5 silver to 1 gold. Bottle roll and column leach recoveries of silver are also reported in the data. Column Silver recoveries ranged from 3% to 16%. Discussion of the silver data contained in this Technical Report is limited to column leach testing of the disseminated low-grade mineralization and to the YellowJacket data, where relatively high silver grades and recovery is indicated by the data.

13.2.1 Metallurgical Testing – 2008 -2013

13.2.1.1 Connection deposit

Connection is a relatively small portion of the mineralization at the NBP. KCA completed bottle roll tests. Gold recovery was 48% and 91% in the two samples. Silver recovery was high at 91% and 73%, respectively. The submitted samples had relatively high Au grades (5.9 and 4.2 g/t). Sodium cyanide and lime consumptions averaged 0.2 kg NaCN /ton and 1.3 kg CaO/ton.

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Table 13-10 2008-2013 Laboratory References

No.	Date	Reference	Tests Conducted
1	2008, 14 July	Hazen Research, Inc.	Hazen bottle roll on six samples
2	2010, March	KCA	Bottle roll tests 11 bags of core, 72 hrs.
3	2011, April	KCA	Reverse Circulation cuttings, 24 samples, 3mm as-received, 72 hrs. bottle rolls at 75 um.
4	2011, May	KCA	Connection 2 bottle rolls on RC cuttings,72 hrs., 75 um.
5	2012 Dec 28	McClelland Laboratories, Inc.	MF, SV, JJ, SB, 157 core, 24 composites, Bottle Rolls 19 mm, 6.3 mm, 1.7 mm, 75 um, and 23 column leach
6	2012, 14 Feb.	McClelland Laboratories, Inc.	Two bulk samples surface SB and JJ at 1.7 mm and 75um and Bulk Columns 50 mm and 12.5 mm.
7	2013, 15 May	McClelland Laboratories, Inc.	JJ, 5 composites, duplicate 19 mm Column Leach Tests, 71-161 days, 19 mm, 62-86% Au
8	2013, 29 April	McClelland Laboratories, Inc.	2 bulk samples Main Dump, and David Adit, comminution, Bottle Roll, 19 mm, 6.3 mm, 1.7 mm, 75 um
9	2013, 29 Jan	McClelland Laboratories, Inc.	12 SV Column Leach Tests, 63-136 days, 19 mm, 71-92% Au
10	2013, 29 May	McClelland Laboratories, Inc.	Sierra Blanca, 14 columns leach tests, 19 mm, 72-133 days, 63-97% Au

13.2.1.2 Jolly Jane Bottle Roll and Column Tests

Jolly Jane deposit gold dissolution has been developed from 54 bottle roll tests and 14 column tests on drill core, RC drill cuttings and bulk surface gab samples. Test results are presented in Table 13-11.

Bottle roll results indicate gold dissolution on Jolly Jane oxide material increases from an average of 68.5% to 87% as the P₈₀ decreases from 19000 to 75 µm. Sodium cyanide consumption ranges 0.1-0.2 kg/ton. Lime consumption ranges from 0.8-1.8 kg/ton.

Bottle roll test result indicate gold dissolution on Jolly Jane Oxide material, using bulk sampling, increases from an average of 74.4% to 80.0% as the P₈₀ decreases from 1700 to 75 µm. Sodium cyanide consumption ranges 0.1-0.2 kg/ton. Lime consumption ranges from 0.8-1.8 kg/ton.

Bottle roll test result indicate gold dissolution on Jolly Jane surface material using RC cuttings increases from an average of 66.2% to 78.8% as the P₈₀ decreases from 1700 to 75 µm. Sodium cyanide consumption ranges 0.1-0.2 kg/ton. Lime consumption averages 1.8 kg/ton.

Column leach test results on Jolly Jane core material indicate gold dissolution at a P₈₀ 19000 µm average 74.8%±7.3% and ranged from 62.3 to 85.9. Sodium cyanide consumption averaged 0.8 kg/ton. Lime consumption averaged 0.8 kg/ton. Leach time ranged from 70 to 161 days and averaged 107 days.

Column leach test results on Jolly Jane surface outcrop indicate gold dissolution at a P₈₀ 50000 µm was 64.0%. Sodium cyanide consumption was 0.9 kg/ton. Lime consumption was 1.6 kg/ton and leach time was 117 days.

Column leach test results on Jolly Jane surface outcrop indicate gold dissolution at a P₈₀ 12500 µm was 63.8%. Sodium cyanide consumption averaged 1.0 kg/ton. Lime consumption averaged 1.6 kg/ton and leach time averaged 104 days.

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Table 13-11 Jolly Jane Bottle Roll and Column Tests 2008-2013

 

13.2.1.3 Mayflower Bottle Roll and Column Tests

Mayflower deposit gold dissolution has been developed from 74 bottle roll tests and 10 column tests on drill core, RC drill cuttings and David adit and main dump surface gab samples. Test results are presented in Table 13-12.

Bottle roll results indicate gold dissolution on Main Dump grab material increases from an average of 37.7% to 98.9% as the P₈₀ decreases from 38000 to 75 µm. Sodium cyanide consumption ranges 0.1-0.2 kg/ton. Lime consumption averages 1.0 kg/ton.

Bottle roll result indicate gold dissolution on David Adit grab material increases from an average of 34.3% to 97.9% as the P₈₀ decreases from 38000 to 75 µm. Sodium cyanide consumption averaged 0.2 kg/ton. Lime consumption averages 1.5 kg/ton.

Column leach test results on Mayflower core material indicate gold dissolution at a P₈₀ 19000 µm average 88.0%±1.6% and ranged from 85.7% to 90.8%. Sodium cyanide consumption averaged 1.2 kg/ton. Lime consumption averaged 1.3 kg/ton. Leach time ranged from 90 to 156 days and averaged 117 days.

Table 13-12 Mayflower Bottle Roll and Column Tests 2008-2013

 

13.2.1.4 Savage Valley bottle Roll and Column Tests

Savage Valley deposit gold dissolution has been developed from 63 bottle roll tests and 12 column tests on drill core, RC drill cuttings for oxide, mixed, and sulphide oxidation materials. Test results are presented in Table 13-13.

Bottle roll test results indicate gold dissolution on oxide material increases from an average of 71.8% to 89.1% as the P₈₀ decreases from 19000 to 75 µm. Sodium cyanide consumption ranges 0.1-0.2 kg/ton. Lime consumption averages 1.2 kg/ton.

Bottle roll test results indicate gold dissolution on mixed oxide material decreases from an average of 70.1% to 64.7% as the P₈₀ decreases from 1700 to 75 µm. Sodium cyanide consumption was 0.1 kg/ton. Lime consumption averaged 1.8 kg/ton.

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Bottle roll results indicate gold dissolution on sulfide material averaged 4% at 75 µm. Sodium cyanide consumption was 0.8 kg/ton. Lime consumption averaged 0.1 kg/ton.

Column leach test results on Savage Valley core material indicate gold dissolution at a P₈₀ 19000 µm average 81.7%±8.4% and ranged from 68.7% to 92.2%. Sodium cyanide consumption averaged 0.7 kg/ton. Lime consumption averaged 0.9 kg/ton. Leach time ranged from 63 to 136 days and averaged 87 days.

Table 13-13 Savage Valley Bottle Roll and Column Tests 2008-2013

 

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13.2.1.5 Sierra Blanca bottle Roll and Column Tests

Sierra Blanca deposit gold dissolution has been developed from 91 bottle roll tests and 18 column tests on drill core, RC drill cuttings for oxide, mixed, and surface bulk grab materials. Test results are presented in Table 13-14.

Bottle roll test results indicate gold dissolution on bulk surface oxide material increases from an average of 83.4% to 85.9% as the P₈₀ decreases from 1700 to 75 µm. Sodium cyanide consumption ranges 0.1-0.2 kg/ton. Lime consumption averages 1.5 kg/ton.

Bottle roll test results indicate gold dissolution on mixed oxide material ranges from 69.6% to 73.1% as the P₈₀ decreases from 1700 to 75 µm. Sodium cyanide consumption was 0.2 kg/ton. Lime consumption averaged 1.8 kg/ton.

Bottle roll result indicate gold dissolution on Sierra Blanca oxide material decreases from 76.1% to 70.5% as the P₈₀ decreases from 19000 to 75 µm. Sodium cyanide consumption averages 0.1 kg/ton. Lime consumption averages 1.1 kg/ton.

Column leach test results on Savage Valley core material indicate gold dissolution at a P₈₀ 19000 µm average 81.7%±8.4% and ranged from 68.7% to 92.2%. Sodium cyanide consumption averaged 0.7 kg/ton. Lime consumption averaged 0.9 kg/ton. Leach time ranged from 63 to 136 days and averaged 87 days.

Table 13-14 Sierra Blanca Bottle Roll and Column Tests 2008-2013

 

13.2.1.6 Initial YellowJacket Sulfide Bottle Rolls

Four sulphide bottle roll test were initially conducted on YellowJacket RC material reference Table 13-15.

Bottle roll test results indicate gold dissolution, at P₈₀ 75 µm with six days of leaching, was 86.0%. Sodium cyanide consumption was 0.6 kg/ton.

Bottle roll test results indicate gold dissolution, at P₈₀ 75 µm with three days of leaching, averaged 16.7%. Sodium cyanide consumption was 0.16 kg/ton.

Table 13-15 YellowJacket - Initial Sulphide Bottle Rolls

 

13.2.1.7 McClelland YellowJacket Metallurgical Testing – 2013-2015

The YellowJacket deposit is a steeply dipping quartz vein, with proximal stockwork, on the eastern edge of the Sierra Blanca resource. A series of metallurgical tests have been performed on YellowJacket mineralization by McClelland Laboratories, which consisted of CN bottle roll tests on core samples ground to 80% - 75µm, CN bottle roll tests on samples with particle sizes ranging from 80% -19 mm to 80% - 75µm , column leach tests on materials with particle sizes of 80% - 6.3 mm and 80% -19mm, E-GRG tests on composite samples of vein material and stockwork material, Gravity concentration with CN leaching of the tail material, and gravity concentration with intensive CN leaching of the gravity concentrate followed by CN leaching of the combined tails. These data were reported by McClelland (2015a) and (2015b) and are discussed in the following sub-sections.

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The two sets of composite samples, prepared to evaluate gravity concentration with CN leach of the tails, were developed for vein and vein stockwork samples from above the oxidation surface (YJ PQ composites) and from below the oxidation surface (YJ JV composites).

13.2.1.7.1 Scoping CN Leach Tests

McClelland Labs performed 24 bottle roll tests on 12 drilling core samples at nominal 80% -75µm feed sizes. Table 13-16 list the results of the bottle roll tests for both gold and silver recovery.

Table 13-16 Summary Metallurgical Results, Bottle Roll Tests, YellowJacket Drill Core Composites, 80%-75 µm Feed Size

Composite	Au Recovery. %	g Au/mt mineralization				Ag Recovery  %	g Ag/mt mineralization				Reagent Requirements  kg/mt mineralization
		Extracted	Tail (1)	Calculated Head	Head Assay (2)		Extracted	Tail (1)	Calculated Head	Head Assay (2)	NaCN Cons.	Lime Added
C226950	96.2	2.27	0.09	2.36	2.16	85.7	8.4	1.4	9.8	11	0.08	1.2
C226950	96.4	2.40	0.09	2.49	2.16	85.6	8.3	1.4	9.7	11	<0.07	1.3
C226985	91.2	11.02	1.06	12.08	12.70	67.8	28.9	13.7	42.6	49	0.09	1.1
C226985	89.9	9.32	1.05	10.37	12.70	65.1	27.4	14.7	42.1	49	0.14	1.2
C226986	78.2	3.08	0.86	3.94	3.90	61.0	26.8	17.1	43.9	49	0.12	1.6
C226986	79.9	3.06	0.77	3.83	3.90	73.0	33.2	12.3	45.5	49	0.21	1.6
C226989	68.7	1.36	0.62	1.98	1.90	69.9	14.6	6.3	20.9	25	0.23	1.1
C226989	73.1	1.41	0.52	1.93	1.90	76.4	15.2	4.7	19.9	25	0.13	1.0
C226990	85.1	6.34	1.11	7.45	6.70	70.7	22.7	9.4	32.1	35	0.20	1.0
C226990	85.6	6.10	1.03	7.13	6.70	72.8	24.3	9.1	33.4	35	0.21	0.9
M612658	85.4	0.70	0.12	0.82	0.80	75.8	2.5	0.8	3.3	3	0.08	1.3
M612658	86.9	0.73	0.11	0.84	0.80	78.8	2.6	0.7	3.3	3	0.07	1.4
M612665	86.5	8.32	1.30	9.62	12.10	80.1	34.7	8.6	43.3	45	0.11	1.1
M612665	89.3	8.64	1.04	9.68	12.10	79.2	33.8	8.9	42.7	45	0.17	1.1
M612674	91.7	0.77	0.07	0.84	1.00	88.2	44.1	5.9	50.0	61	0.12	1.1
M612674	92.9	0.78	0.06	0.84	1.00	85.7	46.2	7.7	53.9	61	0.13	1.1
M612701	96.9	5.86	0.19	6.05	7.10	89.7	35.7	4.1	39.8	49	0.22	1.1
M612701	95.7	5.53	0.25	5.78	7.10	89.1	36.0	4.4	40.4	49	0.15	1.1
M612704	89.1	1.47	0.18	1.65	1.70	83.6	43.3	8.5	51.8	56	0.26	1.1
M612704	85.9	1.40	0.23	1.63	1.70	83.5	43.9	8.7	52.6	56	0.34	1.1
M612716	94.1	1.28	0.08	1.36	1.20	74.2	2.3	0.8	3.1	3	0.13	1.1
M612716	95.1	1.37	0.07	1.44	1.20	75.0	2.4	0.8	3.2	3	0.08	1.2
M612727	90.7	0.49	0.05	0.54	0.60	72.7	0.8	0.3	1.1	1	<0.07	1.2
M612727	91.7	0.66	0.06	0.72	0.60	66.7	0.8	0.4	1.2	1	0.11	1.2
(1) Average of triplicate assays
(2) Head assays were provided by Corvus

13.2.1.7.2 YJ PQ Composite Tests

Extensive core was produced by drilling at YellowJacket during 2013 and 2014 was used to create composite samples designed to be representative of the types of mineralization represented by the Josh Vein and adjacent stockwork zones. Five composite samples (YJ PQ composites) were created from PQ core from above the oxidation surface and submitted to McClelland for testing. A series of tests were performed on the YJ PQ composite samples to identify processing alternatives. Those test series consisted of:

CN bottle roll tests at feed sizes of 80% -75µm, 80% - 0.106mm, 80% - 0.150mm, 80% - 1.7mm, 80% - 6.3mm and 80% - 19mm; CN Column leach tests at 80% - 6.3mm and 80% - 19mm; E-GRG tests at 80% -75µm, 80% - 250µm and 80% - 700µm; Gravity concentration at feed size of 80% -212µm with CN leach of tails at 80% -75µm and 80% -150µm. Results from each of the tests types is presented in the following sub-sections.

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13.2.1.7.2.1 CN Bottle Roll Tests at Various Feed Sizes

The results of bottle roll tests at particle size gradations of 80% -75µm, 80% - 0.106mm, 80% - 0.150mm, 80% - 1.7mm, 80% - 6.3mm and 80% - 19mm are listed in Table 13-17. The tests were performed to test the potential for heap leach processing of YellowJacket mineralization. The results indicated achieving high gold and silver CN recovery would require a mill to grind the mineralization and that good leach recoveries could be achieved with feed sizes in the range of 80% -0.15mm to -75μ.

Table 13-17 Summary of Bottle Roll Tests, YellowJacket YJ PQ Drill Core Composites.

Composite	Feed Size (P80)	Au Recovery, %	g Au/mt mineralization				Ag Recovery, %	g Ag/mt mineralization				Reagent Requirements,  kg/mt mineralization
			Extracted	Tail (1)	Calculated Head	Head (2)		Extracted	Tail (1)	Calculated Head	Head (2)	NaCN Cons.	Lime Added
YJPQ01	19mm	12.7	0.91	6.23	7.14	5.39	10.5	4.5	38.4	42.9	47.5	<0.09	0.5
YJPQ01	6.3mm	30.4	2.70	6.19	8.89	5.39	24.9	13.3	40.1	53.4	47.5	0.12	0.7
YJPQ01	1.7mm	50.0	3.88	3.88	7.76	5.39	46.9	22.5	25.5	48.0	47.5	0.17	0.7
YJPQ01	0.150mm	86.2	4.88	0.78	5.66	5.39	66.8	37.3	18.5	55.8	47.5	<0.12	1.0
YJPQ01	0.106mm	88.4	4.44	0.58	5.02	5.39	68.7	37.1	16.9	54.0	47.5	<0.12	1.0
YJPQ01	0.075mm	88.3	5.00	0.66	5.66	5.39	70.3	32.4	13.7	46.1	47.5	0.17	1.2
YJPQ02	19mm	11.3	0.60	4.69	5.29	9.74	15.0	3.4	19.2	22.6	25.0	0.11	0.7
YJPQ02	6.3mm	32.9	1.62	3.30	4.92	9.74	35.4	7.4	13.5	20.9	25.0	<0.11	0.9
YJPQ02	1.7mm	53.6	3.53	3.06	6.59	9.74	50.2	12.0	11.9	23.9	25.0	0.12	0.9
YJPQ02	0.150mm	92.0	5.29	0.46	5.75	9.74	74.3	18.2	6.3	24.5	25.0	0.10	1.3
YJPQ02	0.106mm	92.9	5.35	0.41	5.76	9.74	75.8	19.1	6.1	25.2	25.0	<0.07	1.4
YJPQ02	0.075mm	90.4	5.64	0.60	6.24	9.74	77.4	19.9	5.8	25.7	25.0	0.18	1.2
YJPQ03	19mm	13.7	0.27	1.70	1.97	1.49	11.4	0.9	7.0	7.9	9.1	0.09	0.6
YJPQ03	6.3mm	27.3	0.59	1.57	2.16	1.49	20.5	1.6	6.2	7.8	9.1	<0.07	0.8
YJPQ03	1.7mm	45.9	0.73	0.86	1.59	1.49	38.4	3.3	5.3	8.6	9.1	0.12	1.1
YJPQ03	0.150mm	88.5	1.62	0.21	1.83	1.49	68.4	5.4	2.5	7.9	9.1	<0.07	1.2
YJPQ03	0.106mm	88.3	1.58	0.21	1.79	1.49	65.9	5.6	2.9	8.5	9.1	<0.07	1.3
YJPQ03	0.075mm	89.6	1.29	0.15	1.44	1.49	76.1	7.0	2.2	9.2	9.1	0.20	1.3
YJPQ04	19mm	23.1	0.09	0.30	0.39	0.62	16.7	0.8	4.0	4.8	4.6	0.07	0.8
YJPQ04	6.3mm	33.3	0.18	0.36	0.54	0.62	41.9	1.8	2.5	4.3	4.6	<0.07	1.1
YJPQ04	1.7mm	57.8	0.37	0.27	0.64	0.62	52.3	2.3	2.1	4.4	4.6	<0.09	1.2
YJPQ04	0.150mm	61.4	0.35	0.22	0.57	0.62	62.0	3.1	1.9	5.0	4.6	0.11	1.5
YJPQ04	0.106mm	67.3	0.37	0.18	0.55	0.62	66.0	3.1	1.6	4.7	4.6	<0.07	1.6
YJPQ04	0.075mm	68.8	0.44	0.20	0.64	0.62	66.7	3.4	1.7	5.1	4.6	<0.08	2.5
YJPQ05	19mm	43.9	0.18	0.23	0.41	0.33	29.4	0.5	1.2	1.7	1.9	<0.07	1.0
YJPQ05	6.3mm	44.2	0.19	0.24	0.43	0.33	36.8	0.7	1.2	1.9	1.9	0.07	1.2
YJPQ05	1.7mm	54.0	0.27	0.23	0.50	0.33	47.4	0.9	1.0	1.9	1.9	0.16	1.3
YJPQ05	0.150mm	76.4	0.42	0.13	0.55	0.33	65.0	1.3	0.7	2.0	1.9	<0.10	1.7
YJPQ05	0.106mm	76.2	0.32	0.10	0.42	0.33	70.0	1.4	0.6	2.0	1.9	0.10	1.8
YJPQ05	0.075mm	74.4	0.32	0.11	0.43	0.33	61.9	1.3	0.8	2.1	1.9	<0.08	2.3

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13.2.1.7.2.2 YJ PQ Column Leach Tests

Although the bottle roll tests on the YJ PQ composites indicated that the mineralization would have to be milled to produce sufficient gold liberation to assure high recoveries, some column leach tests were performed for confirmation. Tests were performed for the relatively small particle size of P80 -6.3mm, with one sample tested at P80 -19mm. The results are listed in Table 13-18and Table 13-19 for gold and silver recoveries, respectively.

Table 13-18 Summary Metallurgical Results, Gold Recovery from Column Percolation Leach Tests, YellowJacket YJ PQ Drill Core Composites (80% -6.3mm and 80% -19mm)

Composite/Feed Size, (P80)	Leach/Rinse Time (days)	Au Recovery %	g Au/mt			Reagent Requirements kg/mt
			Extracted	Tail	Calculated Head	NaCN Cons.	Lime Added
YJPQ01 -6.3mm	181	53.6	5.0	4.36	9.36	7.06	0.8
YJPQ02 -6.3mm	181	60.2	2.21	1.46	3.66	5.62	1.0
YJPQ03 -6.3mm	137	37.8	0.65	1.12	1.77	3.14	0.9
YJPQ04 -6.3mm	137	50.1	0.33	0.32	0.65	2.92	1.3
YJPQ05 -19mm	140	46.5	0.17	0.19	0.36	1.46	1.2
YJPQ05 -6.3mm	137	56.3	0.23	0.18	0.40	2.33	1.4

Table 13-19 Summary Metallurgical Results, Silver Recovery from Column Percolation Leach Tests, YellowJacket YJ PQ Drill Core Composites (80% -6.3mm and 80% -19mm)

Composite/Feed Size, (P80)	Leach/Rinse Time (days)	Au Recovery %	g Au/mt			Reagent Requirements kg/mt
			Extracted	Tail	Calculated Head	NaCN Cons.	Lime Added
YJPQ01 -6.3mm	181	46.2	24.2	28.2	52.4	7.06	0.8
YJPQ02 -6.3mm	181	47.7	10.5	11.6	22.1	5.62	1.0
YJPQ03 -6.3mm	137	34.0	2.4	4.7	7.0	3.14	0.9
YJPQ04 -6.3mm	137	48.4	2.4	2.6	4.9	2.92	1.3
YJPQ05 -19mm	140	38.2	0.8	1.3	2.1	1.46	1.2
YJPQ05 -6.3mm	137	48.7	0.9	1.0	1.9	2.33	1.4

13.2.1.7.2.3 YJ PQ E-GRG Tests

Extended gravity recoverable gold (“E-GRG”) tests were performed on the YJ PQ composite samples to evaluate the grind size requirements to achieve good gravity recovery. This particular test is used as a basis for modeling performance of Knelson^TM concentrators in mill circuits for prediction of recovery performance. The gravity recoverable gold component is measured at 3 progressively finer grind sizes, P80 -700μm, -250μm and -75μm. The E-GRG tests results for the YJ PQ composites are listed in Table 13-20 and Table 13-21 for gold and silver, respectively. Insufficient material was available for the YJPQ05 sample for gravity testing.

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Table 13-20 E-GRG Test Results for Gold Recovery from the YJ PQ composites

Composite	Recovery, % of Total Au Nominal Grind Size					Head Grade g Au/mt mineralization
	700µm	250µm	75µm		Total	Calculated	Average
YJPQ01	33.3	23.2	9.4	65.9		8.13	7.56
YJPQ02	34.5	33.9	11.4	79.8		5.46	5.66
YJPQ03	18.0	17.6	11.1	46.7		1.39	1.66
YJPQ04	12.3	17.8	12.8	42.9		0.74	0.60

Table 13-21 E-GRG Test Results for Silver Recovery from the YJ PQ composites

Composite	Recovery, % of Total Ag Nominal Grind Size					Head Grade g Ag/mt mineralization
	700µm	250µm	75µm		Total	Calculated	Average
YJPQ01	11.5	12.2	2.8	26.5		45.9	49.3
YJPQ02	7.0	8.1	4.1	19.2		19.1	22.9
YJPQ03	2.8	3.0	2.0	7.8		7.6	8.0
YJPQ04	4.5	4.6	3.1	12.2		4.3	4.8

13.2.1.7.2.4 YJ PQ Gravity Concentration with CN Tail Leach tests

Combined gravity concentration with CN leaching of the gravity tail products was conducted on the YJ PQ composites. Based on the results of the E-GRG tests, the composite materials were ground to P80 -212μ (-65 mesh) before being fed into the Knelson^TM concentrator. The gravity gold recovery in the cleaner concentrate (Cl. Conc.) and the CN leach recovery from the gravity tail component are listed in Table 13-22 and Table 13-23 for gold and silver, respectively. The gravity concentrates from the YJ PQ composites were assayed to destruction and therefore the combined recovery assumes that all of the metal in the gravity concentrate is recovered. This is consistent with typical process results where typically have 98% of the gold in the gravity concentrate is recovered in the refinery. Also, since the YJPQ composites were from above the oxidation surface, no gold lockup in sulphide remnants would be expected. The results indicated very high metal recoveries would be possible with a simple gravity with CN leaching plant.

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Table 13-22 Gold Recovered from Gravity Concentrate and CN Leach of Gravity Tail, YJ PQ Composites

Composite/ Feed Size (P80)	Au Distribution, % of total				g Au/mt mineralization					Reagent	Req.
	Au rec. from Cl. Conc.	Au rec. from gravity tail	Comb. Au rec.	Au in Tail	Extracted			Tail	Calc’d Head	NaCN Cons. (kg/mt)	Lime Added (kg/mt)
					Cl. Conc	CN	Comb.
YJPQ01 150µm 75µm	50.7 58.3	43.0 37.3	93.7 95.6	6.3 4.4	4.67 4.67	3.96 3.01	8.63 7.68	0.59 0.36	9.22 8.03	0.26 0.15	2.1 1.4
YJPQ02 150µm 75µm	56.3 64.2	38.6 31.4	94.8 65.6	5.2 4.4	4.34 3.34	3.02 1.63	7.36 4.97	0.40 0.23	7.76 5.20	0.14 0.17	1.7 1.6
YJPQ03 150µm 75µm	25.7 25.4	62.6 67.9	88.2 93.3	11.8 6.7	0.36 0.36	0.88 0.96	1.24 1.32	0.17 0.10	1.40 1.42	0.10 0.13	1.4 1.5
YJPQ04 150µm 75µm	23.6 24.2	49.2 49.9	72.8 74.1	27.2 25.9	0.14 01.4	0.29 0.29	0.43 0.43	0.16 0.15	0.59 0.58	0.15 0.12	1.4 2.0

Table 13-23 Silver Recovered from Gravity Concentrate and CN Leach of Gravity Tail, YJ PQ Composites

Composite/ Feed Size (P80)	Ag Distribution, % of total				g Ag/mt mineralization					Reagent	Req.
	Ag rec. from Cl. Conc.	Ag rec. for gravity tail	Comb. Ag rec.	Ag in Tail	Extracted			Tail	Calc’d Head	NaCN Cons. (kg/mt)	Lime Added (kg/mt)
					Cl. Conc	CN	Comb.
YJPQ01 150µm 75µm	11.7 11.7	63.4 66.5	75.0 78.2	25.0 21.8	5.5 5.5	30.1 31.6	35.6 37.1	11.9 10.4	47.4 47.5	0.26 0.15	2.1 1.4
YJPQ02 150µm 75µm	5.8 5.9	71.2 75.0	76.9 80.8	23.1 19.2	1.2 1.2	15.0 15.5	16.2 16.7	4.9 4.0	21.0 20.6	0.14 0.17	1.7 1.6
YJPQ03 150µm 75µm	2.3 2.2	68.9 74.9	71.1 77.1	28.9 22.9	0.2 0.2	5.6 6.2	5.8 6.4	2.4 1.9	8.2 8.3	0.10 0.13	1.4 1.5
YJPQ04 150µm 75µm	1.9 1.8	62.3 66.5	64.2 68.2	35.8 31.8	0.1 0.1	3.1 3.4	3.2 3.5	1.8 1.6	4.9 5.1	0.15 0.12	1.4 2.0

13.2.1.7.3 YJ JV Samples

PQ3 core drilling during 2014 sampled the Josh vein and vein stockwork mineralization at YellowJacket below the oxidation surface. Seven samples (PQ JV samples) were developed to represent variations of the Josh Vein and stockwork below the oxidation surface where sulphide minerals remained un-oxidized. The total proportion of sulphide sulphur below oxidation was a minor component of the total rock mass, ranging between 1-2%, so any impacts were expected to be relatively small.

13.2.1.7.3.1 YJ JV Bottle Roll Tests

The YJ JV samples were crushed and a split from each was created for blending into two composites for bottle roll testing. Samples YJ JV01 and YJ JV02 were quartz vein and strongly veined stockwork and were blend to produce the Josh Vein plus Stockwork (JV+Stockwork) Composite. Samples YJ JV03, YJ JV04, YJ JV05 and YJ JV06 were blended to produce a Stockwork composite. Sample YJ JV07 was stockwork material from above the oxidation surface and therefore not used in the composite preparation. The bottle roll test results are listed in Table 13-24 and Table 13-25 list the test results for gold and silver, respectively. The tests indicated relatively high recovery of metal by cyanide leaching, although generally lower recovery was achieved than in the YJ PQ tests. Examination of the time-recovery curves indicated that metal dissolution was still increasing at 96 hours, suggesting that coarse gold particles had not been completely recovered in the bottle roll tests.

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Table 13-24 Bottle Roll Tests from YJ JV Composites, Gold Recovery at Various Feed Sizes

Composite	Feed Size (P80)	Test	Au Rec. (%)	g Au/mt mineralization				Reagent Req., kg/mt mineralization
				Ext’d	Tail	Cal’d Head	Head Assay	NaCN Cons.	Lime Added
JV+Stockwork	150μm	Initial	76.2	3.72	1.16	4.88	4.60	0.19	0.6
JV+Stockwork	150μm	Dup	76.4	3.72	1.15	4.87	4.60	0.21	0.6
JV+Stockwork	75μm	Initial	76.7	3.75	1.14	4.89	4.60	0.22	0.7
JV+Stockwork	75μm	Dup	77.0	3.64	1.09	4.73	4.60	0.19	0.9
Stockwork	150μm	Initial	69.2	1.17	0.52	1.69	1.64	<0.07	1.0
Stockwork	150μm	Dup	56.3	0.90	0.70	1.60	1.64	<0.07	1.1
Stockwork	75μm	Initial	75.4	1.50	0.49	1.99	1.64	0.14	1.1
Stockwork	75μm	Dup	73.5	1.39	0.50	1.89	1.64	<0.07	1.3

Table 13-25 Bottle Roll Tests from YJ JV Composites, Silver Recovery at Various Feed Sizes

Composite	Feed Size (P80)	Test	Ag Rec. (%)	g Ag/mt mineralization				Reagent Req., kg/mt mineralization
				Ext’d	Tail	Cal’d Head	Head Assay	NaCN Cons.	Lime Added
JV+Stockwork	150μm	Initial	56.2	34.2	26.7	60.9	61.2	0.19	0.6
JV+Stockwork	150μm	Dup	53.5	33.0	28.7	61.7	61.2	0.21	0.6
JV+Stockwork	75μm	Initial	58.5	35.6	25.3	60.9	61.2	0.22	0.7
JV+Stockwork	75μm	Dup	58.0	34.4	24.9	59.3	61.2	0.19	0.9
Stockwork	150μm	Initial	53.8	3.5	3.0	6.5	7.5	<0.07	1.0
Stockwork	150μm	Dup	51.5	3.4	3.2	6.6	7.5	<0.07	1.1
Stockwork	75μm	Initial	57.1	3.6	2.7	6.3	7.5	0.14	1.1
Stockwork	75μm	Dup	59.1	3.9	2.7	6.6	7.5	<0.07	1.3

13.2.1.7.3.2 YJ JV E-GRG Tests

E-GRG tests were conducted on splits of the two YJ JV composites to characterize metal recovery with increasingly fine grind size. The results are listed in Table 13-26 for both gold and silver.

Table 13-26 E-GRG Test Results for the YJ JV Composites, Gold and Silver Recovery

Composite	Recovery, % of Total Metal Nominal Grind Size				Head Grade g /mt mineralization
	700µm	250µm	75µm	Total	Calculated	Average
Gold
JV+Stockwork	30.4	22.3	15.6	68.3	4.56	4.80
Stockwork	31.6	25.2	11.0	67.8	1.75	1.82
Silver
JV+Stockwork	11.4	12.1	5.5	29.0	56.1	63.8
Stockwork	8.5	7.9	5.5	21.9	6.0	6.4

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13.2.1.7.3.3 Gravity Concentration / Gravity Product Cyanidation Tests

Each YJ JV sample, JV01 through JV06, was fed individually into a Knelson^TM concentrator at a particle size of P80 -212μm (-65 mesh). The individual gravity concentrates were then blended to produce the two composite samples of Josh Vein+Stockwork (YJ JV01 and JV02) and Stockwork (YJ JV03, JV04, JV05 and JV06). Composites of the gravity tail material were also constructed. The gravity concentrates were then re-ground to P80 -45μm (-325 mesh) and then subjected to intense CN leaching for 104 hours. The gravity tail composites were re-ground to produce 3 samples at P80 -45μm (-325 mesh), -75μm (-200 mesh) and -150 micrometres (“μm”) (-100 mesh) to allow characterization of the impact of grind size on tail recovery. The leached gravity concentrate was then blended into the gravity tail material and leached for 144 hours. Gold and silver recoveries for combined tail leach and separate tail leach are listed in Table Table 13-27.

Table 13-27 Gold and Silver Recoveries from YJ JV Composites with Intense CN leaching of Gravity Concentrate and CN Leach of Tail

Composite	Process	Gravity Tail Regrind (P80)	Leach Recoveries (%)		Reagent Requirements, kg/mt mineralization
			Gold	Silver	NaCN	Lime
JV+Stkwrk	Conc. Int. CN w/Pretreat;Combined Tail Leach1	150μm	87.1	66.5	0.22	0.7
JV+Stkwrk	Conc. Int. CN w/Pretreat;Combined Tail Leach1	75μm	89.0	69.7	0.23	1.1
JV+Stkwrk	Conc. Int. CN no/Pretreat;Combined Tail Leach1	150μm	86.2	66.3	0.22	0.7
JV+Stkwrk	Conc. Int. CN no/Pretreat;Combined Tail Leach1	75μm	88.2	69.6	0.23	1.1
JV+Stkwrk	Conc. Int. CN w/Pretreat;SeparateTail Leach2	150μm	83.8	66.5	0.24	3.9
JV+Stkwrk	Conc. Int. CN w/Pretreat;SeparateTail Leach2	75μm	79.0	64.1	0.37	3.7
JV+Stkwrk	Conc. Int. CN w/Pretreat;SeparateTail Leach2	45μm	87.3	63.6	0.29	4.4
JV+Stkwrk	Conc. Int. CN no Pretreat;SeparateTail Leach2	150μm	79.2	66.1	0.24	3.5
JV+Stkwrk	Conc. Int. CN no Pretreat;SeparateTail Leach2	75μm	74.2	63.9	0.37	3.4
JV+Stkwrk	Conc. Int. CN no Pretreat;SeparateTail Leach2	45μm	83.1	63.2	0.29	4.0
Stockwork	Conc. Int. CN w/Pretreat;Combined Tail Leach1	150μm	77.6	64.1	0.15	1.0
Stockwork	Conc. Int. CN w/Pretreat;Combined Tail Leach1	75μm	78.1	64.6	0.11	1.2
Stockwork	Conc. Int. CN no/Pretreat;Combined Tail Leach1	150μm	77.0	64.2	0.14	0.9
Stockwork	Conc. Int. CN no/Pretreat;Combined Tail Leach1	75μm	77.6	64.7	0.10	1.1
Stockwork	Conc. Int. CN w/Pretreat;SeparateTail Leach2	150μm	72.6	60.8	0.11	1.2
Stockwork	Conc. Int. CN w/Pretreat;SeparateTail Leach2	75μm	74.2	64.4	0.10	1.3
Stockwork	Conc. Int. CN w/Pretreat;SeparateTail Leach2	45μm	73.7	68.0	0.11	1.8
Stockwork	Conc. Int. CN no Pretreat;SeparateTail Leach2	150μm	73.0	60.9	0.10	1.1
Stockwork	Conc. Int. CN no Pretreat;SeparateTail Leach2	75μm	74.7	64.6	0.09	1.2
Stockwork	Conc. Int. CN no Pretreat;SeparateTail Leach2	45μm	74.2	68.2	0.10	1.7
1) Combined recoveries and reagent consumptions for gravity concentrate at 80%-212μm, intensive cyanidation of gravity rougher concentrate at 80%-45μm regrind, and leaching of gravity rougher tailings (with the intensive cyanidation residue added for re-leaching) at the indicated regrind size.
2) Combined recoveries and reagent consumptions for gravity concentrate at 80%-212μm, intensive cyanidation of gravity rougher concentrate at 80%-45μm regrind, and leaching of gravity rougher tailings (with the intensive cyanidation residue added for re-leaching) at the indicated regrind size.

Gold and silver recoveries from the YJ JV composite samples were higher than the bottle roll testing confirming that leaching of coarse gold particles was incomplete in the bottle rolls. Total recoveries were high, but less than measured in the YJ PQ composite tests indicating some fall off below oxidation. The tests confirmed that gravity concentration of YellowJacket mineralization with CN leaching of the tails would produce a consistent and high metal recovery from a relatively simple mill circuit.

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13.2.2 Comminution Test Work

Material from the Sierra Blanca, Jolly Jane, and Mayflower Bulk samples were submitted to McClelland for measurements of crusher work index, ball mill grindability and abrasion index. In addition, six composites were developed from PQ3 core holes NB-13-362 and NB-13-363 to develop comminution characteristics of vein and vein stockwork materials from the YellowJacket zone. The YellowJacket composites were developed for materials both above and below the oxidation surface. Table 13-28 lists the results of the both sets of tests.

Table 13-28 Summary of Comminution Test Work on Sierra Blanca, Jolly Jane, and Mayflower Bulk Materials and on YellowJacket PQ3 Core Materials

Sample ID	Location with respect to Oxidation surface	Ball Mill Work Index (kW-hr/tonne)	Crusher Work Index (kW-hr/tonne)	Abrasion Index
Heap Leach
Sierra Blanca (SB 1019)	above	23.27	16.68	0.4577
Jolly Jane (JJ 1019)	above	24.72	22.60	0.6260
Mayflower	above	-	14.05	0.3946
Mill
YJ Comm Comp 1	above	22.8	12.3	0.7154
YJ Comm Comp 1	above	22.1	9.8	0.4259
YJ Comm Comp 1	above	21.5	7.8	0.2718
YJ Comm Comp 1	below	22.1	10.9	0.5729
YJ Comm Comp 1	below	22.5	10.5	0.5766
YJ Comm Comp 1	below	22.0	10.6	0.4778

13.2.3 Gold Deportment Studies

Investigations of gold deportment in NBP sample materials have been performed to add detail on the character of gold occurrence in the YellowJacket mineralization, and to develop preliminary information on gold occurrence in un-oxidized sulphide mineralization.

13.2.3.1 Gold Deportment in Josh Vein and Stockworks

Gold and silver mineralization in the Josh Vein and associated stockwork zones at YellowJacket is predominantly native gold and electrum. This was indicated by the high metal recovery in the gravity concentration with CN leaching of tail materials for the YJ PQ and YJ JV composites. Samples of gravity concentrate from the YJ JV samples JV01 to JV07 were analyzed by QEMSCAN Particle Mineral Analysis (PMA), Trace Mineral Search (TMS), Whole Rock Analysis (WRA) and X-Ray Diffraction Analysis (XDF) by Bureau Veritas Commodities Canada Ltd. (McClelland 2015B). Conclusions and recommendations from the analyses were:

· 95% of the gold in the seven YJ JV concentrates presented as liberated particles or gold adhesion binaries with exposed surfaces. The liberated gold and gold adhesions were probably expected to be leached using further normal cyanidation leaching methodology. Gold leach recovery of 90% to 95% from these concentrates may be expected to be achieved.

· The locked inclusion gold would be unlikely to be leached. The unliberated gold was principally locked with pyrite and non-sulphide gangue either in binary or multiphase forms.

· The silver bearing minerals were identified as gold/electrum, acanthite/argentite, stephanite/pyrargyrite, stromeyerite and tetrahedrite. Unliberated silver was mostly associated with pyrite and non-sulphide gangue either in binary or multiphase forms.

13.2.3.2 Gold Deportment in Sulphide Mineralization

In addition to the substantial oxidized gold mineralization, Josh vein and vein stockwork mineralization, other areas of un-oxidized mineralized material exist at the NBP. Metallurgical testing has indicated that some of the un-oxidized mineralized material are refractory, with gold recovery ranging from <10% - 40%. Some gold deportment work was performed on two samples of the un-oxidized material to better understand the characteristics and occurrence of the gold that was not cyanide recoverable. The gold deportment analysis was performed on the tail material after 72 hours of cyanide leaching by AMTEL. In summary, the primary conclusion from the study was, “The refractoriness to direct cyanidation of the North Bullfrog some un-oxidized mineralization is directly related to the fact that the primary Au carrier is submicroscopic Au in pyrite.”

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The gold deportment as submicroscopic particles were confirmed by metallic screen analysis at the +10 mesh in the study. The +10-mesh material contained 0.1% to 1.8% of the total gold in these tests.

The ICP analysis of all samples indicated that a weak relationship exists between arsenic and gold mineralization.

Bodies of sulphide mineralization with relatively high grades have been detected in drilling in the YellowJacket zone, along fault zones below oxidation and in drilling to the north of Sierra Blanca. A group of samples were developed to confirm previous indications of the metallurgical character of theses sulphide materials. Eleven sulphide composite samples, the YJS composites, were submitted to McClelland Labs for bottle roll testing (McClelland 2014b). Gold head grades ranged from 0.4 g/t to 4 g/t, with gold recoveries ranging from 4% to 35% at a feed size of P80 -75μm. Silver head grades were similar, ranging from 0.9 g/t to 7.4 g/t, with silver recoveries ranging from 28% to 65%. These test results indicate that simple CN leaching of the sulphide ores will not be successful. Section 13.2.5 details gold dissolution results by sulphide flotation followed by AAO and cyanidation.

13.2.4 Crush Size vs Recovery Prediction for Heap Leaching

The metallurgical study was primarily based on column leach tests at a feed size of P80 -19mm (-3/4”) particle size. Projected heap leach recoveries and leach time were the defined using this column leach test data reported here. The test data indicate high solubility, and that ultimate gold recovery is time dependent only. The column test results indicate that there is no relationship between head grade to recovery, with lower grades reporting recoveries as high as the higher grades, albeit in shorter leach time

The time dependence is being investigated at larger particle sizes in vat leach tests on Mayflower and Sierra Blanca mineralized material. The vat leaching was done on select samples at nominal particle sizes of 200-250mm, 150 mm and 75-90mm. The results are listed in Table 13-29 for gold and Table 13-30 for silver.

Mayflower David Adit material at a nominal feed sizes of 75, 150 and 200 mm obtained gold dissolutions of 68.5, 71.9 and 38.7%, respectively, in 197-198 days of leaching. Silver recovery was 9.1, 22.2 and 7.7%, respectively. The material head grade ranged from 0.47-1.22 g/t. Cyanide consumption averaged 1.6 kg/t and lime consumptions ranged from 0.0-0.04 kg/ton.

Mayflower Starlight material at a nominal feed sizes of 75, 150 and 200 mm obtained gold dissolutions of 57.8, 57.0 and 45.7%, respectively, in 197-198 days of leaching. Silver recovery was 23.5, 20.0 and 10.0%, respectively. The material head grade ranged from 1.38-3.08 g/t. Cyanide consumption averaged 1.5 kg/t and lime was not consumed.

Sierra Blanca dump material at a nominal feed sizes of 90, 150 and 250 mm obtained gold dissolutions of 57.1, 31.4 and 21.7%, respectively, in 114 days of leaching. Silver recovery was <12.5, <14.3 and <12.5%, respectively. The material head grade ranged from 0.21-0.35 g/t. Cyanide consumption averaged 0.5 kg/t and lime consumption averaged 0.65 kg/ton.

Table 13-29 Vat Leach Test Measurements of Gold Recovery on Large Particles Sizes from Mayflower and Sierra Blanca Dump Materials

Sample	Nom. Feed Size (mm)	Leach Time, Days	Au Rec., (%)	g Au/mt mineralization			Reagent Requirements Kg/mt mineralization
				Ext’d	Tail	Calc’d Head	NaCN Cons.	Lime Added
MF David Adit	200	198	38.7	0.77	1.22	1.99	2.15	0.00
MF David Adit	150	197	71.9	1.20	0.47	1.67	1.23	0.00
MF David Adit	75	197	68.5	1.02	0.47	1.49	1.42	0.04
MF Starlight	200	198	45.7	0.63	0.75	1.38	1.88	0.00
MF Starlight	150	197	57.0	1.59	1.20	2.79	1.27	0.00
MF Starlight	75	197	57.8	1.78	1.30	3.08	1.40	0.00
SB Dump	250	114	21.7	0.05	0.18	0.23	0.36	0.35
SB Dump	150	114	31.4	0.11	0.24	0.35	0.42	0.67
SB Dump	90	114	57.1	0.12	0.09	0.21	0.71	0.93

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Table 13-30 Vat Leach Test Measurements of Silver Recovery on Large Particles Sizes from Mayflower and Sierra Blanca Dump Materials

Sample	Nom. Feed size (mm)	Leach Time (days)	Ag Rec., (%)	g Ag/mt mineralization			Reagent Requirements Kg/mt mineralization
				Ext’d	Tail	Calc’d Head	NaCN Cons.	Lime Added
MF David Adit	200	198	7.7	0.1	1.2	1.3	2.15	0.00
MF David Adit	150	197	22.2	0.2	0.7	0.9	1.23	0.00
MF David Adit	75	197	9.1	0.1	1.0	1.1	1.42	0.04
MF Starlight	200	198	10.0	0.1	0.9	1.0	1.88	0.00
MF Starlight	150	197	20.0	0.3	1.2	1.5	1.27	0.00
MF Starlight	75	197	23.5	0.4	1.3	1.7	1.40	0.00
SB Dump	250	114	<12.5	<0.1	<0.1	<0.1	0.36	0.35
SB Dump	150	114	<14.3	<0.7	<0.6	<0.7	0.42	0.67
SB Dump	90	114	<12.5	<0.8	<0.7	<0.7	0.71	0.93
*silver head grade below detection resolution

13.2.5 Sulphide Material Processing

Hazen Research, Inc. performed sulphide flotation, AAO and cyanide leach experiments on gold material samples from November 2016 to October 2017. Gold dissolution and sulphide oxidation correlated very well. This study assumes sulphide oxidation will be as good, or better with pressure oxidation of the concentrate instead of AAO oxidation, and that gold dissolution will remain the same as those obtained in AAO tests.

The goals of the bench-scale work were to determine the potential to concentrate gold bearing sulphides by froth flotation, oxidize the sulphide concentrates at atmospheric pressure with oxygen and leaching the oxidized residues with sodium cyanide to recover gold and silver.

The program was conducted in three phases:

a. The first phase, consisted of flotation to produce a sulphide concentrate, concentrate Alkaline Atmospheric Oxidation (AAO) tests, and cyanidation of oxidized material. Flotation was conducted in three stages with 1 kg and 10 kg charges. Flotation amenability tests were completed on Dacite material to establish reagent and material grinding requirements. Variability tests on individual composites were completed after amenability tests. Variability drill hole composites were designated as Sierra Blanca and Pioneer Formation volcanic tuff, Dacite and Rhyolite. Variability tests were followed by 10 kg bulk tests to generate concentrate for AAO amenability testing.

In atmospheric alkaline oxidation (AAO) testing, sulphide in flotation concentrate material was oxidized by pure oxygen with the addition of soda ash or trona and gold and silver dissolution obtained by cyanidation of the oxidized residues.

b. The second phase, consisted of seven additional AAO tests; five on remaining Phase 1 sulphide concentrate and two whole sample oxidation tests. Sierra Blanca Tuff and Dacite materials were used. Gold recovery as a function of sulphide oxidation was evaluated on Sierra Blanca Tuff with soda ash in three tests. Magnesium carbonate as an alternative to soda ash was evaluated on Sierra Blanca and Dacite material in two tests. Whole sample oxidation using AAO conditions was evaluated with magnesium carbonate and soda ash on Sierra Blanca material in two tests.

c. The third phase, consisted of ten flotation tests and five additional AAO and cyanidation tests to evaluate operation at higher temperature, simulate commercial vessel pressure and commercial operation gas composition, 93 vol% oxygen. Sierra Blanca tuff material was used for all Phase III tests. Flotation tests consisted of seven grind sizes versus recovery tests, three duplicate tests at a revised reagent concentration, and three bulk flotation tests to generate concentrate for AAO oxidation.

13.2.5.1 Phase I – Sample and Composite Preparation

Approximately 876 kg of whole sample samples produced from 10 drill holes were received by Hazen in October 2016. Table 13.31 summarizes the received samples. Individual drill holes were identified by o type with seven classified as tuff and two as dacite. The tenth drill hole (NB16S-9) consisted of two ore types, tuff and rhyolite. This increased the number of individual samples received to 11. Previous mineralogical studies indicated gold and silver were mostly bound as submicron particles in solid solution with pyritic mineralization, rendering them refractory to cyanide leaching.

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Sulphur speciation indicated sulfate-sulphur and elemental sulphur assays were below the detection limit. Total sulphur assays were considered to be equivalent to sulphide sulphur. Sulphide sulphur content ranged from 0.9 to 3.2 wt%. Holes NB16S-9 and 10 contain higher levels of thorium and uranium.

Each of the eleven samples was staged crushed to -10 mesh (US) and blended by coning and quartering. Afterwards, each composite was rotary split to generate eight 1 kg charges for the initial round of bench flotation scoping experiments. Analytical head samples were riffle split and analyzed for: gold, silver, the inductively coupled plasma–optical emission spectroscopy (ICP–OES) 19 element suite, uranium, total sulphur, sulfate-sulphur and sulphide by difference. Table 13.32 and Table 13-33 present drill hole composite analyses.

Table 13-31 Drill hole Composite Classification

Hole ID	From Depth Interval, m	To Depth Interval, m	Weight, kg	Ore Type
NB-13-230 (NB16S-1)	128.02	207.26	54.4	Sierra Blanca tuff
NB-13-366 (NB16S-2)	73.21	88.85	36.9	Upper dacite
NB-15-269 (NB16S-3)	94.49	112.78	96.6	Savage Valley dacite
NB-16-295 (NB16S-4)	74.68	86.87	45.3	Sierra Blanca tuff
NB-16-296 (NB16S-5)	74.68	96.01	104.8	Sierra Blanca tuff
NB-16-297 (NB16S-6)	108.2	123.44	85.2	Sierra Blanca tuff
NB-16-298 (NB16S-7)	128.02	144.78	111.7	Sierra Blanca Tuff
NB-16-303 (NB16S-8)	275.84	291.08	74.7	Pioneer tuff
NB-16-300a (NB16S-9a)	217.93	237.74	126.3	Pioneer tuff
NB-16-300b (NB16S-9b)	237.74	254.51	92.1	Rhyolite
NB-16-310 (NB16S-10)	149.35	163.07	48.1	Sierra Blanca tuff

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Table 13-32 Drill hole Composite Analysis for the North Bullfrog Samples 1 of 2

 

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Table 13-33 Drill hole Composite Analysis for the North Bullfrog Samples 2 of 2

 

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Preliminary Economic Assessment – NB-MLP	Page 149

After the initial round of flotation scoping experiments on each drill hole, larger quantities of concentrates were needed for the AAO experimental program. Whole ore composites were blended to obtain four ore type composites. The four composites were V-blended and riffle split into 10 kg quantities as feedstock for lithology flotation experiments. Table 13-34 presents whole ore composite blending information.

Table 13-34 Composites for Lithology Flotation

Composite	Ore Type	Source	Quantity, kg
	Sierra Blanca tuff	NB16S-1	10
	Sierra Blanca tuff	NB16S-4	10
NB16-CompSBT	Sierra Blanca tuff	NB16S-5	10
	Sierra Blanca tuff	NB16S-6	10
	Sierra Blanca tuff	NB16S-7	10
	Sierra Blanca tuff	NB16S-10	10
NB16-CompPFT	Pioneer tuff	NB16S-8	30
		NB16S-9A	30
NB16-CompDAC	Upper Dacite	NB16S-2	10
	Savage Valley Dacite	NB16S-3	10
NB16-CompRHY	Rhyolite	NB16S-9B	20

13.2.5.2 Sulphide Flotation – Development, Variability and Lithology Tests

In Phase I, a series of 25 flotation experiments were performed on nineteen 1 kg drill hole composites and six, 10 kg, lithology composites. In Phase III, ten additional flotation tests were conducted on Sierra Blanca tuff to further evaluate flotation conditions and obtain concentrate material for additional AAO tests.

The mechanical flotation experiments were carried out in a Denver (Metso) sub-aeration flotation machine using cells with capacities of 2.2 and 4.4 L. The impeller speed was 1,500 rpm for the 2.2 L cell and 1,000 rpm for the 4.4 L cell. For the larger lithology flotation experiments, a Denver D-7, 28.3 L capacity cell was used.

Phase I investigations consisted of:

• Development flotation on Composite NB16S-2 (Upper Dacite).

• Variability flotation on 11 drill whole composites.

• Flotation on four lithology composites to produce concentrates for AAO work.

Phase III investigations consisted of:

• Four additional grind-versus-recovery tests on Sierra Blanca tuff.

• Three repeat tests at an adjusted regent concentration.

• Three bulk, 10 kg, flotation tests to produce concentrate for AAO work.

13.2.5.2.1 Flotation – Sierra Blanca Tuff

Sierra Blanca tuff material was tested in thirteen lithology flotation tests and five bulk flotation tests. Flotation pulp densities ranged from 19 to 32% solids. Flotation feed was conditioned for 11–22 min and flotation retention times ranged from 8 to 38 min. The slurry pH averaged 6.7.

Sierra Blanca flotation feed size ranged from P80 48-104 um, flotation mass pull ranged from 7.0-13.8%, gold recovery ranged from 66.9%-96.5% and concentrate gold grades ranged from 6.1-20.1 g/t. The Sierra Blanca tuff reagent schedule and dosages test results are presented in Table 13-35 and Table 13-36.

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Table 13-35 Sierra Blanca Tuff – Flotation Reagents

Flotation Reagent	Reagent	Dosage, kg/t
Modifier	Cyquest 3223	0.12-2.2
Collector	C 3535	0.71
Frother	Polyfroth W22C	0.000 - 0.06

Table 13-36 Sierra Blanca Tuff – Flotation Results

 

13.2.5.2.2 Flotation – Upper and Savage Valley Dacite

Initial development flotation work was performed on Composite NB16S-2 (Upper Dacite). This drill hole sample was collected as core, while the other drill hole samples were rotary bit drilled. A concern was expressed that drilling mud reagents mixed in with rotary cuttings might negatively affect float performance. Test results indicated this was not an issue.

Upper Dacite material was tested in eight development tests, one variability test and one lithology flotation test. Savage Valley Dacite was evaluated in one variability test. Flotation pulp densities ranged from 19 to 39% solids. Flotation feed was conditioned for 6–16 min and flotation retention times ranged from 8 to 30 min. The pH averaged 6.6. The reagent schedule and dosages are provided in Table 13-37.

Table 13-37 Dacite – Flotation Reagents

Flotation Reagent	Reagent	Dosage, kg/t
Modifier	Cyquest 3223	0.0-1.1
Collector	Aerophene 3418 A/ C 3535	0.000 -0.054 /0.20-0.51
Frother	Polyfroth W22C	0.013-0.055

Upper Dacite flotation feed size ranged from P80 73-196 um, flotation mass pull ranged from 9.2-48.7%, gold recovery ranged from 53%-95.6% and concentrate gold grades ranged from 2.7-13.2 g/t. Upper Dacite and Savage Valley Dacite test results are presented in Table 13-38.

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Table 13-38 Upper and Savage Valley Dacite Flotation Results

Name	Material Type	Test No.	Grind, P80	Mass Pull	Recovery Au	Recovery Ag	Concentrate Au Assay
			um	%	%	%	g/t
Development	Upper Dacite	3934-16	73	9.2%	85.2%	n/a	13.0
	Upper Dacite	3934-3	73	41.9%	93.3%	77.4%	2.9
	Upper Dacite	3906-145	93	21.8%	53.0%	64.2%	3.5
	Upper Dacite	3934-6	97	9.6%	88.6%	n/a	13.2
	Upper Dacite	3934-4	97	47.6%	93.6%	86.8%	2.7
	Upper Dacite	3906-142	113	10.5%	59.4%	33.4%	8.1
	Upper Dacite	3906-146	127	13.7%	70.2%	51.1%	7.9
	Upper Dacite	3934-5	191	48.7%	95.6%	n/a	2.8
Variability	Upper Dacite	3934-17	95	13.6%	93.1%	n/a	9.5
Variability	Savage Valley Dacite	3934-60	196	28.7%	81.7%	n/a	3.7
Lithology	Dacite Lithology	3934-105	142	11.5%	86.8%	n/a	10.6

13.2.5.2.3 Flotation – Pioneer Formation and Rhyolite

Pioneer Formation tuff material was tested in two lithology flotation tests and two bulk flotation tests. Flotation pulp densities averaged 30.5% solids. Flotation feed was conditioned for 11 min and flotation retention times ranged from 8 to 16 min. The slurry pH averaged 8.7.

Flotation feed grind sizes ranged from 62 – 91 um. Pioneer Formation flotation mass pull ranged from 4.0-7.7%, gold recovery ranged from 84%-97.6% and concentrate gold grades ranged from 7.0-16.1 g/t. Rhyolite material was tested in one lithology flotation test and one bulk flotation test. Flotation feed grind sizes were 83 and 188 um. Flotation pulp density averaged 32.5% solids. Flotation feed was conditioned for 11 min and flotation retention times were 8 and 16 min. The slurry pH averaged 7.6.

Rhyolite flotation mass pulls were 7.2 and 6.9%, gold recoveries were 89.3% and 76.0% and concentrate gold grades were 17.2 and 16.9 g/t. Pioneer Formation and Rhyolite reagent schedule and dosages and test results are presented in Table 13-39 and Table 13-40.

Table 13-39 Pioneer and Rhyolite – Flotation Reagents

Flotation Reagent	Reagent	Dosage, kg/t
Modifier	Cyquest 3223	1.1
Collector	C 3535	0.03-0.31
Frother	Polyfroth W22C	0.010 - 0.041

Table 13-40 Pioneer and Rhyolite – Flotation Test Results

 

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13.2.5.2.4 Flotation – Uranium Deportment

Uranium deportment was evaluated on Sierra Blanca tuff, Pioneer Formation tuff and Rhyolite material. Uranium concentrations in feed materials were 287, 368 and 281 mg/kg, respectively. Uranium recoveries to concentrates were 29.7, 20.0 and 12.0%, respectively. Concentrate uranium grades were 650, 970 and 490 mg/kg, respectively. Uranium accountability based on mass pull and uranium analysis, [(Concentrate + Tails)/Feed)] were 99.2, 100.6 and 101.7% respectively. Table 13-41 presents flotation uranium deportment test results.

Table 13-41 Flotation Uranium Deportment Test Results

 

13.2.5.3 Phase I – Atmospheric Alkaline Oxidation

The phase I AAO and Cyanidation test results are listed in Table 13-42. Gold bound as submicron particles in pyritic mineralization will leach in cyanide when liberated by sulphide oxidation. Pyrite will be oxidized to goethite or hematite and sulphide oxidizes to sulphuric acid, as shown in Reaction 1.

· 2 FeS₂ + 7.5 O₂ + 4H₂O → Fe₂O₃ + 4H₂SO₄ (1)

The rate of atmospheric pyritic oxidation at temperatures below 100°C increases as the pH is raised. The AAO process targets an alkaline environment by neutralizing the acid generated in Reaction 1 with a base reagent such as soda ash, Na₂CO₃, or trona (Na₂CO₃·NaHCO₃·2H₂O).

Trona dissociates to sodium carbonate, Na₂CO₃, sodium bicarbonate, NaHCO₃ and water, H₂O, according to Reaction 2. Sodium carbonate reacts with the acid generated from the sulphide oxidation to form NaHCO₃. Sodium bicarbonate from trona and generated via Reaction 3 is consumed by Reaction 4 to form sodium sulfate, Na₂SO₄. Thus, the overall reaction when all of the trona is completely decomposed is shown by Reaction 5.

· Na₂CO₃·NaHCO₃·2H₂O → Na₂CO₃ + NaHCO₃ + 2H₂O (2)

· 2FeS₂ + 7.5O₂ + 8Na₂CO₃ + 5H₂O → 2FeOOH + 4Na₂SO₄ + 8NaHCO₃ (3)

· 2FeS₂ + 7.5O₂ + 8NaHCO₃ → 2FeOOH + 4Na₂SO₄ + 8CO₂ + 3H₂O (4)

· 6FeS₂ + 22.5O₂ + 8Na₂CO₃·NaHCO₃·2H₂O → 6FeOOH + 12Na₂SO₄ + 16CO₂ + 17H₂O (5)

13.2.5.3.1 Flotation Concentrate Preparation and AAO Apparatus

The concentrates produced from the six lithology flotation experiments were used as feed for the AAO experiments. The two Sierra Blanca concentrates were combined, V-blended and riffle split into 200 g increments. The two Pioneer Formation tuff concentrates were also prepared in this manner. The dacite and rhyolite concentrates did not need blending because only one lithology float was performed for each ore type. The blended concentrates were ground to a fine particle size, P80 below 20 μm.

The sulphide concentrate was reacted with oxygen in a 2 L glass resin kettle equipped with a Teflon radial agitator, a thermocouple to measure temperature and a reflux condenser on the gas outlet to return evaporated water to the reactor. The resin kettle was placed in heating mantles with automatic temperature controllers. Oxygen was introduced through a 1/8-inch stainless steel tube that discharged directly underneath the mixer blades to better disperse the gas bubbles.

The ground concentrate was slurried with water to 15% pulp density and heated to 75°C. Once the target temperature was reached, dry sodium carbonate, Na₂CO₃, or trona was added and the oxygen turned on. Oxygen flow was manually set with a rotameter to 140–145 cubic centimeters per minute (“cm³/min”). The pH and electromotive force of the slurry were measured periodically. Kinetic samples were collected to measure the sulphide oxidation rate of the solids.

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At the conclusion of the experiment, the oxidized slurry was filtered and the solid cakes washed with three bed displacements of deionized water. The oxidized solids were advanced as a wet filter cake for a 48 hr Carbon-In-Leach (CIL) cyanide bottle roll leach test.

13.2.5.3.2 Phase I – AAO and Cyanidation Results

Table 13-42 present results of AAO and cyanidation tests. Unoxidized Sierra Blanca tuff had a gold dissolution of 9.5%. Cyanide and lime consumptions were 1.0 and 2.5 kg/ton concentrate, respectively.

In soda ash, Sierra Blanca tuff gold dissolution, with 99.4% sulphide oxidation, was 99.2%. Cyanide and lime consumptions were 0.26 and 3.4 kg/ton concentrate, respectively. In trona, Sierra Blanca tuff gold dissolution, with 96.4% sulphide oxidation was 91.9%. Cyanide and lime consumptions were 0.26 and 3.4 kg/ton concentrate, respectively.

In soda ash, Pioneer Formation tuff, Dacite and Rhyolite obtained sulphide oxidation of 98.8, 96.4 and 99.3%, respectively. Gold dissolutions were 99.8-99.9%. Cyanide consumptions were 2.6-3.5 kg/ton concentrate. Lime consumptions were 3.0-5.3 kg/ton concentrate.

In trona, Pioneer Formation tuff, Dacite and Rhyolite obtained sulphide oxidation of 96.7, 96.5 and 99.3%, respectively. Gold dissolutions were 97.6, 96.4 and 98.7%, respectively. Cyanide consumptions were 2.8-3.6 kg/ton concentrate. Lime consumptions were 2.3-8.4 kg/ton concentrate.

13.2.5.3.3 AAO – Soluble Gold

Gold dissolution in the AAO leach solution ranged from 4.2-58.9%. Sulphide sulphur may be partially oxidized in the AAO leach to form sodium thiosulfate, Na2S2O3. Gold is soluble in thiosulfate solutions via the following reaction.

· 4Au+ 8S₂O₃^-2 +O₂ +2H₂O →4[Au(S₂O₃)₂]^-3 + 4OH^-1 (6)

Gold thiosulfate will not be recovered on activated carbon, but is currently recovered commercially on resins. Sodium thiosulfate present in leach solutions also leads to higher lime consumptions by the formation of calcium thiosulfate.

· Na₂S₂O₃- + Ca(OH)₂ → CaS₂O₃ + 2NaOH (7)

Additional test work will be required to eliminate thiosulfate formation in the AAO leach or produce conditions after the leach to naturally decompose the thiosulfate species. Soluble gold in the AAO leach solution is a low risk and manageable issue.

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Table 13-42 Phase I – AAO and Cyanidation Test Results

 

 

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13.2.5.3.4 Phase II – AAO and Results

Phase II AAO leach tests utilized previously floated Sierra Blanca and Dacite concentrate and whole ore Sierra Blanca materials. Table 13-43 presents Phase II AAO test results.

Gold recovery versus sulphide oxidation tests were conducted at 40, 60 and 80% stoichiometric calculated soda ash addition on Sierra Blanca concentrate. Sulphide oxidation was 52, 71 and 78%, respectively. Cyanide leach, CIL, gold dissolutions were 75.6, 88.6 and 79.6%, respectively. Cyanide consumptions were 3.8, 3.5 and 4.5 kg/ton concentrate. Lime consumptions were 10.6, 12.1 and 11.5 kg/ton concentrate. Soluble gold in the AAO leach solutions indicated 6.1, 3.9 and 49.7% gold dissolution.

Magnesium carbonate as an alternative neutralizing agent was evaluated with Sierra Blanca and Dacite concentrate. Test results indicated magnesium carbonate was not a suitable reagent. Sulphide oxidation was 30 and 41%, respectively. Cyanide leach, CIL, gold dissolutions were 35 and 53%, respectively. Cyanide consumptions were 3.4 and 3.6 kg/ton concentrate. Lime consumptions were 4.6 and 6.0 kg/ton concentrate. Soluble gold in the AAO leach solutions was not detected.

Whole ore AAO leach tests at a primary grind P80 of 60 microns were completed on Sierra Blanca material with magnesium carbonate and soda ash. Test results indicate low sulphide oxidation and gold recovery. Sulphide oxidation was 6 and 71.4%, respectively. Cyanide leach, CIL, gold dissolutions were 37 and 77%, respectively. Cyanide consumptions were 2.8 and 2.6 kg/ton whole ore. Lime consumptions were 2.3 and 1.5 kg/ton whole ore. Soluble gold in the AAO leach solutions was not detected.

Table 13-43 Phase II – AAO and Cyanidation Test Results

 

 

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13.2.5.3.5 Phase III – AAO testing and Results

In Phase III, four additional AAO leach tests were completed. The tests were conducted in a two-liter autoclave with 93 vol% oxygen at pressures that ranged from 15-47 psig and temperatures that ranged from 85^oC to 120^oC. Sierra Blanca flotation concentrate material was produced in Phase III bulk flotation tests. Table 13-44 presents results of Phase III AAO and cyanidation tests.

In tests conducted at 85, 100 and 120^oC with oxygen pressures of 33, 15 and 47 psig, sulphide oxidations were 92.5, 85.6 and 98.1%, respectively. Cyanide leach, CIL, gold dissolutions were 92, 84 and 97%, respectively. Cyanide consumptions were 0.4, 1.6 and 0.4 kg/ton concentrate and lime consumptions were 12, 7 and 14 kg/ton, respectively.

In a duplicate test at 85^oC and 33 psig oxygen pressure, the oxidized slurry was immediately neutralized with lime after the AAO leach and the slurry conditioned for 24 hours to allow thiosulfate to decompose. Thiosulfate was monitored and decreased from 5 g/l to 2 g/l after 24 hours. Soluble gold was very low 2%. Cyanide leach, CIL, gold dissolution was 97%. Cyanide consumption was 1.5 kg/ton concentrate. Lime consumption in the conditioning step was very high due to the thiosulfate present, 322 kg/ton concentrate.

Table 13-44 Phase III – AAO and Cyanidation Test Results

 

13.3 Metallurgical Summary

The samples tested are representative of the various types and styles of mineralization and the mineral deposit as a whole at the NBP. The core and bulk samples were augmented by samples from RC holes to project spatial and depth related variability. The Author knows of no known processing factors or deleterious elements that could have a significant effect on economic extraction.

The testing on higher grade vein and stockwork mineralization in the YellowJacket zone indicates it should be processed in a milling system using gravity concentration, intensive CN leaching of the gravity concentrate followed by CN leaching of the combined gravity tail and leached gravity concentrate. A conventional paste tailing management facility would be constructed for the limited volume of milled material. The grades and recoveries of the NBP disseminated mineralized material are suitable for heap leach processing. NBP sulphide material is amenable to sulphide and gold concentration by flotation, oxidative pretreatment by AAO followed by cyanidation. Pressure oxidation tests have not been performed on NBP sulphide material. It is assumed gold recovery will remain the same or increase in with pressure oxidation.

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13.3.1 Mill Process Characteristics and Recovery

The parameters, listed in Table 13-45, assumed for the mill process assumed in the process analysis are based on metallurgical testing results for the gravity concentration, intense cyanide leaching of gravity concentrate and cyanide leaching of the combined gravity tail.

Table 13-45 North Bullfrog Assumed Oxide Mill Process Parameters

Process Parameter	Assumed Value
Primary Cush Size	P80 - 110mm
SAG Mil	P80 – 12 mm
Pebble Crusher Size	P80 - 2.0 mm
Gravity Concentrate Feed	P80 - 212μm
Gravity Concentrate Mass Pull	0.3 - 1%
Gravity Concentrate Re-grind Size	P80 -45μm (-325 mesh)
Ball Mill Cyclone Overflow	P80 -75μm (-200 mesh)
Gravity Concentrate Leach – NaCN consumption	25 .0 kg/tonne
Gravity Concentrate Leach – Lime added	14.0 kg/tonne
Gravity Tail Leach – NaCN consumption	0.13 kg/tonne
Gravity Tail Leach – Lime added	1.0 kg/tonne

Metal recoveries indicated by the gravity concentration/CN leaching test work on YellowJacket vein and vein stockwork samples were high (+90%) across a grade range 1.5 – 9 g/t, then dropped at lower grade (0.5 g/t) for composite samples from above the oxidation surface. The recoveries for composite samples from below the oxidation surface were approximately 5% lower. Mineralization in YellowJacket, associated with the structural zones were subdivided by structure type, grade and location with respect to the oxidation surface to estimate gold and silver recovery. Minor amounts of pyrite (1-2%) were present in the mineralization below the oxidation surface and contained a small fraction of the gold, which probably impacted recovery. Table 13-46 lists the portions of metal projected to occur in each structural zone and the estimated metal recovery used to estimate a weighted average for the YellowJacket mineralization.

Table 13-46 Proportions of Metal and Estimated Gold and Silver Recovery

Oxidation	Structural Zone	Estimated Au Rec. (%)	Proportion of Au (%)	Estimated Ag Rec. (%)	Proportion of Ag (%)
Above	Vein and Vein Stockwork	91.0	36	78.2	52
Above	Minor Stockwork	92.9	4	79.0	2
Above	Faults	76.1	4	68.2	4
Below	Vein and Vein Stockwork	84.7	52	63.3	41
Below	Minor Stockwork	80.8	2	60.7	1
Below	Faults	71.0	1	60.7	1
Weighted Average	All	86.8	100	71.4	100

13.3.2 Mother Lode and North Bullfrog Sulphide Material Characteristics and Recovery

The parameters, listed in Table 13-47, assumed for the mill/flotation/AAO process, are based on metallurgical testing results for flotation concentration, sulphide AAO and CIL cyanidation of the oxidized material.

Gold recoveries are presented in

. Sierra Blanca and Pioneer Formation tuff had overall gold recoveries of 93 and 94%, in Soda Ash, respectively and 91 and 93% when in trona, respectively. Rhyolite gold dissolution was 89 and 87% in soda and trona, respectively. Dacite gold dissolution was 88 and 87% in soda ash and trona, respectively. Silver recovery was based on Sierra Blanca concentrate values back calculated through flotation to a calculated head grade. Silver recovery was estimated at 77% in flotation and 74% in the cyanide leach for a weighted recovery of 57%. Based on a 10% mass pull for all material types, the average bottle roll cyanide and lime consumptions are 0.2 and 0.6 kg/tonne ore.

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Table 13-47 Assumed Mother Lode and NBP Sulphide Mill Process Parameters

Process Parameter	NBP Assume Value	Mother Lode Assume Value
Primary Cush Size	P80 - 110mm
SAG Mill	P80 - 12mm
Peble Crush Size	P80 - 2mm
Ball Mill Cyclone Overflow	P80 -75μm (-200 mesh)
Flotation
Lab Residence Time	20 min	25-30 min
Commercial Residence Time	50 min	50 min
Rougher Flotation Concentrate Mass Pull	10%	26%
Sulphide Recovery to Concentrate	85-95%	85.4-92.6%
Gold Recovery to Concentrate	88-94%	86.9% Tip1 82.3% Tjvs
Silver Recovery to Concentrate	77%	n/a
Flotation Concentrate Re-grind Size	P80 -15μm
PAX	0.25 kg/t	0.025 kg/t
Sodium Metasilicate	n/a	0.25-0.3 kg/t
Aero 208	n/a	0.05 kg/t
Aerofroth 65	n/a	0.005 kg/t
Cyquest 3223	0.1 kg/t	n/a
Polyfroth	0.03 kg/t	n/a
Flocculant	0.01 kg/t	n/a
AAO Leach/POX
Oxygen Concentration	93 vol%
PreAcidification	n/a	Yes
Alkaline Reagent	Soda Ash	n/a
Leach Residence Time	40 hrs.	1 hr.
Operating Temperature, oC	85	225
Operating Pressure, kPa	101.3	3600 (522 psi)
Oxygen Partial Pressure, kPa	~11.5	690 (100 psi)
Cyanide Leach
Gold Dissolution	97%	96.2% Tip1 93.4% Tjvs
Silver Dissolution	74%	n/a
CIL Leach – NaCN consumption	0.4 kg/tonne
CIL Leach – Lime Consumption	2.5 kg/tonne	25 kg/ tonne
Overall Recovery
Overall Gold Recovery	91%	80%
Overall Silver Recovery	57%	N/A

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Table 13-48 AAO Metal and Estimated Gold and Silver Recoveries

Conc. Type	Neutralizing Reagent	Flotation Recovery	CIL Recovery	Overall Recovery
		Au %	Au %	Au %
Sierra Blanca Tuff	Soda Ash	94%	99%	93%
Pioneer Formation Tuff	Soda Ash	94%	100%	94%
Rhyolite	Soda Ash	89%	100%	89%
Dacite	Soda Ash	88%	100%	88%
Sierra Blanca Tuff	Trona	94%	97%	91%
Pioneer Formation Tuff	Trona	94%	99%	93%
Rhyolite	Trona	89%	99%	87%
Dacite	Trona	88%	99%	87%

13.3.3 Heap Leach Process Characteristics and Recovery

Column leach testing was performed on composites samples from Sierra Blanca/Savage Valley, Jolly Jane and Mayflower resource areas. The testing indicated higher gold recovery with decreasing particle size and ultra-high intensity blasting has been assumed during mining of the heap leach mineralization to produce particle size fraction similar to primary crushing at P80 -76 mm.

Size vs recovery data has been used to extrapolate the column leach test results at P80 -19mm to the larger ROM size, and recoveries have been adjusted for time and lift height effects. ROM gold recoveries established by extrapolation of time and feed material size, with the current set of data, carry a risk of being lower because of extrapolation. It is recommended large column tests, at or above, the anticipated ROM size be completed to confirm gold recoveries and allow for interpolation of ROM material recoveries. Table 13-49 lists the projected field leach recoveries assumed after 1000 days of leaching, and the estimated field NaCN and Lime consumptions derived from the tests results.

Table 13-49 Assumed Heap Leach Metal Recoveries and Reagent Consumptions

Resource Area	% of Total Au in Heap Leach Mineralization	Average Column Test  Au Recovery* Au at 360 days (P80 -19mm)	Projected Field Leach Recovery* at 1000 days Assuming ROM at P80 -76 mm		NaCN Cons. (kg/tonne)	Lime Cons. (kg/tonne)
			Au (%)	Ag (%)
Sierra Blanca/Savage Valley	77	83.4	74.4	5.8	0.25	1.0
Jolly Jane	14	79.1	67.3	6.8	0.25	1.0
Mayflower	9	88.0	79.2	9.7	0.45	1.25

The column leach recoveries at 360 days leach is based on column leach tests at P80 -19mm, without consideration of placement, wetting, heap retention, pond retention and process gold recovery. Projected field leach recoveries from the heap leach pad assume 1,000 days leaching and a ROM particle size gradation of P80 -76mm (similar to primary crushing) produced by ultra-high intensity blasting.

13.4 Metallugical Risk, Opporunities and Recommendations

Mother Lode ROM oxide gold heap leach dissolution relies on historical production data for ROM material from the Daisy mine and in one sample showed preg-robbing characteristics, gold dissolution was assumed to be 74%, this assumption is considered valid for the PEA level of study. It is recommended gold recoveries at ROM size be confirmed by large column testing.

North Bull Frog ROM heap leach gold recovery depends on secondary leaching from solutions applied to the next lift stacked over it to complete leaching. Relying on solution applied to an upper lift will cause delays in cash flow from gold in heap inventory and carries significant risk that the ultimate recovery may not be achieved. It is recommended large column ROM recovery test establish the leach cycle with ROM sized material.

It was assumed NBP sulphide material will obtain the same gold recovery in a POX process as in the AAO process. POX testing on NBP sulphide material should confirm this assumption.

Mother Lode sulphide material presented solid/liquid separation difficulties in the test program. Flocculant screening, thickening, and rheology testing should be completed.

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The proposed processing method may at times processing oxide material while processing sulphide concentrate from stockpiles. It was assumed carbonate or dolomite content of the oxide material would be sufficient to neutralize the POX acidic solutions, or sufficient site dolomite or reprocessed rougher tails would be available for neutralization.

Additional pressure oxidation tests should be conducted and hot curing parameters established. Geometallurgical characteristics of the MLP sulphide materials needs more detailed analysis as drilling data increases. This includes the distribution of preg-robbing sulphide mineralization and verification of its passivation by pressure oxidation.

Optimization of the regrind particle size prior to POX should be completed.

Additional flotation optimization with Jameson cell technology, concentrate cleaning and locked cycle tests should be evaluated.

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14 Mineral Resource Estimates

14.1 Summary

RDA has completed the following Mineral Resource Estimates according to the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines. There have been historic resource estimates at MLP, within the context of NI 43-101, including actual mine production. In addition to NBP Resource Estimates, this report is the first Mineral Resource estimate at MLP since the adoption of NI 43-101. Project mineral inventories are reported at various cut-off-grades and classifications. Mineral Resources have been reported in accordance with the disclosure obligations under NI 43-101.

Vulcan® Software was used to estimate and quantify the Project Mineral Resources. Vulcan® software utilizes a block modeling approach to represent the deposit as a series of 3-D blocks to which grade attributes, and other attributes can be assigned. Mineral inventories have been pit-constrained using Whittle® in order to demonstrate the reasonable prospects of eventual economic extraction. General statistics and geostatistics were evaluated using GSLIB®, Sage2000®, Rockworks® Utilities, Excel and a variety of internally developed programs. Maps, cross sections, project layout and other visual aides were evaluated with Vulcan® and ArcGIS®.

The evaluation of Mineral Resources for the Project involved the following procedures:

· Validation of the database and wireframe models developed by Corvus;

· Data Processing (compositing and capping) and statistical analysis;

· Selection of estimation strategies and estimation parameters;

· Block modelling and grade interpolation and validation of the results;

· Classification and tabulation of Mineral Resources;

· Quantifying the reasonable prospects for eventual economic extraction of Mineral Resources.

Table 14-1 summarizes the Mineral Resources, classified according to CIM definitions, for the Project. Reasonable prospects for eventual economic extraction, defined in this section of the report, assume open pit mining, run-of-mine heap leach processing of oxide mineralization, mill processing of un-oxidized mineralization or mill processing of gravity separable gold and silver.

Table 14-1 Mother Lode and North Bullfrog, Pit-constrained, Measured, Indicated and Inferred Mineral Resources at Gold Selling Price of USD $1,250 per Ounce (Effective date September 18, 2018)

	Gold Milling Cutoff Grade 0.63-0.76 Au g/t			Gold Heap Leach Cutoff Grade 0.06-0.15 Au g/t			Gold Total (Milling & Heap Leach)
Resource Classification	Tonnes (x1,000)	Au g/t	Au Ounces (x1,000)	Tonnes (x1,000)	Au g/t	Au Ounces (x1,000)	Tonnes (x1,000)	Au g/t	Au Ounces (x1,000)
Measured	9,311	1.59	475	34,560	0.27	305	43,871	0.55	780
Indicated	18,249	1.68	988	149,374	0.24	1,150	167,623	0.40	2,138
Total M & I	27,560	1.65	1,463	183,934	0.25	1,455	211,494	0.43	2,918
Inferred	2,284	1.61	118	78,742	0.26	549	81,026	0.26	667
	Silver Milling Cutoff Grade 0.63-0.76 Au g/t			Silver Heap Leach Cutoff Grade 0.06-0.15 Au g/t			Silver Total (Milling & Heap Leach)
Resource Classification	Tonnes (x1,000)	Ag g/t	Ag Ounces (x1,000)	Tonnes (x1,000)	Ag g/t	Ag Ounces (x1,000)	Tonnes (x1,000)	Ag g/t	Ag Ounces (x1,000)
Measured	6,019	11.47	2,220	14,525	1.41	658	20,544	4.36	2,878
Indicated	8,315	8.93	2,388	129,251	0.66	2,758	137,566	1.16	5,146
Total M & I	14,334	9.99	4,608	143,776	0.74	3,416	158,110	1.58	8,024
Inferred	116	9.12	34	64,669	0.48	1,008	64,785	0.50	1,042

Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. Numbers may not add up due to rounding. RDA knows of no environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that may materially affect the Mineral Resource estimate in this report. The Qualified Person for the Mineral Resource Estimate is Scott Wilson.

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14.2 Mother Lode

14.2.1 Historic MLP Drilling Validated

In February 2108 RDA sought to validate the accuracy of the historic Goldcorp assay data to ensure resources could be estimated in compliance with generally accepted practices. A review of the Goldcorp exploration-development database, a review of the Corvus exploration-development database and a review of the combined exploration-development database was completed. Data from MLP was a subset of drill holes, near the Mother Lode pit including modern day Corvus drilling from 2017 through February 8, 2018.

Sufficient new information was collected and intermingled geographically throughout the Goldcorp data. The results showed that there was a negligible difference in the two assay databases suggesting that Goldcorp data collection and analysis was carried out with a standard care for the professional analysis of the mineral deposit. Analysis of the combined data and separately the Goldcorp data and the Corvus data revealed nearly identical grade ranges, mean grades, and identical coefficients of variance. Values along section 84385N, which compares four Corvus holes with eight Goldcorp holes (not quite twinning), also demonstrated that there was no statistical difference between the two drilling campaigns.

RDA found no statistical trends which would give concern for the use of the Goldcorp assays in estimating NI 43-101 compliant resources, including classification of the mineral resources. Variances of the data suggest assay reproducibility is not an issue. Coefficients of variance suggest the modern drilling data lay within the same population of assays collected by previous property owners. Comprehensive statistical analysis did not identify anomalies in the Goldcorp data. RDA recommended that it is acceptable to use the historic Goldcorp assays for the estimation of mineral resources at Mother Lode.

14.2.2 Drill hole Database

The data provided to estimate the mineralization inventory consists of data from 263 drill holes, totaling 49,580 meters. There are four channel samples that were collected from the existing pit highwall with a total length of 310 meters bringing the total length of sampling used for this resource estimate to 49,890 meters. In total 24,200 samples were assayed for gold and 10,276 samples were assayed for silver. Silver resources have not been estimated for this report.

Drilling was carried out using various drilling methods as outlined below:

· Diamond Drilling – 8 holes

· Reverse Circulation Drilling – 252 holes

· Rotary Drilling – 3 holes

· Sledgehammer/channel Sampling – 4 holes

Corvus extracts the necessary digital drilling information, from their master database, utilizing specialized scripts. Data is compiled into three separate files, one for collar, one for downhole deviation surveys and one file containing relevant geological, stratigraphic and elemental assay values; referred internally by Corvus as “strat-assay-join” files. RDA has found this to be an acceptable way to receive the drilling data. When errors or discrepancies were discovered, it was very easy to work with Corvus staff to fix such discrepancies. RDA verified that the drilling data used for this resource estimate is professionally maintained and was error free and reproducible with repeated data updates through time.

The drilling cutoff date for this report is August 10, 2018 with assays through hole ML18-078. Figure 14-1 shows drilling added to the mineral deposit by Corvus in their Phase 1 and Phase 2 drilling programs.

14.2.3 Geology Model

The geological model provided as a basis for the 2018 Mineral Resource estimation at Mother Lode was constructed by Corvus geologists using Leapfrog®. The model is an interpretation of the ML stratigraphy and structure as described in sections 7 and 8. The model consists of the elements necessary to contain the resource estimate geologically. Topologically coherent volumes have been created which allowed the block model to be tagged with all appropriate geological properties. Elements such as structural offsets and lithologic contacts are modeled in a way that the resource estimate can confidently be categorized according to CIM definition standards.

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14.2.3.1 Topographic Wireframes

Topography for the project was derived from two sources; detailed survey of the pit and the larger area derived from the drilling collars. The topographic relief at Mother Lode is slight to flat and the drill collar provided a sufficient topographic surface. The topography was converted to a solid model using Vulcan. The topographic surface is used to limit the grade estimate and calculate volumes and quantities of rock and gold for the mineral deposit. The topographic contours in figure 14.1 were derived from this surface.

14.2.3.2 Geological Wireframes

The geology of the deposit was modelled with Leapfrog software. Solid models were provided to RDA in the form of Vulcan triangulation models. The models encompass the geological interpretation of stratigraphy and the offset of the stratigraphy by the Fluorspar Canyon Fault.

Figure 14-1 Corvus Phase1 and Phase2 Drilling (red). Historic holes (black)

 

Mother Lode has been divided into a number of volumes representing topologically coherent geology, structure and stratigraphy. Figure 14-2 shows a geological cross section 4084410N through the Mother Lode Deposit. Figure 14-3 shows the 3D volumetric interpretations on the same section that were used for statistics and block modeling. These volumes were used to flag the block model and drilling composites with stratigraphy to be used in general statistical analysis and grade estimation.

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Figure 14-2 Cross Section 4084410N Looking North through Mother Lode Geological Interpretation

 

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Figure 14-3 Cross Section 4084410N Looking North through Mother Lode 3D Leapfrog Geological Interpretation

  

14.2.4 Domains

In addition to stratigraphy, the mineral deposit was subdivided three domains to aid with the determination mineralization at MLP. This is to add more geologic control to the estimation of mineral resources. Domaining is a common tool that helps yield realistic grade estimates. For the Mother Lode deposit the domains are the hanging wall and foot wall of the Fluorspar Canyon fault and the intrusive dikes of the foot wall domain.

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Figure 14-4 Cross Section 4084410N Looking North through Mother Lode 3D Leapfrog Domain model

 

14.2.5 Exploratory Data Analysis

14.2.5.1 Assay Statistics

A total of 263 drill holes and four trench samples were used in the resource estimate. The database contains 31,381 sample entries, though there are 7,181 null values. There were 24,200 individual samples assayed of which 1,846 samples were below the detection limit of 0.003 ppm. Those samples were set to one-half the detection limit at 0.0015 ppm. Thus 22,354 samples contain measurable au values. Table 14-2 presents the general gold assay statistics for Mother Lode.

Table 14-2 Gold Assay Statistics at Various Cutoff Grades

Zone	Number	Mean Au (g/t)	Stand. Dev.	Min Assay	Max Assay	Coefficient of Variation
Cutoff   0.0015	24,200	0.44	1.09	0.0015	38.74	2.50
Cutoff   0.003	22,354	0.47	1.12	0.003	38.74	2.39
Cutoff   0.100	9,598	1.06	1.53	0.01	38.74	1.44
Cutoff   0.900	3,575	2.26	1.97	0.90	38.74	0.87
Cutoff   1.000	3,282	2.38	2.01	1.00	38.74	0.85

14.2.5.2 Capping

Grade distributions for gold in the mineral deposit were examined to determine if capping was required and, if so, at what value. Assays here are displayed as a lognormal probability plot. Typically, in gold deposits, there are a small number of very high values that strongly affect exploratory statistics such as the coefficient of variance, the mean and the standard deviation. At MLP the qa/qc has shown that these high values are valid so should not be removed. Since there is a clear feeder zone identified it is also helpful to use these statistics to identify the bimodal population of mineralization for the Project and handle these accordingly. These high-grade assays are considered on a deposit by deposit basis. The value of 8 g/t Au was chosen as the capping limit for the deposit gold assays. This effectively takes 66 samples with an average grade of 12 g/t Au, above the 99^th percentile of the gold population and brings the data in line with the overall population of the mineral deposit. Figure 14-5 is the lognormal probability plot of gold assays at MLP and Table 14-3 presents the capped statistics for the mineral deposit. The inflection of the distribution at 0.9 g/t Au was chosen as the feeder zone population. The feeder population is described later in this chapter.

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Figure 14-5 Lognormal Probability Plot Assays Capped at 8 g/t Au. Low Grade and High-Grade Population Identified by Break in line at 0.9 g/t Au.

 

Table 14-3 Gold Assay Statistics at Various Cutoff Grades

Zone	Number	Mean Au (g/t)	Stand. Dev.	Min Assay	Max Assay	Coefficient of Variation
Cutoff   0.0015	24,200	0.42	0.94	0.0015	8.00	2.22
Cutoff   0.003	22,354	0.46	0.97	0.003	8.00	2.11
Cutoff   0.100	9,598	1.03	1.27	0.01	8.00	1.23
Cutoff   0.900	3,575	2.19	1.46	0.90	8.00	0.67
Cutoff   1.000	3,282	2.29	1.47	1.00	8.00	0.64

14.2.5.3 Assay Composites

Capped drill hole assays for MLP were composited using 5 metre down–the–hole composite lengths. Composite lengths were chosen based on the anticipated mine selectivity of 5 metres. A total of 8,295 x 5 metres gold and silver composites were constructed. Intervals with missing assays were ignored and a new composite centroid was generated at that point. A merge tolerance of 2.5 metres was used to limit the number of “short” composite lengths in the database. A high-grade indicator field was added to the composite database and flagged with a 1 to separate the high-grade feeder zone assays from the surrounding low-grade mineralization which was flagged with a 0.

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Composite statistics by high grade and low grade are compiled and listed in Table 14-4.

Table 14-4 MLP Composite Statistics for Gold

Zone	Number	Mean Au (g/t)	Stand. Dev.	Min Assay	Max Assay	Coefficient of Variation
Total Mineral Deposit	8,295	0.388	0.796	0.001	8	2.055
Feeder Zone	1,182	1.982	1.127	0.9	8	0.569
Disseminated	7,113	0.123	0.188	0.001	0.897	1.548

14.2.5.4 Declustering

Declustering is a technique used to smooth out higher drill densities in the bulk of a deposit. Assays are assigned a weight which is directly proportional to the area or volume of interest of each sample (Rossi and Deutsch 2014). Declustering was performed using the nearest-neighbor declustering technique. Each point receives a weight inversely proportional to the number of points that fall in the same cell. The weights are scaled to a mean of 1.

The weights depend on the cell size. When a cell size is very small, each datum is in its own cell and receives an equal weight. When the cell size is very large, all data fall into one cell and are equally weighted (Rossi and Deutsch 2014). To choose the appropriate cell, the declustered mean versus a range of cell sizes is plotted. Figure 14-6 is a graph showing the cell size versus the mean grade for MLP. The lowest mean along the curve is the cell size to be used. A cell size of 400 metres was evaluated for use in grade estimation techniques. The drilling at Mother Lode is not clustered and is fairly evenly spaced throughout the deposit.

Figure 14-6 Mother Lode Cell Declustering Evaluation

 

14.2.6 Block Model

14.2.6.1 Oxidation Model

Oxidation is evaluated on a scale of 1 to 5 with 1 being un-oxidized and 5 being completely oxidized. Levels 2 and 1 are considered to have reasonable prospects of economic extraction with milling processes while levels 3 through 5 are considered ROM heap leachable material. Oxidation was estimated based on indicators. Two variables were added to the composite database based upon oxidation picks:

· Oxide_1 (sulphide) set to 1 if logged sulphidation is 1 or 2, set to 0 if not

· Oxide_5 (oxide) set to 1 if logged sulphidation is 3, 4 or 5, set to 0 if not

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Oxide was estimated throughout the deposit. The probability that the block is oxide or sulphide was stored in the model using inverse distance squared techniques. The resulting variables are evaluated to determine the highest probability of a block being oxide, mixed or sulphide and the block is flagged by the resulting highest probability. Figure 14-7 shows the result of the oxidation estimate for the mineral deposit.

Figure 14-7 Oxidation Model, Section 4084410, Looking North

 

14.2.6.2 Mineralization Model

The MLP model was constructed to correctly model the high-grade feeder zone separately from the surrounding disseminated mineralization. The geological model has been constructed to reflect these two styles of mineralization. The high-grade feeder zone was identified using indicators for mineralization at a cutoff grade of 0.9 g/t Au.

The probability of a block being part of the high-grade zone was stored and evaluated against the assays. If a block has a 40 percent or greater probability of being greater than 0.9 g/t Au, it was flagged as part of the high-grade zone. Figure 14-8 shows the results of the estimation of the high-grade feeder zone. Blocks external to the zone are considered disseminated mineralization.

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Figure 14-8 Section 4084410 Showing High-Grade Indicator Blocks for the Project Looking North

 

14.2.7 Block Model Construction

The Mineral Resource model contains information about the deposit and is stored variably in each block. The information stored includes:

· Estimated characteristics of Au and Oxidation

· Percentage of block below the surface topography

· Specific gravity defined by geologic triangulations

· Stratigraphic Units

· Percentage of a block found within a vein and percentage of a block found within the disseminated material

Table 14-5 outlines the framework for the Mother Lode block model.

Table 14-5 Mother Lode Block Model Framework NAD27 / (UTM Zone 11 North)

Item	Easting	Northing	Elevation
Block Model Minimum Coordinate	530,000	4,083,400	800
Block Model Maximum Coordinate	531,800	4,085,400	1,500
Number of Blocks, X, Y, Z	180	200	140
Model Block Dimensions	10	10	5

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14.2.8 Specific Gravity

There are minimal measurements of the specific gravity of the different geological units. A specific gravity of 2.46 was assumed for all material types at Mother Lode

14.2.9 Grade Interpolation

Gold was estimated in two passes. The high-grade feeder zone was estimated in pass 1. Disseminated mineralization was interpolated in pass 2. A soft boundary technique was utilized where composites within the feeder where all composites within zone, blocks as an example in the previous, were used for the interpolation, excluding composites external to the zone. The second pass utilized all composites within and external to the feeder zone to receive an estimate. This methodology was employed to reflect the interpretation that the mineralization is concentrated in the feeder and then disseminated throughout the stratigraphic units.

14.2.9.1 High Grade Feeder Mineral Estimate

Gold was estimated using inverse distance squared (ID2). Search ellipsoid orientations were determined within the three domains of the deposit. Search ellipsoid orientations were assigned based on the westerly geographic bearing and plunge of the stratigraphy within the domains. Orientations were calculated and stored in each model block. Table 14-6 summarizes the breakdown of the different bearings and plunges used based on these criteria. Figure 14-9 is the typical cross section looking north showing the search ellipses in relation to the mineral deposit.

Table 14-6 High-Grade Estimation Parameters for MLP

Estimation Type	Inverse Distance Squared (ID2)
Search Ellipsoid	Bearing	Plunge	Dip
Domain 1	270	-48	-19
Domain 2	270	-10	-19
Domain 3	270	-65	-19
Search Ellipse	Major Axis	Semi-Major Axis	Minor Axis
Domain 1	180	120	90
Domain 2	180	120	90
Domain 3	180	120	90
Samples	Min	Max
	2	3
Maximum Samples per Drill hole	Max
	1

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Figure 14-9 Search Ellipsoids used for Grade Interpolation

 

14.2.10 Disseminated Mineralization

Gold was estimated into disseminated blocks using Ordinary Kriging as the estimation technique. Mineralization was estimated subsequently to the high-grade estimate. This limited mineralization to the adjacent stratigraphic horizons. Gold was estimated in three passes to ensure complete coverage of mineralization throughout the deposit. Drill hole data was used to define a variogram for the mineralization. Table 14-7 lists the variogram model parameters according to the variogram in Figure 14-10. Variograms were auto-fit and evaluated for accuracy. Table 14-8 lists the grade estimation parameters for the disseminated mineralization at MLP.

Table 14-7 Gold Variogram Model Parameters

	Variogram Model Parameters
	Nugget	0.15	Number of Structures		1	Distance (m)
Domain	Variogram Type	Sill Differential	Bearing	Plunge	Dip	Major Axis	Semi-Major Axis	Minor Axis
1	Exponential	0.857	270	-48	-19	160	90	75
2	Exponential	0.857	270	-10	-19	160	90	75
3	Exponential	0.857	270	-65	-19	160	90	75

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Figure 14-10 Mother Lode Gold Variogram Model

 

Table 14-8 Gold Ordinary Kriging Estimation Parameters in Vulcan® Format

Estimation Type	Inverse Distance Squared (ID2)
Search Ellipsoid	Bearing	Plunge	Dip
Domain 1	270	-48	-19
Domain 2	270	-10	-19
Domain 3	270	-65	-19
Search Ellipse	Major Axis	Semi-Major Axis	Minor Axis
Domain 1	160	90	75
Domain 2	160	90	75
Domain 3	160	90	75
Samples	Min	Max
	8	24
Maximum Samples per Drill hole	Max
	3

14.2.11 Swath Plots (Model Drift)

A swath plot is an analysis which compares estimated block grades to composite grades for a slice taken from the block model. This is a useful tool to help determine whether grade estimation parameters correlate well with expected values based on composite grades. Figure 14-11 and 14-12 show the results of the swath plot analysis. The swath plot analysis indicates that the MLP grade estimates correlate well with sample composites and demonstrate that the mineral estimate has been smoothed. High grade mineralization is controlled and there is confidence with the mineral estimate.

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Figure 14-11 Swath Plot Graphical Analysis by Northing

 

Figure 14-12 Swath Plot Graphical Analysis by Easting

 

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14.2.12 Visual Validation of the Mother Lode resource estimate

Figure 14-13 Plan View of Mother Lode Resource Estimate at 970m elevation

 

Figure 14-14 Long-section Through Mother Lode Resource Model at 531,055 East UTM NAD27 Zone 11 North

 

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14.3 Sierra Blanca

14.3.1 Sierra Banca Geology Model

The geological model provided as a basis for the 2017 Mineral Resource estimation at Sierra Blanca was constructed by Corvus geologists using Leapfrog®. The model is an updated interpretation of the NBP stratigraphy as described in sections 7 and 8. The model consists of the elements necessary to contain the resource estimate geologically. Topologically coherent volumes have been created which allowed the block model to be tagged with all appropriate geological properties. Elements such as structural offsets and lithologic contacts are modeled in a way that the resource estimate has confidence to be categorized into CIM defined resources.

14.3.1.1 Structural Model

Sierra Blanca area has been divided into a number of structural blocks with coherent internal stratigraphy (Figure 14-15). Each of these blocks has an encompassing domain volume which is divided into sub-volumes according to the stratigraphy in the block. Each domain block has been assigned a stratigraphy code to be used in designating estimation domains. Paleozoic Basement (Pz) has been interpreted to be a hard boundary and contains minimal to no mineralization.

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Figure 14-15 Stratigraphy Blocks Defined within Sierra Blanca and YellowJacket (metres).

 

14.3.1.2 Stratigraphic Model

The Project stratigraphic model consists of seven units which from oldest to youngest include:

1. Paleozoic Basement (Pz)

2. Savage Formation (Tsf)

3. North Bullfrog suite (Tnb1)

4. Pioneer Formation (Tpf)

5. Sierra Blanca Tuff (Tsb)

6. Dacite Dikes and Intrusives (Tdi)

7. Savage Valley Dacite (Td)

8. Lithic Ridge and Bullfrog Tuff (Tlr)

9. Rainbow Mountain (Trm)

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Within each structural block, stratigraphic units have a separate volume defined. Vulcan® cross section. (Figure 14.-6).

Figure 14-16 Stratigraphic Domains Modeled in Sierra Blanca (N 4,098,050 Looking North, metres).

 

14.3.1.3 Oxidation Model

With the exception of a few historical holes, the degree of oxidation has been determined for every sample in the drill database. The oxidation is evaluated on a scale of 1 to 5 with 1 being un-oxidized and 5 being completely oxidized. Cyanide shake leach and bottle roll tests have shown that oxidation levels 4 and 5 behave similar to level 3, which has mixed oxide and sulphide. Levels 2 and 1 are dominantly or completely sulphide and will be excluded from the oxide processing the stream in both the mill and heap leach.

Oxidation was estimated based on indicators. Three variables were added to the composite database based upon oxidation picks:

· Oxide_1 (sulphide) set to 1 if logged sulphidation is 1 or 2, set to 0 if not

· Oxide_3 (mixed) set to 1 if logged sulphidation is 3, set to 0 if not

· Oxide_5 (oxide) set to 1 if logged sulphidation is 4 or 5, set to 0 if not

Oxide was then estimated in three runs based on the three above indicators. Each run stores the probability that the block is oxide, mixed or sulphide. The resulting variables are evaluated to determine the highest probability of a block being oxide, mixed or sulphide and the block is flagged by the resulting highest probability.

Sage was used to determine the indicator variograms for each oxidation state. Table 14-9 summarizes the Vulcan® variogram estimation parameters.

Table 14-9 Oxide Indicator Variogram Parameters

Oxidation Type	Nugget	Sill Differential	Major Radius	Semi-Major Radius	Minor Radius	Rotation about Z axis	Rotation about Y axis	Rotation about X axis
Oxide	0.05	0.75	750	390	75	138	0	-3
Mixed	0.76	0.26	430	250	115	152	-9	-6
Sulphide	0.07	0.83	920	475	85	143	-2	-1

Figure 14-17 demonstrates the resulting oxidation estimate, constrained to the 2014 resource constraining pit compared to the 2014 oxidation surface. The resultant 2017 oxidation estimate more accurately reflects logged oxidation in the NBP drilling database.

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Figure 14-17 Comparison of estimated oxidation to the 2014 oxide/sulphide surface (looking north).

 

14.3.2 Mineralization Model

Mineralization at Sierra Blanca occurs in two distinct settings: low-grade disseminated mineralization and higher grade structurally controlled mineralization. The geological model has been constructed to reflect these two styles of mineralization. Two structurally controlled mineralization structures were modeled; the YellowJacket vein (Figure 14-18) and the Liberator vein (Figure 14-19). Two separate grade shells were modeled to represent the higher grades associated with the vein zones. A 0.5g/ton Au cutoff was used to estimate these domain volumes. These two high-grade mineralized zones allow for estimation parameters to be applied to only these vein zones and to limit the effects of high-grades on the surrounding disseminated mineral zones.

Recent drilling identified the Swale Zone which is interpreted as a high-grade lens within disseminated mineralization just to the north and east of the YellowJacket corridor (Figure 14-20). Similar to YellowJacket, a 0.5 g/t Au grade shell was created to characterize this lens and allow for separate estimation parameters to be applied for evaluation.

Disseminated mineralization has been modeled and limited by the newly interpreted stratigraphic models. Paleozoic basement stratigraphy comprising micritic limestone, quartzite and sandstone is considered to be barren of any mineralization. No mineralization was interpreted into this unit. Contact profiles indicated that gold and silver mineralization were related through the remaining stratigraphic units.

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The mineral domains used to control the estimation for mineralization are listed below:

· Sierra Blanca: Pervasive disseminated gold and silver mineralization and represented in all units.

• YellowJacket: High-grade faulting and stockwork veining (Figure 14-18).
• Liberator: High-grade mineralization, faulting and stockwork veining (Figure 14-19).
• Swale: High-grade mineralization (Figure 14-20).

Figure 14-18 Isometric View of Modeled YellowJacket Mineralization Grade Shell (Blue)-metres.

 

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Figure 14-19 Isometric View of YellowJacket (Blue) and Liberator Veins (Green)-metres

 

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Figure 14-20 Isometric View of YellowJacket (Blue) and Disseminated Swale (Yellow)-metres.

 

14.3.3 Exploratory Data Analysis

14.3.3.1 Assay Statistics

A total of 767 drill holes containing 6,003 down-hole surveys and 92,319 assays were provided to be used for the Project mineral estimate. Of these supplied drill holes, a total of 351 drill holes, totaling 78,966 metres, intersected the geologic solid models that define the Sierra Blanca deposit. Table 14.10 presents the gold assay statistics for each of the modeled geologic domains, while Table 14-11 includes the silver assay statistics.

Table 14-10 Gold Assay Statistics Sorted By Zone

Zone	Number	Mean Au (g/t)	Stand. Dev.	Min Assay	Max Assay	Coefficient of Variation
YellowJacket Vein	3,119	2.846	14.352	0.0005	431.0	5.043
Liberator Vein	232	0.836	2.041	0.001	17.299	2.442
Swale	550	0.719	0.512	0.0005	6.390	0.712
Disseminated	88,418	0.121	1.225	0.0005	209.0	10.145
All	92,319	0.218	2.941	0.0005	431.0	13.484

Table 14-11 Silver Assay Statistics Sorted by Zone

Zone	Number	Mean Ag (g/t)	Stand. Dev.	Min Assay	Max Assay	Coefficient of Variation
YellowJacket Vein	3,119	23.751	162.028	0.010	7590.0	6.822
Liberator Vein	232	2.372	5.382	0.030	65.0	2.269
Swale	550	1.892	3.440	0.0005	61.0	1.818
Disseminated	88,418	0.458	2.875	0.0005	308.0	6.276
All	92,319	1.258	30.211	0.0005	7590.0	24.007

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14.3.3.2 CAPPING

Grade distributions for gold and silver within each of the different zones were examined to determine if capping was required and, if so, at what value. Assays for each zone were graphically displayed as histograms and as lognormal probability plots. YellowJacket samples were capped at 100 g/t Au and 550 g/t Ag. Both of these capping levels were determined to be where a natural break occurred in the lognormal distribution curve around the 99.5^th percentile. Figures 14-21 and 14-22 show the lognormal distributions of uncapped gold and silver assays within the YellowJacket Vein. Disseminated samples were capped at 10 g/t au and 200 g/t ag. The lognormal distribution curves both have smooth distributions, so capping levels were determined to be at the 99.9^th percentile. Figures 14-23 and 14-24 show the lognormal distributions of the uncapped gold and silver assays within disseminated mineralization. Assays in the Liberator and Swale zones were unaffected by assay capping. Table 14-12 describes the number of gold and silver assays capped for Mineral Resource estimates.

Figure 14-21 Au Lognormal Graph (within YellowJacket)

 

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Figure 14-22 Ag Lognormal Graph (within YellowJacket)

 

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Figure 14-23 Au Lognormal Graph (Disseminated)

 

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Figure 14-24 Ag Lognormal Graph (Disseminated)

 

Table 14-12 Capped Assays by Zone

Zone	Au Cap (g/t)	Number Capped	Ag Cap (g/t)	Number Capped
YellowJacket	100	11	550	15
Liberator	100	0	100	0
Swale	10	0	200	0
Disseminated	10	27	200	6

The results from capping (Tables 14-13 and 14-14) show reduced standard deviation and coefficients of variation in all groups when compared to the uncapped assays (Tables 14-11 and 14-12).

Table 14-13 Capped Assay Statistics for Gold Sorted By Zone

Zone	Number	Mean Au (g/t)	Stand. Dev.	Min Assay	Max Assay	Coefficient of Variation
YellowJacket	3,119	2.511	8.671	0.0005	100.0	3.453
Liberator	232	0.836	2.041	0.001	17.299	2.442
Swale	550	0.719	0.512	0.0005	6.390	0.712
Disseminated	88,418	0.110	0.312	0.0005	10.0	2.826

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Table 14-14 Capped Assay Statistics for Silver Sorted by Zone

Zone	Number	Mean Ag (g/t)	Stand. Dev.	Min Assay	Max Assay	Coefficient of Variation
YellowJacket	3,119	19.195	59.181	0.010	550.0	3.083
Liberator	232	2.372	5.382	0.030	65.0	2.269
Swale	550	1.892	3.440	0.0005	61.0	1.818
Disseminated	88,418	0.453	2.437	0.0005	200.0	5.376

14.3.4 Assay Composites

Capped drill hole assays for Sierra Blanca were composited using 5 metre down–the–hole composite lengths. Composite lengths were chosen based on the anticipated mine selectivity of 5 metres which corresponds to the block size length in the z-direction. A total of 27,729 x 5 metres gold and silver composites were constructed. Intervals with missing assays were ignored and a new composite centroid was generated at that point. A merge tolerance of 2.5 metres was also used to limit the number of “short” composites lengths in the database. Composite statistics are compiled and listed in Table 14-15 and 14-16.

Table 14-15 Composite Statistics for Gold Sorted By Zone

Zone	Number	Mean Au (g/t)	Stand. Dev.	Min Assay	Max Assay	Coefficient of Variation
YellowJacket Vein	625	1.473	3.152	0.001	37.096	2.140
Liberator Vein	62	0.775	1.556	0.005	8.825	2.009
Disseminated Swale	149	0.721	0.443	0.114	3.537	0.615
Disseminated	26,893	0.106	0.222	0.0005	6.096	2.088
All	27,729	0.142	0.567	0.0005	37.096	4.003

Table 14-16 Composite Statistics for Silver Sorted by Zone

Zone	Number	Mean Ag (g/t)	Stand. Dev.	Min Assay	Max Assay	Coefficient of Variation
YellowJacket Vein	625	11.361	28.733	0.007	349.659	2.529
Liberator Vein	62	1.917	2.660	0.043	12.903	1.387
Disseminated Swale	149	1.930	2.877	0.444	32.104	1.491
Disseminated	26,893	0.413	1.335	0.0005	88.967	3.233
All	27,729	0.671	4.801	0.0005	349.659	7.151

14.3.5 Block Model

The Mineral Resource model contains information about the deposit and is stored variably in each block. The information stored includes:

· Estimated characteristics of Au, Ag, S and Oxide

· Percentage of block below the surface topography

· Specific gravity defined by geologic triangulations

· Stratigraphic Unit

· Percentage of a block found within a vein and percentage of a block found within the disseminated material

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Table 14-17 outlines the framework for the Sierra Blanca block model.

Table 14-17 Sierra Blanca Block Model Framework

Item	Easting	Northing	Elevation
Block Model Reference Point	516760	4096460	720
Number of Blocks	197	259	148
Parent Block Size	10	10	5

14.3.6 Bulk Density

A total of 1,365 specific gravity measurements were used to define the density value of each block based on modeled lithology types. Basic statistics were compiled and tabulated (Table 14-18). The final density values assigned to the model were derived by eliminating 10% of the lowest and highest density values for each lithology type and using the mean value. Table 14.18 is a translation of the old Lithology units to the correlating stratigraphic units used.

Table 14-18 Lithology Types and Corresponding Specific Gravity Values

Lithology	All Samples				Minus 10% of Lowest and Highest Values
	Count	Min	max	mean	Count	min	max	mean
Post SB	462	1.88	2.64	2.33	370	2.15	2.49	2.34
Mélange	40	1.74	2.53	2.34	32	2.2	2.49	2.36
SB Middle	484	2.04	2.63	2.42	387	2.29	2.54	2.43
SB Lower	165	2.19	2.58	2.46	132	2.33	2.55	2.47
Pre SB	1	1.86	1.86	1.86	1	1.86	1.86	1.86
Camb	1	2.56	2.56	2.56	1	2.56	2.56	2.56
Gravel	1	1.85	1.85	1.85	1	1.85	1.85	1.85
Unknown Default	1	1.85	1.85	1.85	1	1.85	1.85	1.85
Rhyolite_9	68	1.85	2.50	2.18	54	2.00	2.41	2.18
Rhyolite	142	2.06	2.60	2.43	114	2.25	2.55	2.44

Table 14-19 Specific Gravity by Stratigraphy

Stratigraphy	Specific Gravity
PZ_Basement	2.56
Tsf	1.86
Tnb1	2.18
Tpf	2.47
Tsb	2.43
Tdi	2.43
Td	2.34
Tlr	2.34
Trm	2.34
YellowJacket	2.36

14.3.7 Contact Profiles

A contact profile analysis investigates the relationships between assay values in relation to the contact of geological units. This analysis was used to identify separate mineral estimation domains based on assay grades in relation to the stratigraphic units. This method takes samples from one stratigraphic unit and pairs it with samples from another geological unit based on a separation distance. The pairs are constructed over an increasing separation distance. The average grade of the first unit is plotted against the average grade calculated with the second unit. Figure 14-25 is the contact profile comparing gold assays within the YellowJacket vein vs all other stratigraphic unit within 30 metres. Figure 14-26 is the contact profile again, but with silver assays. These analyses confirm that the YellowJacket contains much higher gold and silver grades than the surrounding stratigraphy and should be used as its own estimation domain.

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Figure 14-25 YellowJacket Contact Profile (Au)

 

Figure 14-26 YellowJacket Contact Profile (Ag)

 

All stratigraphic units at NBP were evaluated to determine if separate individual estimation domain were required outside of YellowJacket, Liberator and Swale. No additional domains were identified. Figures 14-27 and 14-28 are examples.

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Figure 14-27 Savage Formation vs Pioneer Formation Contact Profile (Au)

 

Figure 14-28 Pioneer Formation vs Sierra Blanca Formation Contact Profile (Au)

 

Contact profile analyses strongly indicate that disseminated mineralization contains consistent grades from one stratigraphic unit to the next. This information allows for the mineral estimation domains to include all of the disseminated mineralization to be estimated in one run.

14.3.8 Declustering

Declustering is a technique used to smooth out higher drill densities in the bulk of a deposit. Assays are assigned a weight which is directly proportional to the area or volume of interest of each sample (Rossi and Deutsch 2014). Declustering was performed using the nearest-neighbor declustering technique. Each point receives a weight inversely proportional to the number of points that fall in the same cell. The weights are scaled to a mean of 1.

The weights depend on the cell size. When a cell size is very small, each datum is in its own cell and receives an equal weight. When the cell size is very large, all data fall into one cell and are equally weighted (Rossi and Deutsch 2014). To choose the appropriate cell, the declustered mean versus a range of cell sizes is plotted. Figure 14-29 is a graph showing the cell size versus the mean grade for YellowJacket. The lowest mean along the curve is the cell size to be used. Figure 14-30 is a graph used for declustering the silver grades. Cell sizes used for gold and silver are 135 and 110 metres, respectively.

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Figure 14-29 YellowJacket Declustering Au

 

Figure 14-30 YellowJacket Declustering Ag

 

Declustering was used for inverse distance estimation techniques at YellowJacket, Liberator and Swale.

14.3.9 Grade Interpolation

Gold and silver were estimated into the four mineralized material domains of YellowJacket, Liberator, Swale and disseminated mineralization. Each material type and their corresponding composites were evaluated independently with unique estimation parameters. Estimation parameters were selected to best represent the style of mineralization and structural attitude of each component. The following sections outline the grade estimation for YellowJacket, Liberator, Swale and Disseminated mineralization.

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14.3.9.1 YellowJacket Mineralization

Gold and silver was estimated using inverse distance squared (ID2). Search ellipsoid orientations were determined by evaluating the structural characteristics and drilling density along the vein. Search ellipsoid orientations were assigned based on the geographic bearing and plunge changes within the YellowJacket mineral shape. Orientations were calculated and stored in each model block. Table 14-20 summarizes the breakdown of the different bearings and plunges used based on these criteria. Figure 14-31 is a horizontal section showing the lateral location of strike measurements. Figure 14-32 is a cross section looking north showing changes in the plunge of the vein with depth. A two-pass estimation was performed to insure mineralization was estimated into the entire shape.

Table 14-20 YellowJacket Vein Estimation Parameters for Au and Ag

YellowJacket Vein Estimation Parameters
Estimation Type	Inverse Distance Squared (ID2)
Search Ellipsoid	Bearing	Plunge	Dip
Northing > 4098500	100		0
Northing 4098275 – 4098500	65		0
Northing 4098225 – 4098275	120		0
Northing 4098175 – 4098225	90		0
Northing < 4098175	95		0
Elevation > 1140		65	0
Elevation < 1140		80	0
Search Ellipse	Major Axis	Semi-Major Axis	Minor Axis
Pass 1	100	60	20
Pass 2	180	130	20
Samples	Min	Max
	4	20
Maximum Samples per Drill hole	Max
	2

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Figure 14-31 YellowJacket Vein Estimation Bearing Changes (Elevation 1,150 m)-metres

 

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Figure 14-32 YellowJacket Vein Estimation Plunge Changes (Northing 4098525)-metres

 

14.3.9.1.1 Liberator

Gold and silver were estimated using inverse distance squared (ID2). Search ellipsoid orientations and dimensions were determined by evaluating the structural characteristics, and drill density along the vein and are listed in Table 14-21.

Table 14-21 Liberator Au and Ag Estimation Parameters

YellowJacket Vein Estimation Parameters
Estimation Type	Inverse Distance Squared (ID2)
Search Ellipsoid	Bearing	Plunge	Dip
	254	80	0
Search Ellipse	Major Axis	Semi-Major Axis	Minor Axis
Pass 1	135	60	20
Samples	Min	Max
	2	20
Maximum Samples per Drill hole	Max
	2

14.3.10 Disseminated Mineralization

Gold, silver and sulphur were estimated into disseminated blocks using Ordinary Kriging as the estimation technique. One variogram was calculated for gold and another for silver. Eight domains were established to ensure that mineralization would not be estimated through the YellowJacket, Liberator and Swale zones. Only composites and blocks in the same domain were allowed to be used during the interpolation. This limited mineralization within the vein domains to have influence on adjacent stratigraphic horizons. Gold was estimated in two passes to ensure complete coverage of mineralization as identified by drilling. Drill hole data was used to define a variogram for the mineralization. Table 14-22 and Table 14-23 lists gold and silver variogram model parameters according to the variograms in Figure 14-33 and 14-34, respectively. Variograms were auto-fit and evaluated for accuracy. Table 14-24 and 14-25 list the estimation parameters used for the gold and silver kriging estimates, respectively.

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Table 14-22 Gold Variogram Model Parameters

Variogram Model Parameters
Nugget	0.153	Number of Structures		1	Distance (m)
Variogram Type	Sill Differential	Bearing	Plunge	Dip	Major Axis	Semi-Major Axis	Minor Axis
Exponential	0.819	330	15	-25	105	54.5	43.4

Figure 14-33 Gold Variogram Model-metres

 

Table 14-23 Gold Ordinary Kriging Estimation Parameters in Vulcan® Format

Disseminated Mineralization
Estimation Type	Ordinary Kriging
Search Ellipsoid	Bearing	Plunge	Dip
	330	15	-25
Search Distance	Major Axis	Semi-Major Axis	Minor Axis
Pass 1	105	54.5	43.4
Pass 2	225	150	50
Samples	Min	Max
	4	20
Maximum  Samples per Drill hole	Max
Pass 1 and 2	2

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Table 14-24 Silver Variogram Model Parameters

Variogram Model Parameters
Nugget	0.331	Number of Structures		1	Distance (m)
Variogram Type	Sill Differential	Bearing	Plunge	Dip	Major Axis	Semi-Major Axis	Minor Axis
Exponential	0.581	160	0	7	140	111.8	91.3

Figure 14-34 Silver Variogram Model-metres

 

Table 14-25 Silver Ordinary Kriging Estimation Parameters in Vulcan® Format

Disseminated Mineralization
Estimation Type	Ordinary Kriging
Search Ellipsoid	Bearing	Plunge	Dip
	160	0	7
Search Distance	Major Axis	Semi-Major Axis	Minor Axis
	140	112	92
Samples	Min	Max
	4	20
Maximum  Samples per Drill hole	Max
	2

Sulphur was kriged into all rock types and domains. Sulphur content will be useful for potential processing evaluations and metallurgical testing. Sulphur content is important for determination of mechanical oxidation potential of various mineralization types. Variograms and kriging estimation parameters for Sulphur are listed in Tables 14-26 and 14-27. The Sulphur variogram model is shown graphically in Figure 14-35.

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Table 14-26 Sulphur Variogram Model Parameters

Variogram Model Parameters
Nugget	0.132	Number of Structures		2	Distance (m)
Variogram Type	Sill Differential	Bearing	Plunge	Dip	Major Axis	Semi-Major Axis	Minor Axis
Exponential	0.299	349	-64	23	104	56	27.5
Exponential	0.368	122	-3	5	408	267	123

Figure 14-35 Sulphur Variogram Model-metres

 

Table 14-27 Sulphur Ordinary Kriging Estimation Parameters in Vulcan® Format

Disseminated Mineralization
Estimation Type	Ordinary Kriging
Search Ellipsoid	Bearing	Plunge	Dip
	122	-3	5
Search Distance	Major Axis	Semi-Major Axis	Minor Axis
	408	267	123
Samples	Min	Max
	4	20
Maximum  Samples per Drill hole	Max
	2

14.3.11 Swath Plots

A swath plot is an analysis which compares estimated block grades to composite grades for a slice taken from the block model. This is a useful tool to help determine whether grade estimation parameters correlate well with expected values based on composite grades. Figure 14-36 is an overview map that shows the spatial relationship of the section line (Northing 4098300) with the YellowJacket, Liberator and Swale grade shells. Figure 14-23 shows the results of the swath plot analysis. Grade shells used for estimation domains are identified. The swath plot analysis indicates that the grade estimates for the Swale, YellowJacket and Liberator grade shells correlate well with sample composites.

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Figure 14-36 North Bullfrog Project Swath Plot Section Line-metres

 

Figure 14-37 Swath Plot Graphical Analysis-metres

 

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14.4 Jolly Jane

The geologic model for Jolly Jane was built by Corvus geologists. Ninety-four (94) reverse circulation drill holes and 5 surface outcrop channel sample lines define the Jolly Jane zone. Within the 9,450 gold assays supplied for Jolly Jane a total of 61 gaps in the assay record were identified. These gaps were filled with values of 0.001 g/t Au which is an acceptable practice and has no impact on the results.

The 3D geology for Jolly Jane was modeled as two surfaces, one describing the lower contact of the mineralized Crater Flat Tuff and the other describing the upper contact. These 3D surfaces were constructed by Corvus geologists and are shown in Figure 14-38. The lower contact is sometimes the original depositional contact on Tertiary sediments or the basement Paleozoic sediments. However, in other places the lower contact is with post-mineral dacite intrusions. The lower contact has been offset by a series of west dipping faults. The upper contact is generally defined by post-mineral dacite intrusions or locally the next stratigraphic unit. Because the dacites are post-mineral they are not offset by the same faults as the lower contact. There are some minor internal dacite intervals. These dacites are a different composition to the post mineral intrusions and they are generally mineralized so they have been included in the volume between the upper and lower contacts. The upper and lower contacts have been extended north and south to the limits that should be modeled. Consequently, the volume to model should be defined by the upper and lower contacts together with the topography and then the ends should just be clipped with vertical planes which coincide with the edge of the triangulated surfaces.

Figure 14-38 Isometric View of Jolly Jane Looking Northeasterly Showing Mineralization Solid in Red, Drill holes in Green and Surface Topography in Grey.

 

14.4.1 Data Analysis Jolly Jane

Drill holes were compared to the geologic solid and the assays were back-tagged with a mineralized code if inside the solid. The sample statistics are tabulated below (Table 14-28).

Table 14-28 Summary of Assay Statistics for Jolly Jane Mineralized Solid

	Inside Solid		Outside Solid
	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)
Number of Samples	4,585	4,417	4,549	4,390
Mean Grade	0.143	0.361	0.033	0.288
Standard Deviation	0.164	0.378	0.075	0.639
Minimum Value	0.001	0.005	0.001	0.005
Maximum Value	1.45	4.46	0.93	17.75
Coefficient of Variation	1.14	1.04	2.29	2.22

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The grade distribution for gold was evaluated using a lognormal cumulative frequency plot for samples within the mineralization solid. Five overlapping lognormal populations made up the gold distribution for the mineralized zone. The highest-grade population, with a mean value of 0.85 g/t Au, represented 0.73% of the data or 33 samples and was not considered erratic high grade. A cap level was chosen at two standard deviations above the mean of this highest-grade population. A cap value of 1.78 g/t Au was used and no assays required capping.

A similar exercise was completed for silver within the mineralization solid. No silver assays within the mineralization zone required capping.

For assays outside the mineralization solid a total of 55 assays were capped at 0.35 g/t Au and 13 assays were capped at 4.4 g/t Ag.

14.4.2 Composites Jolly Jane

Drill holes at Jolly Jane were compared to the mineralization solid and the points at which each hole entered and left the solid were recorded. Uniform down hole composites, 5 metres in length, were formed and made to honor the solid boundaries. Intervals less than ½ the composite length at the solid boundaries were joined with adjoining samples to produce a composites file of uniform support, 5± 2.5 metres in length. The statistics for 5 metres composites are summarized below (Table 14-29). A similar exercise was completed for samples outside the solid.

Table 14-29 Summary of 5 M Composite Statistics for Mineralization Solid Jolly Jane

	Mineralized Solid		Outside Solid
	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)
Number of Samples	1,441	1,408	1,414	1,396
Mean Grade	0.144	0.37	0.031	0.31
Standard Deviation	0.152	0.41	0.057	0.55
Minimum Value	0.001	0.005	0.001	0.005
Maximum Value	1.28	4.46	0.35	4.40
Coefficient of Variation	1.06	1.10	1.85	1.78

14.4.3 Variography Jolly Jane

Pairwise relative semivariograms were used to model the gold continuity at Jolly Jane. The direction of longest continuity for gold in the horizontal plane was along azimuth 0° dipping -30°. In the plane perpendicular to this the longest continuity was along azimuth 90^o dipping -40^o. Nested spherical models were fit to all directions. The nugget-to-sill ratio of 16% for Au and 11% for Ag were very good. For Au and Ag in waste, isotropic spherical models were produced. The parameters are tabulated below.

The parameters are tabulated in Table 14-30.

Table 14-30 Summary of Jolly Jane Gold and Silver Semi-variogram Parameters

Domain	Azimuth	Dip	Co	C1	C2	Short Range (m)	Long Range (m)
Mineralized Solid Au	0o	-30o	0.10	0.10	0.43	20	120
	90o	-20o	0.10	0.10	0.43	30	110
	270o	-70o	0.10	0.10	0.43	20	50
Mineralized Solid Ag	0o	-30o	0.05	0.10	0.30	30	120
	90o	-40o	0.05	0.10	0.30	40	80
	270o	-50o	0.05	0.10	0.30	15	60
Waste Au	Omni Directional		0.15	0.15	0.40	25	90
Waste Ag	Omni Directional		0.10	0.10	0.35	30	100

14.4.4 Bulk Density Jolly Jane

During the 2010 drill campaign on the NBP, a total of 102 samples of RC chips were sent to ALS Minerals for specific gravity measurements by pycnometer (method OA-GRA08b). The average specific gravity from 46 samples within the oxidized Tuff units from mineralization zones drilled in 2010 was 2.60.

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During the 2012 drill campaign 74 specific gravity determinations were made from drill core which is far more representative than RC Chips, as porosity is included. Of these samples 59 were within the Crater Flat Tuff unit which hosts the mineralization at Jolly Jane. The average specific gravity from these samples, listed in Table 14-31, was 2.34.

Table 14-31 Specific Gravity Determinations for Tuff Units-Jolly Jane

HoleID	SampleID	From_m	To_m	SG	StratUnit1
NB-12-130	M610829	17.68	20.73	2.48	fault zone
NB-12-130	M610836	35.97	39.01	2.43	fault zone
NB-12-130	M610838	42.06	45.11	2.44	fault zone
NB-12-130	M610851	79.56	82.76	1.93	fault zone
NB-12-130	M610855	89.31	92.50	1.96	fault zone
NB-12-131	M612269	11.44	13.50	2.38	fault zone
NB-12-131	M612279	36.10	39.01	2.21	fault zone
NB-12-131	M612283	44.40	46.33	2.23	fault zone
Average				2.26	Fault zones
NB-12-131	M612265	0.00	3.05	2.36	Crater Flat Tuff
NB-12-131	M612266	3.05	6.27	2.37	Crater Flat Tuff
NB-12-131	M612267	6.27	9.40	2.41	Crater Flat Tuff
NB-12-130	M610859	101.80	105.55	2.44	lower Crater Flat Tuff
NB-12-130	M610860	105.55	109.42	2.43	lower Crater Flat Tuff
NB-12-130	M610824	2.20	5.38	2.34	middle Crater Flat Tuff
NB-12-130	M610825	5.38	8.45	2.43	middle Crater Flat Tuff
NB-12-130	M610826	8.45	11.58	2.40	middle Crater Flat Tuff
NB-12-130	M610827	11.58	14.63	2.39	middle Crater Flat Tuff
NB-12-130	M610828	14.63	17.68	2.41	middle Crater Flat Tuff
NB-12-130	M610830	20.73	23.77	2.38	middle Crater Flat Tuff
NB-12-130	M610831	23.77	26.82	2.47	middle Crater Flat Tuff
NB-12-130	M610832	26.82	29.87	2.41	middle Crater Flat Tuff
NB-12-130	M610834	29.87	32.92	2.40	middle Crater Flat Tuff
NB-12-130	M610835	32.92	35.97	2.47	middle Crater Flat Tuff
NB-12-130	M610837	39.01	42.06	2.39	middle Crater Flat Tuff
NB-12-130	M610839	45.11	48.11	2.43	middle Crater Flat Tuff
NB-12-130	M610840	48.11	51.21	2.50	middle Crater Flat Tuff
NB-12-130	M610841	51.21	54.25	2.52	middle Crater Flat Tuff
NB-12-130	M610842	54.25	57.30	2.47	middle Crater Flat Tuff
NB-12-130	M610844	57.30	60.35	2.45	middle Crater Flat Tuff
NB-12-130	M610845	60.35	63.74	2.43	middle Crater Flat Tuff
NB-12-130	M610846	63.74	67.18	2.33	middle Crater Flat Tuff
NB-12-130	M610847	67.18	70.30	2.32	middle Crater Flat Tuff
NB-12-130	M610848	70.30	73.34	2.34	middle Crater Flat Tuff
NB-12-130	M610849	73.34	76.48	2.01	middle Crater Flat Tuff
NB-12-130	M610850	76.48	79.56	2.06	middle Crater Flat Tuff
NB-12-130	M610852	82.76	86.26	1.98	middle Crater Flat Tuff
NB-12-130	M610854	86.26	89.31	1.95	middle Crater Flat Tuff
NB-12-130	M610856	92.50	95.52	2.17	middle Crater Flat Tuff
NB-12-130	M610857	95.52	98.63	2.30	middle Crater Flat Tuff
NB-12-130	M610858	98.63	101.80	2.27	middle Crater Flat Tuff
NB-12-131	M612276	27.53	30.29	2.34	middle Crater Flat Tuff

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HoleID	SampleID	From_m	To_m	SG	StratUnit1
NB-12-131	M612277	30.29	33.22	2.39	middle Crater Flat Tuff
NB-12-131	M612278	33.22	36.10	2.32	middle Crater Flat Tuff
NB-12-131	M612280	39.01	42.06	2.35	middle Crater Flat Tuff
NB-12-131	M612282	42.06	44.40	2.31	middle Crater Flat Tuff
NB-12-131	M612284	46.33	49.68	2.36	middle Crater Flat Tuff
NB-12-131	M610862	49.68	52.73	2.28	middle Crater Flat Tuff
NB-12-131	M610863	52.73	56.66	2.21	middle Crater Flat Tuff
NB-12-131	M610864	56.66	60.35	2.34	middle Crater Flat Tuff
NB-12-131	M610865	60.35	64.10	2.38	middle Crater Flat Tuff
NB-12-131	M610866	64.10	67.97	2.40	middle Crater Flat Tuff
NB-12-131	M610867	67.97	71.60	2.49	middle Crater Flat Tuff
NB-12-131	M610868	71.60	75.24	2.45	middle Crater Flat Tuff
NB-12-131	M610869	75.24	78.64	2.37	middle Crater Flat Tuff
NB-12-131	M610870	78.64	81.69	2.42	middle Crater Flat Tuff
NB-12-131	M610872	81.69	84.73	2.22	middle Crater Flat Tuff
NB-12-131	M610873	84.73	87.78	2.30	middle Crater Flat Tuff
NB-12-131	M610874	87.78	90.83	2.29	middle Crater Flat Tuff
NB-12-131	M610875	90.83	93.88	2.31	middle Crater Flat Tuff
NB-12-131	M610876	93.88	97.88	2.22	middle Crater Flat Tuff
NB-12-131	M610877	97.88	101.72	2.50	middle Crater Flat Tuff
NB-12-131	M610878	101.72	105.58	2.47	middle Crater Flat Tuff
NB-12-131	M610879	105.58	109.32	2.43	middle Crater Flat Tuff
NB-12-131	M610880	109.32	112.34	2.22	middle Crater Flat Tuff
NB-12-131	M610882	112.34	116.30	2.13	middle Crater Flat Tuff
NB-12-131	M610883	116.30	119.40	2.34	middle Crater Flat Tuff
NB-12-131	M610884	119.40	120.94	2.10	middle Crater Flat Tuff
Average				2.34	Crater Flat Tuff
NB-12-131	M612268	9.40	11.44	2.16	dacite breccia
NB-12-131	M612270	13.50	16.68	2.33	dacite breccia
NB-12-131	M612272	16.68	19.80	2.51	dacite breccia
NB-12-131	M612273	19.80	23.49	2.48	dacite breccia
NB-12-131	M612274	23.49	26.92	2.45	dacite breccia
NB-12-131	M612275	26.92	27.53	2.25	dacite breccia
NB-12-131	M610885	120.94	122.40	2.32	dacite breccia
Average				2.36	Dacite Breccia

For Jolly Jane, a specific gravity of 2.34 was used to determine tonnage.

14.4.5 Grade Estimation

Grades for gold were interpolated by ordinary kriging into all blocks, with some percentage within the Jolly Jane mineralization solid. Kriging was completed in a series of passes with the dimensions and orientation of the search ellipse for each pass tied to the semi-variogram for gold. The first pass used dimensions equal to ¼ of the semi-variogram range in the three principal directions. If a minimum of 4 composites were found within this ellipse centered on a block, the block was estimated. For blocks not estimated, the search ellipse was expanded to ½ the semi-variogram range. Again, a minimum of 4 composites within the search ellipse were required to estimate any given block. A third pass using the full semi-variogram range was completed for blocks not estimated during the first two passes. Finally, a fourth pass using roughly twice the range was completed. In all cases if more than 12 composites were located in any search, the closest 12 were used. A maximum of three composites from any individual hole were allowed in all passes. The search parameters for the Kriging procedure are tabulated below (Table 14-32).

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A similar procedure was used to estimate silver with the pass four ellipse expanded to the pass four gold search to insure all blocks estimated for gold had a silver value.

Volumes for each block estimated were determined by multiplying the block volume by the percentage of block below topography and within the solid. The tonnage was determined by multiplying the block volume by the S.G. (2.34).

Table 14-32 Summary of Kriging Search Parameters for Jolly Jane

Domain	Pass	Number Estimated	Az/Dip	Dist. (m)	Az/Dip	Dist. (m)	Az/Dip	Dist. (m)
Mineralized Solid Au	1	3,145	0 / -30	30.0	270 / -70	12.5	90 / -20	27.5
	2	35,918	0 / -30	60.0	270 / -70	25.0	90 / -20	55.0
	3	34,632	0 / -30	120.0	270 / -70	50.0	90 / -20	110.0
	4	7,968	0 / -30	240.0	270 / -70	100.0	90 / -20	220.0
Mineralized Solid Ag	1	1,473	0 / -30	30.0	270 / -50	15.0	90 / -40	20.0
	2	28,358	0 / -30	60.0	270 / -50	30.0	90 / -40	40.0
	3	38,531	0 / -30	120.0	270 / -50	60.0	90 / -40	80.0
	4	13,301	0 / -30	240.0	270 / -50	120.0	90 / -40	160.0

14.5 Mayflower

14.5.1 Mayflower Geologic Model

The supplied data for the Mayflower Estimation consisted of 104 drill holes totaling 17,228 m. Of these 104 supplied holes, 79 of them penetrated the mineralization solid and were used for this estimate. A total of 10,950 samples were assayed for gold and 10,492 for silver.

The distribution of mineralization at the Mayflower prospect is complicated. It is clear that mineralization is controlled by a complex fracture network without the clear definition of a simple central vein system. In order to define a volume to be included in the estimation model, a combination of the alteration, trace element geochemistry and gold mineralization were taken into account. The underlying premise for the model was that the form of the zone should be roughly tabular following the main fault zone.

The Mayflower geochemical data show that there is a clear correlation between the higher-grade gold mineralization, potassium feldspar alteration and arsenic mineralization. The first step in building the model was to define the distribution of potassium feldspar alteration using the molar K/Al and Na/Al ratios. The next step was to look at the statistical distribution of arsenic and establish that the mineralized population begins at approximately 10 ppm arsenic in unit Trt2 and 25 ppm in all other lithologies. Finally, the cumulative frequency distribution of gold indicates that 0.04 g/t is the lower limit of the main mineralized population. The “Mayflower Zone” was then defined as continuous drilled intervals which had K-feldspar alteration, high As and Au>0.04 g/t.

The drill hole data with the “Mayflower Zone” designation was loaded into a 3-D view and the “Hanging wall” and “Footwall” was constructed using the top and bottom contacts of the “Mayflower Zone”. Once this was done, the surface was extended and modified by integrating the Barrick drilling data into the model. The same 0.04 g/t cutoff was used to define zones of mineralization in the Barrick holes. The resulting model surfaces were exported to Surpac^TM where the final closed volume was constructed. The base of the modeled volume was arbitrarily cut off at an elevation of 1000 metres.

The overall form of the Mayflower Zone is narrow at depth and widens as it approaches the surface, a configuration that is quite common in near surface fault systems (Figures 14-39 and 14-40). The hanging wall is steeper and more planar than the footwall.

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Figure 14-39 Mayflower Model Looking NW with the Mineralization Solid in Red and Topography in Grey.

 

Figure 14-40 Isometric View Looking W of the Mayflower Geologic Solid in Red with Surface Topography Shown in Grey.

 

14.5.2 Data Analysis

Using the interpreted geologic solid, the assays were back tagged with a mineralization code if inside the mineralized solid. The sample statistics are tabulated below (Table 14-33).

Table 14-33 Summary of Assay Statistics for Mineralization Solid and Waste- Mayflower.

	Inside Mineralization Solid		Outside Mineralization Solid
	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)
Number of Samples	5,408	5,408	5,629	5,171
Mean Grade	0.233	0.274	0.027	0.222
Standard Deviation	1.153	1.147	0.062	0.923
Minimum Value	0.001	0.001	0.001	0.001
Maximum Value	41.50	75.90	1.41	27.43
Coefficient of Variation	4.94	4.18	2.32	4.15

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The grade distributions for gold and silver were evaluated using lognormal cumulative frequency plots for samples within both the mineralization solid and the surrounding waste. In all cases, multiple overlapping lognormal populations were seen. Within the mineralization zone for gold, there were five overlapping populations (Figure 14-41). The highest population with average grades of 37.2 g/t Au, represents 0.07% of the data, and was considered erratic high grade. A cap consisting of two standard deviations above the mean of population of two, a value of 22 g/t was used to cap five gold assays.

Figure 14-41 Lognormal Cumulative Frequency Plot for Au within the Mineralization Solid-Mayflower.

 

A similar exercise completed on silver, resulted in 4 assays capped at 8.0 g/t. Within waste, gold showed 6 overlapping lognormal populations with the upper two populations averaging 1.26 g/t and 0.85 g/t representing a combined 0.23% of the data, considered erratic. A cap level of 0.55 g/t or two standard deviations above the mean of population 3 was used to cap 13 gold assays at 0.55 g/t Au. For silver in waste a cap level of 3.8 g/t Ag was used to cap 27 assays. The results of capping reduce the mean grade slightly and significantly reduce the standard deviation and as a result the coefficient of variation in all variables. The capped assay statistics are listed in Table 14-34.

Table 14-34 Summary of Capped Assay Statistics for Mineralization Solid and Waste-Mayflower

	Inside Mineralized Solid		Outside Mineralized Solid
	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)
Number of Samples	5,408	5,408	5,629	5,171
Mean Grade	0.224	0.257	0.026	0.188
Standard Deviation	0.884	0.385	0.048	0.433
Minimum Value	0.001	0.001	0.001	0.001
Maximum Value	22.00	8.00	0.55	3.80
Coefficient of Variation	3.95	1.50	1.85	2.30

14.5.3 Composites

Drill holes at Mayflower were compared to the mineralization solid and the points each hole entered and left the solid were recorded. Uniform down hole composites, 5 metres in length, were formed and made to honor the solid boundaries. Intervals less than ½ the composite length at the solid boundaries were joined with adjoining samples to produce a composites file of uniform support, 5± 2.5 metres in length. The statistics for 5 metre composites are summarized below (Table 14-35).

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Table 14-35 Summary of 5 M Composite Statistics for Mineralization Solid-Mayflower

	Inside Mineralized Solid
	Au (g/t)	Ag (g/t)
Number of Samples	1,706	1,706
Mean Grade	0.222	0.261
Standard Deviation	0.617	0.296
Minimum Value	0.001	0.001
Maximum Value	13.60	4.74
Coefficient of Variation	2.78	1.14

14.5.4 Variography

Pairwise relative semi-variograms were produced from composites within the mineralization solid for both gold and silver. The semi-variograms were produced along strike, down dip and across dip within the mineralized lens. Nested spherical models were fit to the three directions of anisotropy. The nugget-to-sill ratio was 32% for gold and 24% for silver. The model parameters are summarized below (Table 14-36).

Table 14-36 Summary of Semi-variogram Parameters-Mayflower

Variable	Azimuth	Dip	Co	C1	C2	Short Range (m)	Long Range (m)
Au	315o	0o	0.30	0.40	0.25	50	160
	225o	-70o	0.30	0.40	0.25	40	100
	45o	-20o	0.30	0.40	0.25	30	45
Ag	315o	0o	0.13	0.20	0.22	20	60
	225o	-70o	0.13	0.20	0.22	30	100
	45o	-20o	0.13	0.20	0.22	10	40
Min Ind	315o	0o	0.25	0.40	0.25	30	150
	225o	-70o	0.25	0.40	0.25	10	60
	45o	-20o	0.25	0.40	0.25	15	60
Vein Ind	315o	0o	1.20	0.45	0.35	15	100
	225o	-70o	1.20	0.45	0.35	30	80
	45o	-20o	1.20	0.45	0.35	10	64

14.5.5 Bulk Density

During the 2012 drill program, a total of 271 specific gravity measurements were made from drill core using the weight in air/weight in water method. These determinations came from holes NB-12-132, 133, 140, 141, 142 and 143. The results can be sorted by lithology and by gold grade. While there is a range of specific gravities for the various lithologies sampled, lithology has not been modeled so it is not of any use in assigning density to estimated blocks. There does, however, appear to be a reasonable correlation between gold grade and specific gravity as shown in Table 14-37 with higher densities associated with higher gold grades. As a result, the specific gravity assigned to each block in the model is based on the estimated gold grade as tabulated below in Table 14-38.

Table 14-37 Specific Gravities Sorted by Lithology - Mayflower

Lithology	Number	Min. SG	Max. SG	Average SG
Cz	1			2.58
Fault	6	1.79	2.45	2.20
Tcm	1			2.26
Tdfh	214	2.02	2.63	2.28
Tdfm	4	2.35	2.40	2.38
eTpbx	2			2.38
Trt2	43	1.59	2.56	2.19
Total	271	1.59	2.63	2.27

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Table 14-38 Specific Gravities Sorted by Gold Grade - Mayflower

Gold Grade (g/t)	Average Au (g/t)	Number	Average SG
> 0.0 < 0.1	0.029	135	2.22
≥ 0.1 < 0.5	0.243	83	2.30
≥ 0.5 < 1.0	0.688	36	2.33
≥ 1.0	2.350	17	2.36
Total		271	2.27

14.5.6 Block Model

A block model with blocks 10 x 10 x 5 meters in dimension was superimposed over the mineralization solid. The model was rotated 45^o to better fit the solid. The block model origin was as follows:

Lower left corner of model

518838.0 E	Column width – 10 m	30 columns
4093900.0 N	Row width – 10 m	81 rows

Top of Model

1395 Elevation Level width – 5 m 90 levels

Rotation 45^o counter clockwise

Within each block, the percentage below surface topography and within the mineralization solid was recorded (Figure 14-42).

Figure 14-42 Isometric View Looking NNW Showing Block Model in White and Drill holes in Magenta-Mayflower.

 

14.5.7 Grade Interpolation

Grades for gold and silver were interpolated into all blocks, with some percentage within the mineralization solid, by Ordinary Kriging. Kriging was completed in a series of passes with the dimensions and orientation of the search ellipse for each pass tied to the semi-variogram. The first pass used dimensions equal to ¼ of the semi-variogram range in the three principal directions. If a minimum of 4 composites were found within this ellipse centered on a block, the block was estimated. For blocks not estimated, the search ellipse was expanded to ½ the semi-variogram range. Again, a minimum of 4 composites within the search ellipse were required to estimate any given block. A third pass using the full semi-variogram range was completed for blocks not estimated during the first two passes. Finally, a fourth pass using roughly twice the range was completed. This pass was modified to use the maximum range for both gold and silver to ensure all blocks were estimated for both variables. In all cases, if more than 12 composites were located in any search, the closest 12 were used. In all cases, the maximum number of composites allowed, from a single drill hole, was set to three to ensure all blocks were estimated by a minimum of two drill holes. The search parameters for the Kriging procedure are tabulated below (Table 14-39). Volumes for each block estimated were determined by multiplying the block volume by the percentage of block below topography and within the solid. The tonnage was determined by multiplying the block volume by the block specific gravity.

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In a similar manner, the mineralization indicator and the vein indicator were kriged into all estimated blocks.

Table 14-39 Summary of Kriging Search Parameters - Mayflower

Variable	Pass	Number Estimated	Az/Dip	Dist. (m)	Az/Dip	Dist. (m)	Az/Dip	Dist. (m)
Au	1	12,486	315/0	40.0	225/-70	25.0	45/-20	11.25
	2	13,510	315/0	80.0	225/-70	50.0	45/-20	22.5
	3	7,417	315/0	160.0	225/-70	100.0	45/-20	45.0
	4	6,412	315/0	320.0	225/-70	200.0	45/-20	90.0
Ag	1	3,089	315/0	15.0	225/-70	25.0	45/-20	10.0
	2	14,774	315/0	30.0	225/-70	50.0	45/-20	20.0
	3	12,562	315/0	60.0	225/-70	100.0	45/-20	40.0
	4	9,400	315/0	320.0	225/-70	200.0	45/-20	90.0

14.6 Mineral Resource Classification

Mineral Resources are characterized according to CIM Definitions Standards which are incorporated by reference in NI 43-101. Mineralization at MLP and NBP has been categorized as Inferred Mineral Resources, Indicated Resources and Measured Resources, based upon increasing levels of confidence in various physical characteristics of each Project. Drill hole spacing, search neighborhoods, metallurgical characterization, geological confidence, kriging variance and many other factors are used to give the author confidence in the Mineral Resource estimate for these projects. Appropriate classification criteria should aim to integrate all these concepts. RDA is satisfied that the geological modelling for the Corvus Gold Projects honors their geological information and knowledge of each of the deposits that they control. The location of the samples and the assay data are sufficiently reliable to support resource evaluation.

Classification parameters for Jolly Jane and Mayflower are described in those project technical reports. RDA has chosen to use kriging variance versus distance as the approach to determining the Mineral Resource classification at Sierra Blanca. At MLP a combination of distance and number of samples was used in the classification of resources which were estimated using Inverse Distance techniques. Kriging variance was used for the classification of mineral resources which were derived using ordinary kriging.

14.6.1 Mother Lode

Mineralization within the high-grade feeder was estimated using inverse distance squared interpolation. The linear distance to each sample composite and drill hole has been stored in each model block. For the rest of the deposit, kriged mineralization was evaluated and the kriging variance was stored in each model block for that portion of the mineral deposit.

High-grade feeder classification parameters were:

· Measured mineralization required drilling to be within 15 meters of a model block;

· Indicated Mineralization required drilling to be between 15 and 50 meters;

· Inferred mineralization was any mineralization up to 160 meters.

Kriged classification parameters were:

· Measured mineralization required the kriging variance to be less than 0.25;

· Indicated Mineralization required the kriging variance to be between 0.25 and 0.60;

· Inferred mineralization required the kriging variance to be between 0.60 and 1.145;

· Mineralization with negative kriging variances or variances greater than 1.145 we not considered in the mineral resource estimate.

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Resource classification is subjective across the mining industry but it is good to check visually. Figure 14-43 shows the typical section through the deposit. The classification has a valid “look” to it as well. RDA has confidence as defined by CIM in the Mother Lode classification scheme.

Figure 14-43 Mineral Classification at Mother Lode Section 4084410N

 

14.6.2 Sierra Blanca

Disseminated mineralization at Sierra Blanca has been categorized according to kriging variance. Kriging is an interpolation technique that minimizes the squared error between the estimated value and the unknown true value. The resulting error variance, stored as kriging variance in the block model, is dependent only on the estimation location, the location of samples used in the estimate and the variogram. Kriging variance thresholds were used to differentiate between measured, indicated and inferred categories. The advantage of using kriging variance is the consideration of the spatial structure of the gold and silver grades and redundancy between the samples.

Drill hole NB-16-309 was flagged and determined to have the highest kriging variance in the deposit at 0.406. The highest kriging variance estimated in the deposit is 1.39. A graph of kriging variance versus distance was evaluated. A variance of 0.406 correlated well with a distance of approximately 25 metres. No extraneous holes or islands of mineralization have been interpreted at this threshold. The kriging variance of 0.406 was chosen to be the boundary between measured and indicated classification. (Figure 14-44)

Indicated Mineral Resources has been interpreted for the kriging variance versus distance graph at a threshold of 0.768. This correlates with the near vertical departure of the lower end of the point cloud with the majority of the mineralization being within 50 metres of a drill hole. There are a few model blocks between 50 and 105 metres that fall into the indicated category but the geological confidence is very high for these few blocks. Globally, any model block greater than 0.768 is neither supported by minimal samples nor lays greater than 50 metres from a drill hole. (Figure 14-44)

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Figure 14-44 Kriging Variance vs. Distance, Sierra Blanca Classification-metres

 

14.6.3 Jolly Jane

For the Jolly Jane Deposit, blocks estimated in Pass 1 or Pass 2 using up to one half the variogram ranges for the search ellipse, were classified as Indicated. All other blocks were classified as Inferred. All blocks at Jolly Jane are oxidized.

14.6.4 Mayflower

For the Mayflower Deposit, Blocks estimated during pass 1 and 2 using search ellipses with dimensions up to ½ the variogram range were classified as Indicated. All other blocks were classified as Inferred. All blocks at Mayflower are oxidized.

14.7 Mineral Resources

14.7.1 Pit Constraining Parameters

Mineral Resources must demonstrate reasonable prospects for eventual economic extraction in accordance with the CIM Definition Standards. Pit constraining limits were determined using the Lerchs-Grossman© economic algorithm which constructs lists of related blocks that should or should not be mined. The final list defines a surface pit shell that has the highest possible total value, while honoring the required surface mine slope and economic parameters.

Mineral Resources are not Mineral Reserves and do not demonstrate economic viability. There is no certainty that all or any part of the Mineral Resource will be converted to mineral reserves. RDA knows of no environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that may materially affect the Mineral Resource estimate in this report. Quantity and grade are estimates and are rounded to reflect the fact that the resource estimate is an approximation.

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Economic parameters used in the analysis are listed in Table 14-40 and are based on the following processes:

· Gravity – CIL mill processing of YellowJacket mineralization

· Heap Leach processing of disseminated mineralization

· Ambient Alkaline Oxidation processing of Sierra Blanca sulphide mineralization

· Pressure Oxidation of Mother Lode Mill mineralization

Table 14-40 Pit Constraining Parameters Used for the Corvus Gold September 18, 2018 Mineral Resource Update

Parameter	Unit	Mayflower	Jolly Jane	Sierra Blanca	YellowJacket Vein/Vein Stockwork	YellowJacket Sulphide	Mother Lode Sulphide	Mother Lode Oxide
Mining Cost	US$/tonne	1.64	1.42	1.54	1.54	1.54	1.40	1.40
Au Cut-Off	g/tonne	0.10	0.10	0.10	0.35	0.71	0.63	0.06
Processing Cost	US$/tonne	1.72	1.72	1.15	11.84	25.6	19.50	1.20
Au Recovery	%	70	72	73.8	86.6	91	80.0	74.0
Ag Recovery	%	8	8	6.3	74.3	57.2	NA	NA
Administrative Cost	US$/tonne	0.5	0.5	0.45	0.45	0.45	0.50	0.50
Refining & Sales	US$/oz	5.00	5.00	5.00	5.00	5.00	5.00	5.00
Au Selling Price	US$/oz	1,250	1,250	1,250	1,250	1,250	1,250	1,250
Slope Angle	Degrees	50	50	50	50	50	60	60

The parameters listed in Table 14-40 define a realistic basis to estimate the Mineral Resources for the Project and are representative of similar mining operations throughout Nevada. The Mineral Resource has been limited to mineralized material that occurs within the pit shells and which could be scheduled to be processed based on a defined cut-off grade. All other material within the defined pit shells was characterized as non-mineralized material.

14.8 NB-MLP MIneral Resources

14.8.1 Sierra Blanca Estimated Mineral Resources

Resource tables for the individual NBP mineralization areas are presented in the following sections and are accumulated according by separate classification.

The estimate considers three processing methods; 1) mill processing of oxide and gravity-separable gold and silver; 2) heap leach processing of oxide gold and silver; and 3) mill processing of sulphide gold and silver. Mineralization within the YellowJacket vein, including the surrounding stockwork veining, associated with the YellowJacket structural corridor, is situated within a well-defined zone that holds together at higher cutoff grades within resource constraining pit. Since this part of the mineral deposit contains the highest grades, it is reasonable to expect this part of the Sierra Blanca deposit would be economically extracted prior to economically extracting lower grade, heap-leachable mineralization. This portion of the mineral deposit is referred to as Phase I. The limits of the constraining pit were determined by assuming only mill mineralization and higher grade disseminated mineralization would be processed. Mineralization is reported at higher cutoff grades for this portion of the deposit. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.

Phase I NBP Mineral Resources are estimated according to the following cutoff grades and are reported in Table 14-41:

· Phase I Mill Resources (Oxide, Vein and Stockwork Mineralization) - >= 0.35 g/t Au

· Phase I Mill Resources (Sulphide Disseminated Mineralization) - >= 0.71 g/t Au

· Phase I Heap Leach Resources (Oxide Vein Mineralization) – (>= 0.15 and <0.35) g/t Au

· Phase I Heap Leach Resources (Oxide Disseminated Mineralization) - >= 0.15 g/t Au

Silver is assumed to be recovered as a byproduct of processing and is reported from model blocks that meet the gold cutoff grade criteria.

Mineral Resources which meet the reasonable prospects of eventual economic extraction, based upon the costs and parameters in Table 14-40, are reported at the breakeven cutoff grades for the mineralization and processing methods described. There is no assurance that any or all of the Mineral Resources will be converted to mineral reserves.

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Phase II NBP Mineral Resources are estimated at the following cutoff grades:

· Phase II Mill Resources (Oxide, Vein and Stockwork Mineralization) - >= 0.35 g/t Au

· Phase II Mill Resources (Sulphide Disseminated Mineralization) - >= 0.71 g/t Au

· Phase II Heap Leach Resources (Oxide Vein Mineralization) – (>= 0.10 and <0.35) g/t Au

· Phase II Heap Leach Resources (Oxide Disseminated Mineralization) - >= 0.10 g/t Au

Phase II Mineral Resources are pit constrained and are reported exclusive of Phase I Mineral Resources in Table 14-41.

Table 14-41 North Bullfrog Project Pit Constrained Mineral Resource Estimate. (QP: RDA, Scott Wilson; Effective September 18, 2018)

	Classification	Tonnes (k)	Au (g/t)	Ag (g/t)	Contained Au (000’s)	Contained Ag (000’s)
Phase I	Measured	10,415	1.08	7.59	362	2,540
	Indicated	24,557	0.69	3.70	542	5,459
	Inferred	5,908	0.31	0.74	59	140
Phase II	Measured	10,129	0.26	1.04	84	338
	Indicated	113,009	0.21	0.61	771	2,227
	Inferred	58,887	0.19	0.48	367	902
Total Mineral Resources	Measured	20,544	0.68	4.36	446	2,878
	Indicated	137,566	0.30	1.16	1,314	5,146
	Inferred	64,785	0.20	0.50	426	1,042

Mineral Resources are not mineral reserves and do not demonstrate economic viability. There is no certainty that all or any part of the Mineral Resource will be converted to mineral reserves. Quantity and grade are estimates and are rounded to reflect the fact that the resource estimate is an approximation.

14.8.2 Mother Lode Mineral Resources

Mother Lode Mineral Resources are considered using two processing methods; 1) pressure oxidation processing of sulphide gold and 2) heap leach processing of oxide gold. Mineral Resources for the estimation of sulphide are reported at an Au cutoff grade of 0.63 g/t Au. Oxide Mineral Resources are estimated at 0.06 g/t Au. The results presented Table 14-42 show the resources separated by process method for Measured, Indicated and Inferred resources for Mother Lode. Resources are pit constrained and reported at cutoff grades based on processing methods.

Table 14-42 Mother Lode Mineral Resources (QP: RDA, Scott Wilson; Effective September 18, 2018)

	Milling Cutoff Grade 0.63 Au g/t			Heap Leach Cutoff Grade 0.06 Au g/t			Total (Milling & Heap Leach)
Resource Classification	Tonnes (x1,000)	Au g/t	Au Ounces (x1,000)	Tonnes (x1,000)	Au g/t	Au Ounces (x1,000)	Tonnes (x1,000)	Au g/t	Au Ounces (x1,000)
Measured	3,292	1.41	149	20,035	0.29	185	23,327	0.45	334
Indicated	9,934	1.83	583	20,123	0.37	242	30,057	0.85	825
Total M & I	13,226	1.72	733	40,158	0.33	427	53,383	0.68	1,159
Inferred	2,168	1.60	112	14,073	0.29	129	16,241	0.46	241

Mineral Resources are not mineral reserves and do not demonstrate economic viability. There is no certainty that all or any part of the Mineral Resource will be converted to mineral reserves. Quantity and grade are estimates and are rounded to reflect the fact that the resource estimate is an approximation.

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14.9 NBP Mineral Resources By Deposit

14.9.1 Sierra Blanca Mineral Resources

The results presented Table 14-43 and 14-44, show the resources separated by process method for Measured, Indicated and Inferred resources for Sierra Blanca for Phase I and II, respectively. Resources are pit constrained and reported at cutoff grades based on processing methods.

Table 14-43 Sierra Blanca Mineral Resources Phase I (QP: RDA, Scott Wilson; Effective September 18, 2018)

	Milling			Heap Leach			Sulphide
	0.35 g/t Au cut off			0.15 g/t Au cutoff			0.71 g/t cutoff
Classification	Tonnes (Kt)	Au g/t	Ag g/t	Tonnes (Kt)	Au g/t	Ag g/t	Tonnes (Kt)	Au g/t	Ag g/t	Au Ounces (x1,000)	Ag  Ounces (x1,000)
Measured	4,465	1.87	13.97	5,194	0.37	2.42	756	1.32	5.35	362	2,540
Indicated	4,445	1.80	13.40	13,736	0.34	1.63	1,137	1.56	5.83	464	2,850
Inferred	34	1.83	19.21	5,831	0.30	0.59	15	2.07	16.59	59	140

Mineral Resources are not mineral reserves and do not demonstrate economic viability. There is no certainty that all or any part of the Mineral Resource will be converted to mineral reserves. Quantity and grade are estimates and are rounded to reflect the fact that the resource estimate is an approximation.

Table 14-44 Sierra Blanca Mineral Resources Phase II (QP: RDA, Scott Wilson; Effective September 18, 2018)

	Milling			Heap Leach			Sulphide
	0.35 g/t Au cut off			0.15 g/t Au cutoff			0.71 g/t cutoff
Classification	Tonnes (Kt)	Au g/t	Ag g/t	Tonnes (Kt)	Au g/t	Ag g/t	Tonnes (Kt)	Au g/t	Ag g/t	Au Ounces (x1,000)	Ag  Ounces (x1,000)
Measured	397	0.78	4.07	9,331	0.19	0.85	401	1.24	2.48	84	338
Indicated	1,331	0.89	4.16	91,525	0.18	0.58	1,402	1.18	1.82	625	1,973
Inferred	6	0	5.18	50,939	0.19	0.46	61	1.53	2.04	314	760

Mineral Resources are not mineral reserves and do not demonstrate economic viability. There is no certainty that all or any part of the Mineral Resource will be converted to mineral reserves. Quantity and grade are estimates and are rounded to reflect the fact that the resource estimate is an approximation.

14.9.2 Jolly Jane Mineral Resources

The results presented in Table 14-45, show the Indicated and Inferred mineralization inventory for Jolly Jane. Resources are separated by classification. All material is assumed to be processed by heap leaching methods. Resources are pit constrained and reported at a gold cut off grade of 0.10 g/t Au.

Table 14-45 Jolly Jane Mineral Resources-Phase II (QP: RDA, Scott Wilson; Effective September 18, 2018)

Classification	Tonnes (kt)	Au g/t	Ag g/t	Contained Au (x1,000)	Contained Ag (x1,000)
Indicated	18,571	0.24	0.42	146	254
Inferred	7,871	0.21	0.56	53	142

Mineral Resources are not mineral reserves and do not demonstrate economic viability. There is no certainty that all or any part of the Mineral Resource will be converted to mineral reserves. Quantity and grade are estimates and are rounded to reflect the fact that the resource estimate is an approximation.

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14.9.3 Mayflower Mineral Resources

The results presented in Table 14-46, show the Phase I Indicated and Inferred mineralization inventory for Mayflower. Resources are separated by classification. All material is assumed to be processed by heap leaching methods. Resources are pit constrained and reported at a gold cut-off grade of 0.10 g/t Au.

Table 14-46 Mayflower Resources Phase 1 (QP: RDA, Scott Wilson; Effective September 18, 2018)

Classification	Tonnes (kt)	Au g/t	Ag g/t	Contained Au (x1,000)	Contained Ag (x1,000)
Indicated	5,239	0.46	0.41	78	69
Inferred	28	0.21	0.24	0	0

Mineral Resources are not mineral reserves and do not demonstrate economic viability. There is no certainty that all or any part of the Mineral Resource will be converted to mineral reserves. Quantity and grade are estimates and are rounded to reflect the fact that the resource estimate is an approximation.

14.10 Visual Validation

The portions of the mineralization block models bounded by the Whittle^TM defined open pit mining shells are illustrated in Figures 14-45 through 14-49. Figures 14-46 and 14-46 contain a long section through the YellowJacket corridor and cross-section through Sierra Blanca and YellowJacket, respectively. Figures 14-47 and 14-48 are cross-sections and long sections, respectively, through Mayflower. Figure 14-49 shows a cross section through Jolly Jane.

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Figure 14-45 Long Section through Sierra Blanca/YellowJacket Mineral Resource Model-metres

 

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Figure 14-46 Cross Section across Sierra Blanca/YellowJacket Looking North-metres.

 

Figure 14-47 Cross Section through Mayflower Deposit-metres.

 

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Figure 14-48 Long Section through Mayflower-metres.

 

Figure 14-49 Cross Section through Jolly Jane-metres.

 

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15 Mineral Reserves

There are no Mineral Reserves estimated for the Project.

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16 Mining Methods

This Technical Report assumes conventional surface mining methods using surface drill and blast techniques with off highway haul trucks and front-end loaders will be employed at both the Mother Lode open pit and the North Bullfrog open pits. Higher grade mineralization from the YellowJacket vein and vein stockwork, and any Sierra Blanca sulfide mineralization with sufficient grade would be trucked on Highway 95 to a central mill facility located at Mother Lode.

This PEA is preliminary in nature and is based on technical and economic assumptions which will be evaluated in more advanced studies. The PEA is based on the Mother Lode and North Bullfrog Mineral Resource models which both include Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the PEA will be realized. The basis for the PEA is to demonstrate the economic viability of the Project. The PEA results are only intended as an initial, first-pass review of the Project economics based on preliminary information. There are no advanced studies on the Project that would be impacted by the PEA.

16.1 Mine Configuration

Mineralized material will be delivered to one of two different processes; heap leaching or milling, dependent on the material type and gold grade. Heap leach pads would be constructed at both sites but higher-grade mineralization would only be processed at a mill facility assumed to be constructed at the Mother Lode site. The mill would combine circuits to process either oxide mineralization (YellowJacket vein and vein stockwork) or sulfide mineralization through separate circuits. Structurally controlled mineralization in YellowJacket above 0.35 g/t would be delivered to the mill circuit where it would be crushed to meet mill feed requirements, followed by grinding, gravity concentration and CIL leaching. Sulfide mineralization would be crushed and ground, followed by gravity concentration to produce a concentrate, chemical flotation to produce a concentrate, pressure oxidization of the concentrates, and then CIL leaching. Oxide mineralization between the grades of 0.06 g/t and 0.35 g/t at YellowJacket, above 0.06 g/t from the disseminated mineralization in the Sierra Blanca, Jolly Jane and Mayflower deposits at North Bullfrog, and above 0.06 g/t at the Mother Lode deposit would be delivered to centrally located, run-of-mine heap leaching facilities at each site. Heap leach mineralization would be placed and shaped by haul trucks and dozers. At the heap leach facilities, cyanide solution would be used to dissolve the gold and silver, and then would be processed through standard carbon-in-column leach column arrays. Loaded carbon from the column arrays would be refined into gold doré in-site refinery that would be part of the mill. Overburden material would be trucked to overburden storage facilities proximal to each open pit.

A single leach pad and process facility location was assumed at each site to recover metal from oxide mineralized material. The facilities would average daily leach pad placement of 68.5 k tonnes per day and 13.5 k tonnes per day, for NBP and MLP, respectively. AT NBP, the heap leach pad and ponds would be located near the northwest corner of the Project site. NBP would produce mineralization from both the Company’s federal public land and private land resources. All mineralization at MLP would be produced from federal public land. Each of the Project resource areas at North Bullfrog, Mayflower, Sierra Blanca, Jolly Jane and YellowJacket, and Mother Lode would require a separate surface mine excavation. Material would be mined from each site concurrently, however, the individual deposits at North Bullfrog mine would be mined at different times in the schedule. The Mother Lode deposits and the Sierra Blanca/YellowJacket deposit would produce continuously over the entire LOM to maintain the mill throughput.

In the conceptual mine plans, 10-metre-high benches would be drilled and blasted, then loaded into 133 tonne haul trucks using 20 cubic metre front end loaders. At NBP, the haul trucks would deliver the mineralized material to either the stockpile transfer system for the mill grade mineralization or place the blasted material directly on the leach pad. The stockpile transfer system would consist of a grizzly, a feeder and conveyor belt feeding bins. Highway haul trucks would be loaded from the bin and transport the mill grade mineralization to the MLP mill stockpiles using Highway 95. Mill grade mineralization from MLP would also be stockpiled at the mill location. Mining rates at both NBP and MLP are sufficient to create a mill stockpile that will allow higher grade mineralization to be fed (grade streaming) to the mill in early years of the schedule. The mill stockpile is scheduled to reach a maximum size of 11.7 M tonnes by the end of year 2, and then begin to decline through the LOM.

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The mine plan calls for an average total mining rate of 66 million (“M”) tonnes per year of combined overburden and mineralized material. Of that, an average of 30 M tonnes per year would be mineralized material placed on a heap leach pad at an average rate of 82 k tonnes per day. Another 3 M tonnes per year would be delivered to the mill at an average of 8,200 tonnes per day. The average overburden to mineralized material (strip ratio) would be 1.08. Based on this mining rate, the Project would have an active life of 9 years, with minor gold production for the additional 3-year rinse down period on the leach pad.

16.2 Mine Optimization

Economic mine limits were determined using Gemcom’s Whittle® 4.7.2 software which employs the Lerch-Grossman© economic algorithm. Whittle works on a block model of the mineralized material, and progressively constructs lists of related blocks that should or should not be mined. The final list defines a surface mine outline that has the highest possible total value, while honoring the required surface mine slope parameters.

The mine optimization was performed using the resource block models discussed in Section 14. The models were reviewed by the Michael Cole (QP for mining) and determined to be suitable for use in the mine optimization studies. The individual block models for each mineral deposit were each imported into Whittle®. Model block sizes were re-blocked to 10 m x 10 m x 5 m and defined a minimum mining selection unit of 10 m. The following Table 16-1 lists the parameters used in Whittle® to derive the ultimate pit limits.

Table 16-1 Parameters used in Whittle^TM Analysis

Parameter	Unit	Mayflower	Jolly Jane*	Sierra Blanca	YellowJacket	Mother Lode
Mining Cost	US$/total tonne	1.42	1.42	1.42	1.42	1.42
Au Cut-Off	g/tonne	0.1	0.1	0.06	0.35	0.76
Processing Cost	US$/ process tonne	2.22	2.22	1.20	13.42	19.50
Au Recovery	%	72	72	74	86.8	80%
Ag Recovery	%	6	6	6	74.3	na
Administrative Cost	US$/process tonne	0.5	0.5	0.5	0.5	0.5
Refining & Sales	$/Au oz	5.00	5.00	5.00	5.00	5.00
Au Selling Price	US$/oz	1,250	1,250	1,250	1,250	1,250
Slope Angle	Degrees	50	50	50	50	65

Only oxide mineralized material at NBP and MLP were considered for the Heap Leach circuit. Mineralization above and below the oxidation surface in the structurally controlled YellowJacket zone, where the gold occurs as free gold and electrum, was scheduled for treatment in the oxide milling circuit. Sulfide mineralization at MLP and at NBP were scheduled for treatment in the flotation-pressure oxidation milling circuit. A series of nested excavation shells were created between $500 and $1,500 per gold ounce and used to guide phases within the surface mine plan. For the purposes of scheduling YellowJacket and Sierra Blanca were cut into multiple production phases based on Whittle pit shells.

16.3 Mine Production Schedule

Mining production would begin at an initial rate of 14 k tonnes of mineralized material per day as pre-stripping in year -1, increasing to a peak rate of 106 k tonnes of mineralized material per day by the second production year. Total gold mined would average 450 k ounces contained over the first 4 years and average at approximately 394 k contained ounces per year over the 9 years of production. Grade Streaming (highest grade possible at any point in the schedule considering the mill throughput) was the key factor influencing the production schedule in the Whittle® scheduling software. Stockpile Cut Off Grade Optimization tool within the Whittle ® software was used to optimize the stockpile management. Mill production was assumed at 8,200 tonnes per day. Rates of heap leach mining were set by the requirements of the grade streaming strategy and to deliver the highest grade possible as early in the schedule. YellowJacket would be mined as the start-up pit to provide high grade feed for the mill at commissioning and would continue for 9 years. Sierra Blanca would be mined throughout the life of the operation and would be split into three phases; a starter phase, a mid-grade phase and a low-grade phase. Mayflower would be mined at the beginning of the mine life because of its relatively high grade, and Jolly Jane was scheduled to enhance heap leach grades as the Sierra Blanca heap leach grade declined. The total Project would contain 269 million mineralized tonnes at an average grade of 0.38 g/t gold and a byproduct silver grade of 0.93 g/t. The life of mine strip ratio would be 1.08:1, which includes all heap leach and mill material.

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The LOM mining schedule is presented in Table 16-2, and Table 16-3 lists the process feed schedule from the individual open pits by year. Table 16-4 lists the total mill and heap leach feeds by year. Table 16-5 presents the proportion of the scheduled production by the Resource classifications of Measured, Indicated and Inferred for each of the deposits.

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Table 16-2 Mining production schedule for NB-MLP deposits developed with stockpile scheduler

 

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Table 16-3 Process feed production schedule for individual NB-MLP deposits

 

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Table 16-4 Total process feed production for NB-MLP mill and heap leach

 

Table 16-5 Classification of Mineral Resources in Individual NB-MLP Open Pit Production

Classification	Mayflower	Jolly Jane	Sierra Blanca	YellowJacket	Mother Lode Oxide	Mother Lode Sulfide
Measured	0%	0%	11%	47%	30%	17%
Indicated	100%	74%	58%	52%	40%	70%
Measure & Indicated	100%	74%	69%	99%	70%	87%
Inferred	0%	26%	31%	1%	30%	13%

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16.4 Mine Production

Mining production would come from open pit mining methods. As this is an assessment that is preliminary in nature, unsmoothed Whittle pits were used as the basis of the open pit production quantities. Volumes were extracted from the Vulcan® block models and scheduled out with the Whittle® scheduling software which included the ability to optimize based on the stockpiling of mill feed mineralization. All mineralization in the production plan was selected on a whole block basis and was therefore assumed to include dilution. In all areas of the mine the swell factors used standard assumptions for swell factors of 35% for all bucket and truck body calculations.

The Whittle® process determines the pit volume. Destination of the material within the pit shell was determined by applying cut-off grades based on processing costs. The cut-off grades used for selection of the heap leach or mill were 0.06 g/t and 0.35 g/t, respectively. Mineralization below 0.06 g/t were scheduled for the waste storage facility.

16.4.1 YellowJacket / Sierra Blanca Pits

YellowJacket contains both heap leach mineralization and mill grade mineralization. The pit would be in close proximity to Sierra Blanca such that there would be a sharing of waste between the two pits. No detailed toe/crest mine designs or ramp configurations were created for the ultimate pit. The pit slope was assumed to be 50^o and the pit bottom would be partially below the water table.

For Sierra Blanca, no detailed toe/crest mine designs or ramp configurations were created for the ultimate pit. All resources used for production scheduling were based on the whole-block pits produced directly out of Gemcom’s Whittle® 4.5 software. Sierra Blanca’s ultimate pit depth was limited to the oxidation boundary and the pit slope was assumed to be 50^o.

Yellowjacket and Sierra Blanca share a common block model and were scheduled concurrently in Whittle as the Sierra Blanca Pit, as reflected in the mine production schedule in Table 16-2.

16.4.2 Mayflower Pit

For Mayflower, no detailed toe/crest mine designs or ramp configurations were created for the ultimate pit at the time of this Technical Report. All resources used for production scheduling were based on the whole-block pits produced directly out of Gemcom’s Whittle® 4.5 software.

For Whittle, an ultimate pit slope of 50^o was used

16.4.3 Jolly Jane

For Jolly Jane, no detailed toe/crest mine designs or ramp configurations were created for the ultimate pit at the time of this Technical Report. All resources used for production scheduling were based on the whole-block pits produced directly out of Gemcom’s Whittle® 4.5 software.

For Whittle, an ultimate pit slope of 50^o was used.

16.4.4 Mother Lode

The original Mother Lode open pit was mined in 1989 and extracted near surface oxide material for heap leaching. The mining reached a depth of approximately 60 m where it exposed sulfide mineralization. No detailed toe/crest mine designs or ramp configurations were created for the Whittle® pit at this time. The Gemcom Whittle® 4.5 pit shapes were constrained to honor the extent of Corvus’s claim boundaries at the surface; however, the mineralization block model indicates more mineralization beyond the property boundary. A lay-back agreement may be negotiated in the future which would allow more of the mineralization on Corvus property to be extracted. For Whittle, a pit slope of 65 ^o was assumed.

Resources schedule for production were based on the whole blocks in mineralization block model.

16.5 General Site Layout

A conceptual site layout has been developed to provide a basis to estimate haul distances for equipment specification and cost estimation. The layouts are presented as Figure 16-1 and 16-2 for the Mother Lode site and the North Bullfrog site, respectively.

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Figure 16-1 Conceptual layout of the Mother Lode site showing open pit, TSF, heap leach pad, process facilities and waste storage area

 

Figure 16-2 Conceptual layout of the North Bullfrog site showing open pits, heap leach pad and waste storage

 

A back-country connector road is planned to allow movement of heavy equipment between the two sites to allow maximum efficiency in equipment utilization.

16.6 Mine Fleet and Capital

Preliminary mining fleet definitions were based on material haul routes and cycle times. Haul distances were estimated including distances and road gradients. RPM Global’s Talpac^TM software was used to calculate haul times. Different scenarios were evaluated based on different equipment options. Round-trip cycle times were calculated for 136 tonne haul trucks. These cycle times were used with the mine production schedule to determine fleet requirements.

The mining fleet is expected to contain three 20 cubic metre wheel loaders and up to twenty-one 133 tonne capacity haul trucks, four 306 kW dozers, six rotary drills with 24.6 kilogram per centimeter squared compressors to handle the harder rock types, three 193 kW road graders along with associated support and maintenance equipment. A truck shop at MLP would be equipped for major maintenance with smaller shop at NBP for preventive maintenance. A back-country dirt road was assumed to allow transport of equipment between the two sites based on operational demand and maintenance requirements. Warehouses, fuel depot and operations offices has been assumed in the requirements.

Table 16-6 lists the projected mining mobile equipment requirements for the NBP (Mining Cost Services, Costmine (2018)).

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Table 16-6 Production Equipment Fleet for the NB-MLP

Equipment	Quantity	Unit Cost (000s)	Total Cost (000s)
Front End Loader, 20 cu m bucket	3	$4,673	$14,019
Haul Truck, 133 Tonne	21	$2,593	$54,453
Rotary Blast hole Drill	6	$1,063	$6,378
Dozer	4	$1,085	$4,340
Grader	3	$385	$1,155
Water Truck	2	$639	$1.278
Bulk Explosive Truck	2	$48	$96
Service Truck	3	$73	$147
Tire Truck	1	$172	$344
Fuel Truck	2	$87	$174
Truck & Lowboy Hauler	1	$269	$269
Stockpile Loader	1	$2,050	$2,050
Pumps and Light Plants	10	$10	$100
Light Vehicles	20	$36.5	$730
Backhoe	2	$300	$300
Mill Backhoe	1	$246	$246
Warehouse Forklift	1	$43	$43
Skid Steer Loader	2	$50	$100
Mill Loader	1	$147	$147
Crane Truck	1	$390	$390
Total Mobile Equipment			$87,058

As can be seen in the Table 16-3, the mine production fleet is designed to have up to 21 trucks operating under three-wheel loaders at any time. The availability of 93% and utilization of 87% is normally sufficient to allow normal maintenance on all of the mine equipment as required to meet manufacturer specifications. Based on a 250-hour preventative maintenance schedule, means trucks and loaders would be maintained on day shift only and other mining equipment would be maintained on night shift. The 72 pieces of primary mining equipment would need to be maintained to meet the mining schedule as shown in Table 16-2.

Minimal replacement capital was assumed due to the short Project life. The fleet was projected to meet the mine schedule without any major equipment replacement. All of the operating and major and minor servicing expenses were included in the hourly operating cost.

16.7 Drill and Blast Methods

The drill pattern spacing is based on a pattern spacing of 4.3 metres by 4.3 metres for a 10 metre bench height. The holes would be drilled 1 metres deeper than bench height on average to manage fall back and sub-drill needs. The mill feed and waste would be drilled and shot using an average powder factor of 0.3 kilogram per tonne.

The heap leach material to be stacked on the pads at Mother Lode and North Bullfrog would be blasted using an ultra-high intensity blasting method that has been demonstrated in Australian surface mining as reported by Brent et al, 2014, to increase overall leaching recovery of metals. The basic description of the method would be to increase the powder factor by a factor of 1.9 and 3.8 times more than normal, for the top and bottom sections on any pattern with heap leach material. This method would be used to develop a finer particle size gradation (p80-76mm) and would induce micro fractures in the coarser particles to allow higher heap recoveries than a normal ROM material might achieve. The average powder factor using this high intensity blasting technique would be expected to increase to 0.9 kilograms per tonne of heap leach material. Only the heap leach material would utilize this blasting method. The process would normally load all holes but delay the timing of the second blast with electronic controls. First, the top hole of the bench would be blasted and the particles allowed to settle from the blasting energy. Then, the bottom of each hole and intermediate holes would be blasted on a second initiation of the shot pattern. This would allow a very high powder factor to provide finer rock particle sizes and develop micro-fractures to allow the cyanide on the heap leaching facility to infiltrate into the core of each coarse rock fragment. The plan would be to develop a square pattern that allows a center hole to be drilled and very heavily loaded with ANFO to provide ultra-high intensity blasting in the heap leach mineralized zones. This would require about one more hole for each four holes in the leach mineralization zones of the pits to be drilled and double initiation of shots. All heap zones holes would be double primed and then shot a second time in a short sequence to allow quick turnaround of these heap leach patterns.

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16.8 Manpower

The NBP mining group would be expected to operate on a 24-hour day and 365 days per year. The mining and maintenance operations plus support would be expected to be spread over a four-crew rotation planned as 12-hour day and night shifts with rotation between days and nights as defined by the operations group. The engineering and geology groups would work a four-day schedule and would have sufficient coverage over the weekends to limit any delays caused by manning schedules. The technical staff schedule would be based on rotating each month. This would provide overlap between the two survey, engineering and geology groups during the week when the workload would be heaviest. The projected staff requirements are shown in the Table 16-7.

Table 16-7 NB-MLP Mining Personnel Estimates

`	Function	No. of Mining Personnel
Salaried	Mine Manager	1
	Shift Foreman	8
	Chief Engineer	1
	Mine Engineer	4
	Surveyor	2
	Asst. Surveyor	2
	Environmental Tech	2
	Chief Geologist	2
	Ore Control Geologist	4
	Maintenance Foreman	4
	Subtotal Salaried	30
Operations Hourly	Drill Operators	48
	Loader Operators	16
	Truck Drivers	84
	Dozer Operators	16
	Water Truck Operator	8
	Grader Operators	12
	Blasters	16
	Pump Operators	8
	Training/Absentee	8
	Subtotal Operations	216
Maintenance Hourly	Electrical	16
	Mechanics	54
	Welder/other	8
	Subtotal Maintenance	78
Total	Total Mining	254

The mine operation would be scheduled based on a 4-days on, 4-days off schedule on a 12-hour shift basis. This would produce about 8 hours of overtime in an average week, which has been utilized for cost estimates. The burden rate was estimated at 38% above the base rate to cover the cost of insurance, social security and other benefits to be determined by the owners of this project.

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17 Recovery Methods

Figure 17-1 presents a block flow diagram for the integrated Project. The sulphide mill material and energy balance were modeled with the METSIM™ process simulator. Results are derived from the information described in Section 13.

17.1 Mother Lode and NBP Recovery Methods

The PROJECT Oxide materials evaluated in this Technical Report will be processed in a standard heap leach recovery circuit commonly used in the region. The cyanide solutions from the heap leach will be treated in a Carbon-in-Columns (“CIC”) gold adsorption and desorption circuit.

The MLP mill will process NBP oxide, NBP sulphide and MLP sulphide materials separately from stockpiles. NBP Oxide material will be processed in a conventional gravity/CIL mill. Sulphide materials from both MLP and NBP will be processed through a milling and flotation plant to produce a concentrate. The flotation concentrate will be thickened, stockpiled, and processed through a pressure oxidation circuit followed by a CIL circuit. Oxide and sulphide CIL tails will be detoxified through an Inco/SO₂ type process. Tailings will be pumped to a paste tailings thickener. Slurry will be deposited as a paste in a tailings facility. Detoxified, non-cyanide containing paste thickener overflow solution, will be returned to the mill.

Carbon from the CIC and CIL circuits will be processed in a carbon handling circuit located at the mill. The carbon handling will include carbon stripping, acid washing and reactivation circuits. The process solutions from carbon stripping will be sent to electrolytic cells for gold recovery in a precipitate sludge. The sludge will be retorted for mercury removal and then melted into doré bars in a furnace.

17.1.1 Gold Recovery

Mother Lode ROM oxide material gold recovery and reagent consumptions for the heap leach materials are based on bottle roll tests and historical Daisy mine ROM gold production data. The Au recovery for the heap leach is assumed to be 74% and cyanide and lime consumptions are 0.05 kg NaCN/t and 2 kg CaO/t.

The NBP disseminated mineralization in the Sierra Blanca/Savage Valley, Mayflower and Jolly Jane deposits, gold recovery and reagent consumptions for the heap leach materials are based on +100 bottle roll tests at varying sizes and 40 column tests In the test program, Savage Valley ores were independently tested, but because of the metallurgical response, similar geology and overlapping location of the two pits, the Savage Valley and Sierra Blanca ores were combined and described hereafter as Sierra Blanca. The average Au recovery for the heap leach, considering the three mineralization resources was 74%. Silver recoveries were assumed to range from 5.8% to 9.7% for the different resource areas, averaging 6%.

Mother Lode sulphide material processed in a flotation/concentrate pressure oxidation/CIL process indicated a weighted gold dissolution, for Tip1 and Tjvs materials, of 80%. Sodium cyanide and lime consumptions are estimated to be 0.2 kg/t and 25 kg/t.

NBP sulfide material gold dissolution in this study assumes AAO and POX oxidized processes will provide similar sulphide oxidation and gold dissolution. Sulphur oxidation and gold dissolution correlated very well in the AAO tests. Gold recovery based on a flotation and sulphur oxidation by AAO, indicated Sierra Blanca and Pioneer Formation tuff had overall gold recoveries of 93 and 94%, in Soda Ash, respectively and 91 and 93% when in trona, respectively. Rhyolite gold dissolution was 89 and 87% in soda and trona, respectively. Dacite gold dissolution was 88 and 87% in soda ash and trona, respectively. Silver recovery was based on Sierra Blanca concentrate values back calculated through flotation to a calculated head grade. Silver recovery was estimated at 77% in flotation and 74% in the cyanide leach for a weighted recovery of 57%. The average bottle roll cyanide and lime consumptions are 0.2 and 0.6 kg/ton ore.

17.2 Heap Leaching

The Mother Lode and NBP properties have main electrical power and water located nearby, in quantities sufficient for the operation. Process reagents are common commodities utilized in the Nevada gold mining industry.

17.2.1 Mother load Heap Leaching facility

The Mother Lode heap leach materials would be trucked to the heap leach pad at a nominal rate of 13,600 tonnes per day. The mineralized material would be stacked on the heap in 9.1 m (30’) lifts. Lime will be added to individual truck loads. The heap will be leached using buried emitters with a weak cyanide solution.

The barren solution from the carbon columns would be adjusted with makeup water, sodium cyanide, anti-scalant, and will be pumped to the heap leach pad. The pH will be adjusted with additional lime (CaO), added to pad, if necessary. The solution distribution on the heap will use buried emitters. Solution application rates will range from 12 l/hr-m² to 6.0 l/hr-m². Provision will be made to recycle a portion of the pregnant solution back to the heap to extend the leach retention times.

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Pregnant solution from the heap will be collected in a pond and pumped to the CIC Plant located in the MLP mill. The plant would treat 680 m³/hr and will consist of a single train of five cascading carbon columns. The columns would be equipped with a feed box and trash carbon screen on the solution discharge. Carbon will advance on a weekly basis countercurrent to the solution flow, in a 3.0 tonne batch. Loaded carbon will be acid washed, stripped, and regenerated in a kiln and returned to the CIC columns

The strip solution will be sent to electrolytic cells for gold recovery in a precipitate sludge. Sludge will be retorted for mercury removal and melted into doré bars in a furnace.

17.2.2 NBP Heap Leaching

NBP heap leach materials would be trucked to the heap leach pad at a nominal rate of 61,000 tonnes per day. The mineralized material will be stacked on the heap in 9.1 m (30’) lifts. Lime will be added to individual truck loads. The heap will be leached using buried emitters with a weak cyanide solution.

The barren solution from the carbon columns will be adjusted with makeup water, sodium cyanide, anti-scalant, and will be pumped to the heap leach. The pH will be adjusted with additional lime (CaO) added to pad if necessary. The solution distribution on the heap will use buried emitters. Solution application rates will range from 12 l/hr-m² (0.005 gpm/ft²) to 6.0 l/hr-m² (0.0025 gpm/ft²). Provision will be made to recycle a portion of pregnant solution to the heap leach pad to extend the leach times. The makeup water requirement for the heap leach pad loss was estimated to be 2800 l/min (750 gpm).

Pregnant solution from the heap will be collected in a pond and pumped to the CIC Plant. The CIC Plant will consist of two parallel trains, comprising five cascading carbon columns, processing approximately 1,818 m³/hr., (8,000 gpm) of solution. The column trains will be equipped with feed boxes and trash carbon screens on the solution discharge. Carbon will be advanced on a weekly basis countercurrent to the solution flow in two x 3.5 tonne lots. Loaded carbon will be trucked to the Mother Lode mill for acid washing, stripping, and regeneration. Regenerated carbon will be trucked back to the NBP CIC plant as needed to maintain sufficient carbon inventory.

17.3 Project Milling operation

NBP oxide and sulphide mineralization will be trucked to the MLP mill and stockpiled in feed piles as needed. MLP sulphide material will be fed to the mill from stockpiles as needed. The stockpiled materials will be processed at a rate of 8,200 tpd. Process time for this study assumed the mill will process oxide material 50% of the time and sulfide material 50% of the time. The flotation plant will remain idle while oxide material is processed. The pressure oxidation plant will be fed from the flotation plant, or stockpiled flotation concentrate, and will run continuously.

17.3.1 YellowJacket Material Processing

YellowJacket oxide material will be processed using a mill/gravity/cyanide leach circuit. The material will be primary crushed, at a rate of 8200 tpd, followed by primary SAG milling with pebble crushing, and secondary ball milling, gravity separation of the cyclone underflow. Gravity tails at a P₈₀ -size of 74 µm will be leached in a CIL circuit for 24 hours. CIL tails slurry will be treated in a cyanide detoxification process utilizing Inco/SO₂ air process methods, followed by pumping to a paste tailings thickener. The slurry would be pumped from the thickener underflow, as a paste, into a storage facility. Paste thickener overflow solution will be returned to the mill or heap leach process. Loaded carbon from the CIL circuit will be acid washed, stripping, and regenerated for return to the CIL tanks. The strip solution, from the carbon, will be combined with solution from the intense cyanide leaching of gravity concentrate, and gold precipitated in electro winning cells. Gold precipitate will be retorted to remove mercury and melted into gold doré bars.

17.3.2 MLP and NBP Sulfide Milling

17.3.2.1 Crushing and Grinding

Mother Lode and NBP sulphide materials will be delivered from the mines via truck and stockpiled, or directly dumped into the primary crusher feed bin. Reclaimed materials will be blended to control the sulphide content in feed to the flotation/POX plant. Material will either be direct dumped and or pushed with a small track dozer into the feeder. The feeder will control the material processing rate to an average of 8,200 tonne/day. The mineralized materials will be metered on to the primary crusher feed conveyor and carried to a primary jaw crusher. Product from the jaw crusher will be conveyed to the milling circuit. Milling will be performed using a primary SAG mill with recycle pebble cone crushing, and a secondary ball mill in closed circuit with cyclones. The ball mill product will be pumped to the cyclones with cyclone overflow, at a P₈₀ of 74 µm exiting the grinding circuit to the flotation circuit. Material coarser than 74 µm will report to the cyclone underflow and may by-pass gravity separation returning to the ball mill or may be processed in gravity separation and return to the ball mill. The gravity circuit would consist of the equipment described in oxide ore processing.

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17.3.2.2 Sulphide Flotation, Concentrate Storge, Reclaim, and Regrind

The cyclone overflow will flow by gravity to a trash screen and to the flotation circuit. Gold and sulphide minerals predominately in pyrite, will be concentrated by bulk sulphide flotation in rougher and scavenger flotation cells. The flotation concentrate will be thickened and either pumped to the regrinding mill or, filtered and stored in covered storage buildings for reclaim, to feed the POX circuit when oxide ore is processed.

Flotation concentrate will be reclaimed from covered storage by a front-end loader that will fill a feeder bin. Reclaimed concentrate will be fed by feeder and conveyors to the regrind mill circuit at rate of 410-1066 tpd. The concentrate will be ground to P₈₀ 10-15 µm and stored in agitated tanks for feed to the POX circuit.

17.3.2.3 Concentrate Pre-Acidification, Pressure Oxidation, Hot Cure and Neutralization

Concentrate slurry from storage tanks will be pumped to the Pre-Acidification tank and combined with acidic overflow solution from the pressure leach autoclave discharge. The acidic solution will react with carbonate minerals in the feed material and liberate carbon dioxide prior to entering the pressure leach autoclave. Pre-acidified slurry will be pumped into the pressure leach autoclave by high pressure piston positive displacement pumps. Sulphide minerals in the slurry will be oxidized by oxygen at a temperature of 225 ^oC and 3600 kPa. The oxygen partial pressure will be maintained at 690 kPa. Oxygen at 93% v/v will be supplied from a vacuum swing adsorption oxygen plant.

POX slurry will be let-down to atmospheric pressure through flash vessels and retained for several hours in brick lined agitated Hot Cure tanks. Slurry from the Hot Cure tanks will flow to a thickener. Thickener underflow will enter a two stage CCD wash circuit. Thickener overflow will be split to pre-acidification and solution neutralization. The washed slurry from the second CCD thickener will be pumped to the CIL circuit and lime added to the slurry prior to entering the CIL tanks. Sulphuric acid in solution from the CCD tanks will be neutralized with flotation tailings followed by lime and thickened. Thickener underflow will be pumped to the tailings sump. Thickener overflow solution will be pumped to the CCD wash stage.

17.3.2.4 Cyanide Leaching and Carbon Absorption

The oxide mill material and POX neutralized slurry will be pumped to the CIL circuit. The CIL leach tanks will provide approximately 24 hours of leach time. Activated carbon will be added to the Carbon in Leach (CIL) tanks, each containing seven tonnes of carbon. Stripped carbon will be added to the circuit in the last CIL tank and will be pumped once a day upstream into the fourth CIL tank and so on. Loaded carbon from the first CIL tank will be taken in seven-ton batches, for processing in the carbon handling circuit, where it will be screened, stripped of gold, acid washed, and regenerated in a reactivation kiln prior to being brought back to the sixth tank. The slurry exiting the last CIL tank would be screened to remove carbon, and gravity flow to the Cyanide Detoxification circuit.

17.3.2.5 Tailings detoxification and Storage

Material exiting the CIL circuit would be detoxified with an Inco/SO₂ air process and be pumped to a paste tailings thickener. The thickener underflow would be pumped to the storage facility (TSF). Solution overflow from the paste thickener would be pumped back to the mill and heap leach pad for reuse.

17.3.2.6 Carbon Handling

Carbon recovered from the CIC circuit and the CIL circuit will be processed in one carbon handling facility located at the mill. Carbon from both the mill and heap leach processes will be batch stripped, in seven-ton batches, in one of two strip vessels. Upon leaving the strip vessel, the carbon will be screened to have fine carbon removed from the circuit. After screening the carbon will be acid washed in a weak hydrochloric acid solution to remove calcium and other contaminants. The washed carbon will then be pumped to the regeneration kiln for reactivation.

Carbon processed in the kiln will be heated to 700°C in a reducing atmosphere. This process will open the pores on the carbon and raise the carbon’s affinity for gold and silver. Following reactivation, the carbon will be sized once again using a screen to remove any fines generated in the process. The screened reactivated carbon will then be returned as needed to the last CIL tank, to the Mother Lode CIC circuit, or trucked to the NBP CIC plant.

17.3.2.7 Refinery

The refinery will have electro winning cells with sludge tanks, a mercury retort, flux mixer, and furnace. The refinery will also contain the high intensity leach circuit for the gravity concentrate. Pregnant solutions from the high intensity leach will be combined with the pregnant strip solution from the strip circuit and will be processed in one of two electro winning cells. Periodically, the electro winning cells will be drained, cathodic sludge will be pressure washed from the cathodes, and the sludge drained from the cell, and placed into a mercury retort for mercury removal.

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After retorting, the sludge will be placed into a furnace with fluxing agents and melted. The melted metal will be poured into doré bars for shipping to an outside refinery for final processing and sale. The refinery will be enclosed with a pre-engineered steel building. The capital cost of the carbon handling plant and refinery included all valves, samplers, and security cameras necessary to operate in an efficient manner. The State of Nevada requires that all processes with temperatures exceeding 79 ^oC (175^oF) be equipped with mercury abatement equipment. The NBP would comply with this requirement.

17.3.3 Assay Laboratory

A full-scale Assay Laboratory with Fire Assay and ICP capabilities will built for the project. Because of the low cut-off grades, a 30g Fire Assay, followed by an ICP finish will be utilized on mine samples. The Assay Laboratory will contain equipment for sample preparation, screening of crushed products, and for carbon activity testing.

A metallurgical test laboratory within the assay laboratory will have screening, bottle roll, column testing, flotation, and pressure oxidation facilities. As the project developed and bulk samples become available, column and vat leach tests will be conducted on crushed and ROM samples.

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Figure 17-1 Block Flow Diagram for Integrated Processing of NB-MLP Mineralized Materials.

 

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17.4 Process Manpower

Manpower requirements for the mill and heap leach process facilities is listed in Table 17-1.

Table 17-1 Process Manpower

Class	Function	No. of Mill Personnel	Function	No. of Heap Leach Personnel ML+NBF
Salaried	Process Manager	1	Operations Planner	1
	Chief Metallurgist	1	Maintenance Foreman	1
	Metallurgist	1	Metallurgist	1
	Shift Foreman	4	-	-
	Maint,/Elect. Foreman	1	-	-
	Chief Assayer	1	-	-
	Refiner	1	-	-
	Subtotal	10	Subtotal	3
Operations Hourly	Crusher Operators	4	Assay Prep	4
	Grinding Operators	4	Refiner	2
	Mill Control Operators	4	Feeder/Conveyor Operators	N/A
	Leach/CIP/Gravity Operators	4	Leach Pad Operators	16
	Flotation, Concentrate Handling and Regrind	4
	POX Pre-Acidification, Hot Cure, and Neutralization	4
	Stripping Operators	4	Helpers	8
	Paste Tails Operators	4	-	-
	Assayers	2	-	-
	Subtotal	34	Subtotal
Maint. Hourly	Mechanics/Electricians	12	Mechanics/Electricians	8
	Total Mill	46	Total Heap Leach	38
	Total Process	97	-	-

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18 Infrastructure

18.1 General Area Resources

Much of the primary infrastructure required to develop a surface mine is available in close proximity to both the Mother Lode and North Bullfrog properties. The availability of these key infrastructure elements is described in the following sections on location and access, human resources, electrical power, water resources and project infrastructure elements. Location of the properties with respect to Beatty, NV with the open pits and access roads are shown in Figure 18-1.

Figure 18-1 Map showing the extent of NB-MLP Property Boundary and locations of mining resources, and infrastructure

 

18.1.1 Location and Access

The MLP is located approximately 10 km (6 miles) east of the community of Beatty, in Nye County. The property can be reached by two gravel roads that connect with Highway 95, which lies to the west. An unnamed gravel road begins at the south edge of Beatty and runs along Fluorspar Canyon (shown on Figure 18-1). This road was historically the main access to the Mother Lode and Daisy mining operations. A second unnamed road connects to Highway 95 just 3 km (2 miles) north of Beatty and connects with the road in Fluorspar Canyon just below the historic Secret Pass open pit mine.

The NBP is located approximately 16 kilometres (9.3 miles) north of the community of Beatty, in Nye County Nevada. The property is immediately west of Nevada Highway 95, which connects the major cities of Las Vegas and Reno. Access to the property from the highway is currently by several dirt roads that are maintained and provide access for an existing commercial aggregate producer, as well as, cattle grazing operations further to the west. A dirt road at the north end of the property, Strozzi Ranch Road, would provide access to the major facilities location and would connect to the powerline running along Highway 95.

An overland, dirt road (shown in Figure 18-1 as Transportation Route-Mother Lode to NBP)) would be constructed to connect the two project sites and provide the ability to move major equipment between sites to maximize mining equipment efficiency. The road would be constructed on public land and would require a Right-of-Way permit with the BLM. The equipment moves would be intermittent, limiting the amount and frequency of road maintenance required. The road would be single lane, and equipment moves would be accompanied by leading and following hazard vehicles to control any encounters with public traffic along the road or at the required crossing of Highway 95.

Major mining and construction equipment sales and service are readily available throughout Nevada; however, most major mining operations are located in the northern part of the State and are serviced from the cities of Reno and Elko. Las Vegas, 200 kilometres (125 miles) south of NB-MLP, also has a major construction industry and heavy equipment sales and service are available there.

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Beatty is a small residential community with motels, restaurants and stores.

18.1.2 Human Resources

Human resources are available within the community of Beatty, which has a population of approximately 1,100 people, and historically had provided a substantial portion of the workforce for the Bullfrog Mine, which operated between 1989 and 1998 as both an open pit and underground gold mining operation. The community has a long association with the mining industry and could contribute some experienced personnel to a mining project. The community has schools, a medical clinic, motels, fuel service and food stores.

Pahrump, approximately 110 kilometres (68 miles) to the southeast of Beatty, is a larger community with a population of 36,000. Pahrump is a local regional center, with a hospital and emergency medical services, a college campus with technical training for industrial support and expanded service sectors. Pahrump has traditionally provided human resources for the Nevada Test Site, which had numerous high technology and underground construction projects. The Test Site is approximately 40 miles from Pahrump, so locals are used to relatively long commutes on a daily basis. Similarly, Tonopah, NV is a regional center to the north of Beatty, with the capability to contribute to the available labor supply.

18.1.3 Electrical Power

Electrical power is provided to the immediate area of THE PROJECT by the Valley Electric Association (VEA), Inc., which is headquartered in Pahrump, Nevada. A 130-kV line runs along Highway 95 to a location several miles to the south of the access road to the MLP in Fluorspar Canyon then connects into the Beatty Substation. Current substation capacity at Beatty is projected to be sufficient to meet the needs of the required pressure oxidation processing plant at Mother Lode. A single-phase powerline currently runs along Fluorspar Canyon to provide power to a communications tower located in the Bare Mountains above the Secret Pass open pit. A high capacity powerline would need to be constructed to meet the Mother Lode requirements. To provide power at the NBP, a 25-kV line runs north from Beatty along US Highway 95. VEA has recently upgraded the main powerline, which now exceeds the projected requirements for NBP with 15 Mw capacity. Corvus’s NBP requirements were considered in the upgrade of the powerline.

At NBP, two electrical feeder lines run west from the main line, one to the perimeter of the NBP property to power an aggregate crushing plant operating in the southern portion of NBP and a second line traverses the property to power a centrally located communication station and the Company’s weather station which has been installed on Corvus controlled patented mining claims near Mayflower.

18.1.4 Water Resources

Water resources for mining at MLP and NBP must be obtained from the ground water in the Crater Flats (Basin 229), Oasis Valley (Basin 228) and Sarcobatus (Basin 146) hydrographic basins.

The MLP access road in Fluorspar Canyon is located across Highway 95 from the termination of the water network operated by the Beatty Water and Sanitation District (BWSD). It is anticipated that MLP process water will be provided by a utility agreement with BWSD, which has substantial, existing and unused water rights in the Oasis Valley and Amargosa hydrographic basins. Historic ground water permits from the Mother Lode and Daisy mining operations are being re-activated by Corvus to provide for site water and for dust suppression for future exploration and operations at the MLP.

In 2014, Corvus purchased a 430-acre property located 48 kilometres (30 miles), along Highway 95, to the north of NBP in the Sarcobatus hydrographic basin which included a 1600-acre-foot water right. In 2018, Corvus applied for and received conversion of the water right to “mining application” by the Nevada Division of Water Resources. Historic production testing of a single well at the property indicated that a 270-foot-deep well could produce water at the permitted flowrate. Electrical power is available at the Sarcobatus property, and a pipe line could be constructed through the Sarcobatus alluvial valley to NBP, which lies 32 km (20 miles) to the south. Alternatively, the northeast corner of NBP lies within the Sarcobatus hydrographic basin, and Corvus would explore for a water production well field in the basin nearer the Project and make application to the Nevada State Engineer for a temporary relocation of the production location.

Water wells, booster stations and pipelines would be developed to transport water to the mines, process facilities and ancillary structures.

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18.2 Project Specific Infrastructure Elements - MLP

A conceptual configuration of the project facilities at MLP is shown in Figure 18-1.

18.2.1 Mill Facility

A mill facility would be constructed north of the Mother Lode open pit, and would consist of primary, secondary and tertiary stage crushing equipment, a gravity concentrator, a flotation plant, a pressure oxidation vessel, oxygen plant and leach tanks. Minimal plant buildings would be required due to the relatively mild climate at MLP. The mill facility would be open air, with some portions of the plant enclosed. The refinery and laboratory facilities would require climate-controlled buildings.

As described in Section 17, the plant would have the capacity to bypass the flotation and pressure oxidation stages and would process YellowJacket vein and vein stockwork oxide mineralization approximately half of the year and would process Mother Lode sulfide mineralization through the pressure oxidation circuit the other half of the year.

18.2.2 Tailing Storage Facility

Locations with relatively flat open terrain to the north of the MLP open pit have sufficient area for both the heap leach pad and a tails storage facility (TSF). The MLP conceptual configuration would locate the TSF north of a low ridge and to the next to the heap leach pad as shown in Figure 18-1. Up-stream raising of the embankments is projected for incremental increases of the storage capacity. A thickened tail, using high rate thickeners, is projected to favorably impact water consumption and shorten the TSF reclamation period. It will also reduce the construction requirements to build the embankments.

18.2.3 Heap Leach Pad and Ponds

The heap leach pad would be located north of the Mother Lode open pit and is shown in Figure 18-1 with an approximate total storage capacity of 40 M tonnes of mineralized material. The final design of the leach pad would include construction specifications for a prepared subgrade, soil underliner, HDPE liner, solution collection system and overliner layer. The leach pad would be constructed in phases with three expansions added to the initial leach pad constructed.

The final design would include a solution management system that connects the solution collection system within the leach pad, to the process pond and CIC array. It is intended that the final design would show the compartmentalization of the leach pad into solution collection cells and the individual cells would have flow measuring points and sampling locations. From this point, the solution would be directed to the pond or the CIC array, as necessary for maximizing recovery. Adjacent to the process pond would be an event pond and the two ponds would be connected by a spillway. The size of the ponds would be based on the 100-yr, 24-hr event and the 24-hr drain down of the leach pad, as well as, 0.6 metres (two feet) of freeboard per state regulations. Similar to the heap leach pad, the ponds would have a prepared subgrade, soil underliner and two layers of HDPE liner for leak detection and primary containment, respectively.

18.2.4 Ancillary facilities

The MLP would require the design and the construction of several ancillary structures for the day to day operations of the mine. These facilities would include a truck/tire shop, warehouse, administration, and clinic, wash bay, fuel island, assay/MET lab, security and maintenance facilities. These structures would be centrally located near the process plant and leach pad with access points to each of the facilities from the existing site access roadway. The aforementioned would be constructed in the first phase of mine development and infrastructure construction. To the extent possible, the facilities would be of a temporary nature consistent with the currently envisioned short life of the Project.

18.2.5 Surface Water Management Facilities

The location of the heap leach pad, ponds, and process plant would require diversion structures to redirect surface runoff around these facilities and discharge into natural drainage channels. For ease of construction, the channels were assumed to be 3 metres (10 feet) wide at the bottom of the channel, 1.5 m (5 feet) deep, and have side slopes of 3:1. The channel would be lined with a geotextile fabric and riprap along the entire length of the channel and up the side slopes to protect the channel for a flow depth of 1.2 m (4 feet).

18.2.6 Waste Rock Management Facilities

Based on the most recent mine production schedule, 178 M tonnes of waste rock would need to be stockpiled. Figure 18-1 shows the conceptual location of the waste rock stockpile. A portion of the stockpiles may require underdrain material with neutralizing characteristics, which would be comprised of the dolomite rocks that form a part of the stripped waste material.

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18.3 Project Specific Infrastructure Elements - NBP

A conceptual configuration of the NBP is shown in Figure 18-1.

18.3.1 Site Access Road

Access to the NBP would be from US 95 just north of the Town of Beatty and east along the Strozzi Ranch Road. The site access would follow this existing roadway corridor for approximately 7.5 kilometres (4.6 miles) and would require improvement to allow two-lane traffic. The access to the operations site from the access road would be designed so it would be secured with a perimeter fence and security gate, but historical access along the Strozzi Ranch Road by others would not be impeded.

The current concept envisions hauling the mill grade mineralization from North Bullfrog to the Mother Lode mill for processing. An intermediate stockpile would hold the mineralization with a feeder, conveyor and bin system, located to the east of the NBP Heap Leach Pad, to load contracted 40 tonne highway trucks with tandem trailers. A dozer would push mineralization from the stockpile into the feeder. The mineralization would be trucked to Highway 95 along Strozzi Ranch Road, then on the highway through Beatty, and then to the Mother Lode process facility via the access road up Fluorspar Canyon. Stockpiles located at the MLP mill site would be used to feed the mill with the scheduled mineralization.

18.3.2 Haul Roads

The conceptual plan for loading the leach pad would be by haul trucks and dozers. The Mayflower mine requires the longest haul road with an approximate haul distance of 4.5 kilometres (2.8 miles) along the western edge of the property to the leach pad location. The mine plan calls for early mining of the Sierra Blanca/YellowJacket Resource prior to mining of the Jolly Jane Resource. A haul road within the Sierra Blanca pit would be planned for hauling from the Jolly Jane pit to the heap leach pad.

18.3.3 Leach Pad Access Road

A leach pad perimeter road would be constructed for access to the leach pad and process ponds.

18.3.4 Heap Leach Pad and Ponds

The heap leach pad would be located northwest of Sierra Blanca and is shown in Figure 18-1 with an approximate total storage capacity of 200 M tonnes of mineralized material. The final design of the leach pad would include construction specifications for a prepared subgrade, soil underliner, HDPE liner, solution collection system and overliner layer. The leach pad would be constructed in phases with three expansions added to the initial leach pad constructed.

The final design would include a solution management system that connects the solution collection system within the leach pad, to the process pond and CIC array. It is intended that the final design would show the compartmentalization of the leach pad into solution collection cells and the individual cells would have flow measuring points and sampling locations. From this point, the solution would be directed to the pond or the CIC array, as necessary for maximizing recovery. Adjacent to the process pond would be an event pond and the two ponds would be connected by a spillway. The size of the ponds would be based on the 100-yr, 24-hr event and the 24-hr drain down of the leach pad, as well as, 0.6 metres (two feet) of freeboard, per state regulations. Similar to the heap leach pad, the ponds would have a prepared subgrade, soil underliner and two layers of HDPE liner for leak detection and primary containment, respectively.

18.3.5 Ancillary facilities

The NBP would require the design and the construction of several ancillary structures for the day to day operations of the mine. These facilities would include a minimal truck/tire shop, warehouse, administration, clinic, wash bay, fuel island, security and maintenance facilities. Major equipment maintenance would be performed at the Main Truck Shop located at Mother Lode. These structures would be centrally located near the leach pad with access points to each of the facilities from the existing site access roadway from US 95. The aforementioned would be constructed in the first phase of mine development and infrastructure construction. To the extent possible, the facilities would be of a temporary nature consistent with the currently envisioned short life of the Project.

18.3.6 Surface Water Management Facilities

The location of the heap leach pad, ponds, and leach columns would require diversion structures to redirect surface runoff around these facilities and discharge into natural drainage channels. For ease of construction, the channels were assumed to be 3 metres (10 feet) wide at the bottom of the channel, 1.5 m (5 feet) deep, and have side slopes of 3:1. The channel would be lined with a geotextile fabric and riprap along the entire length of the channel and up the side slopes to protect the channel for a flow depth of 1.2 m (4 feet).

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18.3.7 Waste Rock Management Facilities

Based on the most recent mine production schedule, 112 M tonnes of waste rock would need to be stockpiled. Figure 18-1 shows the conceptual locations of the waste rock stockpile on the east side of the Sierra Blanca pit and west of the Jolly Jane pit, and east of the Mayflower pit.

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19 Market studies and contracts

No market studies have been undertaken by the Project at this time, and no contracts have been discussed for the sale of the gold which may be produced at the Project. It is assumed that the process facilities at the Mother Lode mill would produce a gold doré with high purity, which will be shipped to a commercial refiner such as Johnson Matthey in Salt Lake City. All-in charges from such refiners are currently in the range of US $1.50-2.00/Oz, based on a minimum one-year contract at quantity levels consistent with this Project. Sales price would be based on the spot price of gold. A gold price of $1,250 per ounce has been assumed for the life of the mine.

Gold is readily sold on the spot market, and historically has not been a demand limited commodity. This PEA assumes that gold with be sold at spot price and this assumption is considered to be reasonable by the Author.

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20 Environmental Studies, Permitting and Social or Community Impact

Corvus currently has permits to conduct exploration activities at NBP with both the Nevada Division of Environmental Protection (NDEP)-Bureau of Mining Regulation and Reclamation (BMRR) and the Bureau of Land Management (BLM). Those permits allow 20 acres and 120 acres of surface disturbance on the private and public land, respectively. The permits for activities on the public lands are based on an Environmental Assessment that contains environmental baseline data on wildlife, climate and local physical characteristics.

In addition, Corvus has permits to conduct exploration with the BLM at three areas with the MLP (Sawtooth, Willys, and Mother Lode). These permits are Notices for exploration disturbance of less than five acres each. Corvus is also in the process of obtaining permits from BLM and BMRR for 145 acres of exploration related disturbance. The permits for activities on the public lands are based on an Environmental Assessment that contains environmental baseline data on wildlife, climate and local physical characteristics.

20.1 Existing Environmental Baseline Characterization Activities

Corvus has developed characterization plans which describe the on-going and future collection of baseline environmental data that will be required to support a future mine permitting process at NBP and MLP. Current baseline characterization activities include:

Geochemical characterization of waste rock geochemistry – At NBP, Acid-Base Accounting (ABA) characterization of waste rock as defined in the current mining plans using static tests have been completed and a first phase of Humidity Cell Tests (HCT) have been completed for the waste rock associated with mining the heap leach disseminated mineralization. The second phase of HCTs are being defined to characterize waste associated with the mining of the YellowJacket mineralization.

At NBP, hydrologic characterization testing has been performed during installation of 12 ground water monitoring wells. Collection of water quality data began in Q4 2012 on a quarterly basis for the ground water monitoring wells and a group of 12 surface springs surrounding NBP. NDEP Profile I parameters are reported for each ground water and spring sample.

Surveys of plants and wildlife have been conducted on a large portion of NBP and the Mother Lode area of MLP, including special surveys for bats and desert tortoise. Desert tortoise range is currently limited to the eastern portion of NBP, outside of the areas containing currently defined Mineral Resources. Recent habitat studies of the eastern portion of NBP have indicated that the area is not critical tortoise habitat.

A meteorological station has been in operation on the NBP site since August 2012.

These studies address some of the baseline data whose collection is time critical to production of a mining plan of operations which would serve as the basis to initiate the BLM’s National Environmental Policy Act (NEPA) process (likely through the preparation of an environmental impact statement (EIS) that is required for the processing of a Plan of Operations). There will be additional baseline data collection requirements of which timing could affect the schedule for completion of the EIS. In addition, other baseline characterization activities would be required but would not control the schedule for completion of the EIS.

No known environmental issues have been identified at the NBP or MLP sites that would materially affect the current mine design or scope of the needed environmental permits. Geochemical characteristics of the waste rock suggest that no acid generation and only minor metals leaching would be expected from the waste materials associated with the heap leach mineralization. Ground water quality is typical of the regional data and drilling activities suggest minimal water inflow because only a small portion of mining would be below the water table.

20.2 Permits Required for Future Mining Activities

This section of the Technical Report summarizes the permits that will likely be required to conduct mining activities at the NBP and MLP. The details of the mine area and activities are not well defined at this time. However, some general design criteria are known. Both NBP and MLP will be open pit mining operations, and each would have associated waste rock dumps and a heap leach ore processing facility. At MLP there will also be milling operations with an associated tailings storage facility. The mill will have both a carbon-in-leach (CIL) process and a pressure oxidation process.

In order to conduct mining and processing activities, NBP and MLP will each need specific permits from the NDEP BMRR and the BLM. The following is a list of the major permits that will be required followed by a brief discussion of each. None of the permits are currently in application stage.

· Plan of Operations/Nevada Reclamation Permit;

· Water Pollution Control Permit;

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· Air Quality Operating Permits;

· Water Rights; and

· Industrial Artificial Pond Permit.

20.3 Plan of Operations / Nevada Reclamation Permit

A Plan of Operations/Nevada Reclamation Permit (Plan) is a joint application that is submitted to the BLM and NDEP-BMRR that utilizes a format accepted by the BLM and BMRR. The application will describe the operational procedures for the construction, operation and closure of the Project. As required by the BLM and BMRR, the Plan will include a waste rock management plan, quality assurance plan, a storm water spill contingency plan, reclamation plan, a monitoring plan and an interim management plan. In addition, the Plan includes a Reclamation Cost Estimate for the closure of the Project. The mine design must be completed prior to submittal of the Plan.

20.4 Water Pollution Control Permit Application

The Water Pollution Control Permit (WPCP) application must address the open pit, waste rock dump, heap leach pad, mining activities and the water management system, as well as the potential for these facilities to degrade waters of the state. The application includes an engineering design for the waste rock dump, a waste characterization report and a modeling report for the closure of the waste rock dump, as well as an engineering design for the water management system.

A Tentative Permanent Closure Plan must also be completed and submitted to the NDEP-BMRR in conjunction with the WPCP. A Final Permanent Closure Plan will need to be developed two years prior to Project closure.

20.5 Air Quality Operating Permits

An Application for a Class II Air Quality Permit for those portions of the stationary source that have the potential to emit pollutants must be prepared using Bureau of Air Pollution Control (BAPC) forms. The Application includes a description of the facility and a detailed emission inventory. The Application also includes locations, plot plans and process flow diagrams. The Application must also include a fugitive dust control Plan to be used during construction and operation of the Plan. If the facility will process loaded carbon or electrowinning precipitate, then a Mercury Operating Permit application and a Title V Operating Permit application will also be necessary, which will have to address the necessary state and federal mercury controls, respectively.

20.6 Water Rights

Water rights will need to be obtained from the Nevada Division of Water Resources (NDWR) to remove and utilize the water from the mining operation and to provide water for the public water system.

20.7 Industrial Artificial Pond Permit

The development of the water storage pond, which is part of the water management system, will require an Industrial Artificial Pond Permit (IAPP) from the Nevada Department of Wildlife.

20.8 Minor Permits and Applications

In addition to the above noted permits, Table 20-1 lists other notifications or ministerial permits that will likely be necessary to conduct the mining operations.

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Table 20-1 Required Minor Permits and Applications

Notification/Permit	Agency	Comments
Mine Registry	Nevada Division of Minerals	-
Mine Opening Notification	State Inspector of Mines	-
Solid Waste Landfill	Nevada Bureau of Waste Management	-
Hazardous Waste Management Permit	Nevada Bureau of Waste Management	-
General Storm Water Permit	Nevada Bureau of Water Pollution Control	-
Hazardous Materials Permit	State Fire Marshall	-
Fire and Life Safety	State Fire Marshall	-
Explosives Permit	Bureau of Alcohol, Tobacco, Firearms and Explosives	Mining contractor may be responsible for permit
Notification of Commencement of Operation	Mine Safety and Health Administration	-
Radio License	Federal Communications Commission	-
Public Water Supply Permit	NV Division of Environmental Protection	-
MSHA Identification Number and MSHA Coordination	U.S. Department of Labor Mine Safety and Health Administration (MSHA)	-
Septic Tank	NDEP – Bureau of Water Pollution Control	-
Petroleum Contaminated Soils	NV Division of Environmental Protection	-

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21 Capital and Operation Costs

Capital and operating costs used for the PEA are based on evaluation of multiple sources of information. Active Nevada mining operations based on exclusive heap leach processing and combined heap leach processing and mill processing of gold ores were visited in the period November-December, 2011, again in October of 2014. Operating costs, consumable costs and construction costs were discussed with mine site staff during the visits and were used as bench marks for establishing the PEA estimates. Capital and operating costs were also developed using information available from the CostMine cost data service for 2018 by InfoMine USA, Inc. In addition, all available Project technical data and metallurgical process related test work were considered to build up a processing operating cost estimate.

Preliminary site infrastructure (mill, tailing storage facility, heap leach pad, overburden storage facility, roads, shops, offices, etc.) have been evaluated and a conceptual arrangement was defined as the basis of capital costs estimates, as described in Section 18. Capital costs were developed based on the nominal mining rates of 66 M tonnes of total material per year. The capital cost includes estimates of sustaining capital and all facilities and equipment needed for all phases of the Project over its projected 9 years of active mining. The mobile equipment was assumed to be financed over its first 5 years of use with 20% down payment and assuming 6% annual interest. All costs are in constant USD from 2018.

Cost accuracy is estimated to be + or – 30%, in the opinion of the Author.

21.1 Capital Cost Estimate

Capital cost estimates are described in the general categories of initial capital and sustaining capital.

21.1.1 Initial Capital Cost Estimate

The initial capital cost estimates are listed in Table 21-1 and consist of go-forward costs to be incurred after approval of a Plan of Operations, and after construction/operating permits have been received. Capital estimates are based on the project configuration at both Mother Lode and North Bullfrog as described in Section 18. It covers the schedule period year -1, and includes all construction costs up to the start of production, which is defined as when the first mineralized material is placed on the leach pad and the first mineralization is processed in the mill. The scope of the initial capital includes direct capital costs, indirect costs, Owner’s costs and EPCM and contingency. Direct capital costs include the, heap initial leach pad construction, heap leach pumps and ponds, YellowJacket mineralization loading facility, carbon columns, mining equipment, shops, offices, mill facilities, and site infrastructure. A mill facility with 8,200 tonne/day crushing and grinding circuit, flotation, pressure oxidation, leach and CIL circuits, carbon ADR plant, tailings storage facility and mill water storage ponds/tanks was assumed to constructed at Mother Lode, and higher-grade mineralization produced from Sierra Blanca/YellowJacket at North Bullfrog was assumed to be hauled to the mill facility at Mother Lode using Highway 95. Pre-production mining would occur at both mining sites in the construction year -1 to ensure that mill feed would be available in year +1, and to allow the project to develop a mill stockpile sufficient to allow higher grade mill mineralization to be supplied to the mill selectively during the early portion of the schedule (grade-streaming). Indirect costs include Engineering, Procurement and Construction Management (“EPCM”) which is applied as a proportion of direct costs ranging from 18% to 7% depending on the complexity of the capital item. Owner’s costs include an allowance for property maintenance, owner project management, and the expansion and training of the mine management and labor force. Contingency was set at proportions of the capital item cost ranging from 25% to 10% depending on the complexity of the capital item excluding owner’s cost, mobile equipment and EPCM.

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Table 21-1 NB-MLP – Initial Capital Cost Estimate

Capital Item	Estimated Capital Cost (USD $M)
Initial Direct Capital Cost	$323.9
EPCM	$38.9
Contingency	$54.7
Owner’s Cost	$6.0
Total	$423.5

The direct initial capital costs are further subdivided in Table 21-2. Only the first year, down payment on the mobile equipment was included in Table 21-2.

Table 21-2 NB-MLP Project – Initial Direct Capital Cost Estimate (including Owner’s Cost, EPCM &

Contingency)

Capital Item	Estimated Capital Cost (USD $M)
Mill	$193.4
Heap Leach	$13.4
Mobile Equipment	$17.4
Infrastructure & Facilities	$42.9
Capitalized Mining	$56.8
Total	$323.9

For indirect initial capital costs, listed in Table 21-3, EPCM was ranged from 18-7% of the initial capital costs depending on complexity of the item, and Contingency ranged between 25-10%.

Table 21-3 NB-MLP Project - Initial Indirect Capital Cost Estimates

Capital Item	Estimated Capital Cost (USD $M)
EPCM	$38.9
Owner’s Cost	$6.0
Contingency	$54.7
Total	$99.6

21.1.2 Sustaining Capital Cost Estimate

Sustaining capital cost estimates included all capital costs that would be incurred after production starts (year +1 to year 9). It included estimated capital for expansion of production capability (leach pad expansion, Tails Storage Facility, haul road construction, etc.) and replacement capital (mobile equipment overhaul or replacement of worn out equipment). Annual principal payments for the mobile equipment were also included in sustaining capital. Sustaining capital cost estimates are listed in Table 21-4, along with the remaining EPCM and contingency depending on the complexity of cost items (the mobile equipment and working capital). EPCM is greater than contingency in Table 21-4 because it has been applied to the mobile equipment principal, but contingency has not. Mining costs for mineralization stockpiled has been transferred to capital in the year it is mined, and credited back to operating cost in the year it is processed, to that the next effect is neutral over the LOM.

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Table 21-4 NBP – Sustaining Capital Cost Estimates and Remaining Contingency

Capital Item	Estimated Capital Cost (USD $M)
Sustaining LOM	51.9
Remaining EPCM	3.3
Remaining Contingency	4.9
Total	$60.1

21.1.3 Working Capital and Initial Fills

Working capital was estimated to be equivalent to operating costs for the first 3 months of production, or $40.1 M., and is not included in Table 21-4. Working capital is credited out of the project capital costs at the end of production year 2.. Initial fills costs were estimated to be $0.6 M and were recovered by transferring to operating costs in the last year of active operation.

21.1.4 Life of Mine Capital Estimate

Total estimated capital costs are listed in Table 21-5 and reflect costs transferred to opex and recovered from accounting cash flow.

Table 21-5 NB-MLP PEA - Total Capital Cost Estimate

Capital Item	Estimated Capital Cost (USD $M)
Total Initial Capital	423.5
Sustaining Capital	157.6
Total Estimated Expenditure	581.2
Capitalized Mining Transferred to Opex in Year 1	($56.8)
Working Capital Recovery End Year 2	($40.1)
Initial Fills Recovery End of Life	($0.6)
Total Project Capital	$483.6

21.2 Operating Cost Estimates

Operating cost estimates were developed from spreadsheet cost models built up from first principles. These spreadsheets considered the schedule of production physicals, assumed equipment productivities, consumables and operating maintenance and production labor. The projected unit costs were compared to benchmark information for the Nevada heap leach operations visited in late 2014 and in public reports of Nevada operating mines. The process cost estimation spreadsheets were developed separately for heap leach processing and for the mill processing.

Operating costs were estimated for the categories of mining, processing, administration and reclamation. Table 21-6 list the operating cost estimates used in the economic analysis.

Table 21-5 NB-MLP Project - Average Unit Operating Cost Assumptions

Operating Cost Area	LOM Average Operating Cost ($/ tonne processed)	LOM Average Operating Cost ($/Au oz)
Mining(1)	2.95	$312
Heap Leach Processing(2)	1.20	$114
Mill Oxide Processing(2)	16.74	$65
Mill Sulfide Processing(2)	18.89	$132
Administration(2),(3)	0.53	$56
Reclamation(2)	0.12	$13
Total	na	$692
(1)-total operating cost/total processed tonnes
(2)-operating cost per process specific tonnes
(3)- includes royalties, financing interest, transport & refining charges

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Mining costs were estimated at $1.42 per unit tonne of mineralized material or overburden material. The mill processing cost estimate averaged $16.74 per oxide tonne processed (includes transport from NBP to Mother Lode), $18.89 per sulfide tonne processed and heap leach processing cost was estimated to average $1.20 per tonne processed.

Reclamation costs are applied over a 3-year period (year 8, 910,11 and 12). No salvage income is assumed.

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22 Economic Analysis

This PEA is preliminary in nature and is based on technical and economic assumptions which will be evaluated in more advanced studies. The PEA is based on the Mineral Resource Estimate in Chapter 14 (effective date September 18, 2018). The PEA is based on a production plan that includes material in Measured, Indicated and Inferred classifications from both Resource models (NBP and MLP). This report assumes that the Project would support an open pit mining operation that recovers gold and silver metals from the ground to be sold at a profit.

This PEA is preliminary in nature, it includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the preliminary economic assessment will be realized. The current basis of Project information is not sufficient to convert the in-situ Mineral Resources to Mineral Reserves, and Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. The PEA results are only intended as an initial, first-pass review of the Project economics based on preliminary information.

The economic analysis of the Project assumed constant 2018 US dollars and was performed on an annual basis beginning at the start of year -1 when operating permits are assumed to have been issued. Construction was assumed to require 1 year with placement of mineralized material on a heap leach pad and mill processing to start at the beginning of year +1. This PEA assumes a central mill processing facility constructed at the MLP which has a circuit that will be configured to process oxide vein and vein stockwork mineralization from the Sierra Blanca/YellowJacket deposit at NBP and sulfide mineralization from the MLP and NBP. The mill grade mineralization at NBP would be loaded on to highway trucks and hauled to the mill site at Mother Lode using Highway 95. Mining would be concurrent at both sites. Lower grade, oxide mineralization at both locations would be processed on heap leach pads. Gold would be captured from the leach solutions using carbon-in-pulp columns, and the carbon from North Bullfrog would be hauled to the Mother Lode mill for metal stripping and refining.

The PEA estimated mill gold recoveries based on bottle roll testing and gravity concentration testing performed on the YellowJacket vein and stockwork materials which indicate high gold and silver recovery using simple CN leaching. Sulfide mineralization exists at both Mother Lode and North Bullfrog and would require the addition of a flotation and a pressure oxidation circuit to the mill. Gold recoveries for the mill are estimated to vary from 80% - 87% for Mother Lode and North Bullfrog mill mineralization, respectively. Metallurgical testing programs were carried out by McClelland Laboratories, Hazen Research Inc. and Resource Development Inc. on composite samples from both deposit to provide the basis of the mill circuit configuration. Heap leach recoveries were based on column leach testing data for composite samples constructed from Mayflower, Jolly Jane, Savage Valley and Sierra Blanca 2012 PQ core drilling for North Bullfrog. Historical data from the operating heap leaches at Mother Lode and the Daisy Project in 1988-1999 were combined with laboratory bottle roll and shake leach testing from samples from Corvus drilling at Mother Lode were used to estimate heap leach recoveries for Mother Lode oxide mineralization.

The mill process recovery assumptions reflected overall gold and silver recoveries based on the flow diagram shown in Section 17. The average mill recoveries for gold were 82.9% and 74.3% for silver.

The heap leach process recovery assumptions reflected consideration of the particle size resulting from ultra-high intensity blasting to produce a gradation similar to primary crushing (P80 -76mm) and the leach pad placement schedule. The leach pad production model predicts an average gold recovery of 73.8%, and an average silver recovery of 6% of fire assay grade. The production model assumes a 3 year buildup of gold in solution inventory which would require 3 years of rinsing after the final leach pad placement to recover inventory. No cost escalation was included in the calculations, and the cash flows were presented after-royalty and after-tax. A gold price of $1,250 per ounce was assumed for all years (1-8+) for the base case. All economic projections were made on an after-royalty and after-tax basis.

The analysis included Measured, Indicated and Inferred Mineral Resources in the mining and economic study. Measured and Indicated Resources make up 81% of the gold ounces in the total production plan. The remaining 19% of the gold ounces in the production plan are classified as Inferred Mineral Resources. Subdivided by process, 41% of the gold ounces in the mill production schedule are classified as Measured and Indicated Mineral Resources, and 40% of the gold ounces in the heap leach production schedule are classified as indicated Mineral Resources. The Inferred Mineral Resources are subdivided as 3% processed by the mill and 16% processed by Heap Leach.

22.1 Key Performance Parameters

Mining physicals in the production schedule presented in Table 16-2 were used in conjunction with unit operating cost assumptions to estimate OPEX costs on an annual basis. Estimated capital costs were input on an annual basis from a conceptual schedule that included initial capital associated with pre-mining construction of the Project in year -1 and sustaining capital over the LOM. Mobile equipment was assumed to be financed with 20% down payment and a five-year term at 6% interest. Interest costs were transferred to administrative operating cost.

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Key performance parameters are listed in Table 22-1; Figure 22-1 shows the projected annual gold and silver production from the Project.

Table 22-1 Projected Key Performance Parameters from the NB-MLP Preliminary Economic Assessment (Constant $US, No Escalation, Constant $1,250 per Ounce Gold Price, after-Royalty and after-Tax)

Parameter	Data Value
Measured Resource within Whittle Shells*	43.9 M t at 0.55 g/t Au for 779 kozs and at 2.04 g/t Ag for 2,880 kozs
Indicated Resource within Whittle Shells*	167.6 M t at 0.40 g/t Au for 2,133 kozs and at 0.95 g/t Ag for 5,138 kozs
Inferred Resource within Whittle Shells*	81.0 M t at 0.26 g/t Au for 668 kozs and at 0.40 g/t Ag for 1,040 kozs
Post-Tax and Royalty NPV at 5%	$US 586M
Post Tax and Royalty IRR	38 %
Pre-tax cashflow ; IRR	$970 M ; ≈45%
Overall Strip Ratio	1.08:1 (overburden:mineralized material)
Average Annual Payable Gold Production years 1-4	347 kozs/year
Average LOM Payable Gold Production years	282 kozs/year
Average Gold Recovery - mill	82.9%
Average Gold Recovery- heap leach	74.3%
Average Cash Cost	$US 692/Au Oz*
Average Silver Recovery-mill	74.3%
Average Silver Recovery – heap leach	6%
Average Total Mining Rate	170 k tonne/day
Average Mineralized Material Mining Rate	82 k tonne/day
* Mill selection COG = 0.35-0.76g/t, Heap Leach selection COG= 0.06-0.1 g/t Au

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Figure 22-1 Estimated Annual Gold and Silver Production from NB-MLP (81% Measured and Indicated Resources, 19% Inferred Resources)

 

Estimated physical data for the Project PEA are listed in Table 22-2, for the estimated mine life of 9 years of active mining (followed by 3 years of leach pad rinse down).

Table 22-2 Summary of Physical Data from the NB-MLP PEA Production Schedule

Key Physical Data	Units	Value
Heap Leach Feed Mined	M tonnes	241.4
Mill Feed Mined	M tonnes	27.6
Overburden Mined	M tonnes	289.5
Total Material Mined	M tonnes	558.9
Mine Life*	Years	9
Contained Gold**	M Oz	3.26
Recovered Gold Payable	M Oz	2.54
Contained Silver	M oz	8.00
Recovered Silver	M Oz	3.49
Average Strip Ratio	Overburden/Process Feed	1.08
Average Diluted Gold Grade Heap Leach	g/t	0.23
Average Diluted Gold Grade Mill	g/t	1.69
Average Gold Recovery	%	78
Annual Process Feed Mined	M tonnes/yr	29.8
Annual Gold Produced	K Oz/yr	282
* active mining, excludes leach pad rinse period at end of mine life
** 81% Measured and Indicated Mineral Resource; 19% Inferred Mineral Resource

LOM unit costs for OPEX and capex are listed in Table 22-3.

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Table 22-3 Projected LOM Unit Operating Cost and Capital Cost per Process Tonne and per Produced Au Ounce for the Project (Constant 2018 $US, No Escalation).

Cost Area	Cost per Process tonne ($/tonne)	Cost per Recovered Gold Oz ($/Oz)
Mining	$2.95	$312
Processing	$2.94	$311
Administration	$0.53	$ 56
Reclamation	$0.12	$13
Total Operating Cost	$6.54	$692
Capital Cost	$1.80	$191
Projected Total Cost	$8.33	$883

22.2 Cash Flow

The projected annual production and cash flow (after-royalty and after-tax) for the NBP are listed in Table 22-4. The estimated payback period assuming the average gold price of $1,250 and average silver price of $15.08 per ounce is 2.1 years.

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Table 22-4 Projected Annual Production and Cash Flow (after-Royalty and after-Tax) for the Integrated Project – Base Case (Gold Price $1,250; Silver Price $15.08)

Year	Over-burden Mined (M t)	Process Feed Mined (M t)	Contained Au* (k Oz)	Payable Au1 (k Oz)	Gold Revenue1  (US $M)	Payable Ag1 (k Oz)	Silver Revenue1 (US $M)	Operating Cost (US $M)	Capital Cost   (US $M)	Pre-Tax, After Royalty Cash Flow  (US $M)	Federal Income Tax (US$M)	Nevada NPT Tax (US$M)	Cash Flow After Tax, After Royalty (US $M)
(1)	28.6	5.0	-	-	-	-	-	-	($423.5)	($423.5)	-	$0.0	($423.5)
1	49.2	16.2	434.8	317.7	$397.2	965.2	$14.6	($227.1)	($13.7)	$170.9	($3.2)	($10.5)	$157.2
2	27.2	34.0	516.8	378.4	$473.0	768.9	$11.6	($198.6)	$4.3	$290.2	($26.4)	($12.8)	$251.0
3	29.5	37.3	464.8	364.6	$455.8	500.9	$7.6	($208.3)	($12.9)	$242.2	($24.0)	($11.6)	$206.6
4	34.0	32.2	406.8	319.5	$399.4	355.7	$5.4	($205.0)	($16.8)	$182.9	($16.0)	($8.7)	$158.2
5	35.8	31.6	334.5	264.3	$330.4	206.3	$3.1	($206.4)	($25.4)	$101.7	($9.4)	($5.1)	$87.3
6	32.4	35.2	356.7	272.7	$340.9	228.6	$3.4	($203.0)	$2.3	$143.7	($11.9	($5.8)	$126.0
7	36.2	32.0	305.8	229.4	$286.8	234.8	$3.5	($198.2)	$0.1	$92.2	($7.4)	($3.4)	$81.4
8	16.1	40.1	333.1	254.6	$318.3	193.1	$2.9	($196.6)	($3.3)	$121.3	($10.7)	($1.4)	$109.3
9	0.5	4.8	87.7	92.7	$115.9	31.3	$0.5	($68.5)	$3.8	$51.6	($4.4)	-	$47.2
10	-	0.7	16.1	24.8	$31.0	2.2	$0.0	($24.7)	$1.6	$7.9	($0.3)	-	$7.6
11	-		-	6.0	$7.5	1.1	$0.0	($10.8)	-	($3.3)	-	-	($3.3)
12	-	-	-	2.4	$2.9	0.5	$0.0	($10.8)	-	($7.8)	-	-	($7.8)
LOM	289.5	269.0	3,257.1	2,527.2	$ 3,156.	3,488.6	$52.6	$(1,757.7)	$(483.7)	$970.2	($113.8)	$ (59.2)	$797.2
1-Less royalty ounces										$728.7	NPV 5%		$585.6
										45%	IRR		38%

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22.3 Sensitivity

The sensitivity of the Preliminary Economic Assessment for the Project has been evaluated for variations in the gold price assumption, gold recovery assumption, operating cost and capital cost. These sensitivities are evaluated around the base case price assumptions of an average gold price of US $1,250 per ounce, and the average gold recovery, OPEX and capex price assumptions listed in Tables 22-2 and 22-3. Table 22.5 lists the estimated Net Present Value (NPV) at discount rates of 0%, 5%, 7.5% and 10%, and the estimated Internal Rate of Return (“IRR”) for the gold price assumptions between US $1,050 and $1,450- per ounce.

Table 22-5 Projected Sensitivity of Net Present Value and Internal Rate of Return to Variation in Gold Price (after-Royalty and after-Tax)

Gold Price ($/Oz)	Total Cash Flow (US $M)	NPV @ 5% (US $M)	NPV @ 7.5%  (US $M)	NPV @ 10% (US $M)	IRR (%)	Payback (years)
$1,050	$366.6	$236.7	$184.2	$138.3	20%	2.9
$1,150	$584.4	$413.2	$344.5	$284.6	30%	2.4
$1,250	$797.2	$585.6	$501.0	$427.5	38%	2.1
$1,350	$1,007.8	$756.0	$655.8	$568.8	46%	1.8
$1,450	$1,202.5	$914.0	$799.3	$699.9	53%	1.7

Sensitivity to the proportional change from the base case economic projection, derived at an average gold price of US $1,250 per ounce and gold recovery, OPEX and capex unit costs listed in Tables 22-2 and 22-3, were estimated for a nominal range of + 25% to - 25% from the base case assumptions. The sensitivity is shown graphically for NPV @ 5% and for IRR in Figures 22.2 and 22.3, respectively.

Figure 22-2 Sensitivity of Estimated NPV @ 5% (after-Royalty and after-Tax) for Changes in Cost, Gold Recovery or Gold Price as a Percent of the Base Case at a Gold Price of $1,250 per Ounce, Gold:Silver Price Ratio of 82.9, 78% Gold Recovery and Cost as Defined in Table 22-3

 

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Figure 22-3 Sensitivity of Estimated IRR (after-Royalty and after-Tax) for Changes in Cost, Gold Recovery or Gold Price as a Percent of the Base Case at a Gold Price of $1,250 per Ounce, Gold:Silver Price Ratio 0f 82.9, 78% Gold Recovery and Cost as Defined in Table 22-3.

 

The sensitivity analysis indicates that the Project would be most sensitive to gold price and gold recovery assumptions. The PEA was less sensitive to changes in cost, with changes in capex having a greater effect than changes in OPEX.

22.4 Taxes, Royalties and other Interests

Corvus would be subject to the following taxes as they relate to the Project:

· Nevada Net Proceeds Tax

· Federal Income Tax

Corvus would also be subject to royalties as described in Section 22.4.3.

Estimates of these taxes and royalties were made based on the production schedule in Table 22-4 and operating and capital cost estimates described in Section 21.

22.4.1 Nevada Net Proceeds Mineral Tax

In Nevada, if the net proceeds of a mine in the taxable year totals US$4 million or more the tax rate is 5%. The gross proceeds from the sale of the minerals minus certain allowable deductions were used to estimate the taxable net proceeds. The Nevada net proceeds tax is calculated before deductions of Federal income tax. In general, all operating costs and capital costs directly related to the mining operation are deductible, using Nevada depreciation and depletion schedules.

22.4.1.1 Federal Income Tax

Corporate Federal income tax was estimated by computing the regular federal income tax as modified in 2018. Regular tax was estimated by subtracting Nevada Net Proceeds Mineral tax, all allowable operating expenses, overhead, depreciation, amortization and depletion from revenues on an annual basis to estimate the taxable income. The highest effective corporate income tax was 21%.

22.4.2 Depletion

Generally speaking, depletion, like depreciation, is a form of cost recovery. Just as the owner of a business asset is allowed to recover the cost of an asset over its useful life, a miner would be allowed to recover the cost of the mineral property. Depletion was taken over the projected period that minerals would be extracted.

For federal income tax purposes, two forms of depletion are allowed: cost depletion and percentage depletion. The taxpayer is required to use the method that will result in the greatest deduction.

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22.4.2.1 Cost Depletion

Cost depletion was estimated based on the adjusted basis of the depletable property multiplied by the units of mineralized material projected to be produced over the production schedule in Table 16-1.

22.4.2.2 Percentage Depletion

Under the percentage depletion method, a flat percentage of 15% of adjusted gross income from gold mining was used to estimate the depletion allowance. However, the deduction for depletion cannot exceed 50% of the adjusted taxable income from the activity. This limitation was computed without regard to the depletion allowance. The amount of the deduction allowable under percentage depletion is not limited by the basis of the property, except for AMT purposes. Thus, even though the basis of the property would be reduced by the amount of depletion taken, if the basis becomes zero, the depletion based on the percentage of adjusted gross income may continue to be claimed for tax purposes.

22.4.2.3 Depreciation

Cost recovery for capital invested was estimated using standard depreciation schedules specified for different types of investment. The estimated cost recovery for calculation of Federal income tax included the 7 years 200% declining balance calculation, a expense 73% with 6% for the next 4 years and the final 3% in the last year calculation, and a units of production depreciation schedule. Both an alternative minimum tax and regular tax depreciation was estimated.

22.4.3 Royalties

The calculation of estimated royalties was based on projected mining production underlying individual leases to which royalties apply. The royalty status of the various patented claim blocks at North Bullfrog is discussed in Section 4.1, Table 4.1. Mother Lode has a 1% royalty at the gold price of $1,250 per Au ounce. Where lease agreements provide for royalties and had gold production in the plan (specifically Mayflower, Jolly Jane, [Sussman and Kolo Corp.] and Mother Lode), annual gold and silver production from those claim blocks has been projected and used to estimate royalty ounces of gold and silver. Total royalty ounces were estimated to be 14,570 Au ozs, or 0.6% of the total estimated gold production. Those royalty ounces were deducted from the produced gold to the calculate payable gold and silver production. In this study, there was no silver production estimated and attributable to any of the royalty properties.

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23 Adjacent Properties

There are two adjacent properties that are relevant to the NBP; the Sterling Mine, recently acquired by Coeur Mining, Inc. and the Reward Gold Project owned by Waterton Resources Limited (previously owned by Atna Resources, Limited). Both properties have published NI 43-101 technical reports that can be found under the respective owner’s profiles on SEDAR.

The overall author, Scot Wilson, has been unable to verify the information available with respect to the adjacent properties, and such information is not necessarily indicative of the mineralization at the NBP.

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24 Other Relevant Data and Information

There is no other relevant data or information for the Project.

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25 Interpretations and Conclusions

This report was prepared by a group of independent consultants, all Qualified Persons as defined by NI 43-101, to demonstrate the economic viability of open pit mining and processing, based upon the estimated Mineral Resources at NB-MLP. This report provides a summary of the results and findings to the level that should be expected for a preliminary economic assessment. Standard industry practices and assumptions have been applied in this study.

This report is based on all available technical and scientific data available as of September 18, 2018, the effective date of the report. Mineral Resources are considered by the QP to meet the reasonable prospects of eventual economic extraction due two main factors; 1) cutoff grades are based on scientific data and assumptions related to the project and 2) Mineral Resources are estimated only within pit limits derived by the scientific data as well as by using generally accepted mining and processing costs that are similar to many projects in Nevada. Additionally, many of the costs were derived specifically for this Project. Confidence in the Mineral Estimate was used to classify Mineral Resources based upon drill hole spacing, geological knowledge of the deposits, metallurgical studies and a proper QA/QC program.

25.1 Preliminary Economic Assessment

Estimated Mineral Resources were assumed to be processed with commonly utilized recovery methods which are in operation throughout the state of Nevada. Processing of lower grade oxide materials on heap leach pads is widely utilized. High grade gravity-separable material as well as sulphide mineralization would be processed through the ML mill. The ML mill would process NBP oxide as well as NBP and ML sulphide materials from multiple stockpiles. NBP Oxide material would be processed in a conventional gravity/CIL mill. Sulphide materials from the Mother Lode and NBP would be processed from stockpiles through a milling and flotation plant to produce a concentrate. The flotation concentrate would be thickened, stockpiled, and processed through a small pressure oxidation circuit followed by the CIL circuit. The Mill is anticipated to process 8,200 tpd continuously, 365 days per year. Again, these are all generally accepted metal recovery methods in Nevada.

Under the base case assumptions for the project, the preliminary economic assessment indicates an undiscounted pre-tax cash flow of $970 million, and a post-tax NPV 5% of $586 million.

The results of sensitivity analyses of post-tax cash flow and post-tax IRR show that the project is most sensitive to recovery and gold price while the Project is least sensitive to changes in capital costs. A 20% decrease in gold price results in a positive NPV5% of $149 million, at a gold price of $1,450 per ounce the NPV5% results in $914 million.

The base case assumptions demonstrate that the Project may produce an average of 347 thousand ounces of gold in years 1 through year 4. The Project may produce 282 thousand ounces per year averaged through the entire life of the mine.

25.2 Geology and Exploration

The Phase I and Phase II drilling programs at Mother Lode not only added known mineralization but also greatly increased the stratigraphic and mineralization knowledge at the Project. These successful programs grew the Resource throughout 2017 and 2018 and added continuity to the mineral estimate.

MLP drilling and subsequent block modeling have indicated that the interpolated mineralization is open and has not been “closed off” by exploration drilling. Exploration drilling in these areas will add confidence in the mineral estimate. It is critical that the extent of the high-grade feeder zone be delineated with continued exploration drilling.

During 2012-2104 Corvus delineated the YellowJacket zone of Sierra Blanca. This represents a completely blind discovery of a previously unrecognized style of mineralization at NBP. A 3D IP survey, a gravity survey in the eastern Steam-heated Zone along with more detailed structural/geologic mapping in early 2015 have provided the basis for target generation on the rest of NBP. These structural targets and the general Jolly Jane and Sierra Blanca areas are the priority for future work at the NBP. There are also other alteration and geochemical anomalies throughout in the Eastern Steam-heated Alteration Zone that the Project should evaluate.

Jolly Jane, Sierra Blanca, Yellow Jacket and the Savage Valley areas contain significant low-grade gold-silver deposits. Drilling has shown that the Sierra Blanca Tuff is a primary host rock for broad areas of disseminated semi-stratabound gold mineralization. Much of the mineralization is near surface, oxidized, and metallurgical testing has indicated good heap leaching characteristics.

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25.3 Metallurgy

Mother Lode metallurgical test work has been performed on samples from core and RC drill cuttings generated during the 2017-2018 exploration programs at Mother Lode and on a “grab” sample from the Mother Lode pit in 2017. The metallurgical test work addresses both the oxidized portion of the mineralization, suitable for heap leach processing and sulphide portion, which is amenable to cyanidation after concentration by flotation and pressure oxidation of sulphides.

Mother Lode ROM oxide gold heap leach dissolution relies on historical production data for ROM material from the Daisy Mine and bottle roll testing of drilling samples for 2017-2018 at Mother Lode. Based upon the production data a gold recovery of 74% was assumed to be appropriate for a PEA.

The testing on higher grade vein and stockwork mineralization in the YellowJacket zone indicates it should be processed in a milling system using gravity concentration, intensive CN leaching of the gravity concentrate followed by CN leaching of the combined gravity tail and leached gravity concentrate. A conventional paste tailing management facility would be constructed at Mother Lode for the limited volume of milled material. The grades and recoveries of the NBP disseminated mineralized material are suitable for heap leach processing. NBP sulphide material is amenable to sulphide and gold concentration by flotation, oxidative pretreatment by Presssure Oxidation or AAO followed by cyanidation. Pressure oxidation tests have not been performed on NBP sulphide material. It is assumed gold recovery will remain the same or increase in with pressure oxidation

25.4 Mining

Conventional surface mining methods using surface drill and blast techniques with off highway haul truck and front-end loaders were assumed for this report and are standard amongst similar mines in Nevada. Equipment requirements were based on typical equipment performance in similar mines. Haul truck estimates were based on estimated cycle times for each year in each pit’s current mining phase. Detailed mine designs with haul roads incorporated will provide more detailed cycle time information for better understanding of hauling requirements.

Pit slope angles were assumed to be 50 degrees for all pits except for Mother Lode, which was assumed to be 65 degrees. These slope angles may be aggressive and further geotechnical studies are needed to understand the stability of the pit walls. Detailed pit designs with haul roads my also contribute to flatter overall wall slope angles depending on geometric configuration for each pit.

The mine production schedule was created using Gemcom’s Whittle® 4.5 software. Nested whittle pit shells were generated between $500 and $1,500 gold price and used as phases for scheduling each pit area. No crest/toe pit designs with haul roads were created for interim phases nor ultimate pits. This method of mine design and production scheduling is adequate for a preliminary technical report to provide the initial viability of an economic mine schedule. That nature of these nested pit shells can lead to phases that are not physically possible. Detailed mine designs incorporating haul roads for each phase of each pit will provide more attainable mine production schedule.

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26 Recommendations

The PEA results, for the integrated NB-MLP Project with a mill facility at MLP and heap leach operations at both NBP and MLP, indicate the substantial financial impact on project performance of the higher grade sulphide and vein/vein stockwork types of mineralization that occurs at each site. Therefore, it is recommended that future exploration should focus on the identification and development of other deposits and sources of the higher-grade mineralization. In addition, RDA recommends that Corvus define and continue to execute programs that enhance and develop the supporting technical information, particularly as related to metallurgical processing of sulphide resources at both locations. Specifically, promising approaches to heap leach processing of sulphide mineralization have recently been tested at different locations and have the potential for substantial and favorable impacts on the conceptual project economic performance. These data would be used to refine the preliminary economic assessment of the combination of North Bullfrog with Mother Lode, at the time of the next resource update. These recommended activities are:

· Exploration drilling at identified targets at both Mother Lode and North Bullfrog;

· Resource definition drilling at Mother Lode;

· Metallurgical testing of gold bearing mineralization, particularly novel sulphide heap leaching technology;

· Continued environmental baseline characterization;

The projected costs for the next phase of this program are outlined in Table 26-1.

Table 26-1 Proposed Work Program to Advance NB-MLP

Activity	Amount
Exploration Drilling Resource Definition Drilling and Data Management	US$ 0.80 M
Baseline Data Collection	US$ 0.05 M
Metallurgical Testing	US$ 0.35 M
Resource Model, Geologic Model and PEA Update	US$ 0.15M
Total	US$ 1.35M

RDA has not recommended successive phases of work for the advancement of the Project.

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