DARNLEY BAY RESOURCES LTD. PINE POINT PEA

NI 43-101 PRELIMINARY ECONOMIC ASSESSMENT TECHNICAL REPORT ON THE PINE POINT ZINC PROJECT, NORTHWEST TERRITORIES, CANADA

Prepared for: Qualified Persons Company

Garett Macdonald, P.Eng. JDS Energy & Mining Inc.

Kelly McLeod, P.Eng. JDS Energy & Mining Inc. DARNLEY BAY RESOURCES LTD. Dino Pilotto, P.Eng. JDS Energy & Mining Inc. Suite 400, 365 Bay Street Ken Embree, P.Eng. Knight Piésold Ltd. Toronto, ON, M5H 2V1 Albert Daniel Siega, P.Eng Independent Consultant Paul Gann, P.Geo Independent Consultant Prepared by: JDS ENERGY & MINING INC. Suite 900, 999 W Hastings Street Vancouver, BC V6C 2W2

EffectiveEFFECTIVE Date D: AprilATE :18, A 2017PRIL 18, 2017 i REPORT DATE: JUNE 1, 2017

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

NOTICE JDS Energy & Mining, Inc. prepared this National Instrument 43-101 Technical Report, in accordance with Form 43-101F1, for Darnley Bay Resources Ltd. The quality of information, conclusions and estimates contained herein is based on: (i) information available at the time of preparation; (ii) data supplied by outside sources, and (iii) the assumptions, conditions, and qualifications set forth in this report. Darnley Bay Resources Ltd. filed this Technical Report with the Canadian Securities Regulatory Authorities pursuant to provincial securities legislation. Except for the purposes legislated under provincial securities law, any other use of this report by any third party is at that party’s sole risk.

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Contents Contents ...... iii Tables and Figures ...... ix 1 Executive Summary ...... 1-1 1.1 Introduction ...... 1-1 1.2 Project Description and Location ...... 1-1 1.3 Access and Ownership ...... 1-1 1.4 History, Exploration and Drilling ...... 1-2 1.5 Geology & Mineralization ...... 1-3 1.6 Metallurgical Testing and Mineral Processing ...... 1-4 1.7 Mineral Resource Estimates ...... 1-5 1.8 Mineral Reserve Estimate ...... 1-6 1.9 Mining ...... 1-6 1.10 Recovery Methods ...... 1-10 1.11 Infrastructure ...... 1-10 1.12 Environment and Permitting ...... 1-11 1.13 Capital Cost Estimate ...... 1-12 1.14 Operating Cost Estimate ...... 1-13 1.15 Economic Analysis ...... 1-14 1.15.1 Main Assumptions ...... 1-14 1.15.2 Results ...... 1-16 1.15.3 Sensitivities ...... 1-16 1.16 Conclusions ...... 1-17 1.16.1 Risks ...... 1-17 1.16.2 Opportunities ...... 1-18 1.17 Recommendations ...... 1-18 2 Introduction ...... 2-1 2.1 Basis of Technical Report ...... 2-1 2.2 Scope of Work ...... 2-1 2.3 Qualifications and Responsibilities ...... 2-2 2.4 Site Visit ...... 2-2 2.5 Units, Currency and Rounding ...... 2-3 2.6 Sources of Information ...... 2-3 3 Reliance on Other Experts ...... 3-1 4 Property Description and Location ...... 4-1 4.1 Area of the Property ...... 4-1 4.2 Location of the Property ...... 4-1 4.3 Type and Details of Mineral Tenure ...... 4-2 4.4 Issuer’s Title To and Interest in the Property ...... 4-2 4.4.1 Surface Rights ...... 4-2 4.4.2 Obligations for Retention of Property ...... 4-3 4.4.3 Expiration Date of Claims, License, or Other Property Tenure Rights ...... 4-3

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4.5 Agreements and Encumbrances ...... 4-3 4.5.1 Environmental Liabilities ...... 4-3 4.6 Permitting ...... 4-3 4.6.1 Land and Environmental Permits ...... 4-3 4.7 Other Factors and Risks ...... 4-4 5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ...... 5-1 5.1 Topography, Elevation, and Vegetation ...... 5-1 5.1.1 Physiography...... 5-1 5.1.2 Vegetation and Wildlife ...... 5-1 5.2 Means of Access to the Property ...... 5-1 5.3 Proximity to Population Centers ...... 5-2 5.3.1 Climate ...... 5-2 5.4 Infrastructure ...... 5-2 5.4.1 Exploration and Operating Season ...... 5-3 6 History ...... 6-1 6.1 Early History ...... 6-1 6.2 Timeline ...... 6-2 6.3 Prior Ownership...... 6-5 6.4 Previous Exploration ...... 6-5 6.5 Previous Production ...... 6-5 6.5.1 List of Historical Remaining Resources ...... 6-7 6.6 Production History ...... 6-14 7 Geological Setting and Mineralization ...... 7-1 7.1 Regional Geology ...... 7-1 7.2 Local Geology ...... 7-3 7.3 Property Geology ...... 7-4 7.4 Mineralization ...... 7-4 7.4.1 Distribution of Resources ...... 7-7 7.4.2 Controls on Mineralization ...... 7-8 7.4.3 Mineralogy and Paragenesis...... 7-9 7.4.4 Deposit Morphology and Size ...... 7-12 8 Deposit Types ...... 8-1 8.1 Mississippi Valley-Type Deposits ...... 8-1 8.2 Geological Model(s) ...... 8-1 8.3 Deposit Nomenclature at Pine Point ...... 8-4 8.4 W-85 (North Trend Cluster Pit) Open Pit Deposit ...... 8-4 8.5 X-65 (North Trend Cluster Pit) Open Pit Deposit ...... 8-5 8.6 The (Main Trend) Cluster Pits ...... 8-7 9 Exploration ...... 9-1 9.1 Exploration History ...... 9-1 9.1.1 Drilling ...... 9-1 9.1.2 Geophysics ...... 9-1 9.1.3 Geochemistry ...... 9-2

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9.2 Historical Exploration Methods and Drill Core ...... 9-2 9.2.1 Past Exploration ...... 9-2 10 Drilling ...... 10-3 10.1 N-204 Drilling Summary ...... 10-3 10.1.1 Tamerlane Drilling ...... 10-3 10.1.2 Pine Point Mines Ltd. (Cominco)...... 10-11 10.1.3 Westmin Resources ...... 10-11 10.1.4 Sampling Methodology & QA/QC ...... 10-11 10.1.5 QA/QC Check Samples ...... 10-12 10.1.6 Data Verification ...... 10-12 10.1.7 QP’s Opinion ...... 10-15 10.2 Cluster Pits ...... 10-15 10.2.1 Tamerlane Drilling ...... 10-16 10.2.2 Tamerlane’s Core Logging Procedures ...... 10-18 10.2.3 Pine Point Mines Ltd. (Cominco)...... 10-19 10.2.4 Geology of Cluster Pits ...... 10-19 10.2.5 Darnley Bay Resources 2017 Drilling Update ...... 10-24 10.3 Sampling Method and Approach ...... 10-26 10.3.1 Density Measurements ...... 10-26 10.3.2 Pine Point Mines Ltd. (Cominco) Practice ...... 10-26 10.3.3 Westmin Practice ...... 10-27 10.3.4 Tamerlane Practice ...... 10-27 11 Sample Preparation, Analyses, and Security ...... 11-1 11.1 Cominco Samples ...... 11-1 11.2 Westmin Samples ...... 11-1 11.3 Tamerlane Samples, 2005 ...... 11-1 11.4 QP’s Opinion ...... 11-2 12 Data Verification ...... 12-1 12.1 Cominco Drilling ...... 12-1 12.2 Westmin Drilling ...... 12-3 12.3 Data Verification Procedures ...... 12-4 12.4 QP’s Opinion ...... 12-5 13 Mineral Processing and Metallurgical Testing ...... 13-1 13.1 Metallurgical Test Work for R-190 and O-556 / Z-155 ...... 13-2 13.1.1 Heavy Liquid Separation Tests ...... 13-2 13.1.2 Flotation Test Work Development ...... 13-4 13.1.3 Large Scale Heavy Liquid Test Work ...... 13-5 13.1.4 Flotation Process Verification...... 13-5 13.1.5 Comminution and Liberation Results ...... 13-8 13.2 Metallurgical Test Work for N-204...... 13-9 13.2.1 Sample Composition ...... 13-9 13.2.2 DMS Test Work Results ...... 13-9 13.2.3 Flotation Test Work Results ...... 13-10 13.2.4 DMS Discussion ...... 13-11

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13.3 Process Selection ...... 13-12 13.4 Recovery Projections ...... 13-13 14 Mineral Resource Estimates ...... 14-1 14.1 National Instrument 43-101 Qualified Resources ...... 14-1 14.1.1 R-190, X25, G-O3, P-499, O-556, Z-155 ...... 14-1 14.1.2 N-204 ...... 14-2 14.2 Methods, Cluster Pit Deposits ...... 14-4 14.2.1 Density ...... 14-7 14.2.2 Domaining ...... 14-8 14.2.3 Compositing ...... 14-8 14.2.4 Capping ...... 14-8 14.2.5 Statistical Analysis for the CP Deposits ...... 14-8 14.2.6 Variography ...... 14-13 14.2.7 Tabular Deposits ...... 14-13 14.2.8 Prismatic type Deposits ...... 14-14 15 Mineral Reserve Estimates ...... 15-1 16 Mining Methods ...... 16-1 16.1 Open Pit Mine Design ...... 16-1 16.2 Optimization ...... 16-1 16.2.1 Input Parameters ...... 16-1 16.2.2 Optimization Results ...... 16-3 16.2.3 Pit Stages ...... 16-11 16.3 Mine Production Schedule ...... 16-12 16.4 Mine Equipment ...... 16-18 16.5 Mine Personnel ...... 16-20 17 Process Description/Recovery Methods ...... 17-1 17.1 Introduction ...... 17-1 17.2 Plant Design Criteria ...... 17-2 17.3 Process Plant Description ...... 17-5 17.3.1 Crushing and DMS ...... 17-5 17.3.2 Grinding ...... 17-6 17.3.3 Flotation ...... 17-6 17.3.4 Reagents ...... 17-7 17.3.5 Plant Air ...... 17-8 17.3.6 Assay Laboratory ...... 17-8 17.4 Process Personnel ...... 17-9 18 Project Infrastructure and Services ...... 18-1 18.1 Summary ...... 18-1 18.2 General Site Arrangement ...... 18-1 18.3 Site Access ...... 18-1 18.4 Power ...... 18-3 18.5 Ancillary Facilities ...... 18-3 18.5.1 Truck Shop & Warehouse Building ...... 18-3 18.5.2 Mine Dry and Office Complex ...... 18-3

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18.5.3 Camp ...... 18-4 18.5.4 Fuel Storage ...... 18-5 18.5.5 Explosives Storage ...... 18-5 18.6 Mine Water Management ...... 18-5 18.6.1 Hydrogeological Summary ...... 18-5 18.6.2 Water Management Plan ...... 18-12 18.7 Waste Rock Management ...... 18-14 18.8 Tailings Management Facility ...... 18-16 19 Market Studies and Contracts ...... 19-1 19.1 Market Studies ...... 19-1 19.2 Contracts ...... 19-1 19.3 Royalties ...... 19-2 19.4 Metal Prices ...... 19-2 20 Environmental Studies, Permitting and Social or Community Impact ... 20-1 20.1 Environment ...... 20-1 20.2 Baseline Studies ...... 20-1 20.3 Closure ...... 20-2 20.4 Permitting ...... 20-2 20.5 Social Engagement ...... 20-2 21 Capital Cost Estimate ...... 21-1 21.1 Capital Cost Summary ...... 21-1 21.2 Capital Cost Profile ...... 21-2 21.3 Key Estimate Assumptions ...... 21-2 21.4 Key Estimate Parameters ...... 21-3 21.5 Basis of Estimate ...... 21-3 21.5.1 Labour Rates ...... 21-3 21.5.2 Fuel & Energy Supply ...... 21-3 21.5.3 Mine Capital Cost Estimate ...... 21-3 21.5.4 Site Development & Earthworks ...... 21-4 21.5.5 Processing Cost Estimate ...... 21-5 21.5.6 On-Site Infrastructure ...... 21-5 21.6 Indirect Cost Estimate ...... 21-6 21.7 Owners Cost Estimate ...... 21-7 21.8 Closure Cost Estimate ...... 21-7 21.9 Contingency ...... 21-8 21.10 Capital Estimate Exclusions ...... 21-8 22 Operating Cost Estimate ...... 22-1 22.1 Operations Labour ...... 22-2 22.2 Basis of Estimate ...... 22-4 22.2.1 Mine Operating Cost Estimate ...... 22-4 22.2.2 Processing Operating Cost Estimate ...... 22-5 22.2.3 General and Administration Operating Cost Estimate ...... 22-5 23 Economic Analysis ...... 23-7 23.1 Summary of Results ...... 23-7

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23.2 Key Assumptions ...... 23-7 23.3 Mine Production ...... 23-8 23.4 Processing & Concentrate Terms ...... 23-9 23.5 Taxes ...... 23-11 23.6 Royalties ...... 23-11 23.7 Results ...... 23-11 23.8 Sensitivities ...... 23-13 24 Adjacent Properties ...... 24-1 25 Other Relevant Data and Information ...... 25-1 26 Interpretations and Conclusions ...... 26-1 26.1 Risks ...... 26-1 26.2 Opportunities ...... 26-3 26.2.1 Addition of Underground Resources ...... 26-3 26.2.2 Other Potential Opportunities ...... 26-3 27 Recommendations ...... 27-1 28 References ...... 28-1 29 Units of Measure, Abbreviations and Acronyms ...... 29-5

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Tables and Figures

Table 1.1: Recovery Projections for Each Deposit ...... 1-5 Table 1.2: Mineral Resource Estimate ...... 1-6 Table 1.3: Optimization Parameters ...... 1-7 Table 1.4: PEA Proposed Mine Plan Summary ...... 1-8 Table 1.5: Mine and Mill Production Schedule...... 1-9 Table 1.6: Capital Cost Summary ...... 1-12 Table 1.7: Summary of Operating Cost Estimate ...... 1-13 Table 1.8: Main OPEX Assumptions ...... 1-14 Table 1.9: Metal Prices and F/X Rate ...... 1-14 Table 1.10: Net Smelter Return Assumptions...... 1-15 Table 1.11: Economic Results* ...... 1-16

Table 1.12: Sensitivity Results (After-Tax NPV8%) ...... 1-17 Table 2.1: QP Responsibilities ...... 2-2 Table 2.2: QP Site Visits ...... 2-2 Table 4.1: Resources Lease Surface Area ...... 4-1 Table 6.1: Historical Production at Pine Point...... 6-6 Table 6.2: Historical Tamerlane 2014 Open Pit Resources ...... 6-8 Table 6.3: Summary of Unexploited Historical Resources ...... 6-9 Table 6.4: Historical Indicated Prismatic Deposits ...... 6-9 Table 6.5: Historical Indicated Open Pit Tabular Deposits ...... 6-10 Table 6.6: Pine Point Area Historical Indicated Underground Resources ...... 6-11 Table 6.7: Historical Inferred Open Pit Prismatic Resources ...... 6-12 Table 6.8: Pine Point Area Historical Inferred Underground Resources ...... 6-13 Table 6.9: Historical Inferred Open Pit Tabular Resources ...... 6-14 Table 7.1: Mineralized Middle Devonian Units at Pine Point ...... 7-5 Table 8.1: Summary of Drilling on the Cluster Deposits ...... 8-8 Table 10.1: Drilling Summary at N-204 ...... 10-4 Table 10.2: Confirmatory Drill Holes and Mineralized Intercepts ...... 10-7 Table 10.3: Comparison: The Tamerlane Confirmation Drill Holes vs. the Historic Cominco Drill Holes ...... 10-9 Table 10.4: Intersections of Significance Drilled by Tamerlane between 2007 and 2008 ...... 10-18 Table 10.5: Tamerlane Drilling 2007 - 2008 ...... 10-22 Table 10.6: 2017 Ongoing Drilling Results ...... 10-25 Table 13.1: Material Samples Used in the Pine Point Metallurgical Test Program ...... 13-2 Table 13.2: Calculated Head Grades for the R-190 DMS Composite ...... 13-3 Table 13.3: DMS Results for R-190 ...... 13-3 Table 13.4: Locked Cycle Flotation Test Results for R-190 ...... 13-4 Table 13.5: Head Grades for O-556 and Z-155 ...... 13-5 Table 13.6: DMS Test Results for O-556 / Z-155 ...... 13-5 Table 13.7: Locked Cycle Flotation Test Results for O-556 / Z-155 ...... 13-6 Table 13.8: Finalized Flotation Reagent Scheme Pb Circuit ...... 13-7 Table 13.9: Finalized Flotation Reagent Scheme Zn Circuit ...... 13-7 Table 13.10: Final Concentrate Impurity Analysis ...... 13-8

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Table 13.11: N-204 Composite Head Assays ...... 13-9 Table 13.12: DMS Test Results for N-204 ...... 13-10 Table 13.13: Locked Cycle Flotation Test Results for N-204 ...... 13-11 Table 13.14: Recovery Projections for Each Deposit ...... 13-15 Table 14.1: Mineral Resource Estimates, Underground Deposits: R-190, X-25, G-03, O-556 & Z-155 . 14-2 Table 14.2: Criteria - N204 ...... 14-3 Table 14.3: Resource Estimates, N-204 ...... 14-4 Table 14.4: Criteria - Cluster Deposits ...... 14-5 Table 14.5: Resource Estimates, Cluster Pits ...... 14-7 Table 14.6: Descriptive Statistics for HZ ...... 14-9 Table 14.7: Descriptive Statistics for J-68 ...... 14-9 Table 14.8: Descriptive Statistics for K-68 ...... 14-10 Table 14.9: Descriptive Statistics for X-65 ...... 14-10 Table 14.10: Descriptive Statistics for W-85 ...... 14-11 Table 14.11: Descriptive Statistics for M-67 ...... 14-11 Table 14.12: Descriptive Statistics for M-62/63 ...... 14-12 Table 14.13: Descriptive Statistics for O-53...... 14-12 Table 14.14: Descriptive Statistics for L-65 ...... 14-13 Table 16.1: Pit Shell Summary ...... 16-1 Table 16.2: Input Parameters OP Optimization ...... 16-2 Table 16.3: Pit Shell Summary ...... 16-4 Table 16.4: Pit Stage Summary ...... 16-11 Table 16.5: LOM Production Schedule ...... 16-14 Table 16.6: Mobile Mine Equipment Fleet (average number of units) ...... 16-19 Table 16.7: Total Mine Personnel Requirements ...... 16-21 Table 17.1: Process Design Criteria ...... 17-3 Table 17.2: Reagents and Consumptions ...... 17-8 Table 17.3: Total Process Personnel ...... 17-9 Table 18.1: Water Management Summary ...... 18-13 Table 18.2: Waste Rock and DMS Reject Storage Summary ...... 18-14 Table 18.3: Capacity of Waste Rock Storage Facilities ...... 18-16 Table 18.4: Pit Backfill Tailings Management Facilities ...... 18-17 Table 19.1: Concentrate Terms ...... 19-1 Table 19.2: Metal Price and Exchange Rate ...... 19-4 Table 21.1: Capital Cost Summary ...... 21-1 Table 21.2: Mining Capital Costs ...... 21-4 Table 21.3: Process Plant Basis of Estimate ...... 21-5 Table 21.4: Infrastructure Basis of Estimate ...... 21-6 Table 21.5: Indirect Cost Basis of Estimate ...... 21-7 Table 22.1: Breakdown of Estimated Operating Costs ...... 22-1 Table 22.2: Main OPEX Assumptions ...... 22-2 Table 22.3: Annual Manpower Numbers by Discipline ...... 22-3 Table 22.4: Mine Operating Costs by Category ...... 22-4 Table 22.5: Processing Operating Cost by Category ...... 22-5 Table 22.6: Summary of G&A Costs ...... 22-6 Table 23.1: Metal Price & Exchange Rates used in Economic Analysis ...... 23-8

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Table 23.2: LOM Plan Summary ...... 23-8 Table 23.3: Concentrate Terms ...... 23-9 Table 23.4: Summary of Results ...... 23-13 Table 23.5: After-Tax Sensitivity Results on NPV @ 8% Discount Rate ...... 23-14 Table 23.6: Discount Rate Sensitivity ...... 23-15 Table 26.1: Main Project Risks ...... 26-2 Table 27.1: Feasibility Study Cost Estimate ...... 27-1

Figure 1.1: Darnley Bay’s Leases and Claims ...... 1-2 Figure 1.2: Local Geology ...... 1-3 Figure 4.1: Darnley Bay’s Leases and Claims ...... 4-2 Figure 5.1: Topographic Map of Pine Point Area ...... 5-4 Figure 6.1: Annual Ore Production at Pine Point (Mt) 1964 - 1987 ...... 6-15 Figure 7.1: Geological Provinces of Canada ...... 7-1 Figure 7.2: Middle Devonian Sedimentary Facies in West-Central Canada ...... 7-2 Figure 7.3: Local Geology ...... 7-3 Figure 7.4: Middle Devonian Sedimentary Facies ...... 7-5 Figure 7.5: Sub-crop Geologic Map of Pine Point District ...... 7-6 Figure 7.6: Cross Section through Southwest Area of Pine Point Property ...... 7-7 Figure 7.7: Mineralized Trends in the Pine Point District ...... 7-8 Figure 7.8: Mineral Paragenesis - Pine Point Lead-Zinc District ...... 7-10 Figure 7.9: Textures of Mineralized Rock, Pine Point District ...... 7-12 Figure 8.1: Schematic of Pine Point MVT Deposits ...... 8-2 Figure 8.2: Formation of Karst Topography ...... 8-3 Figure 8.3: Location Map of W-85 and X-65 ...... 8-5 Figure 8.4: North Trend Section of X-65 Deposit ...... 8-6 Figure 8.5: North Trend Geological Plan Including X-65 Deposit ...... 8-7 Figure 8.6: Cluster Pit Area Showing Historic Mining, Proposed Mining and Existing Haul Roads ...... 8-9 Figure 8.7: Cross Section of the Main Trend Showing the Relationship of Mineralization to Geology .. 8-10 Figure 10.1: Location of the Confirmatory Drilling Program at N-204 ...... 10-5 Figure 10.2: Tamerlane DDH Locations Overlying the Historic Cominco Drill Sites ...... 10-6 Figure 10.3: N-204 Deposit - Main Zone Percent Variance in Original vs. Check Pulp Samples (N=16) ...... 10-13 Figure 10.4: N-204 Deposit - Upper Zone Percent Variance in Original vs. Check Pulp Samples (N=39) ...... 10-14 Figure 10.5: Cluster Pit Area Showing Historic and Proposed Mining with Existing Haul Roads ...... 10-15 Figure 10.6: Location Map of W-85 and X-65 ...... 10-16 Figure 10.7: Deposit Location for K-68, M-67 and L-65 ...... 10-19 Figure 10.8: Location and Historical Resource* of K-68, M-67 and L-65 ...... 10-20 Figure 10.9: Plan View of K-68 and M-67 Deposits and Associated Drill Holes ...... 10-23 Figure 10.10: Plan View of K-68 and M-67 Deposits with Aerial Photograph ...... 10-23 Figure 10.11: K-68 West-East Longitudinal Section ...... 10-24 Figure 10.12: M-67 West-East Long Section (1% Pb+Zn Grade Shell) ...... 10-24 Figure 12.1: Selected Cominco Lithologic Unit Codes ...... 12-2 Figure 12.2: Selected Westmin Lithologic Unit Codes ...... 12-4 Figure 13.1: Locked Cycle Flotation Flowsheet for N-204 ...... 13-10

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Figure 13.2: DMS Feed Rates vs. Feed Grades...... 13-12 Figure 13.3: Lead and Zinc Grade Recovery Curves ...... 13-14 Figure 16.1: Overview of Deposit Locations ...... 16-5 Figure 16.2: J68 Pit Shell ...... 16-6 Figure 16.3: Hinge Zone Pit Shell ...... 16-6 Figure 16.4: W85 Pit Shell ...... 16-7 Figure 16.5: X65 Pit Shell ...... 16-7 Figure 16.6: N-204 Pit Shell ...... 16-8 Figure 16.7: M67 Pit Shell ...... 16-8 Figure 16.8: O53 Pit Shell ...... 16-9 Figure 16.9: K68 Pit Shell ...... 16-9 Figure 16.10: L65 Pit Shell ...... 16-10 Figure 16.11: M62/63 Pit Shell ...... 16-10 Figure 16.12: Pit Stages ...... 16-12 Figure 16.13: LOM Schedule ...... 16-17 Figure 16.14: Mineralized Material by Pit Phase ...... 16-18 Figure 17.1: Process Flowsheet ...... 17-4 Figure 17.2: DMS Flowsheet ...... 17-6 Figure 18.1: Overall Site Lay-out ...... 18-2 Figure 18.2: Regional Hydrogeology ...... 18-7 Figure 18.3: Geologic and Hydrogeologic Stratigraphy of the Presqu’ile Barrier Reef Complex at the A70 Pit Area ...... 18-9 Figure 18.4: Pine Point Area – Groundwater Contour Map ...... 18-11 Figure 18.5: Pit Backfill Tailings Management Facilities ...... 18-17 Figure 19.1: Historical Lead Price ...... 19-2 Figure 19.2: Historical Zinc Price ...... 19-3 Figure 19.3: Historical US$:C$ F/X Rates ...... 19-3 Figure 21.1: Capital Cost Profile (Excluding Closure Years) ...... 21-2 Figure 22.1: Operating Cost Distribution ...... 22-2 Figure 23.1: Payable Lead Production by Year ...... 23-10 Figure 23.2: Payable Zinc Production by Year ...... 23-10 Figure 23.3: Revenue Distribution ...... 23-11 Figure 23.4: Annual After-Tax Cash Flow ...... 23-12 Figure 23.5: Sensitivity, Post-Tax NPV @ 8% Discount Rate ...... 23-15 Figure 23.6: Post-Tax NPV Sensitivity to Discount Rate ...... 23-16 Figure 23.7: Economic Cash Flow Model ...... 23-1

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1 Executive Summary

1.1 Introduction Darnley Bay Resources Ltd. (Darnley Bay) commissioned JDS Energy & Mining Inc. (JDS) to complete a Preliminary Economic Assessment (PEA) for the Pine Point Project. The purpose of this study is to complete a review and compilation of the resources, mining designs and preliminary economics.

This report was prepared in order to fulfill the reporting requirements stipulated by Canadian National Instrument 43-101, Standards of Disclosure for Mineral Projects (NI 43-101).

It must be noted that this PEA is preliminary in nature and includes the use of Inferred Mineral Resources for which the geology is considered too speculative to be categorized as Mineral Reserves. Therefore, there is no certainty that the PEA will be realized.

1.2 Project Description and Location Darnley Bay’s Pine Point properties, including the 10 subject deposits described in this report, are located about 800 km north of Edmonton, near the south shore of Great Slave Lake, in the Mackenzie Mining Division of the Northwest Territories (NWT) of Canada.

The Darnley Bay leases range between 42 and 110 km from Hay River, NWT. The lease areas are situated about 10 km south of the Great Slave Lake and lie about 60 m above the lake level, which is at an elevation of 156 m above sea level (masl). Geographic coordinates are from 114° to 115° 15’ West longitude and from 61° 0’ to 61° 45’ North latitude.

1.3 Access and Ownership The mineral leases lie north of the Territorial Highways 5 and 6 which connect Hay River and the Pine Point town site, and extend intermittently from 25 to 80 km east of the town of Hay River. An all- weather year-round highway parallels the southern boundary of the Darnley Bay lease block (Figure 1.1).

The mineral deposits at Pine Point are covered by 40 leases in total. The leases are held by Darnley Bay as a result of transfer of title on December 21, 2016 from Pine Point Holding Corp (PPHC) to Darnley Bay Resources Limited. As of the effective date of this report, the leases are in good standing.

Darnley Bay has a 100% interest in the mineral leases discussed in this report, subject to a 3% Net Smelter Return (NSR) payable to Karst Investments LLC (Karst). The project is not subject to any other royalties, back-in rights, payments, or other agreements or encumbrances other than the territorial royalty (calculated as a tax but called a royalty).

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Figure 1.1: Darnley Bay’s Leases and Claims

Source: Northwest Territories Mining Recorder’s Office Database (2017)

1.4 History, Exploration and Drilling Geologic exploration of lead-zinc showings south of Great Slave Lake dates from 1898. Prospectors heading for the Klondike gold rush staked claims on outcrops of oxidized sulphide. Intermittent exploration continued until 1948 when Cominco Ltd. began major exploration programs that continued until 1955.

Drilling in the district by Westmin Resources Ltd. (Westmin), Pine Point Mines Ltd (Cominco), and Tamerlane Ventures Inc. (Tamerlane) totalled 1,308,742 m (approximately 4,293,751 ft) in an estimated 18,406 core holes. There was some additional exploration drilling by junior companies in the 1970s, mainly on the eastern fringes of the district, and by Pine Point Mines Ltd (Cominco) west of the district in the Hay River area.

Prior to 1963, exploration at Pine Point was accomplished by grid diamond drilling, which was used to define a historical resource estimate. Stratigraphic and structural models were developed for prioritizing the drill targets.

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Pine Point Mines Ltd (Cominco) commenced production on several identified deposits in 1964. Large-scale mining by Pine Point Mines Ltd (Cominco) commenced in 1964 with reported reserves of 21.5 Mt averaging 4% lead and 7.2% zinc (a historical figure, not in accordance with NI 43-101, reported by Giroux and McCartney, 2004). The Cominco "Pine Point Mine" was actually an assemblage of 46 separate open pits and two underground deposits, lying along a 35 km trend, which also included a mill at the Pine Point town site. No mining took place on Westmin's Great Slave Reef area, located west of Cominco's Pine Point property, which extended to about 5 km west of the Buffalo River.

1.5 Geology & Mineralization The Middle Devonian rocks of interest in the Pine Point district are overlain by non-mineralized Middle and Upper Devonian rocks. These in turn are overlain by Pleistocene glacial deposits and post-glacial swamp deposits, which cover nearly all the surface of the district. Outcrops are rare.

Figure 1.2 shows the facies relationships, centering on the Presqu'ile Barrier Reef in the middle, with an open marine environment to the northwest, and the Elk Point lagoon or sea of restricted circulation to the southeast. A minor uplift after Devonian time exposed the main reef to karsting, and ended the barrier-reef regime.

Figure 1.2: Local Geology

Source: Modified from Rhodes et. al. 1984

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The Middle Devonian sedimentary rocks which host the lead-zinc deposits in the district are little- deformed. There is also extensive karstification (the dissolution of the soluble carbonate rocks), which occurred during a period of uplift of the sedimentary rocks. Folding is minor and generally associated with differential compaction of the original sediments, or karst subsidence features. Southwesterly trending lineaments in the underlying Precambrian basement have a 0.9 m/km slope to the west beneath the sedimentary rocks of the district, but these do not have visible expressions in the Devonian rocks.

Specific formations within the Middle Devonian sedimentary sequence host the mineralization at Pine Point. The most controversial and confusing facies issue was the status of the "Presqu'ile Formation". This coarsely-crystalline dolomite unit was formerly assigned a great stratigraphic and facies range, but was eventually recognized to be a diagenetic alteration of several different carbonate depositional facies. Thus, it has been discontinued as a stratigraphic unit, but is still referred to as a rock type (e.g. in drill logs).

Mineralization occurs in carbonate (limestone and dolomite) host rocks and is typically associated with the ingress of mineralizing brines into structurally prepared cavities associated with dissolution of the carbonate rocks by karst processes. In the case of Pine Point, this dissolution and mineralization process occurred on an extensive scale and was controlled by the distribution of a major carbonate barrier reef complex of Devonian age. A very detailed study of the geology for the Cominco Pine Point Property was compiled in a 1984 paper in Economic Geology Vol. 79 (Rhodes et al., 1984). The Rhodes paper built on early pioneering work by H. Skall of Cominco who first described the Pine Point Reef Complex in modern geological terms (Skall, 1975).

Pine Point lead-zinc mineralization was derived from a host some distance from the present position of the deposits. Circulating epithermal fluids from a possible magmatic source may have been responsible for the mineralization which occurs in the carbonate sediments as open-space cavity fillings and local replacements.

Shale within the Lower Keg River Member is a useful marker horizon, as it is continuous across the district. The “E” facies, consisting of un-mineralized, brown, sucrosic dolomite, underlies nearly all mineralization, and thus most drill holes terminate in this rock unit.

1.6 Metallurgical Testing and Mineral Processing Metallurgical test programs were conducted on samples from across the Pine Point District including R-190, O-556, Z-155 and N-204 deposits. Heavy media separation and flotation test work indicated that standard zinc and lead flotation preceded by dense media separation will likely yield excellent recoveries.

The recovery projections for each deposit are based on the metallurgical results as summarized in Table 1.1. Test work program 2007 SGS Project 11633-001 and historical operating data was used to predict recoveries and grades for the cluster pits (CP). These are preliminary estimates and more test work might be completed in the next stage of engineering to confirm these values.

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Table 1.1: Recovery Projections for Each Deposit

Head Grade Recovery Stage Pb (%) Zn (%) Pb (%) Zn (%) Cluster Pits DMS Stage Recovery 6.91 13.95 94.4 93.5 Flotation Stage Recovery 70.0 61.9 90.0 94.2 Overall Recovery 70.0 61.9 85.0 88.1 LOM Head 2.31 4.71 100 100 N-204 DMS Stage Recovery 2.47 9.32 87.1 88.4 Flotation Stage Recovery 55.7 55.3 82.0 88.2 Overall Recovery 55.7 55.3 71.4 78 Calculated Head 0.7 2.6 100 100 Source: SGS (2007), G&T (2012), Historical Operating Data

1.7 Mineral Resource Estimates The Pine Point Barrier Reef, which hosts the Pine Point deposits, plunges gently south-west at a rate of 0.9 m/km. Deposits on the east side of the Buffalo River are considered to be shallow enough to be mined by open pit methods and 48 deposits were historically mined by open pit in this area; a large number of potential open pit deposits remain. The main open pit deposits that Darnley Bay is focused on at this time are the N-204 located on the South Trend, the W-85 and X-65 located on the North Trend and the Main Trend deposits which include J-68, K-68, M-62/63, M-67, O-53, L- 65 and HZ (Hinge Zone). In addition to these there are a number of other deposits that currently remain in the historical resources category. Confirmatory drilling was conducted at deposit N-204; the Cluster Pit deposits were modeled using Pine Point Mines historic drilling intersections and additional drilling completed by Tamerlane at the HZ, M-67, W-85 and L-65 deposits.

Table 1.2 documents the Measured and Indicated Mineral Resources used in this PEA.

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Table 1.2: Mineral Resource Estimate

Measured & Indicated Inferred Pit Feed Pb Zn Pb Zn Feed Pb Zn Pb Zn (kt) (%) (%) (M Lb) (M Lb) (kt) (%) (%) (M Lb) (M Lb) J68 304 2.25 4.90 15 33 - - - - - HZ 1,549 2.76 3.58 94 122 - - - - - W85 3,983 1.98 3.46 174 304 - - - - - X65 5,430 0.80 2.41 95 288 56 0.00 3.67 0 5 O533 81 0.91 2.90 2 5 121 0.83 2.79 2 7 N-204 10,005 0.80 2.98 177 657 3,527 0.78 2.89 61 224 M67 1,365 0.84 3.55 25 107 - - - - - K68 1,193 0.81 2.77 21 73 - - - - - L65 1,578 0.70 1.95 25 68 - - - - - M62/63 352 1.42 2.13 11 16 - - - - - Total 25,841 1.12 2.94 640 1674 3,703 0.77 2.90 63 236 Notes: The effective date for the open pit mineral resource is April 18, 2017. Mineral Resources which are not mineral reserves do not have demonstrated economic viability. The estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, taxation, sociopolitical, marketing, or other relevant issues. The quantity and grade of reported Inferred Resources in this estimation are uncertain in nature and there has been insufficient exploration to define these Inferred Resources as an Indicated or Measured Mineral Resource and it is uncertain if further exploration will result in upgrading them to an Indicated or Measured Mineral Resource category. Source: Siega (2017)

1.8 Mineral Reserve Estimate This PEA does not state a Mineral Reserve.

1.9 Mining This Plan covers mining 10 deposits at Pine Point via conventional open pit (OP) methods. Mining of the deposits will produce a total of 29.5 Mt of plant feed, along with 117.2 Mt of waste (4.0:1 overall strip ratio) over a 13-year mine life (excluding pre-production period). The mine design process for the deposits commenced with the development of OP optimization design parameters. These parameters included estimates of metal prices, mining dilution, process recovery (both dense media separation (DMS) and flotation), off-site costs, geotechnical constraints (slope angles), and royalties.

The block value is based on the NSR derived from using DMS plants followed by traditional grinding and flotation to upgrade mineralized material for the shipment of lead & zinc concentrates. Table 1.3 summarizes the NSR inputs and optimization parameters.

The current life of mine (LOM) plan focuses on achieving consistent plant feed production rates, and early mining of higher value material, as well as balancing grade and strip ratios.

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Table 1.3: Optimization Parameters

Parameter Unit Value Revenue, Smelting & Refining Lead Price US$/lb Pb 0.90 Zinc Price US$/lb Zn 1.05 Exchange Rate US$:C$ 0.75 Royalty (3%) % NSR 3.0 OPEX Estimates OP Waste Mining Cost C$/t waste mined 2.85 OP Mill Feed Mining Cost C$/t feed mined 2.85 Strip Ratio (estimated) Waste t:Mill feed t 4.0 OP Mining Cost C$/t processed 14.25 Processing Cost (DMS and Milling) (CP) C$/t processed 9.20 Processing Cost (DMS and Milling) (N-204) C$/t processed 7.43 Haulage from DMS to Mill C$/t processed 1.57 Additional Waste Rock and Tailings Reclamation C$/t processed 0.25 G&A C$/t processed 7.13 Total OPEX (excluding Mining) (CP) C$/t processed 18.15 Total OPEX (excluding Mining) (N-204) C$/t processed 16.38 Recovery and Dilution OP External Mining Dilution % 12 OP Mining Recovery % 95 Metal Recovery (DMS & Flotation) Lead (CP) % 85.0 Zinc (CP) % 88.1 Lead (N-204) % 71.4 Zinc (N-204) % 78.0 Other Overall Pit Slope Angles degrees 52 Mill Production Rate t/d 1,800 Smelter Terms and Offsite Costs Concentrate Grades %

Lead concentrate (CP) % 61.0 Lead concentrate (N-204) % 55.7 Lead concentrate moisture content % 8.00 Zinc concentrate (CP) % 62.0 Zinc concentrate (N-204) % 55.3 Zinc concentrate moisture content % 8.00 Payables Lead % 95.0 Zinc % 85.0 Unit deductions Lead % 3.0

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Parameter Unit Value Zinc % 8.0 Smelting & Refining Costs Lead concentrate smelter cost (TC) US$/dmt $135.00 Zinc concentrate smelter cost (TC) US$/dmt 135.00 Freight & Marketing Lead Concentrate US$/wmt 167 Zinc Concentrate US$/wmt 154 Note: *The values in this table vary from those used in the economic model as parameters were further refined in the economic model as the Project progressed. The differences are not considered material to pit shape definition. wmt = wet metric tonnes/ dmt = dry metric tonnes Source: JDS (2017)

Pit optimization software, utilizing the Lerch-Grossman algorithm, and the parameters listed above were used to determine the optimal mining shells of the various deposits with the assumed overall slope angles. Preliminary mining stages were selected and mine planning and scheduling were then conducted from the selected optimal shells. The PEA mineral resources for the project are presented in Table 1.4.

Table 1.4: PEA Proposed Mine Plan Summary

Description Unit Value Mine Production Life years 13 Total Process Feed Material Mt 29.5 Total Diluted Lead Grade % 1.08 Total Diluted Zinc Grade % 2.93 Total Contained Lead Mlbs 703 Total Contained Zinc Mlbs 1,910 Total Waste Mt 117.2 Total Material Mt 146.7 Overall Strip Ratio Waste t:Mill feed t 4.0 Note: PEA Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. Source: JDS (2017)

The mine production schedule is summarized in Table 1.5.

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Table 1.5: Mine and Mill Production Schedule

Year Parameter Unit Total Pre-prod 1 2 3 4 5 6 7 8 9 10 11 12 13 Mine Feed (to DMS) kt 29,544 304 1,632 2,156 2,481 2,307 2,507 2,456 2,396 2,512 2,363 2,324 2,267 2,462 1,378

Lead Grade % 1.08 2.25 2.84 2.21 1.32 0.75 0.86 0.64 0.77 0.91 0.80 0.83 0.71 0.81 1.07

Zinc Grade % 2.93 4.90 3.49 3.30 3.27 2.31 3.53 2.56 2.43 3.13 2.59 3.30 2.67 2.41 3.13

Total Waste kt 117,178 385 9,142 10,147 12,066 7,995 7,720 5,598 6,850 8,615 10,029 10,553 11,794 11,463 4,821

Total Mined kt 146,722 689 10,774 12,303 14,547 10,302 10,227 8,054 9,246 11,128 12,392 12,877 14,060 13,925 6,199

Strip Ratio w:mf 4.0 1.3 5.6 4.7 4.9 3.5 3.1 2.3 2.9 3.4 4.2 4.5 5.2 4.7 3.5

Mill Feed kt 8,252 - 657 657 657 657 657 657 657 657 657 657 657 657 368 Source: JDS (2017)

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1.10 Recovery Methods The process design criteria and flowsheets have been developed based on both the recent metallurgical test work results detailed in Section 13, and historical operating data. The test data indicates that the Pine Point mineralized material can be treated using conventional mineral processing techniques, applying conventional differential flotation techniques for the recovery of lead and zinc concentrates.

The two crushing plants and DMS plants will each process between 5,700 and 7,300 t/d of material from the cluster pits (CP) and N-204 pit, respectively. The facilities will include 3-stage crushing and DMS operations.

The concentrator plant will consist of a grinding mill, lead rougher flotation and one stage of lead cleaner flotation and zinc rougher flotation, and two stages of zinc cleaner flotation. Both concentrates will be dewatered in concentrate thickeners and pressure filters.

The LOM average metal recoveries are expected to be in the range of 71 to 85% for lead at a grade of 56 to 70% lead. Zinc recoveries are expected to be in the range 78 to 88% with concentrate grades of 55 to 62%.

The crushing circuits will operate on average 18 hours a day. The process plant will operate 24 hours per day, 365 days per year at an availability of 92%.

The processing facilities will consist of the following unit operations:

 Two plants - 3-stage crushing and screening;  Two DMS plants;  Primary grinding;  Lead flotation using conventional and column cells;  Zinc flotation using conventional and column cells;  Lead concentrate dewatering, filtration and truck load-out facility;  Zinc concentrate dewatering, filtration and truck load-out facility; and  Tailings disposal.

1.11 Infrastructure The project envisions the upgrading or construction of the following key infrastructure items:

 Primary power supply from the NWT power grid;  Secondary and back-up power supply from LNG generators;  Tailings management facilities in existing and mined out open pits;

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 Waste rock storage areas (WRSA) in existing open pits or external stockpiles;  Pit dewatering wells and reinjection wells;  Two crushing and dense media separation circuits;  Lead/Zinc process facility with grinding, flotation, thickening and dewatering;  Assay laboratory;  Administration, mine dry and first aid facilities;  Mine maintenance facilities;  Plant warehouse facilities;  Bulk fuel storage and distribution; and  Temporary worker housing complex.

1.12 Environment and Permitting The PEA shows the Pine Point project has key positive features that will minimize environmental impacts, namely:

 No tailings pond, no tailings dam: the tailings management concept that is developed for the Pine Point project is based on “in-pits” disposal (refer to Section 18.8 – Tailings Management); as the Project does not include a tailings pond facility with dams, this approach thus eliminates the geotechnical risks associated with long term operation of tailings dams, while also eliminating the need for long term treatment and discharge of tailings pond water;  Minimal surface discharge of mine dewatering water: most water from the dewatering of the open pits will be returned, by gravity in most instances, to adjacent historical open pits or recently mined pits (refer to Section 18.6 – Mine Water Management);  No discharge of liquid effluents from the process plant: the process plant will operate under a zero discharge operating plan whereby water will recycled/reused from a basin, available or constructed (refer to Section 17);  Geochemistry of waste rock and DMS rejects: Previous available results from geochemistry tests by Tamerlane where reviewed by the consultant “pHase Geochemistry” as part of the PEA. Phase Geochemistry concluded that the available results (static ABA) bring evidence that the waste rock samples tested based on the Tamerlane mine plan at the time were non-PAG. However, pHase Geochemistry concluded that the representativeness of the samples tested at this point will need further evaluation at the Feasibility Study level as a function of for the proposed mine plan under the PEA; and  Small footprint: The management concept for the Waste Rock and the DMS rejects is to minimize the surface storage footprint.

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1.13 Capital Cost Estimate Estimated project capital costs total $271 M, consisting of the following distinct stages:

 Initial Capital Costs – includes all pre-production costs to develop the cluster pit resources to a 1,800 t/d operation. Initial capital costs total $154 M and are expended over a 24-month pre- production construction and commissioning period;  Sustaining Capital Costs – includes all costs related to the acquisition, replacement, or major overhaul of assets during the mine life required to sustain operations. This includes the DMS plant and infrastructure required to develop the N-204 resource. Sustaining capital costs total $100 M and are expended in operating Year 1 through 13; and  Closure Costs – includes all costs related to the closure, reclamation, and ongoing monitoring of the mine post operations. Closure costs total $17 M, and are primarily incurred in Year 14, with costs extending into Year 18 for ongoing monitoring activities.

The capital cost estimate was compiled using a combination of quotations, database costs, and database factors.

Table 1.6 presents the capital estimate summary for initial, sustaining, and closure capital costs in Q1 2017 Canadian dollars with no escalation.

Table 1.6: Capital Cost Summary

Pre-Production Sustaining/Closure Total Capital Costs (C$M) (C$M) (C$M) Open Pit Mining 17.1 14.4 31.5 Mine Water Management 2.7 17.0 19.8 Site Development 3.6 2.7 6.3 Processing 52.7 29.0 81.6 Infrastructure 24.8 15.8 40.6 Indirect Costs 32.8 9.9 42.7 Closure - 17.5 17.5 Subtotal 133.7 106.3 240.1 Contingency @15% 20.1 11.2 31.2 Total Capital Costs 153.8 117.5 271.3 Source: JDS (2017)

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1.14 Operating Cost Estimate The operating cost estimate (OPEX) is based on a combination of experience, reference project, budgetary quotes and factors as appropriate with a preliminary study.

Preparation of the operating cost estimate is based on the JDS philosophy that emphasizes accuracy over contingency and utilizes defined and proven project execution strategies.

The operating cost estimate in this study includes the costs to mine, process the mineralized material to produce lead and zinc concentrates, and general and administrative expenses (G&A). These items total the mine operating costs and are summarized in Table 1.7.

The target accuracy of the operating cost is -25/+30%. The operating costs are based on leased mine production equipment.

The operating cost estimate is broken into three major sections:

 Open Pit Mining;  Mineral Processing; and  General & Administrative.

The total operating unit cost is estimated to be $37.64/t processed (or $38.03/t processed excluding pre-production tonnes). Total LOM and unit operating cost estimates are summarized in Table 1.7.

Table 1.7: Summary of Operating Cost Estimate

Operating Costs C$/t milled** LOM (C$M) Mining* 21.32 630 Processing 11.49 339 G&A 4.83 143 Total 37.64 1,112 *Average LOM mining cost amounts to $4.29/t mined at a 4:1 strip ratio. *includes pre-production tonnes milled. Source: JDS (2017)

Operating costs are expressed in Canadian dollars. No allowance for inflation has been applied.

The main OPEX assumptions are outlined in Table 1.8.

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Table 1.8: Main OPEX Assumptions

Item Unit Value Electrical power cost (NWT Power) $/kWh 0.11 Electrical power cost (LNG generators) $/kWh 0.14 Average power consumption MW 6.3 Overall power consumption (all facilities) kWh/t processed 20.4 Diesel cost (delivered) $/litre 0.88 LOM average manpower (including contractors, excluding corporate) employees 321 Source: JDS (2017)

1.15 Economic Analysis 1.15.1 Main Assumptions

Table 1.9 outlines the metal prices and exchange rate used in the economic analysis.

Table 1.9: Metal Prices and F/X Rate

Parameter Unit Value Lead Price US$/lb 1.00 Zinc Price US$/oz 1.10 Exchange Rate US$:C$ 0.75 Source: JDS (2017)

No preliminary market studies were completed on the potential sale of lead concentrate and zinc concentrate from the Pine Point Project. The terms selected are in-line with current market conditions and were reviewed and found to be acceptable by QP Garett Macdonald, P.Eng.

No contractual arrangements for shipping, port usage, or refining exist at this time. Table 1.10 outlines the terms used in the economic analysis.

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Table 1.10: Net Smelter Return Assumptions

Assumptions & Inputs Unit Value Lead Concentrate Metal Recovery to Pb Con (CP) % 85.0 Metal Recovery to Pb Con (N-204) % 71.4 Average Pb Recovery to Concentrate % 80.4 Average Pb Concentrate Grade Produced % 65 Minimum Deduction % Pb/t 3.0 Pb Payable % 95.0 Pb Treatment Charge C$/dmt conc. 200.00 Pb Concentrate Transportation Cost C$/wmt 167.00 Zinc Concentrate Metal Recovery to Zn Con (CP) % 88.1 Metal Recovery to Zn Con (N-204) % 78.0 Average Zn Recovery to Concentrate % 83.4 Average Zn Concentrate Grade Produced % 58.9 Minimum Deduction % Zn/t 8.0 Zn Payable % 85.0 Zn Treatment Charge C$/dmt conc. 200.00 Zn Concentrate Transportation Cost C$/wmt 167.00 Source: JDS (2017)

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1.15.2 Results

The economic results for the Project based on the assumptions outlined in Section 1.15.1 are shown in Table 1.11.

Table 1.11: Economic Results*

Parameter Unit Value Pre-Tax Operating Income C$M 908 Pb (Straight Cash Cost) US$/lb 2.51 Pb Cash Cost (Net of By-Products) US$/lb (0.27) Zn (Straight Cash Cost) US$/lb 0.99 Zn Cash Cost (Net of By-Products) US$/lb 0.60 Cash Flows Working Capital C$M 20 LOM C$M 637 Pre-Tax Cash Flow C$M/a 49 Taxes LOM C$M 214 LOM C$M 423 After-Tax Cash Flow C$M/a 33 Economic Results

Pre-Tax NPV8% C$M 341 Pre-Tax IRR % 48 Pre-Tax Payback Years 1.4

After-Tax NPV8% C$M 211 After-Tax IRR % 35 After-Tax Payback Years 1.8 *This preliminary economic assessment is preliminary in nature and includes the use of inferred mineral resources that are considered too speculative geologically to have 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. Source: JDS (2017)

The pre-tax break-even zinc price for the project is approximately US$0.76/lb, based on the LOM plan presented herein, a lead price of US$1.00/lb and an FX rate of 0.75 US$:C$.

1.15.3 Sensitivities

A simplistic sensitivity analysis was performed to determine which factors most affect the project economics and is discussed in Section 23. Each variable evaluated was tested using the same sensitivity values, although some may be more likely to experience significantly more fluctuation in value over the LOM (i.e. CAPEX versus metal prices).

Sensitivity analyses were performed on metal prices, exchange rate, mill feed grade, recovery, capital costs, and operating costs as variables. The value of each variable was changed plus and

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Table 1.12: Sensitivity Results (After-Tax NPV8%)

Parameter -40% -30% -20% -10% Base +10% +20% +30% +40%

Price (271) (123) 6 111 211 310 410 508 607

C$:US$ FX 870 635 459 321 211 120 41 (30) (104)

Mill Feed Grade (135) (31) 54 133 211 287 364 441 517

Recovery (334) (163) (11) 108 211 310 410 508 607

OPEX 377 335 294 252 211 169 127 84 39

CAPEX 298 276 254 233 211 189 167 145 123

Source: JDS (2017)

The analysis revealed that the project is most sensitive to exchange rate, followed by recovery, metal prices, mill feed grade, and operating costs. The Project showed the least sensitivity to capital costs.

1.16 Conclusions It is the conclusion of the QP’s that the PEA summarized in this technical report contains adequate detail and information to support the positive economic result herein contained. The PEA proposes the use of industry standard equipment and operating practices. To date, the QP’s are not aware of any fatal flaws for the Project.

Using the assumptions highlighted in this report, the Pine Point Project offers sufficient economic potential to be advanced to the next stage of study (Feasibility Study).

1.16.1 Risks

The most significant potential risks associated with the Project are:  Ground water management;  Operating and capital cost escalation;  Unforeseen schedule delays;  The ability to attract and retain experienced professionals;  The ability to raise financing; and  Metal prices.

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These risks are common to most mining projects, many of which may be mitigated, at least to some degree, with adequate engineering, planning and pro-active management.

1.16.2 Opportunities

There are significant opportunities that could improve the economics, timing, and/or permitting potential of the Project. The major opportunities that have been identified at this time, excluding those typical to all mining projects, such as changes in metal prices, exchange rates, etc., are:

 The addition of underground resources into the mine plan. Pine Point has two significant underground deposits, R190 and X25, that have potential to contribute higher grade feed to the process plant. Resources at these two deposits include; o R190: 1.0 Mt (Measured & Indicated) grading 5.28% Pb and 10.98% Zn; and o X25: 2.1 Mt (Indicated) grading 2.32% Pb and 6.73% Zn.  Conversion of historical resources into Measured and Indicated Resources, through confirmation drilling. Historically there are 46 undeveloped deposits on the Pine Point Project and of these, 10 are included in the PEA;  Explore for additional deposits - Although more than 100 distinct deposits have been outlined historically in the Pine Point area, there are several portions of the mineralized trends that remain under-explored;  Additional metallurgical testing could increase either recoveries, concentrate grades, or both;  Additional analysis to determine ways to lower the costs of water handling; and  The PEA envisages the use of electric power sourced from a combination of local grid power and natural gas generators on the property. Upgrades to the Taltson River hydroelectric dam would reduce capital and operating costs significantly.

1.17 Recommendations In the opinion of JDS, financial analysis of this PEA demonstrates that the Pine Point Project has positive economics and warrants consideration for advancement to feasibility-level engineering by Darnley Bay. This more advanced study will further detail:

 A Mineral Reserve estimate;  Process engineering based on the PEA flowsheets;  Project scheduling;  Capital and operating cost estimation; and  Economic results. The study will improve the confidence in the Project design and execution and will result in an improved accuracy of project economics.

It is estimated that a FS and its supporting work programs would cost approximately $2.6 M.

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2 Introduction

2.1 Basis of Technical Report Darnley Bay Resources Ltd. (Darnley Bay) commissioned JDS Energy and Mining Ltd. (JDS) to complete a Preliminary Economic Assessment (PEA) for the Pine Point Project. The purpose of this study is to complete a review and compilation of the resources, mining designs and preliminary economics.

This report was prepared in order to fulfill the reporting requirements stipulated by Canadian National Instrument 43-101, Standards of Disclosure for Mineral Projects (NI 43-101).

It must be noted that this PEA is preliminary in nature and includes the use of Inferred Mineral Resources (100% of mine plan tonnes) for which the geology is considered too speculative to be categorized as Mineral Reserves. Therefore, there is no certainty that the PEA will be realized.

2.2 Scope of Work The following companies contributed to this Technical Report and provided Qualified Person (QP) sign-off for their respective sections: JDS Energy & Mining Inc. (JDS):  Compile the technical report including information provided by other consulting companies;  Establish an economic framework for PEA mineral resources;  Select mining equipment;  Develop a conceptual flowsheet, specifications and selection of process equipment;  Design required site infrastructure, identify suitable sites for plant facilities and other ancillary facilities;  Estimate mining, process plant and infrastructure OPEX and CAPEX for the project;  Prepare a financial model and conducting an economic evaluation including sensitivity and project risk analysis;  Interpret the results and make conclusions that lead to recommendations to improve value and reduce risks; and  Evaluate existing project studies and data and provide recommendations for additional studies and data collection required to advance the project to a feasibility study. Knight Piésold Ltd. (KP)  Tailings and Water Management. Albert Daniel Siega, P.Eng, Independent Consultant  Mineral Resource estimate.

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Paul Gann, P.Geo, Independent Consultant  Data verification of geology, sample collection and sample processing.

2.3 Qualifications and Responsibilities The results of this PEA are not dependent upon any prior agreements concerning the conclusions to be reached, nor are there any undisclosed understandings concerning any future business dealings between Darnley Bay and the respective QP’s. The QP’s are being paid a fee for their work in accordance with normal professional consulting practice.

The following individuals, by virtue of their education, experience and professional association, are considered QP’s as defined in the NI 43-101, and are members in good standing of appropriate professional institutions/associations. The QP’s are responsible for the specific report sections as follows in Table 2.1.

Table 2.1: QP Responsibilities

QP Company QP Responsibility/Role Report Section(s) Overall Responsibility, 1 to 3, 18 (except 18.6 and Garett Macdonald, P. Eng. JDS Energy & Mining Inc. Infrastructure, Costs And 18.7), 19, 20, 21 to 27 Economics Mineral Reserves Estimate, Dino Pilotto, P.Eng. JDS Energy & Mining Inc. 15, 16 Mining Tailings and Water Ken Embree, P. Eng. Knight Piésold Ltd. 18.6, 18.8 Management Kelly McLeod, P. Eng. JDS Energy & Mining Inc. Process & Metallurgy 13, 17 Albert Daniel Siega, P.Eng Independent Consultant Mineral Resource Estimate 14 Paul Gann, P. Geo Independent Consultant Geology 4 to 12 Source: JDS (2017)

2.4 Site Visit QP site visits were conducted as per Table 2.2.

Table 2.2: QP Site Visits

Accompanied Description of Qualified Person Company Date by Inspection JDS Energy & Stanley Pine Point Project Dino Pilotto, P.Eng. March 3, 2017 Mining Inc. Clemmer Inspection Stanley Pine Point Project Ken Embree, P. Eng. Knight Piésold Ltd. March 3, 2017 Clemmer Inspection Independent Pine Point Project Albert Daniel Siega, P.Eng September 5, 2013 Consultant Inspection Source: JDS (2017)

Garett Macdonald, Kelly McLeod and Paul Gann did not perform site visits but relied on the other QP’s for information.

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2.5 Units, Currency and Rounding The units of measure used in this report are as per the International System of Units (SI) or “metric” except for Imperial units that are commonly used in industry (e.g., ounces (oz.) and pounds (lb.) for the mass of precious and base metals).

All dollar figures quoted in this report refer to Canadian dollars (C$ or $) unless otherwise noted.

Frequently used abbreviations and acronyms can be found in Section 29. This report includes technical information that required subsequent calculations to derive subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, the QP’s do not consider them to be material.

2.6 Sources of Information This report is based on information provided by Darnley Bay throughout the course of JDS’s investigations. Other information was obtained from the public domain. JDS has no reason to doubt the reliability of the information provided by Darnley Bay. This technical report is based on the following sources of information:

 Discussions with Darnley Bay personnel; o John Key, Chief Operating Officer (COO); o Tim Smith, Vice President, Operations; and o Judy Dudley, Vice-President, Environment.  Review of geochemistry, hydrogeology and hydrology data collected by Darnley Bay;  Review of geological exploration data collected by Darnley Bay; and  Review of metallurgical data collected by Darnley Bay.

Engineering and geological information from historical documents were used in this report after determination by JDS that the work was performed by competent persons or engineering firms. Data derived from engineering companies, consultants and authors listed in the reference section of this report. The documentation received and the sources of information are listed in Section 28.

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3 Reliance on Other Experts

The QP’s opinions contained herein are based on information provided by Darnley Bay and others throughout the course of the study. The QP’s have taken reasonable measures to confirm information provided by others and take responsibility for the information.

Non-QP specialists relied upon for specific advice are:

 Marcel Pineau, P.Eng., PhD. for environmental studies;  Wentworth Taylor, CPA for taxation guidance; and  Shannon Shaw, M.Sc., P.Geo (BC, NWT), pHase Geochemistry Inc. for geochemistry data review.

The QP’s used their experience to determine if the information from previous reports was suitable for inclusion in this Technical Report and adjusted information that required amending.

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4 Property Description and Location

4.1 Area of the Property The mineral claims and mining leases that comprise Darnley Bay’s property in the Pine Point District encompass a total of 18,009 hectares (ha) comprising two mineral claims and 40 mining leases.

The Pine Point deposits discussed in this report are on nine mining leases with a total area of 3,582 ha (see Table 4.1). Deposits on Darnley Bay’s remaining 31 leases are not covered in this report. It should be noted that the actual project footprint within each lease will be small relative to the total lease area.

Table 4.1: Resources Lease Surface Area

Area of Lease Deposit Lease Name Lease Number Expiration Date (Hectares) K-68 M11 5255 8/25/2033 536 M-67 M11 5255 8/25/2033 536 J-68 M11 5255 8/25/2033 536 M-62/63 M10 5254 8/25/2033 389 Hinge Zone (HZ) M10 5254 8/25/2033 389 L-65 M10 5254 8/25/2033 389 W-85 N9 4871 7/16/2028 692 O-53 M8 5252 8/25/2033 592 X-65 N3 5241 8/25/2033 251 N-204 N17 5245 8/25/2033 65 N-204 N-18 5244 8/25/2033 39 N-204 S1 5246 8/25/2033 82 Source: Siega & Gann (2014)

4.2 Location of the Property Darnley Bay’s Pine Point properties, including the 10 subject deposits described in this report, are located about 800 km north of Edmonton, near the south shore of Great Slave Lake, in the Mackenzie Mining Division of the Northwest Territories (NWT) of Canada. The mineral leases lie north of the Territorial Highways 5 and 6 which connect Hay River and the Pine Point town site, and extend intermittently from 25 to 80 km east of the town of Hay River. An all-weather year-round highway parallels the southern boundary of the Darnley Bay lease block (Figure 4.1).

The Darnley Bay leases range between 42 and 110 km from Hay River, NWT. The lease areas are situated about 10 km south of the Great Slave Lake and lie about 60 m above the lake level, which is at an elevation of 156 m above sea level (masl). Geographic coordinates are from 114° to 115° 15’ West longitude and from 61° 0’ to 61° 45’ North latitude.

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Figure 4.1: Darnley Bay’s Leases and Claims

Source: Northwest Territories Mining Recorder’s Office Database (2017)

The ten deposits evaluated in this document include: J-68, HZ, M-67, K-68, M-62/63, X-65, W-85, O- 53, L-65 and N-204.

4.3 Type and Details of Mineral Tenure The mineral deposits at Pine Point are covered by 40 leases in total. The leases are held by Darnley Bay as a result of transfer of title on December 21, 2016 from Pine Point Holding Corp (PPHC) to Darnley Bay Resources Limited. As of the effective date of this report, the leases are in good standing.

4.4 Issuer’s Title To and Interest in the Property 4.4.1 Surface Rights

The surface rights to Darnley Bay’s leases at Pine Point are administered by the Government of the Northwest Territories. Surface rights may be used in conjunction with valid mineral tenure. All the land within the project area is unoccupied.

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Surface access to the land is granted through the Mackenzie Valley Land Use Regulations for short term use of the land (Land Use Permits) or the Territorial Lands Act for long term use (Surface Leases).

4.4.2 Obligations for Retention of Property

Mining leases are granted for 21-year periods. They can be renewed for a further 21-year period for a fee of $25/acre. Yearly rental fees for leases are $1/acre/year for the first 21-year period and $2/acre/year for subsequent 21-year periods.

4.4.3 Expiration Date of Claims, License, or Other Property Tenure Rights

All leases discussed in this report are in good standing until May 9, 2028 (Table 4.1).

4.5 Agreements and Encumbrances Darnley Bay has a 100% interest in the mineral leases discussed in this report, subject to a 3% Net Smelter Return (NSR) payable to Karst Investments LLC (“Karst”). The project is not subject to any other royalties, back-in rights, payments, or other agreements or encumbrances other than the territorial royalty (calculated as a tax but called a royalty).

4.5.1 Environmental Liabilities

Existing site liabilities are minimal, due mainly to the “brownfield” (i.e. previously disturbed) nature of the project area. Potential liabilities for the project are mainly focused around ongoing water management and the need to manage groundwater discharge required during the dewatering and mining phases of the project.

Other resource areas studied for impact and mitigation measures as part of the process for obtaining permits for the Pine Point Pilot Project included: archaeological resources, noise, air quality, geology and soils, surface water, groundwater, wildlife, vegetation and habitat, species at risk, and socioeconomics. In all cases, the potential for adverse impacts associated with the project were determined to be minimal. A number of commitments were agreed to by Tamerlane, assumed by Darnley Bay, as part of the Environmental Impact Assessment process in order to reduce potential adverse impacts, and they are detailed in Appendix B of the Mackenzie Valley Environmental Impact Review Board's (MVEIRB) 2008 report.

Cumulative impacts of the project in a regional context are considered minimal.

4.6 Permitting 4.6.1 Land and Environmental Permits

Darnley Bay must obtain both a Water License and a Land Use Permit to proceed with mine development.

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Tamerlane held a Mineral Exploration Permit for all of the leases in the Pine Point area. The permit was issued on July 2, 2009, with an expiry date of July 1, 2016. Pine Point Holding Corp. received renewal of the permit on December 19, 2016, and the permit was formally assigned to Darnley Bay Resources Limited on February 16, 2017.

Pine Point Holding Corporation held both a Type-A Land Use Permit for the Construction and Operation of an Underground Lead/Zinc Mine at the R-190 pilot project and a Type-A Water License for dewatering, mining, milling, and wastewater discharge purposes at the R-190 pilot project. Both permits expired in March 2017. Those permitted activities have been Pre-Screened by the MVLWB.

4.7 Other Factors and Risks The majority of the Darnley Bay leases are surrounded by the Akaitcho and the Dehcho Interim Land Withdrawals. New mineral claims and mining leases are prohibited on withdrawn lands. The mining leases and claims issued to Tamerlane and PPHC predate the Land Withdrawals, and as such the Land Withdrawals have no effect on the sub-surface rights of the mineral claims and mining leases discussed in this report.

There are no known access problems to this mining site. It can be accessed via highway, boat, and by plane.

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5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1 Topography, Elevation, and Vegetation The Pine Point area is generally low lying boggy ground with karst features in the vicinity of the deposits. Surface topography slopes gently northward toward Great Slave Lake.

5.1.1 Physiography

Pine Point is located in a sub-arctic environment near 61° North latitude. Wetlands, primarily swampy muskeg, and shallow, marl-bottomed ponds characterize the project area topography. A blanket of glacial till and outwash gravel, sand and clay produce a low-lying, hummocky terrain that slopes gently northward toward the Great Slave Lake. The Great Slave Lake, the 10th largest lake on earth, with a surface area of 27,200 km2, has a surface elevation of 156 masl (metres above sea level).

The north-flowing Buffalo River is the principal drainage in the Project area; the Little Buffalo River, Birch Creek, Hanbury Creek and Twin Creek also drain portions of the Pine Point area. All are tributaries to the Great Slave Lake. Total topographic relief in the project area is approximately 60 m. In general, the surface elevation is approximately 200 masl along the Main Trend and 180 masl along the North trend (Figure 5.1). There are some surface expressions of modern karst features, including intermittent creeks, natural springs, and sinkholes. Many sinkholes have been filled by glacial debris.

5.1.2 Vegetation and Wildlife

Pine Point is located in the Mackenzie and Slave Lowland Mid-Boreal Ecoregion. The area is characterized by short, cool summers and long, cold winters. Surface vegetation is a mosaic of upland and wetland plant communities. The dominant trees in the uplands are scrub jack pine, black spruce and willows, which are interspersed between large wetland areas. The ground around the former open pits is disturbed and the abandoned pits are mostly filled with surface water

A variety of game and non-game animal species are found at Pine Point including moose, black bear, deer, wolf, beaver, fox, and hares. The area is south of the normal caribou migration areas and northwest of the predominant wood buffalo range, although both species occasionally occur at Pine Point.

5.2 Means of Access to the Property Access to the Pine Point property is by 1,100 km of paved highway north from Edmonton, Alberta to Hay River and then a paved highway to the Pine Point District. Access to the deposits within the project area is via gravel haul roads, most of which are of a good quality.

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The town of Hay River has all major services including an airport with scheduled jet service from Edmonton and Yellowknife, a rail terminal and a port from which barge traffic traverses the Mackenzie and Slave Rivers. An airstrip suitable for small aircraft exists near the former Pine Point town site. The airstrip in Fort Resolution has year-round access. Many parts of the project claim area can be accessed using snowmobiles, tracked vehicles and other all-terrain vehicles during the winter months. A series of all-season haul roads built to service the various open pit and underground mines that were worked by Pine Point Mines Ltd. (Cominco) can be used to access the resource areas. Many of these roads remain serviceable, and summer light-truck access could reestablish access to outlying resource areas with repair of culverts and breaches.

5.3 Proximity to Population Centers The town of Hay River, with a population of about 3,500 people, is approximately 65 km west of the Cluster Pit area. The remainder of Darnley Bay’s leased deposits is further east of the Cluster Pits with the most distant deposit located 120 km from Hay River. Hay River has all major services including an airport with scheduled service from Edmonton and Yellowknife, a rail terminal, and a port from which barge traffic traverses the Mackenzie and Slave rivers. There is a bridge over the Buffalo River that connects the east and west sides of the claim block areas.

The Hamlet of Fort Resolution, with a population of about 500 residents is approximately 160 highway km east of Hay River. Services in the town include an RCMP station, a general store, a motel, and a small airport.

5.3.1 Climate

Winter conditions are sub-arctic, with average January temperatures of -17° C Summers are hot and relatively dry with an average July temperature of +16° C with long daylight periods. The mean annual precipitation is 34 cm, which includes 18 cm of rainfall, and the snowfall contains an additional equivalent 16 cm of water. Snowfall is relatively light, with maximum snow pack of 1 m or slightly more. The frost-free season averages 95 days. Discontinuous permafrost exists in the region.

5.4 Infrastructure Surface rights are discussed in Section 4.4.1.

Currently, infrastructure near the project area includes a power line that connects the Taltson Dam to nearby communities, and an all season, two-lane, paved highway that is located within 1 km of the proposed plant site. All of the land within the project area is unoccupied.

This project can be considered a brownfield-type project. Historically, some of the leased areas had open pit and underground mining, railroad lines, power lines, processing facilities, buildings and a tailings dam.

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In 1965, an 18 MW hydroelectric plant was built on the Taltson River to supply power to Hay River, Fort Resolution, Fort Smith and the former Pine Point Mines Ltd. (Cominco) mine site at Pine Point. A 274 km long single-circuit 115 kV transmission line connects the hydro plant to the mine site. The capacity of the electrical system was increased in 1976 to allow for an electric powered dragline. At present the power line is routed from the main Hay River road to the N-81 pit and from there, along an old haul road to the Pine Point Substation.

Mining personnel are available, both in the nearby communities of Hay River and Fort Resolution, and through the air and road transportation facilities at Hay River.

There are sufficient potential mine waste disposal areas available in the vicinity of the proposed new open pits.

5.4.1 Exploration and Operating Season

The exploration season is year round as well as 24/7. The only drawback is that during the short summer season, swampy ground is hard to access, therefore for certain areas of the property the best time for exploration drilling is the winter season.

Darnley Bay started exploration drilling in April 2017 and has recently begun geophysics studies. Drilling consists of exploration drilling on new targets as well as confirmation drilling of historical deposits.

Historically, Cominco operated Pine Point Mines 365 days per year 24 hours per day. There are no reasons that subsequent operations could not also follow this production schedule.

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6 History

6.1 Early History The following is adapted from Neil Campbell’s 1957 report, Stratigraphy and Structure of the Pine Point Area, NWT (Campbell, 1957).

“The occurrence (at the Pine Point Property) of lead and zinc was reported in 1899 by R. Bell of the Geological Survey of Canada, and in the following years several claims were staked and some prospecting was done. In 1928 Atlas Exploration Co. took over the property containing the discovery occurrences. In 1929, Cominco and Ventures Ltd. participated with Atlas in forming the Northern Lead- Zinc Co., which, under the direction of J. M. Bell, carried out 7,467 m of drilling in the discovery and surrounding areas. Two shafts and many surface pits were also sunk. This work showed that the discovery deposits contained about 450,000 t of lead-zinc mineralization, but no new deposits was found. In 1936, D. F. Kidd drew attention to the highly significant space relations between the base- metal discoveries at Pine Point and the projections of certain major Precambrian faults mapped earlier by C. H. Stockwell. Geological work by Cominco in 1940 and subsequent years suggested the presence of potentially economic deposits within a belt of land including part of the area of earlier field work and extending far beyond it. Between 1946 and 1954, exploration financed by Cominco and Ventures Ltd. included 55,763 m of diamond drilling, 117 m of shaft sinking, and 73 m of drifting, raising, and underground development. This work was successful in finding substantial quantities of new mineralization in places where bedrock is completely concealed. The topography is one of low relief, characterized in many places by swampy muskeg and shallow, marl-bottomed ponds marl or marlstone is a calcium carbonate. The area where mineralization has been found is generally higher ground which, though not well drained, is comparatively dry. It is also marked by raised beaches and a few sand dunes overgrown by small spruce and pine trees. Bedrock is generally overlain by a variable thickness of hard, glacial boulder clay. This in turn is covered by silt, sand, and gravel, probably deposited from the waters of Great Slave Lake, which formerly extended many miles beyond its present southern shore.” The CNR Great Slave Lake Railway was completed in 1964 connecting Hay River to the former Pine Point town site. The tracks between Hay River and the Pine Point area were removed in the early 1990’s when the former Pine Point mine ceased shipments of concentrates.

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6.2 Timeline The following section is adapted from PAH (2008), based on a report written by Ross Burns (Burns and Kent, 2001).

Geologic exploration of lead-zinc showings south of Great Slave Lake dates from 1898. Prospectors heading for the Klondike gold rush staked claims on outcrops of oxidized sulphide. Intermittent exploration continued until 1948 when Cominco Ltd. began major exploration programs that continued until 1955.

A chronological history of the Pine Point area is adapted from Burns and Kent (2001):  1898: A fur trader at nearby Fort Resolution staked eight claims. When it was discovered that the mineralization contained no gold or silver the claims were allowed to lapse.  1899: The deposits were reported by Dr. Robert Bell, Director of the Geologic Survey of Canada.  1908: A few claims were staked but allowed to lapse.  1914: Additional claims were staked by a mining engineer but these were allowed to lapse. Dr. Charles Camsell reported on the deposits to the Geologic Survey of Canada.  1916: Dr. A.E. Cameron of the Geologic Survey made positive reports on the geology of the area.  1920 : Dr. J. McIntosh Bell, who had been with Dr. Robert Bell in 1899 on site,, sent an engineer, D.B. Dawson, to examine the deposits. This enterprise was assisted financially by certain Boston interests. Claims were again staked.  1921: The work of 1920 was continued and reported favorably. Dr. Bell evidently considered these deposits to be geologically similar to the famous Tri-State lead-zinc deposits of southeastern Missouri.  1926: W.M. Archibald, Cominco’s Manager of Mines, became interested in the property and began forming an exploration staff.  1927: W.L. McDonald and Ted Nagle (son of the original claim staker) of Cominco’s new group visited and reported favorably on the deposits.  1928: Again under Dawson, and now joined by the Atlas Exploration Company, a much more extensive program of churn drilling was begun. Shaft sinking was undertaken. Ted Nagle, noting the new activity, secured an option on 16 adjoining claims for Cominco.  1929: The Boston interests, Atlas Ventures Ltd. and Cominco combined to form the Northern Lead-Zinc Company. A 1929 - 1930 work program under Dr. Bell included sinking pits and shafts plus drilling. With surface showings fully explored, the resource estimate was a disappointing half-million tons of 15% lead-zinc. Widespread surface work failed to find further indications of mineralization. Northern Lead-Zinc would not advance more funds, but Dr. Bell’s optimism was shared by Archibald, who obtained Cominco financing to hold the ground. In that period Cominco, through Mr. Archibald, developed a corporate flying organization to explore the north with aircraft.

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 1930: From this year until 1948, only enough work was carried out to maintain 104 claims. This was undertaken and financed by Cominco.  1940s: Cominco’s geological staff extended geological theories and concluded that a number of deposits might occur in the vicinity.  1948: Cominco obtained a 500 square mile concession surrounding the area of known mineralization. A second concession was obtained the following year. The ownership and the exploration and development program, which required seven years of work, was shared with Ventures Ltd.  1951: Pine Point Mines Limited was formed to finance the continuing work. Cominco obtained a 78% interest.  1961: It was in 1961 under the “Roads to Resources” program that the agreement of a railway to Great Slave Lake became a reality. The agreement was reached between the Federal Government, Pine Point Mines Limited, and Canadian National Railways whereby the Government undertook to construct the railway and Cominco undertook to bring Pine Point Mines into production. Total investment, including the railroad, mill and hydroelectric plant exceeded $130,000,000. The Northern Canada Power Commission agreed to build a 25,000 horsepower (approximately 18.6 megawatt) hydroelectric plant on the Taltson River to supply power to Pine Point, Fort Smith and other future developments. The cost was underwritten by Pine Point Mines Limited.  1962: Railroad construction began, and mining preparations were commenced with Cominco continuing as manager and agent of Pine Point Mines Limited.  1963: A town site was laid out and services installed, in collaboration with the Department of Northern Affairs and National Resources. Modern bunkhouses, a dining room, recreation hall and 53 homes were built. Stripping and construction of a 5,000 ton per day concentrator commenced as part of a $23,000,000 project.  1964: The CNR Great Slave Lake Railway was completed in 1964, connecting Hay River to the Cominco mine site near the town of Pine Point. The railway reached Pine Point late in the year, well ahead of schedule. Construction of the power plant on the Taltson River and construction of the ore concentrator were not due for completion until the end of the following year.  1965: Early transportation and the availability of high grade ore averaging 50% combined lead- zinc, made possible ore shipments to Cominco’s treatment plants in British Columbia. Ore shipments to treatment plants continued after the concentrator was completed in November. High mining interest in the Pine Point area generated a major staking rush. Late in the year the Pyramid Mining Co. Ltd. found a major deposit to the east of Pine Point’s ground. An 18 MW hydroelectric plant was built on the Taltson River to supply power to Fort Smith and a 274 km long single-circuit, 115 kV transmission line connected the hydro plant to the Cominco mine site. The capacity of the electrical system was increased in 1976 to allow for an electric powered dragline.  1966: Pine Point Mines Ltd. acquired Pyramid’s mineral claims in the area and expanded the mill to 10,000 tons per day (t/d).

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 1970: A test program was undertaken to determine the feasibility of mining some of the ore by underground methods.  1975: Western Mines (later known as Westmin Resources, Ltd, or "Westmin"), acquired claims west of Cominco's property, which is essentially west of the Buffalo River. Westmin then proceeded to conduct an extensive IP survey and drilling program referred to as, “The Great Slave Reef (GSR) Project”. This project was a joint venture of Westmin, controlled by Boliden of Sweden, DuPont Exploration Canada.  1976-1979: Western Mines Ltd. commences exploratory drilling.  1979: Pine Point Mines Ltd. starts an intense exploration program to delineate more reserves.  1983: D. Rhodes et al publish a paper on the geology of Pine Point.  1985: In January 1985, Cominco reduced its 69% interest in Pine Point to approximately 51%.  1986: Pine Point Mines Ltd. decides to close the mine in an orderly fashion and commences permanent layoffs. An attempt is made to form a grout curtain to eliminate or reduce pumping at the N81 ore body, containing 2.6 million tons at a grade of 7% lead and 14% zinc. The attempt fails due to technical problems and it is decided to mine the deposit very quickly and stockpile the ore alongside the pit then haul it to the mill as needed. The ore was milled and the concentrate stored and shipped as required over the next few years. The decision was taken to remove the town as well as the mill facilities and this was completed in the early 1990s; the tailings pond was covered with coarse material to prevent windblown tails. From the original rail line between Hay River and Pine Point, only the Hay River Railway Bridge and the rail bed remains. The four concrete pylons for the Buffalo River railway bridge are still in place. The track beds were contoured and set back from smaller stream crossings such as Twin Creek and Birch Creek to reestablish normal drainage.  2000: Pine Point Mines Ltd. (Cominco) allows some claims to lapse.  2001: All remaining Pine Point mining leases expire and the Westmin (Boliden)-Dupont mining leases expire.  2001: Kent Burns Group (now Karst) stake all the Pine Point and Westmin claims.  2004: Karst Investments options claims to Tamerlane Ventures.  2006: Tamerlane acquires 100% interest in the Pine Point and Westmin claims, subject to a 3% NSR to Karst.  2008: PAH completes an NI 43-101 Technical Report on the Pine Point Project.  2011: PAH completes an NI 43-101 Technical Report on the N-204 deposit.  2012: MineTech completes an NI 43-101 Technical Report on the N-204 deposit and another Technical Report on deposits R-190, X-25, P-499, O-556, Z-155 and G-03.  2014: Siega and Gann authored the 2014 TAM Technical Report on March 14, 2014.  2016: Paul Gann NI 43-101 Summary Technical Report Update of the Pine Point Mine Development Project, NWT, Canada

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 2017: Siega & Gann; NI 43-101 Technical Report Resource Summary, Pine Point, NWT, Canada

6.3 Prior Ownership When the mining leases expired in 2001, Mr. Ross Burns staked a large part of the former Pine Point Mine property and the Great Slave Reef property, which covers the down-plunge extension to the west formerly owned by Westmin/Dupont/Boliden. The property position was transferred to Karst. Shortly thereafter, the property was optioned to Terrastar Incorporated, which became Pine Point Mines Incorporated. In 2003, the property was returned to Karst Investments, and in 2004 Tamerlane acquired an option to earn 60% interest in the Pine Point property from Karst (CAM, 2007). In September 2006, Tamerlane acquired the remaining 40% interest in the Pine Point property, making Tamerlane a 100% owner (subject to a 3% NSR back to Karst). In 2011, for financing purposes, Tamerlane created a 100% owned subsidiary, Pine Point Holding Corp., and transferred the leases and permits into this subsidiary.

6.4 Previous Exploration Drilling in the district by Westmin, Pine Point Mines Ltd. (Cominco), and Tamerlane totaled 1,308,742 m (approximately 4,293,751 ft) in an estimated 18,406 core holes. There was some additional exploration drilling by junior companies in the 1970s, mainly on the eastern fringes of the district, and by Pine Point Mines Ltd. (Cominco) west of the district in the Hay River area.

Prior to 1963, exploration at Pine Point was accomplished by grid diamond drilling, which was used to define a historical resource estimate. Stratigraphic and structural models were developed for prioritizing the drill targets.

6.5 Previous Production Historical zinc and lead production at Pine Point is summarized in Table 6.1.

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Table 6.1: Historical Production at Pine Point

Metal Produced Ore Head Grade Deposit Production (tonnes) (tonnes) % Pb % Zn % Pb + % Zn Pb Zn NORTH TREND – listed from Northeast to Southwest X-17 44,910 1.5 6.3 7.8 674 2,829 T-37 358,960 2.1 6.3 8.4 7,538 22,614 X-51 1,203,980 2.2 6.7 8.9 26,488 80,667 Y-53 967,710 1.5 5.6 7.1 14,516 54,192 X-52 1,104,080 1.6 6.3 7.9 17,665 69,557 X-53 1,231,940 2.7 9.2 11.9 33,262 113,338 Z-53 380,520 1.4 5 6.4 5,327 19,026 A-55 1,550,830 3 7.6 10.6 46,525 117,863 Y-54 263,840 1.3 4 5.3 3,430 10,554 X-54/X-55 216,130 2.1 6.7 8.8 4,539 14,481 X-56/X-57 1,319,580 1.6 6.3 7.9 21,113 83,134 Z-57 827,870 1.1 4.2 5.3 9,107 34,771 Y-65 149,770 7 12.9 19.9 10,484 19,320 Y-60 512,490 2.1 7.3 9.4 10,762 37,412 Y-61 549,040 3.5 9.3 12.8 19,216 51,061 Z-64 913,470 1.4 5.1 6.5 12,789 46,587 A-70 2,289,360 4.5 10.4 14.9 103,021 238,093 MAIN TREND – listed from Northeast to Southwest P-24 496,640 3.5 7.6 11.1 17,382 37,745 L-30 262,170 1.1 2.8 3.9 2,884 7,341 O-28 1,483,870 2 3.7 5.7 29,677 54,903 P-29 476,120 1.6 3.3 4.9 7,618 15,712 N-31 505,200 1.6 4.1 5.7 8,083 20,713 P-31 604,760 2.2 3.6 5.8 13,305 21,771 P-32 694,980 3.2 3.5 6.7 22,239 24,324 O-32 375,970 2.8 6.4 9.2 10,527 24,062 N-32 1,862,070 3.4 8.4 11.8 63,310 156,414 L-37 3,417,550 1 3.4 4.4 34,176 116,197 N-38 1,182,110 4.9 7.4 12.3 57,923 87,476 N-42 2,959,680 5.3 9.5 14.8 156,863 281,170 O-42 2,742,720 8.8 11.6 20.4 241,359 318,156 P-41 196,140 2.1 8.3 10.4 4,119 16,280 M-40 350,870 2.2 5.5 7.7 7,719 19,298

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Metal Produced Ore Head Grade Deposit Production (tonnes) (tonnes) % Pb % Zn % Pb + % Zn Pb Zn J-44 1,282,230 5.9 9.8 15.7 75,652 125,659 I-46 389,870 5.1 4.2 9.3 19,883 16,375 M-52 455,260 3.5 7.6 11.1 15,934 34,600 K-53 468,900 3.7 9.3 13 17,349 43,608 K-57 1,564,540 6.5 5.2 11.7 101,695 81,356 K-62 1,001,590 3.6 4.8 8.4 36,057 48,076 I-65 194,510 3.8 11.1 14.9 7,391 21,591 M-64 178,460 4.9 8 12.9 8,745 14,277 R-61 1,034,540 1.6 5.2 6.8 16,553 53,796 J-69 854,770 1.2 5.2 6.4 10,257 44,448 K-77 511,120 6.4 6.4 12.8 32,712 32,712 N-81 2,699,950 7 14.1 21.1 188,997 380,693 SOUTH TREND – listed from Northeast to Southwest X-15 17,474,260 2 6.2 8.2 349,485 1,083,404 W-17 3,515,400 2 6.1 8.1 70,308 214,439 T-58 563,310 4.5 12.6 17.1 25,349 70,997 R-61 1,034,540 1.6 5.2 6.8 16,553 53,796 S-65 575,550 1.2 5.7 6.9 6,907 32,806 TOTAL 64,294,130 2,023,000 4,569,671 Source: Siega (2013)

6.5.1 List of Historical Remaining Resources

Upon the closure of the Pine Point Mine, a list of remaining deposits was compiled by Pine Point Mines Ltd. (Cominco) (given below in Table 6.2 through 6.9). These deposits were tabulated according to the mining method by which they were expected to be extracted and given the appropriate dilution. Currently, those deposits mentioned in these lists are considered as historical resources, however a large number of them have been qualified as NI 43-101 compatible reserves and resources by qualified people.

Care should be taken in examining these historical resources as some have been duplicated in tables according to the mining method and grade being considered. The N-204 deposit is a prime example of this as two separate mining scenarios are considered.

The estimates presented below as historic information and have not been verified or relied upon for economic evaluation by the Issuer or the writer. These historical mineral resources do not refer to any category of sections 1.2 and 1.3 of the NI-43-101 Instrument such as mineral resources or mineral reserves as stated in the 2010 CIM Definition Standards on Mineral Resources and Mineral

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Reserves. The authors have read the documents pertaining to the description of the different methods used in the historical evaluation of the resources/reserves.

As the authors have not done sufficient work to classify the historic estimates as current mineral resources or mineral reserves, they are of the opinion that the above quoted resources cannot be relied upon. The authors are also acting as QP’s for this report.

Table 6.2: Historical Tamerlane 2014 Open Pit Resources

Lead Grade Zinc Grade Combined Pb + Zn Deposit Total Tonnes Category % % % CP North Trend Deposits W-85 2,326,514 2.82 4.58 7.4 Measured W-85 1,125,598 1.47 3.14 4.61 Indicated X-65 2,510,448 1.45 3.65 5.1 Measured CP Main Trend Deposits J-68 265,516 2.68 5.8 8.48 Measured J-68 2780 0.63 2.34 2.97 Indicated HZ 911,600 3.1 3.65 6.75 Measured HZ 773,796 2.21 3.67 5.88 Indicated M-67 473,465 1.35 4.57 5.92 Measured M-67 210,419 0.89 5.2 6.09 Indicated K-68 262,800 1.09 3.27 4.36 Measured K-68 769,126 0.76 2.61 3.37 Indicated M-62/M-63 803,721 1.01 2.25 3.26 Measured O-53 274,812 0.83 2.71 3.54 Indicated South Trend Deposit N-204 11,740,000 0.8 2.9 3.7 Indicated Note: The historical estimates presented above are not in accordance with the mineral resources or mineral reserves classifications contained in the CIM Definition Standards on Mineral Resources and Mineral Reserves, as required by National Instrument 43-101 ("NI 43-101"). Accordingly, the Company is not treating these historical estimates as current mineral resources or mineral reserves as defined in NI 43-101 and such historical estimates should not be relied upon. A Qualified Person has not done sufficient work to date to classify the historical estimates as current mineral resources or mineral reserves. Source: Siega (2013)

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Table 6.3: Summary of Unexploited Historical Resources

Previous Owner Resource Type Tonnes Pb% Zn% WESTMIN INDICATED*--Prismatic 7,307,000 3.59 7.26

PINE POINT INDICATED*--Prismatic Open Pit 8,440,080 2.8 4.5

PINE POINT INDICATED*--Tabular Open Pit 13,858,031 1.3 4.2 PINE POINT INDICATED*--Underground 3,820,220 2.9 7.6 PINE POINT INFERRED*-- Prismatic Open Pit 5,379,700 1.1 3.2 PINE POINT INFERRED*-- Tabular Open Pit 22,995,200 0.8 2.8 PINE POINT INFERRED*-- Underground 8,962,000 1.2 4.1 GRAND TOTAL HISTORICAL RESOURCE 70,762,231 1.59 4.19 Note: *The historical estimates presented above are not in accordance with the mineral resources or mineral reserves classifications contained in the CIM Definition Standards on Mineral Resources and Mineral Reserves, as required by National Instrument 43-101 ("NI 43-101"). Accordingly, the Company is not treating these historical estimates as current mineral resources or mineral reserves as defined in NI 43-101 and such historical estimates should not be relied upon. A Qualified Person has not done sufficient work to date to classify the historical estimates as current mineral resources or mineral reserves. Source: Siega (2013) Table 6.4: Historical Indicated Prismatic Deposits

Pine Point Area, Historical Indicated* Prismatic Deposits, 20% Dilution Zone Tonnes %Pb %Zn Tonnes Pb Tonnes Zn G-03 3,444,140 3 4.1 103,324 141,210 J-68 203,670 3.9 8.1 7,943 16,497 K-51 154,500 1.1 4.4 1,700 6,798 M-48 256,360 2.1 4.1 5,384 10,511 O-53 248,700 1.9 8.1 4,725 20,145 O-556 861,000 4.3 4.3 37,023 37,023 P-499 876,000 2.88 6.5 25,229 56,940 R-67 372,150 3 9.7 11,165 36,099 R-190 1,013,000 6.3 12.1 63,819 122,573 V-46 522,000 3.01 5.5 15,712 28,710 W-19 141,000 0.44 5.9 620 8,319 W85 3,760,000 2.3 4 86,480 150,400 X-25 3,289,000 2.6 7.1 85,514 233,519 Z-155 605,000 5.5 7.2 33,275 43,560 TOTAL 15,746,520 3.1 5.8 481,913 912,303 Note: *The historical estimates presented above are not in accordance with the mineral resources or mineral reserves classifications contained in the CIM Definition Standards on Mineral Resources and Mineral Reserves, as required by National Instrument 43-101 ("NI 43-101"). Accordingly, the Company is not treating these historical estimates as current mineral resources or mineral reserves as defined in NI 43-101 and such historical estimates should not be relied upon. A Qualified Person has not done sufficient work to date to classify the historical estimates as current mineral resources or mineral reserves. Source: Siega (2013)

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Table 6.5: Historical Indicated Open Pit Tabular Deposits

Pine Point Area, Historical Tabular Deposits – Indicated* Open Pit With 30% Dilution Zone Tonnes %Pb %Zn Tonnes Pb Tonnes Zn K-32 207,450 2.9 5.2 6,016 10,787 K-35 666,280 1.3 3.9 8,662 25,985 L-27 1,845,330 0.7 3.5 12,917 64,586 L-36 1,598,840 1.5 4.1 23,982 65,552 N31ext 119,280 0.8 2.9 954 3,459 N-204 4,897,480 1.1 4.1 53,872 200,795 T-37N 71,690 2.8 3 2,007 2,079 X-49 301,710 2.2 5.9 6,638 17,801 X-52 102,470 0.8 4.2 820 4,304 X-57 86,080 0.5 3.7 430 3,185 X-61 475,060 2.8 5.1 13,302 24,228 X-65 1,273,570 1.6 4.7 20,377 59,857 Y-53 827,280 1.6 5.4 13,236 44,673 Y-55 483,880 0.8 3.3 3,871 15,968 Z-60 367,760 0.8 4.9 2,942 18,020 Z-64 533,870 0.6 3.1 3,203 16,550 TOTAL 13,858,031 1.3 4.2 173,230 577,828 Note: *The historical estimates presented above are not in accordance with the mineral resources or mineral reserves classifications contained in the CIM Definition Standards on Mineral Resources and Mineral Reserves, as required by National Instrument 43-101 ("NI 43-101"). Accordingly, the Company is not treating these historical estimates as current mineral resources or mineral reserves as defined in NI 43-101 and such historical estimates should not be relied upon. A Qualified Person has not done sufficient work to date to classify the historical estimates as current mineral resources or mineral reserves. Source: Siega (2013)

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Table 6.6: Pine Point Area Historical Indicated Underground Resources

Historical Indicated* Underground Resources 15% Dilution And 75% Extraction Zone Tonnes %Pb %Zn Tonnes Pb Tonnes Zn K-60 200,130 2.3 13.2 4,603 26,416 K-68 496,920 1.9 6.8 9,441 33,791 L-35 221,550 5.2 6.9 11,516 15,281 L-65 535,510 2.1 5.4 11,246 28,917 M-40 367,680 1.4 5.5 5,147 20,222 M-62 476,950 3 7.6 14,309 36,248 M-63 312,250 1.8 6.7 5,621 20,921 M-67 483,340 1.5 7.4 7,250 35,767 X-64 152,500 3.6 6.9 5,494 10,529 X-71 303,050 7.1 11.2 21,517 33,942 Y-62 163,690 5.6 8.1 9,167 13,260 Y-65 106,550 6.9 14.9 7,352 15,876 TOTAL 3,820,220 2.9 7.6 112,661 291,169 Note: *The historical estimates presented above are not in accordance with the mineral resources or mineral reserves classifications contained in the CIM Definition Standards on Mineral Resources and Mineral Reserves, as required by National Instrument 43-101 ("NI 43-101"). Accordingly, the Company is not treating these historical estimates as current mineral resources or mineral reserves as defined in NI 43-101 and such historical estimates should not be relied upon. A Qualified Person has not done sufficient work to date to classify the historical estimates as current mineral resources or mineral reserves. Source: Siega (2013)

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Table 6.7: Historical Inferred Open Pit Prismatic Resources

PINE POINT AREA, HISTORICAL INFERRED* RESOURCES OPEN PIT PRISMATIC-20% DILUTION ZONE TONNES %Pb %Zn K-57 648,900 1.1 2.6 K-62 232,200 0.7 2.2 M-64 108,400 1.3 2.9 N-33 67,100 0.6 1.8 N-99 814,600 1.2 3.9 O-28 74,800 1.5 1.6 O-32 177,400 0.9 2.2 P-31 103,600 1.3 2.2 T-58 385,600 1.4 4.2 W-17 49,900 1.7 5.6 X-15 2,514,800 0.9 2.9 YBM 220,000 1.7 5.9 TOTAL 5,379,700 1.1 3.2 Note: *The historical estimates presented above are not in accordance with the mineral resources or mineral reserves classifications contained in the CIM Definition Standards on Mineral Resources and Mineral Reserves, as required by National Instrument 43-101 ("NI 43-101"). Accordingly, the Company is not treating these historical estimates as current mineral resources or mineral reserves as defined in NI 43-101 and such historical estimates should not be relied upon. A Qualified Person has not done sufficient work to date to classify the historical estimates as current mineral resources or mineral reserves. Source: Siega (2013)

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Table 6.8: Pine Point Area Historical Inferred Underground Resources

Historical Inferred* Resources Underground - 15% Dilution And 75% Extraction Zone Tonnes %Pb %Zn HINGE ZONE 1,909,500 0.9 3.8 K-48 492,300 1.2 2.4 K-66N 311,600 0.8 3.3 K-66S 253,000 0.7 4.1 K-68 667,900 1.4 5.8 L-65 1,389,400 2.1 3.7 M-40 1,384,500 1.1 5.4 M-67 1,074,600 1.1 4.1 N-50 140,800 0.9 4.9 V-90 478,700 0.7 2.8 X-59N 337,800 2.1 5.4 X-68N 271,000 0.7 3.2 Z-61N 250,900 0.3 2.8 TOTAL 8,962,000 1.2 4.1 Note: *The historical estimates presented above are not in accordance with the mineral resources or mineral reserves classifications contained in the CIM Definition Standards on Mineral Resources and Mineral Reserves, as required by National Instrument 43-101 ("NI 43-101"). Accordingly, the Company is not treating these historical estimates as current mineral resources or mineral reserves as defined in NI 43-101 and such historical estimates should not be relied upon. A Qualified Person has not done sufficient work to date to classify the historical estimates as current mineral resources or mineral reserves. Source: Siega (2013)

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Table 6.9: Historical Inferred Open Pit Tabular Resources

Pine Point Area, Historical* Inferred Resources, Open Pit Tabular—30% Dilution Zone Tonnes %Pb %Zn K-32 759,900 0.5 2 K-35 1,935,800 0.5 2.9 L-30 248,400 1.1 6.1 L-36 1,222,400 1.5 2.4 N-204 16,111,300 0.8 2.5 X-51N 517,000 0.7 3.2 X-58N 690,600 0.9 3.7 Y-61N E 1,509,800 1 2.4 TOTAL 22,995,200 0.7 2.8 Note: *The historical estimates presented above are not in accordance with the mineral resources or mineral reserves classifications contained in the CIM Definition Standards on Mineral Resources and Mineral Reserves, as required by National Instrument 43-101 ("NI 43-101"). Accordingly, the Company is not treating these historical estimates as current mineral resources or mineral reserves as defined in NI 43-101 and such historical estimates should not be relied upon. A Qualified Person has not done sufficient work to date to classify the historical estimates as current mineral resources or mineral reserves. Source: Siega (2013)

6.6 Production History Pine Point Mines Ltd. (Cominco) commenced production on several identified deposits in 1964. Large-scale mining by Pine Point Mines Ltd. (Cominco) commenced in 1964 with reported reserves of 21.5 Mt averaging 4% lead and 7.2% zinc (a historical figure, not compliant with NI 43-101, reported by Giroux and McCartney, 2004). The Cominco "Pine Point Mine" was actually an assemblage of 46 separate open pits and two underground deposits, lying along a 35 km trend, which also included a mill at the Pine Point town site. No mining took place on Westmin's Great Slave Reef area, located west of Cominco's Pine Point property, which extended to about 5 km west of the Buffalo River. The deposits mined by Cominco are shown in Table 6.1.

Figure 6.1 is a graphical representation of annual ore production by Pine Point Mines Ltd. (Cominco) at Pine Point.

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Figure 6.1: Annual Ore Production at Pine Point (Mt) 1964 - 1987

Source: Siega (2013)

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7 Geological Setting and Mineralization

The discussion of geology below is largely summarized from Campbell (1957), Rhodes, et al. (1984) and Hannigan (2007), with some additions from other sources.

7.1 Regional Geology The Pine Point district area lies within the Western Canada Sedimentary Basin (see Figure 7.1 and Figure 7.2). At Pine Point, 350 to 600 m of Ordovician to Devonian sandstones, shale, carbonate rocks were deposited in a shallow sea which inundated the lowland between the outcropping Precambrian Shield to the east and mountains to the west.

Figure 7.1: Geological Provinces of Canada

Source: Adapted from Hannigan (2007)

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Figure 7.2: Middle Devonian Sedimentary Facies in West-Central Canada

Source: Modified from Rhodes et al, 1984

During Middle Devonian time, a large barrier reef complex nearly 1,000 km in length developed along a southwest-northeast basement ridge, restricting sea water circulation between the south side, (back reef) and the open sea on the north side (fore reef). As shown in Figure 7.2 , the complex was more than 10 km wide and 200 m in thickness separating a closed evaporite basin in the south (the Elk Point Basin in Alberta, Saskatchewan, Manitoba, and North Dakota) from a shallow-water open-marine environment (the Mackenzie Basin) in the north.

The restricted circulation caused evaporation of water from the back reef, resulting in increased salinity and causing the precipitation of gypsum (CaSO4.2H2O), anhydrite (CaSO4) and sodium salt (NaCl), thereby increasing the ratio of magnesium to calcium in the back reef water column. As the magnesium enriched water migrated from the back reef basin through the barrier reef, the more porous calcium carbonate rocks were altered to dolomite.

The various litho-facies deposited during Devonian time are remarkably consistent along the northeasterly strike of the reef complex, but show abrupt north-south variability across the reef from fore-reef to back-reef environments. There is abundant literature with details of the Pine Point sedimentary complex by Rhodes, et al. (1984) and Hannigan (2007). A geologic map of the rocks

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The Rhodes paper built on early pioneering work by H.Skall of Cominco who first described the Pine Point reef complex in modern geological terms (Skall, 1975).

7.2 Local Geology The Middle Devonian rocks of interest in the Pine Point district are overlain by non-mineralized Middle and Upper Devonian rocks. These in turn are overlain by Pleistocene glacial deposits and post-glacial swamp deposits, which cover nearly all the surface of the district. Outcrops are rare.

Figure 7.3 shows the facies relationships, centering on the Presqu'ile Barrier Reef in the middle, with an open marine environment to the northwest, and the Elk Point lagoon or sea of restricted circulation to the southeast. A minor uplift after Devonian time exposed the main reef to karsting, and ended the barrier-reef regime. The lead-zinc deposits shown in Figure 7.3 are subsequent to the depositional events described above.

Figure 7.3: Local Geology

Source: Modified from Rhodes et. al. (1984)

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7.3 Property Geology The Middle Devonian sedimentary rocks which host the lead-zinc deposits in the district are little- deformed. There is also extensive karstification (the dissolution of the soluble carbonate rocks), which occurred during a period of uplift of the sedimentary rocks. Folding is minor and generally associated with differential compaction of the original sediments, or karst subsidence features. Southwesterly trending lineaments in the underlying Precambrian basement have a 0.9 m/km slope to the west beneath the sedimentary rocks of the district, but these do not have visible expressions in the Devonian rocks.

7.4 Mineralization Specific formations within the Middle Devonian sedimentary sequence host the mineralization at Pine Point (see Figure 7.3 and Table 7.1). The most controversial and confusing facies issue was the status of the "Presqu'ile Formation". This coarsely-crystalline dolomite unit was formerly assigned a great stratigraphic and facies range, but was eventually recognized to be a diagenetic alteration of several different carbonate depositional facies. Thus, it has been discontinued as a stratigraphic unit, but is still referred to as a rock type (e.g. in drill logs). As discussed below, in Figure 7.4, the Presqu'ile diagenetic event is of great importance to mineralization.

Mineralization occurs in carbonate (limestone and dolomite) host rocks and is typically associated with the ingress of mineralizing brines into structurally prepared cavities associated with dissolution of the carbonate rocks by karst processes. In the case of Pine Point, this dissolution and mineralization process occurred on an extensive scale and was controlled by the distribution of a major carbonate barrier reef complex of Devonian age. A very detailed study of the geology for the Cominco Pine Point Property was compiled in a 1984 paper in Economic Geology Vol. 79 (Rhodes et al., 1984). The Rhodes paper built on early pioneering work by H. Skall of Cominco who first described the Pine Point Reef Complex in modern geological terms (Skall, 1975).

Pine Point lead-zinc mineralization was derived from a host some distance from the present position of the deposits. Circulating epithermal fluids from a possible magmatic source may have been responsible for the mineralization which occurs in the carbonate sediments as open-space cavity fillings and local replacements.

Descriptions of the mineralization below are taken from the PAH (2008) report, which in turn derived the section primarily from Hannigan (2007), Rhodes et al (1984), as well as from various Westmin and Tamerlane internal documents, and the observations of the PAH (2008) authors.

Table 7.1 indicates how the lead-zinc mineralization is limited to the specific listed stratigraphic units, with most deposits transgressing lithologic boundaries.

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Table 7.1: Mineralized Middle Devonian Units at Pine Point

Thickness Major Unit Sub-Unit Lithology Environment (Typical) Slave Point limestone, micritic (muddy) -- open-marine platform 50 m Formation shale Subsidence Watt Mountain micritic limestone, minor -- post-uplift tidal flat 20 m Formation gypsum Minor uplift, karstification, and disconformity 0-60 m Buffalo R. Shale calcareous shale fore-reef basin in north only Sulphur Point back-reef debris, dolomitized limestone 0-20 m Formation bioherms Pine Point Group Muskeg dolostones, anhydrite/gypsum, back-reef evaporitic 0-50 m Formation halite basin in south only Upper Keg River organic reef, isolate dolomitized limestones 40 m Member reefs, reef mounds B-Spongy vuggy, clay-rich dolomitized shallow water 12-18 m Member limestone Source: MineTech (2012)

Figure 7.4: Middle Devonian Sedimentary Facies

Source: Modified from Hannigan (2007)

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Shale within the Lower Keg River Member is a useful marker horizon, as it is continuous across the district. The “E” facies, consisting of un-mineralized, brown, sucrosic dolomite, underlies nearly all mineralization, and thus most drill holes terminate in this rock unit.

Figure 7.5 depicts the local subsurface geology located below the mine sites.

Figure 7.5: Sub-crop Geologic Map of Pine Point District

Source: Adapted from Hannigan (2007)

Figure 7.6 shows a cross-section through the southwest part of the Darnley Bay Property, where five deposits are located. Note the vertical exaggeration. The section shows the lateral persistence of units along strike with the Main Trend of the barrier reef, in contrast to Figure 7.7, which shows variations across the reef trend.

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Figure 7.6: Cross Section through Southwest Area of Pine Point Property

Source: Tamerlane Ventures Inc. (2011), modified from Westmin (1982)

7.4.1 Distribution of Resources

The Pine Point District contains 97 known lead-zinc mineralized areas. These are distributed along three trends designated the North, Main and South trends which extend over 68 km along strike and 6 km in width. Cominco mined 49 of these deposits during the period from 1964-1988. The distribution of these deposits is shown on Figure 7.7. The Darnley Bay Leases and Claims cover 72 of the 97 mineralized areas, including 46 of the 49 known mineralized areas in the district, which have not been mined.

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Figure 7.7: Mineralized Trends in the Pine Point District

Source: Adapted from Rhodes et al (1984)

7.4.2 Controls on Mineralization

As summarized by Hannigan (2007) mineralization occurs as sphalerite, galena, marcasite, and pyrite, replacing karst-filling sediments and breccias, open-space filling, and peripheral disseminations in vuggy porosity. The mineralization occupies the most permeable portion of the highly porous Presqu'ile barrier complex (i.e. the Pine Point and Sulphur Point units). The highest- grade mineralization seems to be restricted to karstic structures hosting the greatest amount of infilling sediments. The most typical host lithology for ore-grade mineralization is the coarse-grained vuggy Presqu'ile dolomite, which has replaced permeable reef facies.

Collapse breccia is formed by the upward sloping of the roofs of karst caverns and is seen as a diluent in many of the prismatic deposits and distinct horizons within the stratigraphy that can often be traced as distinct zones of rubble within the deposit. It is not unusual to find areas of mineralized rock having two or more collapse structures. Mineralized rock is occasionally incorporated as fragments into later stages of collapse breccia. Since these multiple collapses represented very large caverns, usually of the prismatic type, the largest mineralized zones are most frequently of this type.

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The grade of mineralization varies widely, with the highest grade zones being in open-space-filled deposits where the mineralization formed without dilution by the host rock. The deposits lie along three well-defined linear trends (North, Main, and South trends). These trends appear to coincide with southwesterly striking lineaments, forming a “hinge-line”. These Precambrian basement rocks, outcrop in the northeast.

Subtle tectonic adjustments along these lineaments contributed to the evolution of numerous paleo- environmental facies within and adjacent to the barrier complex. These lineaments are approximate projections of the underlying Precambrian McDonald-Great Slave Lake fault system.

Minor differential vertical movements on these breaks during Paleozoic times apparently created the "hinge zones" in the Devonian sediments, which controlled sedimentation (e.g. the position of the barrier reef on a slight topographic high), as well as post-Devonian alteration (karst development and Presqu'ile dolomitization) and mineralization (lead-zinc deposits). These basement lineaments may well have controlled several aspects of the mineralization:

 Creation of the carbonate reef complex as a host;  Preparation of that host by increasing the porosity of the rock through dolomitization and increased permeability through uplift & fracturing leading to karst development; and  Circulation of metalliferous fluids through a combination of emission of fluids of crustal origin and migration of metalliferous formation waters from shales though the receptive carbonates during differential movements.

All Pine Point deposits are hosted in Middle Devonian carbonate strata within or adjacent to the dolomitized barrier-reef complex. The mineralization occurs in interconnected paleo-karst cavity networks that are directly related to certain lithofacies and their behavior during episodes of uplift and karst formation. The deposits are strongly controlled by individual stratal horizons, although many deposits are discordant. The coarsely-crystalline dolomite alteration of the Presqu'ile dolomite event is also a key control for many deposits.

Epigenetic sulphide mineralization occurs as open-space fillings and local replacement of carbonate strata. The main mineralization stage at the Pine Point property witnesses a repetition of diagenetic processes, such as karst formation (i.e. solution, fracturing, collapse), dolomitization, calcite deposition, and introduction of sulphides and organic derivatives (bitumen and sulphur).

7.4.3 Mineralogy and Paragenesis

The sequence of events leading to the emplacement of mineralization at Pine Point began with karst development, and diagenetic alteration of the original rocks, which were mostly composed of calcium carbonate.

Karst development apparently occurred at several periods, beginning with the slight uplift after Devonian and before deposition of the Watt Mountain Formation. Later karst events occurred after Watt Mountain time.

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The diagenetic history began with several stages of calcite recrystallization and dolomitization adding magnesium (Mg), as shown in Figure 7.8.

Figure 7.8: Mineral Paragenesis - Pine Point Lead-Zinc District

Source: PAH (2008); adapted from Hannigan (2007)

This part of the history is typical of marine limestone in general. The conversion of calcite to dolomite by circulating fluids commonly involves a volume loss of as much as 13%, resulting in increased porosity and permeability of the rock, thus further enhancing further fluid flow.

Several pre-ore stages of dolomite are recognized, including the coarse-grained rhombic dolomite replacement which is referred to as Presqu'ile dolomite. The Presqu'ile type of dolomitization has affected 60 to 70% of the barrier-reef rocks (Hannigan, 2007). The ore-stage events resulted in deposition of sphalerite, galena, and other iron sulphides (pyrite and marcasite) in open spaces, and as a replacement of carbonates.

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Apart from minor surficial oxidation products, zinc is present only as sphalerite, which is typically coliform, i.e. as crusts and radiating microcrystalline masses of light to dark brown sphalerite lining former vugs and in massive replacements of individual beds; however, sphalerite may also be present as coarse, subhedral crystals.

Galena is the only important lead mineral, and tends to form crystalline masses nested inside sphalerite, toward the centers of former vugs (Figure 7.9) or as massive replacement of individual beds. Large, euhedral galena crystals are rare. Marcasite and/or pyrite occur in lesser amounts with occasional pyrrhotite, celestite, barite, gypsum anhydrite, and fluorite. Where lead-zinc mineralization occurs as large open space fillings, it is often very pure and finely crystalline.

There is ample evidence that the carbonate rocks of the barrier reef complex were host to oil and gas deposits prior to and/or during the formation of the lead-zinc mineralization. Tarry bitumen and brittle pyro-bitumen occur within the upper portions of the deposits, especially in areas capped by impermeable shale, such as the North Trend tabular deposits. In other areas, bitumen can be found in vugs and sometimes filling fractures in the collapse breccia. All the mineralized cores at Pine Point contain some volatile hydrocarbon compounds. Hydrogen sulphide gas was reported by Westmin in a few drill holes (e.g. Westmin, 1981, page 4).

Based on fluid-inclusion data, the ore-stage depositional temperatures were 90-100°C higher than could be accounted for by burial alone. Thus, a source of epithermal fluids is suggested, although no source of hot basinal fluids or magmatic heat has been noted in or near the district.

Post-PEM events include the deposition of minerals in remaining vugs. These minerals include early dolomite and later calcite, in addition to crystalline sulphur, bitumen, and rare fluorite. In some intervals, sulphur and bitumen may each comprise 1% of the rock volume. Native sulphur apparently formed by bacterial reduction of sulphur compounds related to the hydrocarbons.

Because most of the PEM-stage and post-PEM minerals fill open spaces or massively replace host rock, much of the porosity which existed in the pre-PEM stage due to natural deposition or early dolomitization has been eliminated in lead-zinc mineralized rock. This is visible in Figure 7.9.

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Figure 7.9: Textures of Mineralized Rock, Pine Point District

Source: PAH (2008)

7.4.4 Deposit Morphology and Size

According to a study of historical production by Hannigan (2007), the 49 lead-zinc deposits mined by Cominco averaged 1.32 Mt of lead-zinc ore, grading nearly 7.0% zinc and 3.0% lead. The deposits ranged in grades between 0.4 to 8.8% lead and 2.8 to 14.1% zinc. The iron content of individual deposits ranged from less than 0.5% to nearly 10.5%, averaging 3.5%.

Four distinct deposit morphologies have long been recognized in the district: tabular, normal prismatic, abnormal prismatic, and B-spongy. Each of these morphologies is inherited from different types of karst cavities. These types are discussed at great length by Rhodes, et al. (1984), and are summarized by Hannigan (2007).

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All six of the Pine Point Project underground deposits and W-85 are of the normal prismatic type. The other deposits are generally tabular in nature, while the abnormal prismatic and B-Spongy are developed in portions of the reef where deposits are not normally found.

7.4.4.1 Prismatic Deposits Normal prismatic deposits occur as fillings of collapse breccia pipes formed by intense karstification. These bodies are sub-circular in outline, and typically extend through 50 to 100 m of stratigraphy, of which 30 to 50 m may be mineralized. The collapse pipes tend to end at the level of tabular channels, i.e. at a paleo-aquifer. Sulphide minerals replace fine-grained post-karst infill sediments and fill cavities. To a lesser degree, sulphides may replace large breccia fragments. Normal prismatic deposits comprise most of those in the Main and South Trends, and only a few of those on the North Trend.

There is some suggestion of zoning, both within individual deposits, and within the district, although no simple pattern has emerged. According to Hannigan (2007), Pb/(Pb+Zn) ratios of Pine Point district deposits range from 0.06 to 0.56 giving a district average of 0.26. Weighted averages of lead and zinc in various deposits from the upper and lower portions of the Presqu'ile barrier indicate weak vertical metal zoning. In the upper barrier, a weighted average Pb/(Pb+Zn) ratio of 0.20 indicates that these deposits are slightly richer in zinc than deposits occurring in the lower barrier (0.26). Kyle (1981) reported a district-wide pattern of metal zoning as determined from ore-reserve data at that time. He indicated that the Pb/(Pb+Zn) ratio increases from about 0.2 in the southeast to about 0.5 in the northwest in zones parallel to the Main and North trends.

However, a weighted average calculation of the Pb/(Pb+Zn) ratio on deposits along the entire length of the Main Trend, reveals a consistent ratio of 0.36, thus indicating the lack of a district-wide lateral zoning pattern.

Deposits of the normal prismatic variety are generally lead-rich with a ratio of 0.35%. There is also a distinct lead high associated with the Main trends normal prismatic deposits whose ratio average nearly 0.37%, while normal prismatic deposits in the North and South trends whose ratio average nearly 0.25%. Abnormal prismatic deposits, present in all trends, are relatively zinc-rich whose ratio average nearly 0.25%. Some of these bodies are zoned with a lead-rich core, passing outward into a zinc-rich zone and enveloped by an iron-rich aureole. The areal distribution of iron-sulphides is wider than galena and sphalerite and the Fe/(Fe+Zn+Pb) ratio increases outward from prismatic deposits. There are relatively small barren sulphide bodies in the district, consisting mostly of iron sulphides with minor lead and zinc.

7.4.4.2 Tabular Deposits Tabular deposits occur in underground stream channels formed at the water table level at the time of karsting. The tabular deposits are characteristically long, narrow and sinuous and are often found at the base level of the prismatic deposits. The deposits can occur as open space filling of caves or as disseminated deposits within porous dolomites.

On the Main Trend, the tabular deposits tend to be low grade and disseminated through a larger volume of rock, while the North Trend tabular deposits, which are capped by the Buffalo River shale,

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The North Trend tabular and prismatic deposits are known to contain germanium in recoverable amounts, however; the main trend mineralized zones do not contain this metal. The North Trend tabular deposits tend to be much more focused due to the shale cap, than the Main Trend tabular deposits, with grades up to 35% Pb+Zn. The Z-64 deposit however, which lies slightly south of the X-65 - Y-65 high grade underground tabular deposits, was a diffuse low grade tabular deposit similar to that found on the Main Trend.

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8 Deposit Types

The Pine Point camp belongs to the class of carbonate hosted lead-zinc sulphide deposits known as Mississippi Valley Type (MVT), and is probably the most famous and best known Canadian example of this type of deposit. Nanisivik, Polaris and the Grey River deposit are other Canadian examples of the same geological class. MVT deposits are among the most important sources of lead and zinc in the world, exhibiting moderate to high grades, simple metallurgy, and easy processing and beneficiation characteristics.

Bell (1929) recognized the similarity between the Pine Point area and the lead-zinc deposits of the Tri- State area of Missouri, Kansas and Oklahoma. The characteristic features of Mississippi Valley- Type mineralization include development of karst in the carbonate host rocks, dolomitization, and the genetic association of evaporite deposits and hydrocarbons. All of these features are present in the Pine Point District (Hannigan, 2007).

8.1 Mississippi Valley-Type Deposits A concise definition of Mississippi Valley-type deposits is the following: “Mississippi Valley-type (MVT) lead-zinc deposits are a varied family of epigenetic deposits that form predominantly in dolostone and in which lead and zinc are the major commodities.” (Leach et. al, 2010).

8.2 Geological Model(s) Lead-zinc mineralized rock of the Pine Point District is hosted in paleokarst cavities within a Middle Devonian carbonate barrier complex (Figure 8.1). Karst topography is developed by the dissolution of the carbonate rocks thus creating caves and channels (Figure 8.2). These openings were enlarged as more rock was dissolved by a fluctuating water table.

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Figure 8.1: Schematic of Pine Point MVT Deposits

Source: Ginette Carter Procedures Manual, Cominco

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Figure 8.2: Formation of Karst Topography

Source: Adapted from image on www.visualphotos.com

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8.3 Deposit Nomenclature at Pine Point During the exploitation of Pine Point, a total of 49 deposits were mined totaling 64 Mt of ore. These deposits included 50 deposits that were mined as open pits and two that were mined using underground mining methods. Two main terms were used to designate the deposit types.

Prismatic deposits referring to vertical cave fill deposits that have a large vertical axis relative to their diameter.

Tabular deposits refer to underground stream filled deposits that have a long horizontal axis relative to their vertical axis.

8.4 W-85 (North Trend Cluster Pit) Open Pit Deposit The W-85 deposit is located at 60°50’54”N and 114°50’15”W at an elevation of 572 ft above sea level. The deposit is currently accessed by a 4-wheel drive exploration road which connects to the main haul road at the N-81 pit, however, the north haul road comes within about 100 m from the deposit but is not accessible by foot or vehicle due to swampy ground (Figure 8.3).

The W-85 deposit lies at the west end of the North Trend. Surface outcrops are rare in the region. Overburden consists of clayey glacial till with occasional gravel beds and areas of varved clays and cemented fine sands, and varying in thickness from 3 to 45 m (Durston 1979), generally becoming finer with depth. The glacial till forms an effective confining layer and has a shallow water table established in it (Stevenson, 1983). This material is often overlain by an organic, peaty layer of muskeg that is up to 3 m thick. Bedrock in the vicinity of the W-85 project site consists of a sequence of gently folded, west-dipping, stratified rocks, associated with a Devonian reef complex. The main stratigraphic units (from top to bottom) are as follows:

 The Slave Point Formation limestone/dolomite is a light beige dolomite or limestone which has pitted and scattered vuggy zones and contains shaly beds called the Amco Shale which is a dark grey, argillaceous, micritic limestone sequence up to 15 m thick. The Slave Point dips gently westward. The Watt Mountain Formation is bright blue green shale with interbeds of grey limestone that may be gritty, fossiliferous or micritic in nature. The Watt Mountain Formation. dips gently to the west;  The Buffalo River Formation is a grey- green shale which laps the northern face of the reef and dips and thickens to the north; and  The Pine Point Formation is the main reef which plunges gently to the southwest. It has a large variation of rock types ranging from the beige dolomitic E facies sandstone that forms the main part of the reef to coarsely fossiliferous dolomite and limestone which have localized the mineralization into trends.

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8.5 X-65 (North Trend Cluster Pit) Open Pit Deposit The X-65 Deposit is a North Trend tabular deposit located 15 km by existing roads, north of the Provincial Highway 6, which is the paved road from Pine Point to Hay River. It is 10 km north of the Cluster Pit area by existing haul roads (Figure 8.3).

Figure 8.3: Location Map of W-85 and X-65

Source: Google Earth

The mineralization is hosted in the B-Reef Substrate (BRS) facies of the Pine Point Formation (Fm) and may persist in collapse zones up to the overlying Watt Mountain Fm. The Buffalo River Fm overlaps the Pine Point Fm and is itself overlain by the Watt Mountain Fm (Figure 8.4). The Buffalo River Fm shales have formed a barrier to karsting and mineralizing fluids which intensified the mineralizing process by forcing the fluids to move laterally. The BRS facies consists of abundant white fossils encased in a dirty brown dolomite and is often extremely vuggy due to the solution of the fossil material.

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Drilling often encounters poor recoveries in the BRS although it is the main mineralized horizon. In spite of the poor recovery in some areas the reserves in prior pits on the north trend have reconciled well. This is thought to be due to the fact that when mineralization is present it cements the host formation sufficiently to allow good core recovery in mineralization. The overlying Buffalo River Fm is a grey-green shale with short limestone or dolomite layers. The Buffalo River Fm. can be very muddy and unconsolidated in karsted zones. The Watt Mountain Fm is a shaley limestone which contains a very distinctive blue-green shale that can be traced through collapse zones due to its unique color.

Figure 8.4: North Trend Section of X-65 Deposit

Source: Based upon Rhodes et al. (1975)

The North Trend geological section for the X-65 deposit is shown in Figure 8.4 and the plan view and location for X-65 relative to other North Trend deposits is shown in Figure 8.5.

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Figure 8.5: North Trend Geological Plan Including X-65 Deposit

Source: After Rhodes et al. (1975)

8.6 The (Main Trend) Cluster Pits The Cluster Pits are located along the main Pine Point haul road and hydro power from the main transformers at the old Pine Point mill site is located along the side of the haul road. The main haul road is easily accessed in the Cluster Pit area at K-57 where a gravel road connects the haul road to the main paved road from Hay River. The deposits are about 4.0 km north of Provincial Highway 6 which links the towns of Hay River and Fort Resolution. The Cluster deposits are located approximately 65 km east of Hay River.

These deposits are located on the Pine Point Main Trend and are about 45 m below the surface.

Tamerlane Ventures Inc. conducted confirmation drilling between 2007 and 2008 on M-67 and L-65 and the HZ (Table 8.1).

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Table 8.1: Summary of Drilling on the Cluster Deposits

Original Drilling (Pre-1988) Confirmatory Drilling (2007 & 2008) Deposit # of # of Company Meters Feet Company Meters Feet Holes Holes K-68 PPML 45 3,600 11,807 Tamerlane 0 0 0 M-67 PPML 68 6,018 19,740 Tamerlane 3 354 1,161 L-65 PPML 29 2,588 8490 Tamerlane 6 653 2,142 J-68 PPML 63 4,569 14993 Tamerlane 0 0 0 M-62 PPML 121 10,760 35,303 Tamerlane 0 0 0 M-63 PPML 122 11,278 37,002 Tamerlane 0 0 0 HZ PPML 183 4,461 14637 Tamerlane 5 650 2,133 Total 631 43,274 141,972 9 1,007 5,436

Source: Siega (2014)

Inclined drill holes were drilled on M-67 in a fan pattern, designed to intersect and confirm significant mineralization in three historical drill holes drilled by Pine Point Mines Ltd. (Cominco). Six drill holes were drilled on L-65, three of which were inclined drill holes drilled to confirm prior mineralization discovered by Pine Point Mines Ltd. The three other drill holes on M-67 were drilled as exploration holes to determine if mineralization was continuous between M-67 and L-65. Six holes were completed on the Hinge Zone HZ08-TV1 to HZ08-TV-08. Holes TV05 and 06 were not drilled.

A total of 653 m, 354 m, and 650 m were drilled on L-65, M-67 and HZ, respectively. The results of the drilling (M-67, L-65) suggest that both deposits are contiguous and represent a long tabular style deposit. More drilling is necessary to confirm that these deposits are connected and currently they are being assessed as separate deposits. No confirmation drilling was done on the K-68, M-62, M-63, O-53, J-68 deposits, as the original Pine Point Mines Ltd. (Cominco) or Westmin drilling has been confirmed in every deposit to date with holes twinned by Tamerlane. It is the author’s opinion that no further drilling is necessary on the known deposits to include them into the reserve category. It is recommended that any future drilling be used to examine the potential continuity between Cluster deposits.

The Cluster Pit area contains several prismatic deposits and a long semi-continuous string of tabular deposits which are located south of prismatic deposits that have been mined by Pine Point Mines Ltd. (Cominco). The prismatic deposits that have been mined include J-69, M-64, I-65, K-57 and K-62, and as a result the haul road and power line pass just north of most of the existing Cluster Pit deposits which remain (Figure 8.6). The tabular Cluster Pit deposits are typical sinuous underground stream hosted deposits. These tabular deposits weave amongst the prismatic deposits with the prismatic deposits representing vertical cave formation periodically along the length of the underground stream. The deposits are hosted by the Sulphur Point Formation which overlies the Pine Point Formation in the area. The Sulphur Point is a white limestone/dolomite which has been karsted in this area and is the main host.

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D2, which is the fossiliferous facies of the Sulphur Point (FM), forms bioherms (a circumscribed mass of rock exclusively or mainly constructed by marine sedimentary organisms), which are the focus for the karsts that host mineralization in the Cluster Pit area.

Figure 8.6: Cluster Pit Area Showing Historic Mining, Proposed Mining and Existing Haul Roads

Source: Siega and Gann (2014)

Sections of the Cluster Pits showing the mineralization with respect to the host formations are in Figure 8.7.

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Figure 8.7: Cross Section of the Main Trend Showing the Relationship of Mineralization to Geology

Source: Adapted from Rhodes et al. (1984)

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9 Exploration

9.1 Exploration History Between the 1960’s to 1980’s, major exploration programs were carried out in the Pine Point district, principally by Pine Point Mines Ltd. (Cominco) and Westmin. Due to the extensive (99 percent) glacial cover, varying up to 50 m in thickness, exploration surface methods (geologic mapping, rock geochemistry) were ineffective as there are few outcrops in the district. Three exploration methods were utilized: drilling, geophysics and geochemistry.

9.1.1 Drilling

Prior to 1963, exploration in the district was accomplished by grid drilling, followed by shaft sinking. By 1963, drilling had successfully delineated nearly 8 Mt of Potential Economic Material (PEM) averaging 2.6% lead and 5.9% zinc contained in several PEM bodies. In addition, stratigraphic and structural models had been developed for prioritizing drill targets. Drilling in the district by Westmin, Pine Point Mines Ltd. (Cominco) and Tamerlane has totaled approximately 1,308,742 m (4,293,811 ft) in 18,422 (PAH, 2011) diamond drill holes.

Drilling is discussed in Section 10 and Darnley Bay’s current database includes 18,406 drill holes totaling 4,293,751 ft or 1,308,742 m. Current 2017 drilling costs, range between $250 - $300 /m. Using the mean of $275, this represents about $358 million, that has been completed on the Pine Point Property.

9.1.2 Geophysics

The PEM at Pine Point consists of sphalerite, galena, marcasite and pyrite. This combination of minerals is not responsive to normal electromagnetic survey techniques however they do respond to induced polarization (IP) techniques, which was responsible for most of the discoveries to date. Induced polarization (IP) geophysical methods were introduced to the Pine Point district in 1963 and proved successful and increased the mineral inventory of the district.

The mineralization contains sufficiently high concentrations of the conductive minerals galena, pyrite, and marcasite to make the IP method feasible. Sphalerite is a non-conductive mineral. However, in these deposits the electrical continuity among conducting minerals is poor, making standard electromagnetic (EM) methods ineffective. The galena is often surrounded by nonconductive and non-polarized sphalerite, and abundant limestone and dolomite gangue, which interrupt the conducting paths.

Seismic surveys were successful in identifying karstic collapse structures on certain horizons of interest. At least one seismically identified brecciated collapse structure was confirmed by drilling. Gravity surveys also were able to confirm the presence of mineralization but were not widely used.

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Tamerlane completed no ground based geophysical surveying; however, in 2004 an Aeroquest Modern detailed helicopter borne TEM and Magnetic surveys were conducted over the entire property totaling 3,050 km.

The survey was conducted at 200 m spacing’s and was interpreted by Stratagex, resulting in the location of 34 anomalies of interest, which have not been followed up. Should this method prove effective, it would significantly decrease future exploration costs.

9.1.3 Geochemistry

Attempts to locate Mississippi Valley-type mineralization in the district by rock and soil geochemistry did not prove to be economically effective.

9.2 Historical Exploration Methods and Drill Core 9.2.1 Past Exploration

The Pine Point Barrier Reef, which hosts the lead-zinc deposits is extensive and plunges gently westward from Pine Point where its erosional edge is found. The reef continues southwest several hundred miles during which the deposits gradually become deeper as the reefs hosting them, plunge deeper toward the Rocky Mountains. Lead and Zinc mineralized carbonates have been intersected during the exploration for hydrocarbons at over 609.6 m (2,000 ft) in depth.

Pine Point Mines Ltd. (Cominco) restricted its primary exploration west of the mill to a zone where ore haulage to the mill was deemed economic. Beyond the N-81 deposit, which was the farthest west and last deposit to be mined, exploration was curtailed to wide spread IP surveys looking for large deposits. Within this area only one deposit, G-03, was discovered.

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10 Drilling

The early drilling by Pine Point Mines Ltd. (Cominco) utilized churn drilling, which proved to be unsatisfactory and was replaced by AX diamond drilling and subsequently by AQ sized drill core. BQ diamond drilling replaced AQ in the late 1970’s and was itself replaced by NQ coring in the 1980’s, with a significant improvement in core recovery and fewer problems caused by down-hole caving. Core recoveries varied and both AQ and BQ core had poor recoveries in karsted zones. NQ core plus the use of drill mud recovered in excess of 98% and rarely fell below 85% recovery from the early 1980s until Pine Point Mines Ltd. (Cominco) closed. The core recoveries were shown to be adequate for reserve estimation as the deposits mined by Pine Point Mines Ltd. all reconciled positively with respect to the reserve calculations based on drill core. Westmin Resources used BQ drill core. Tamerlane used NQ core exclusively.

The Pine Point Mines Ltd. (Cominco) drilling program was based on the premise that shallow deposits in the eastern part of the district were more valuable than deeply-buried deposits further west. As the mill was located in the eastern part of the district, haul distances were consequently greater for deposits located in the west. Therefore, the target size for identifying resources increased toward the west, resulting in greater drill spacing.

Westmin Resources also modeled their geologic practices on those observed at Pine Point, such as the use of vertical drilling to reflect the true thickness of the stratigraphy and avoid blocky drilling. Every collar survey record in the historic drilling was completed using vertical holes which were not down hole surveyed. This method avoided caving in the drill holes and promoted more efficient and faster drilling. For all holes with a vertical inclination, drill intercepts were assumed to represent true stratigraphic thickness, since the strata dip less than one degree towards the southwest.

The core from the Pine Point Mines Ltd. (Cominco) drilling is stored in an area called the Back 40 and can be reconstructed, less the mineralized zones, if care is used. Drilling completed by Westmin Resources was laid out on the ground, resulting in the core boxes rotting and the core being scattered on the ground and currently in very poor condition.

The drilling carried out by Tamerlane during 2010, utilized both vertical and angle holes to provide confirmation data on the deposits that it examined. Angle holes drilled proved difficult due to down- hole caving problems.

10.1 N-204 Drilling Summary 10.1.1 Tamerlane Drilling

Confirmation drilling by Tamerlane during the 2010 field work on N-204, consisted of 23 inclined cored drill holes, totaling 1,433.01 m. Hay River based ProCore Drilling, carried out the program. Drilling was carried out between February and March, 2010. Noteworthy mineralization was intersected in all 23 drill holes (Table 10.1).

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Table 10.1: Drilling Summary at N-204

Easting Northing Depth Depth Claim Claim Hole Number (NAD 83 (NAD 83 RL (m) Azimuth Dip (meters) (feet) Number Name V11) V11)

N204-10-TV2 659471.00 6760906.50 182.86 70.71 232 245 -57 F75690 N17 N204-10-TV3 659471.00 6760906.50 182.86 72.24 237 65 -55 F75690 N17 N204-10-TV4 659607.00 6760899.25 180.63 60.05 197 243 -62 F75690 N17 N204-10-TV5 659607.00 6760899.25 180.63 60.05 197 63 -64 F75690 N17 N204-10-TV6 659830.94 6760811.82 180.65 66.14 217 27 -62 F75690 N17 N204-10-TV7 659830.94 6760811.82 180.65 60.05 197 206 -53 F73123 S1 N204-10-TV8 659698.00 6760713.00 183.17 60.05 197 20 -57 F73123 S1 N204-10-TV9 659698.00 6760713.00 183.17 60.05 197 199 -55 F73123 S1 N204-10-TV10 659505.56 6760696.67 185.58 66.14 217 3 -55 F73123 S1 N204-10-TV11 659505.56 6760696.67 185.58 72.24 237 48 -66 F73123 S1 N204-10-TV12 659505.56 6760696.67 185.58 75.29 247 183 -57 F73123 S1 N204-10-TV13 659596.00 6760609.00 186.38 60.05 197 244 -52 F73123 S1 N204-10-TV14 659596.00 6760609.00 186.38 55.57 182 64 -51 F73123 S1 N204-10-TV15 659596.00 6760609.00 186.38 66.14 217 183 -58 F73123 S1 N204-10-TV16 659711.19 6760559.43 181.46 60.05 197 293 -64 F73123 S1 N204-10-TV17 659711.19 6760559.43 181.46 60.05 197 112 -63 F73123 S1 N204-10-TV18 659382.55 6760410.95 188.99 60.05 197 20 -63 F73123 S1 N204-10-TV19 659382.55 6760410.95 188.99 60.05 197 200 -64 F73123 S1 N204-10-TV20 659580.70 67603.84.02 183.77 60.05 197 36 -64 F73123 S1 N204-10-TV21 659580.70 67603.84.02 183.77 60.05 197 215 -67 F73123 S1 N204-10-TV22 659490.64 6760268.68 184.47 59.99 187 286 -61 F73123 S1 N204-10-TV23 659946.05 6760009.36 181.60 60.05 197 63 -50 F75732 N18 N204-10-TV24 660085.00 6760433.00 177.63 50.90 167 308 -55 F75732 N18

Source: Siega & Gann (2014)

All drill sites were reclaimed following the end of the drilling program. All the holes drilled were angle holes which ranged between -50 to -67° dip. Inclined drilling provided the ability to drill multiple holes in several different directions from the same drill pad, thus minimizing environmental disturbance. Water for the drilling was accessed from a borrow pit and trucked to the various sites. The drilling was designed to verify the grade and extent of the mineralization and to provide core for RQD measurements as well as metallurgical testing. Good drilling practices also provided better core recovery. Figure 10.1 shows a plan view of the drilling completed to date at N-204. Figure 10.2 shows the historic Cominco exploration holes with the Tamerlane holes and section lines overlain. Table 10.2 lists the mineralized intersections.

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Table 10.3 compares these intersections to the historic holes drilled by Pine Point Mines Ltd. (Cominco).

Tamerlane conducted a drilling program that consisted of 23 angle holes totaling 1,433.01 m (4,701 ft) which were used to:

 Provide metallurgical samples for the Dense Media Separation;  Confirm historic grades and tonnages;  Test the recoverability of core using good drilling practices; and  Provide RQD on the drill core data for the geotechnical studies.

Figure 10.1: Location of the Confirmatory Drilling Program at N-204

Source: Siega & Gann (2014)

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Figure 10.2: Tamerlane DDH Locations Overlying the Historic Cominco Drill Sites

Source: Siega & Gann (2014)

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Table 10.2: Confirmatory Drill Holes and Mineralized Intercepts

Interval (feet) Thickness Thickness Average Combined Drill Hole No. % Core Recovery From To Feet Meters (feet) %Zn %Pb %Zn + %Pb N-204-10-TV2 47 50.36 3.38 1.03 2.83 2.68 0.73 3.41 100 N-204-10-TV2 73.85 77 3.14 0.96 2.63 1.22 0.47 1.69 100 N-204-10-TV2 147 212 65 19.81 54.51 3.98 0.88 4.85 92 N-204-10-TV3 77 92 15 4.57 12.27 1.9 0.64 2.54 102 N-204-10-TV3 147 212 65 19.81 53.17 2.84 0.53 3.37 99 Containing: 147 177 30 9.14 24.54 5.44 0.97 6.41

N-204-10-TV4 57 77 20 6.1 17.7 1.5 0.55 2.05 91 N-204-10-TV4 122 187 65 19.81 57.4 1.66 0.29 1.95 96 Containing: 122 137 15 4.57 13.2 5.31 0.42 5.73

N-204-10-TV5 122 172 50 15.24 44.9 2.34 0.4 2.74 100 122 137 15 4.57 13.47 4.01 0.44 4.45

Containing: 147 157 10 3.05 8.08 3.07 0.8 3.87

162 172 10 3.05 8.08 2.3 0.45 2.81

N-204-10-TV6 37 42 5 1.52 4.41 2.65 0.64 3.29 91 N-204-10-TV6 137 177 40 12.19 35.3 3.07 0.7 3.77 100 Containing: 137 167 30 9.14 20.48 3.85 0.93 4.78

N-204-10-TV7 142 197 55 16.76 43.9 5.66 1.5 7.16 85 Containing: 147 167.57 20.57 6.27 16.42 8.81 1.97 10.78

N-204-10-TV8 122 197 75 22.86 52.85 4.43 0.9 5.33 97 Containing: 132 147 15 4.57 12.57 8.64 1.76 10.4

N-204-10-TV9 122 187 65 19.81 53.24 3.99 1.04 5.03 94 Containing: 127 148.2 21.2 6.46 17.37 7.34 1.8 9.14

N-204-10-TV10 147 172 25 7.62 20.45 4.31 0.91 5.22 100 N-204-10-TV11 132 152 20 6.1 18.28 6.45 1.78 8.23 100 N-204-10-TV12 152 237 85 25.91 71.24 7.4 2.28 9.68 63 157 187 30 9.14 25.15 11.33 3.36 14.69

197 212 15 4.57 12.57 8.84 3.05 11.89

147 167 20 6.1 15.76 3.4 1.02 4.42 92 Containing: 147 165.43 18.43 5.62 14.34 6.9 1.3 8.2 99 152 162 10 3.05 7.78 10.78 2.09 12.87

132 172 40 12.19 33.92 6.45 1.16 7.61 98 132 152 20 6.1 16.96 11.37 2.12 13.49

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Interval (feet) Thickness Thickness Average Combined Drill Hole No. % Core Recovery From To Feet Meters (feet) %Zn %Pb %Zn + %Pb 162 172 10 3.05 8.48 3.04 0.3 3.34

137 152 15 4.57 13.47 3.72 0.78 4.5 73 187 197 10 3.05 8.92 1.89 0.13 2.02

167 192 25 7.62 22.3 2.73 0.45 3.18 85 167 182 15 4.57 13.38 4.08 0.67 4.75

162 192 30 9.14 26.94 1.32 0.55 1.87 91 152 182 20 6.1 17.96 1.64 0.76 2.4

122 177 55 16.76 49.39 3.34 0.71 4.05 99 122 152 30 9.14 26.94 4.83 1.11 5.94

127 172 50 15.24 46 2.47 0.49 2.96 87 122 147 25 7.62 23 4.29 0.85 5.14

47 57 10 3.05 8.74 0.84 0.2 1.04 98 137 167 30 9.14 26.22 3.56 0.99 4.55 100 102 137 35 10.67 30.64 2.78 0.74 3.52 91 102 112 10 3.05 7.66 3.64 1.23 4.87

127 137 10 3.05 7.66 5.36 1.22 6.58

32 42 10 3.05 8.2 2.09 0.65 2.74 98 122 167 45 13.72 36.9 3.72 0.65 4.37 98 122 157 35 10.67 28.7 4.52 0.78 5.3

Source: MineTech (2011)

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Table 10.3: Comparison: The Tamerlane Confirmation Drill Holes vs. the Historic Cominco Drill Holes

Tamerlane 201- Holes Historic Cominco Holes True True Comments From From BHID To (m) %Zn+Pb %Pb %Zn Thickness Zone BHID To (m) %Zn+Pb %Pb %Zn Thickness Zone (m) (m) (m) (m) N-20410TV2 12.32 19.99 0.56 0.13 0.43 7.67 Upper N-204-346 12.19 31.39 0.39 0.15 0.24 19.2 Upper N-204-10-TV2 drilled at -57 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-346 (-90) N-20410TV2 38.32 54.93 4.86 0.88 3.98 16.61 Main N-204-346 36.27 52.12 4.1 0.62 3.48 15.85 Main N-20410TV3 16.39 22.63 1.62 0.42 1.2 6.24 Upper N-204-324 10.36 20.72 2.14 0.61 1.52 10.36 Upper N-204-10-TV3 drilled at -55 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-324 (-90). Drilled from same pad as TV2 N-20410TV3 35.67 49.4 3.76 0.59 3.17 13.73 Main N-204-324 34.74 49.07 1.2 0.35 0.86 14.33 Main N-20410TV4 15.66 22.39 1.67 0.45 1.22 6.73 Upper N-204-326 0 0 0 Upper N-204-10-TV4 drilled at -62 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-326 (-90) N-20410TV4 33.55 48.34 2.18 0.34 1.84 14.79 Main N-204-326 33.22 48.46 2.49 0.51 1.98 15.24 Main N-20410TV5 15.85 22.7 1.49 0.35 1.14 6.85 Upper N-204-353 8.83 21.34 1.17 0.5 0.67 12.5 Upper N-204-10-TV5 drilled at -64 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-353 (-90). Drilled from same pad as TV4 N-20410TV5 37.19 50.88 2.74 0.4 2.34 13.69 Main N-204-353 33.22 48.46 2.98 0.57 2.41 15.24 Main N-20410TV6 8.59 14.96 1.05 0.2 0.85 6.37 Upper N-204-038 17.68 20.12 3.25 0.85 2.4 2.44 Upper N-204-10-TV6 drilled at -62 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-038 (-90) N-20410TV6 31.07 41.82 3.77 0.7 3.07 10.75 Main N-204-038 38.4 47.24 7.32 1.27 6.05 8.84 Main N-20410TV7 9.07 11.59 7.29 1.54 5.75 2.52 Upper N-204-290 15.24 17.98 4 1.35 2.65 2.74 Upper N-204-10-TV7 drilled at -53 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-290 (-90). Drilled from same pad as TV6 N-20410TV7 33.86 47.22 7.16 1.5 5.66 13.36 Main N-204-290 35.51 46.63 10.19 2.63 7.55 11.13 Main N-20410TV8 11.23 16.4 0.67 0.16 0.51 5.17 Upper N-204-12B 14.33 17.98 1.15 0.33 0.83 3.66 Upper N-204-10-TV8 drilled at -57 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-12B (-90) N-20410TV8 29.25 44.54 6.22 1.08 5.14 15.29 Main N-204-12B 29.87 48.16 8.81 1.47 7.34 18.29 Main N-20410TV9 12.33 15.29 2.43 0.23 2.2 2.96 Upper N-204-057 12.49 16.76 1.5 0.2 1.3 4.27 Upper N-204-10-TV9 drilled at -55 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-057 (-90). Drilled from same pad as TV8 N-20410TV9 30.01 46.2 5.03 1.04 3.99 16.19 Main N-204-057 32.31 49.38 4.78 0.92 3.87 17.07 Main N-20410TV10 0 0 0 Upper N-204-148 9.45 10.97 1.3 0.5 0.8 1.52 Upper N-204-10-TV10 drilled at -55 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-148 (-90) N-20410TV10 35.02 49.96 2.54 0.46 2.09 14.94 Main N-204-148 35.36 52.73 4.5 0.93 3.57 17.37 Main N-20410TV11 0 0 0 Upper N-204-05B 0 0 0 Upper N-204-10-TV11 drilled at -66 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-05B (-90). Drilled from same pad as TV10 N-20410TV11 35.23 50.5 3.57 0.76 2.82 15.27 Main N-204-05B 34.44 60.05 3.56 0.79 2.78 25.6 Main N-20410TV12 0 0 0 Upper N-204-082 0 0 0 Upper N-204-10-TV12 drilled at -57 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-82 (-90). Drilled from same pad as TV10 N-20410TV12 38.38 60.06 9.68 2.28 7.4 21.68 Main N-204-082 39.01 60.96 6.42 1.4 5.02 21.95 Main N-20410TV13 17.14 19.81 2.33 0.24 2.1 2.67 Upper N-204-21B 0 0 0 Upper N-204-10-TV13 drilled at -52 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-21B (-90) N-20410TV13 35.46 40.25 4.42 1.02 3.4 4.79 Main N-204-21B 35.05 41.45 7.59 1.51 6.07 6.4 Main N-20410TV14 18.74 19.92 0.18 0.08 0.09 1.18 Upper N-204-22B 19.51 20.42 0.9 0.3 0.6 0.91 Upper N-204-10-TV14 drilled at -51 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-22B (-90). Drilled from same pad as TV13 N-20410TV14 33.55 37.91 8.19 1.3 6.89 4.36 Main N-204-22B 30.48 39.62 6.98 1.6 5.38 9.14 Main N-20410TV15 0 0 0 Upper N-204-065 19.2 19.51 1.1 0.6 0.5 0.3 Upper N-204-10-TV15 drilled at -58 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-65 (-90). Drilled from same pad as TV13 N-20410TV15 34.19 44.54 7.6 1.16 6.44 10.35 Main N-204-065 35.66 42.06 7.77 1.6 6.17 6.4 Main N-20410TV16 17.28 18.13 0.41 0.18 0.23 0.85 Upper N-204-374 0 0 0 Upper N-204-10-TV16 drilled at -64 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-374 (-90) N-20410TV16 36.81 40.9 4.5 0.78 3.72 4.09 Main N-204-374 40.23 47.24 7.19 1.59 5.6 7.01 Main N-20410TV17 16.41 19.12 0.35 0.24 0.12 2.71 Upper N-204-375 0 0 0 Upper N-204-10-TV17 drilled at -63 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-375 (-90). Drilled from same pad as TV16 N-20410TV17 36.32 40.07 1.52 1.26 0.26 3.75 Main N-204-375 37.19 41.15 8.61 2.26 6.35 3.96 Main N-20410TV18 0 0 0 Upper N-204-158 0 0 0 Upper N-204-10-TV18 drilled at -64 to undercut mineralization in Upper and Main Zone in

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Tamerlane 201- Holes Historic Cominco Holes True True Comments From From BHID To (m) %Zn+Pb %Pb %Zn Thickness Zone BHID To (m) %Zn+Pb %Pb %Zn Thickness Zone (m) (m) (m) (m) Cominco hole N-204-158 (-90) N-20410TV18 44 48.07 4.75 0.67 4.08 4.07 Main N-204-158 45.72 51.82 5.5 1.8 3.7 6.1 Main

N-20410TV19 0 0 0 Upper N-204-411 0 0 0 Upper N-204-10-TV19 drilled at -64 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-411 (-90). Drilled from same pad as TV18 N-20410TV19 42.92 51.13 1.87 0.55 1.32 8.21 Main N-204-411 45.11 50.92 3.66 1.23 2.43 5.18 Main N-20410TV20 16.28 20.11 0.88 0.22 0.66 3.83 Upper N-204-431 18.29 21.34 0.23 0.09 0.14 5.18 Upper N-204-10-TV20 drilled at -64 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-431 (-90) N-20410TV22 32 47.03 4.05 0.72 3.34 15.03 Main N-204-431 33.53 41.15 6.9 1.98 4.92 4.27 Main N-20410TV21 0 0 0 Upper N-204-168 14.63 16.76 0.19 0 0.19 2.13 Upper N-204-10-TV21 drilled at -67 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-168 (-90). Drilled from same pad as TV20 N-20410TV21 33.47 47.46 2.96 0.49 2.47 13.99 Main N-204-168 33.53 50.29 4.22 1.1 3.12 16.76 Main N-20410TV22 12.09 14.75 1.04 0.2 0.84 2.66 Upper N-204-161 0 0 0 Upper N-204-10-TV22 drilled at -61 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-161 (-90) N-20410TV22 35.23 43.21 4.55 0.99 3.56 7.98 Main N-204-161 36.58 45.72 4.4 0.92 3.48 9.14 Main N-20410TV23 0 0 0 Upper N-204-028 0 0 0 Upper N-204-10-TV23 drilled at -50 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-028 (-90) N-20410TV23 23.14 31.29 3.52 0.74 2.78 8.15 Main N-204-028 24.69 31.09 13.76 3.6 10.16 6.4 Main N-20410TV24 8.19 10.68 2.74 0.65 2.1 2.49 Upper N-204-278 7.92 10.52 2.2 0.3 1.9 2.59 Upper N-204-10-TV24 drilled at -55 to undercut mineralization in Upper and Main Zone in Cominco hole N-204-278 (-90) N-20410TV24 31.09 42.32 4.37 0.64 3.72 11.23 Main N-204-278 30.78 39.17 7.91 1.57 6.34 8.38 Main

Source: MineTech (2011)

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10.1.2 Pine Point Mines Ltd. (Cominco)

To increase the recovery of the core from the drill holes, Pine Point Mines Ltd. (Cominco) drilled all the exploration holes vertically, so no down-hole surveys had been completed. As the regional dip of the stratigraphy was approximately 1°/km, all intersections were therefore assumed to be of true thickness.

Both sample preparation and assaying were undertaken in the in-house Pine Point mill laboratory building. Pine Point Mines Ltd. (Cominco) used an X-ray fluorescence (XRF) technique in which capsules of pulverized sample were irradiated and the resultant fluorescent radiation was measured. This was a commonly used method during the period between 1960 and 1980. Samples out of normal range were re-assayed using different calibrations. Results were reported as % Zn, % Pb and % Fe.

10.1.3 Westmin Resources

Westmin did not drill any exploration holes on N-204.

10.1.4 Sampling Methodology & QA/QC

The logging and sampling methods described below were applied to all the exploration that was carried out on all the deposits that Tamerlane worked on in the district. The logging and sampling procedures were setup by Pine Point Mines Ltd. (Cominco). Sampling intervals were based on 1.524 m (5 ft.) lengths. These varied depending on lithology, natural breaks or style of mineralization. Generally the project geologist determined the sample lengths. These were at times problematic as recovery varied depending on the core consisting of solid intervals. In an effort to compensate for lost core, intervals consisting of vugs and dissolved sections, Tamerlane assumed that an interval with less than 100% core recovery represented a complete run. A ratio was calculated, which was based on the actual amount of recovered core, that reflected what an actual length of core represented vs the total core run.

A technician then split the core and placed one half in the core box and the other into a numbered bag and tag. A duplicate tag was stapled to the core box marking the sampled interval. Filled sample bags and transmittal forms, were then placed into 5-gallon buckets and labeled. These were then sealed and labeled for shipping to ALS Chemex in North Vancouver via Buffalo Air Express.

All samples were prepared and assayed at the lab. Samples were weighed, oven dried, then crushed to 70% <2 mm, then split and pulverized to 85% < 75 um. All the samples were assayed for zinc, lead, and iron by Chemex method AA-62.

This involved a four acid digestion and atomic-absorption analysis. All assays above 30% were re- run using titration (methods ZN-VOL50, PB-VOL50 and FE-VOL %). All were reported to the nearest 0.01 % Zn, Pb or Fe.

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10.1.5 QA/QC Check Samples

Sample checks are regularly carried out as a matter of practice for all drilling projects. Number of intervals resampled is determined before-hand and is carried out as the drilling progresses. Of the 450 samples that were originally assayed, 55 pulp samples were re-assayed comprising approximately 12% of the assayed samples. The check samples were split between the Upper and Main Zones in the proportion of 30% and 70%. These check sample pulps were shipped by Express courier to ACME Labs in Vancouver from Tamerlane’s secure office location in Hay River, NWT. Assays were run for zinc, lead and iron using the ACME 7TD method, which involved a four acid digestion and ICP-ES finish for which the detection limits for Zn, Pb and Fe are 0.01 %, 0.02 % and 0.01 % respectfully.

ACME, as a major assay lab, is ISO 9000:2000 and ISO 17025 certified in its North American facilities. Routine internal QA/QC check analysis on blacks, duplicates and standard reference materials (SRM) are carried out. Verification of the check samples revealed no discrepancies for the internal blanks and duplicate pulps. Analysis of SRM samples, were all within acceptable ranges for the values for Zn, Pb and Fe.

Figures 10.3 and 10.4 compare the percent variances between the original analyses from Tamerlane’s 2010 drilling by Chemex versus the analysis of duplicate check samples conducted by ACME Labs. For both the Upper and Main Mineralized Zones, the paired samples demonstrate a precision well within the envelope of +/- 10% for Zn, Pb and Fe.

10.1.6 Data Verification

Within the Upper Zone, variances ranged between -2.0%, -2.4% and +5.2%, respectfully. These excluded a single high value (#48) suggesting incomplete sample digestion during the analysis of the check samples (Figure 10.3).

For the Main Zone, the range of variance was between -1.4%, -2.4% and +3.7%, respectfully. These excluded an anomalous value (#24) also suggesting incomplete sample digestion during the analysis of the check samples (Figure 10.4). Excluded also was the anomalous value (#45) for lead, indicating an analytical issue in sample preparation.

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Figure 10.3: N-204 Deposit - Main Zone Percent Variance in Original vs. Check Pulp Samples (N=16)

Source: MineTech (2011)

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Figure 10.4: N-204 Deposit - Upper Zone Percent Variance in Original vs. Check Pulp Samples (N=39)

Source: MineTech (2011)

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10.1.7 QP’s Opinion

For the N-204 deposit, the logging, splitting, logistics, as well as safety of the core, were carried out in a manner utilized as a standard practice within the industry, following the regulations that were in- force at that time. Processing of these by the selected laboratory and their internal QA/QC procedures and the requirement for these laboratories to work to specific international standards, have all been met.

Check analysis confirming the assay results for the 2010 drilling program, are within acceptable limits of analytical precision. Correlation coefficients (R2) for zinc, lead, and iron are all high at 0.99, 0.94 and 0.99, respectfully. As a result, this information can be utilized for purposes of resource modeling.

It is recommended that Darnley Bay maintain continuous and systematic QA/QC procedures for all future drilling programs on all the deposits.

10.2 Cluster Pits The Cluster Pits consist of the following deposits: X-65, J-68, K-68, M62/63, M67, W-85, O-53, L-65 and HZ. Their locations are found in Figures 10.5 and 10.6. Details are described before each listing of drill coordinates in the sections below.

Figure 10.5: Cluster Pit Area Showing Historic and Proposed Mining with Existing Haul Roads

Source: Siega & Gann (2017)

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Figure 10.6: Location Map of W-85 and X-65

Source: Siega (2013)

The Cluster Pits were discovered by Pine Point Mines Ltd. (Cominco). The tables below list the coordinates of the various drill holes drilled by Cominco and re-drilled by Tamerlane (listed at the bottom of the tables as TV numbered holes) confirmation holes, as described below.

More detailed information (plan views and sections) is provided for X-65, HZ, M-67, K-68 and L-65 with the Tables that list the drill hole coordinates. The remaining Cluster Pits, O-53, M-62/63, W-85 with drill hole coordinates and down-hole surveys are provided below.

10.2.1 Tamerlane Drilling

This is a blanket statement that covers all the Cluster Pits (CP) described below. These procedures were followed through all the drilling carried out by Tamerlane.

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Between the period 2005 and 2008 Tamerlane conducted confirmatory and infill drilling on the PPP deposits to confirm historic grades and tonnages and to provide samples for metallurgical studies. The 30 completed holes totaled 5,162 m drilled.

The holes were drilled to fulfill three objectives:  Confirm historic grade and tonnage;  Provide metallurgical samples to determine the applicability of Dense Media Separation (DMS) process; and  To prove the recoverability of core with good drilling practice.

The confirmation drilling included both vertical and inclined core holes. Downhole surveys were performed in most, using a Reflex EZ-Shot Drill Hole survey device. Measurements of down-hole inclination were made every 60.9 m (200 ft) down-hole and at the end of the drill hole. Measurements varied +/- 2° from the original design and inclination at the start of the hole.

Three drilling contractors were utilized onsite. These were:

 Discovery Diamond Drilling of Morinville, Alberta  Titan Drilling of Yellowknife, NWT  ProCore Drilling, of Hay River, NWT

Tamerlane conducted confirmation drilling between 2007 and 2008 on eight deposits including M-67 and L-65. Three inclined drill holes were drilled on M-67 in a fan pattern, and were designed to intersect and confirm significant mineralization in three historical drill holes drilled by Pine Point Mines Ltd. (Cominco). Six drill holes were drilled on L-65, three of which were inclined drill holes drilled to confirm prior mineralization drilled by Pine Point Mines Ltd. (Cominco). The three other drill holes on M-67 were drilled as exploration holes to determine if mineralization was continuous between M-67 and L-65.

A total of 653 m and 354 m were drilled on L-65 and M-67 respectively. The results of the drilling suggest that both deposits are contiguous and represent continuous mineralization along a tabular style deposit. No confirmation drilling was done on the K-68 deposit.

Table 10.4 lists significant intersections drilled by Tamerlane between 2007 and 2008.

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Table 10.4: Intersections of Significance Drilled by Tamerlane between 2007 and 2008

Interval (Feet) Interval True Thickness Average Combined Drill Hole ID From To (Feet) Feet Meters %Zn %Pb (%Zn + %Pb) M6708TV1 249.60 282.40 32.80 30.49 9.29 8.14 3.08 11.21 249.60 266.00 16.40 15.25 4.65 9.04 5.16 14.20 Containing 274.90 282.40 7.50 6.97 2.13 13.53 1.99 15.52 M6708TV2 287.30 333.00 45.70 35.56 10.84 2.96 0.41 3.37 Containing 289.90 298.40 8.50 6.61 2.02 6.21 0.77 6.98 M6708TV3 316.20 359.00 42.80 31.60 9.63 5.67 0.85 6.52 Containing 347.70 359.00 11.30 8.34 2.54 10.86 1.75 12.61 L6508TV1 327.00 337.00 10.00 10.00 3.05 5.17 0.25 5.41 L6508TV1 352.20 367.00 14.80 14.80 4.51 9.51 0.94 10.45 L6508TV1 389.00 391.30 2.30 2.30 0.70 38.20 2.30 40.50 L6508TV2 347.00 359.20 12.20 8.63 2.63 6.01 2.37 8.38 L6508TV2 374.00 387.00 13.00 9.19 2.80 6.13 0.63 6.76 L6508TV3 217.00 247.00 30.00 29.11 8.87 7.75 1.08 8.82 L6508TV4 196.50 237.50 41.00 41.00 12.50 2.19 0.37 2.56 L6508TV5 226.60 242.40 15.80 15.80 4.82 6.46 1.79 8.26 L6508TV6 244.60 262.00 17.40 17.40 5.30 9.42 1.51 10.94

Source: Siega & Gann (2017)

10.2.2 Tamerlane’s Core Logging Procedures

At first, the logging procedures were found to be deficient in recording of core recoveries and geotechnical data. But the methods were corrected and consisted of geotechnical and recovery data for every 3 m interval, detailed lithologic, facies and alteration logging of the core, density measurements on selected lithologies and mineralized intervals or suspected changes in geology.

Sample intervals were usually 1.5 m, however these varied depending on natural breaks structure or mineralization. All core was photographed prior to being split for sampling.

Rock Quality Designation (RQD) techniques were completed following standard practices. These were recorded on the logging forms and then transferred to Excel spreadsheets which calculated the percent recovery and the RQD factor for the selected core intervals. Due to the brecciated nature of the mineralized rocks, average core recovery was about 85%; however, this varied between 30 to 90%, depending on the degree of alteration present and the type of lithologies intersected.

It is the opinion of the QP, Paul Gann that great improvements were made by Tamerlane in their logging and sampling procedures between 2005 and 2008, and reflect a high degree of technical competency and completeness.

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10.2.3 Pine Point Mines Ltd. (Cominco)

All drilling, assaying and sampling procedures would have been carried out on an equal basis throughout the entire period that Pine Point Mines Ltd. (Cominco) carried out exploration in this area. Please refer to the drill hole tables and previous descriptions for greater detail.

10.2.4 Geology of Cluster Pits

The K-68, M-67, and L-65 deposits lie on the mid portion of the main trend, and on the east side of Buffalo River catchment, north of Highway No. 5. The land is low-lying and poorly-drained, with a gentle north-westerly slope toward the southern shore of Great Slave Lake. Elevations in the region range from approximately 212 m in the northwest to 224 m in the southeast. Swamp, muskeg, and low gravel ridges are the main topographic features of the area. Figures 10.7 and 10.8 depict plan views of the general area for the Cluster Pits.

Figure 10.7: Deposit Location for K-68, M-67 and L-65

Source: Siega & Gann (2017)

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Figure 10.8: Location and Historical Resource* of K-68, M-67 and L-65

*The historical estimates presented above are not in accordance with the mineral resources or mineral reserves classifications contained in the CIM Definition Standards on Mineral Resources and Mineral Reserves, as required by National Instrument 43-101 ("NI 43-101"). Accordingly, Darnley Bay is not treating these historical estimates as current mineral resources or mineral reserves as defined in NI 43-101 and such historical estimates should not be relied upon. A Qualified Person has not done sufficient work to date to classify the historical estimates as current mineral resources or mineral reserves. Source: Siega & Gann (2017)

Surface outcrops are rare in the region. It was determined from the 2007 to 2008 drilling that overburden consists of clayey glacial till with occasional gravel beds and areas of varved clays and cemented fine sands, which vary in thickness from 8 to 20 m, generally becoming finer with depth. The glacial till forms an effective confining layer and has a shallow water table established within it (Stevenson, 1983). This material is often overlain by an organic, peaty layer of muskeg up to 3 m thick. Bedrock in the vicinity of the project area consists of a sequence of gently folded, west- dipping, stratified rocks, associated with a Devonian reef complex

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10.2.4.1 Deposit Type The Project area belongs to the class of carbonate hosted lead-zinc sulphide deposits known as Mississippi Valley Type (MVT) deposits, and is probably the most famous and best known Canadian example of this type of deposit. Nanisivik and Polaris are other Canadian examples of the same geological class. MVT deposits are among the most important sources of lead and zinc in the world, exhibiting moderate to high grades, simple metallurgy, and easy processing and beneficiation characteristics. They occur in carbonate (limestone and dolomite) host rocks and are typically associated with the ingress of mineralizing brines into structurally prepared cavities associated with dissolution of the carbonate rocks by karst processes. In the case of Pine Point this dissolution and mineralization process occurred on an extensive scale and was controlled by the distribution of a major carbonate barrier reef complex of Devonian age.

10.2.4.2 Mineralization Epigenetic sulphide mineralization occurs as open-space cavity fills and local replacement of internal sediments in carbonate strata. The main mineralization stage at the Pine Point property encompasses a complex overlap and repetition of diagenetic phases or processes. These include deposition of sulphides, hydrothermal dissolution, fracturing and collapse of host rocks, internal clastic sedimentation, saddle dolomite cementation, and thermal alteration of host rock and associated bitumen (P. Hannigan, 2006.). Mineralization consists of galena and sphalerite. Metal grades vary from sub economic levels on all fringes of the tabular body to as high as 19.86% combined lead and zinc.

Lead and zinc occur as fine crystals on the rims of vugs or within calcite +\- dolomite in thin veinlets on the margins of a deposit. They increase in abundance, filling vugs, larger veins, and occur in zones of massive mineralization, with skeletal to massive galena and coliform sphalerite common.

Sulphur and bitumen commonly occur in vugs, veins, and pore spaces in and around the lead-zinc mineralization. Marcasite +\- pyrite occur as disseminations within the massive mineralized sequence or as massive horizons near the base of the lead-zinc mineralization and continuing below the lead-zinc mineralization. Gypsum needles and their casts are common in association with the mineralized horizons.

10.2.4.3 Geology of the L-65 Deposit The L65 deposit lies within the Main trend of the Pine Point mineralized trends (Figures 10.7 and 10.8). Overburden thicknesses encountered in the 2007 and 2008 drill programs ranged from 8 to 20 m. No outcrop exposures were seen in the area of the deposits during these drill campaigns. The deposits are a tabular type occurring within a karst, which terminates just above the E-facies (sandy dolomite) in the Pine Point Formation. Figure 7.7 provides a visual reference to the individual mineralized trends.

Tamerlane Ventures Inc. conducted confirmation drilling between 2007 and 2008 on eight deposits including L-65. Six drill-holes were drilled on L-65, three of which were inclined drill-holes drilled to confirm prior mineralization drilled by Cominco. Three other drill-holes on M-67 were drilled as exploration holes to determine if mineralization was continuous between M-67 and L-65.

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A total of 653 m and 354 m were drilled on L-65 and M-67 respectively. The results of the drilling suggest that both deposits are contiguous and represent continuous mineralization along a tabular style deposit. The 30 completed holes totaled 5,162 m drilled.

Table 10.5 lists significant intersections drilled by Tamerlane between 2007 and 2008.

Table 10.5: Tamerlane Drilling 2007 - 2008

Interval (Feet) Interval True Thickness Average Combined Drill Hole ID (%Zn + % From To (feet) Feet Meters %Zn %Pb Pb) 249.6 282.4 32.8 30.49 9.29 8.14 3.08 11.21 M6708TV1 249.6 266 16.4 15.25 4.65 9.04 5.16 14.2 Containing 274.9 282.4 7.5 6.97 2.13 13.53 1.99 15.52 M6708TV2 287.3 333 45.7 35.56 10.84 2.96 0.41 3.37 Containing 289.9 298.4 8.5 6.61 2.02 6.21 0.77 6.98 M6708TV3 316.2 359 42.8 31.6 9.63 5.67 0.85 6.52 Containing 347.7 359 11.3 8.34 2.54 10.86 1.75 12.61 L6508TV1 327 337 10 10 3.05 5.17 0.25 5.41 L6508TV1 352.2 367 14.8 14.8 4.51 9.51 0.94 10.45 L6508TV1 389 391.3 2.3 2.3 0.7 38.2 2.3 40.5 L6508TV2 347 359.2 12.2 8.63 2.63 6.01 2.37 8.38 L6508TV2 374 387 13 9.19 2.8 6.13 0.63 6.76 L6508TV3 217 247 30 29.11 8.87 7.75 1.08 8.82 L6508TV4 196.5 237.5 41 41 12.5 2.19 0.37 2.56 L6508TV5 226.6 242.4 15.8 15.8 4.82 6.46 1.79 8.26 L6508TV6 244.6 262 17.4 17.4 5.3 9.42 1.51 10.94

Source: Siega & Gann (2017)

10.2.4.4 Geology of the K-68 & M-67 Deposits The K-68 and M-67 deposits lie within the Main trend of the Pine Point mineralized trends (Figures 10.15 and 10.16). Overburden thicknesses encountered in the 2007 and 2008 drill programs ranged from 8 to 20 m. No outcrop exposures were seen in the area of the deposits during these drill campaigns. The deposits are a tabular type occurring within a karst, which terminates just above the E-facies (sandy dolomite) in the Pine Point Formation.

Figures 10.9 and 10.10 provide plan views for M-67 and K-68. Figure 10.11 provides an East-West longitudinal section for K-68. Figure 10.12 provides an East-West longitudinal section for M-67.

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Figure 10.9: Plan View of K-68 and M-67 Deposits and Associated Drill Holes

Source: Siega & Gann (2014)

Figure 10.10: Plan View of K-68 and M-67 Deposits with Aerial Photograph

Source: Siega & Gann (2014)

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Figure 10.11: K-68 West-East Longitudinal Section

West K-68 West – East Long Section (1.0% Pb + Zn Grade Shell) East 0

50

100 Horizontal scale (m)

0 50 100 150 200 150 Slave Point Pine Point 1% Pb + Zn Overburden Formation Formation Grade Shell

Depth below surface (m) surface below Depth 200

Source: Siega & Gann (2014)

Figure 10.12: M-67 West-East Long Section (1% Pb+Zn Grade Shell)

M-67 West – East Long Section (1.0% Pb + Zn Grade Shell) West East

0

50

100 Horizontal scale (m)

0 50 100 150 200 150 Slave Point Pine Point 1% Pb + Zn Overburden Formation Formation Grade Shell

Depth below surface (m)surfacebelow Depth 200

Source: Siega & Gann (2014)

10.2.5 Darnley Bay Resources 2017 Drilling Update

Exploratory drilling commenced on March 9, 2017. To date (end of May 2017), 15 DDH’s were drilled on the property, two on W85 and 13 on L65 with the 14th in progress. Table 10.6 shows the specific details of this drilling. To-date, a total of 1,572.7 m has been drilled.

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Table 10.6: 2017 Ongoing Drilling Results

Darnley Bay Resources - Pine Point Project (Drill Program Progress)

Area Hole # Easting Northing Elevation (m) Azimuth Inclination Date Start Date End Final Depth (m) W85 W85-17-DBL-001 619645 6746311 184 0 -90 17-03-09 17-03-25 117

W85 W85-17-DBL-002 619681 6746350 176 0 -90 17-03-28 17-04-02 109.7

L65 L65-17-DBL-001 629279 6744243 220 0 -90 17-05-01 17-05-03 110

L65 L65-17-DBL-002 629320 6744219 216 180 -80 17-05-03 17-05-04 101

L65 L65-17-DBL-003 629221 6744240 202 0 -90 17-05-04 17-05-05 110

L65 L65-17-DBL-007 629252 6744208 211 0 -90 17-05-09 17-05-10 101

L65 L65-17-DBL-004 629248 6744264 205 0 -90 17-05-05 17-05-06 101

L65 L65-17-DBL-005 629280 6744295 214 0 -90 17-07-06 17-05-08 104

L65 L65-17-DBL-009 629158 6744241 0 -90 17-05-12 17-05-13 101

L65 L65-17-DBL-006 629191 6744209 215 0 -90 17-05-08 17-05-09 104

L65 L65-17-DBL-008 629160 6744177 212 0 -90 17-05-10 17-05-12 110

L65 L65-17-DBL-011 629215 6744179 0 -90 17-05-15 17-05-16 104

L65 L65-17-DBL-012 NA NA 0 -90 17-05-16 17-05-17 95

L65 L65-17-DBL-013 NA NA 0 -90 17-05-17 17-05-17 113

L65 L65-17-DBL-010 629219 6744149 0 -90 17-05-14 17-05-14 92

L65 L65-17-DBL-014 NA NA 0 -90 17-05-18

Source: Darnley Bay (2017)

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10.3 Sampling Method and Approach As a requirement for legal declaration, it must be noted that neither Darnley Bay Resources Ltd., nor Tamerlane Ventures Inc., are in anyway connected to ALS CHEMEX or ACME. Both resource companies are independent of either laboratory.

Tamerlane's core logging and sampling procedures were conducted by drilling contractors who were supervised by Tamerlane Geologists (2010 drilling program). The procedural and logging protocols followed closely those of Pine Point Mines Ltd. (Cominco) except that Tamerlane used a somewhat simplified version of Pine Point Mines Ltd.’s (Cominco) complex lithological coding system, due to the small number of Tamerlane holes and the restricted geographic/geologic range of Tamerlane's drilling program.

The drill geologist determined sample intervals, which were recorded on the drill logs and marked on the core boxes. Sample intervals were determined visually and measured to the nearest 0.03 m (0.1 ft.). Nearly all the sampled intervals contain +2% Zn+Pb. A technician split the core and placed half back in the core box, the other half being placed in a numbered bag with a sample tag.

A duplicate numbered tag was stapled in the core box to mark the sample interval. Sample bags and a transmittal form were placed in a shipping box and labeled. The box was then taped shut and labeled with the lab address and Tamerlane’s return address, plus a bar code for rapid tracking. The boxes were shipped to ALS Chemex in North Vancouver for analysis via ground or air transport.

Later, Tamerlane used the remaining half of mineralized core for metallurgical tests; thus no strong mineralization remains to be seen in the core. Tamerlane also submitted six drill core samples to Terra Testing to be tested for unconfined compressive strength and stress/strain measurements. The Tamerlane drilling was professionally executed, and should accurately reflect the nature of the mineralization at the drilled sites, such that the samples can be used for mineral resource estimation purposes.

10.3.1 Density Measurements

The term “bulk density” is used here in preference to "density" or “specific gravity” since the measurement of importance is the overall density of a mass of more or less porous mixture of mineral grains, rather than the density of any particular solid mineral grain. Both “bulk density” and “specific gravity” are measured in relation to the weight of an equivalent volume of water; i.e. tonnes/m3 (equal to gm/cm3). Because assays of mineralized material are determined in the laboratory on oven-dried samples, the result used in resource calculations must also be in dry tonnes. Most of the historical tonnage estimates in the Pine Point district have used calculated rather than measured bulk densities of mineralized rock.

10.3.2 Pine Point Mines Ltd. (Cominco) Practice

Neither Pine Point Mines Ltd.’s (Cominco) historical tonnage estimations nor those of Giroux (1982) utilized original measured rock bulk densities. Pine Point Mines Ltd.’s (Cominco) bulk densities during the operational period were calculated from the density of dolomite, adjusted by the amount

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DARNLEY BAY RESOURCES LTD. PINE POINT PEA of sphalerite, galena, and marcasite/pyrite as determined by metal assays. A porosity of 5% was assumed. Reserve-to-mined reconciliations were sufficiently satisfactory that the method was not changed.

10.3.3 Westmin Practice

According to Westmin (1983b), a Kilborn Engineering study in 1979 (Kilborn, 1979) used a tonnage factor of 9.0 cubic feet per short ton (SG = 3.55) for Westmin deposits, while Wright Engineers in 1979 calculated bulk densities for each drill hole composite, based on metal assays, and an assumed porosity of 5%. Various reports by Giroux (1982 a, 1982 b, 1982 c) used a calculated bulk density which included a porosity estimate of 5%. An internal Westmin study in 1982-1983 assumed 9.3 cubic feet per short ton (SG = 3.44). None of the Westmin resources were used by Tamerlane in their resource estimates.

10.3.4 Tamerlane Practice

Direct density measurements of core samples during Tamerlane’s drill programs were taken during the time of the reserve estimation and found to be isolated and variable, therefore, a calculated density was estimated based on knowledge of the chemistry of samples and a 5% porosity utilized, confirmed and reconciled over 25 years of mining by Pine Point Mines Ltd.

The variable sulphide content exerts a dramatic variability to density of mineralized zones. Principal sulphides are pyrite, marcasite, sphalerite and galena. The distribution of the sulphides, by nature, was precipitated through the karst hosted deposits by open space filling. The weight abundances of these sulphides can be calculated from assays from Fe, Zn and Pb providing simplifying assumptions are made. These assumptions are as follows:

 Mineralization consists of dolomite (dl), pyrite (py), sphalerite (sph) and galena (gn) with an assumed porosity of 5%;  Weight percentages of pure minerals (FeS2, ZnS, PbS and CaMg(CO3)2) are determined as follows: o Wt.% py = %Fe x 2.147 o Wt.% sph = %Zn x 1.490 o Wt.% gn = %Pb x 1.155 o Wt.% dl = 100-Porosity-Wt% py –Wt% sph – Wt% gn  Ideal densities of mineral components are: dolomite -2.85, pyrite - 5.02, sphalerite - 3.90 and galena - 7.50;  Weight percent minerals are related to volume percent by the following relation: o Vol.% mineral = Wt% Mineral x Mineralized zone density/ Mineral density.

Based on the above assumptions, the calculation used to determine the bulk density using metal grades for each deposit is as follows:

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ퟏ − 퐎 퐏 = 퐓퐨 − 퐓퐟퐩퐛 퐓퐨 − 퐓퐟퐳퐧 퐓퐨 − 퐓퐟퐟퐞 퐓퐨 − {[( ) × 퐀퐩퐛] + [( ) × 퐀퐳퐧] + [( ) × 퐀퐟퐞]} 퐙퐠퐧 퐙퐬퐩퐡 퐙퐩퐲

Where: P = Bulk Density O = Porosity (Equations) To = Host Rock Tonnage Factor A (Pb, Zn, Fe) = Assays Tf (Pb, Zn, Fe) = Tonnage Factor Z(gn, sph, py) = Wt. Conc. of mineral in its pure mineral species

The following density formula was used: ퟏ − ퟎ. ퟎퟓ 퐏 = ퟎ. ퟑퟓퟎퟗ − {[ퟎ. ퟎퟎퟐퟓ × 퐀퐩퐛] + [ퟎ. ퟎퟎퟏퟓ × 퐀퐳퐧] + [ퟎ. ퟎퟎퟑퟑ × 퐀퐟퐞]}

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11 Sample Preparation, Analyses, and Security

11.1 Cominco Samples Pine Point Mines Ltd.’s (Cominco) sample preparation and assaying were undertaken in Pine Point Mines Ltd. assay laboratory located in the Pine Point mill building. Pine Point Mines Ltd. (Cominco) used an X-ray fluorescence technique in which capsules of pulverized sample were bombarded and the resultant fluorescent radiation was measured. This method was commonly used world-wide during the 1960 to 1980 period. Samples outside a nominal range of contents for Zn, Pb, or Fe were assayed again using different machine calibrations. Results were reported as %Zn, %Pb, and %Fe.

11.2 Westmin Samples No information was found relating to sample preparation or assaying methods conducted on Westmin samples. However, Westmin was a large and respected mining company in Canada and all Tamerlane confirmation drilling did confirm the values reported by Westmin.

11.3 Tamerlane Samples, 2005 Split core samples from Tamerlane's drilling were shipped commercially to ALS Chemex in North Vancouver, B.C. There the samples were prepared and assayed. Most samples weighed between 2 kg and 5 kg as received. After oven-drying, samples were crushed to 70% < 2 mm, then split and pulverized to 85% < 75 um. Analyses were performed for Zn, Pb, and Fe by Chemex method AA-62, which involves a four-acid digestion and atomic-absorption analysis. Any Zn, Pb, or Fe values above 30% were re-determined by titration (methods ZN-VOL50, PB-VOL50 and FEVOL %), and reported to the nearest 0.01% Zn, Pb, or Fe.

ALS Chemex assay results were transmitted to CAM as secured .pdf files. ALS Chemex are a major assay lab, with ISO 9001:2000 and ISO 17025 certification of North American facilities. They routinely undertake QA/QC measures that are of world-standard class.

QA/QC procedures were strictly followed by company geologists and specified staff. For the 450 samples taken during the 2010 drilling season, 12% of the selected sample intervals were duplicated creating 55 repeated sample results. Standards and blanks were also inserted into the sample stream. All sampled core was stored in a secure facility, where access was controlled to qualified staff members. Company geologists controlled the samples right up to the delivery to the ALS Chemex sample reception, thus maintaining a strict chain-of-custody order. In the case where samples were shipped by unsupervised commercial carriers to ALS Chemex, the samples were shipped in sealed, tamper-proof containers.

The Westmin data and, to a lesser degree, the Pine Point Mines Ltd. (Cominco) data, are incompletely documented in terms of these phases. Documentation is complete for the Tamerlane samples. It is known that company employees logged and split core in all three drilling campaigns, and that Pine Point Mines Ltd. (Cominco) prepped and assayed samples in-house. Zinc and lead are the only payable minerals in these deposits, and their abundances can be readily estimated visually. The Tamerlane verification drilling was sufficient to validate the historic sampling.

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The author is not aware of any drilling, sampling, or recovery factors that could materially impact the accuracy or reliability of the results. The intersections of the drill holes with the mineralization in each deposit are shown in Chapter 8 on the Plan View and Cross Section figures (green +’s and green vertical lines, respectively).

ALS Chemex and Acme Laboratories are two labs that process field rock samples for assaying the various values of included minerals. The certificate below states the particulars for the certification of ALS Chemex.

11.4 QP’s Opinion The author is not aware of any drilling, sampling, or recovery factors that could materially impact the accuracy or reliability of the results.

The logging, splitting, logistics, as well as safety of the core, were carried out in a manner utilized as a standard practice within the industry, following the regulations that were in-force at that time. Processing of these by the selected laboratories and their internal QA/QC procedures and the requirement for these labs to work to specific international standards, have all been met.

Check analysis confirming the assay results for the 2010 drilling program, are within acceptable limits of analytical precision. Correlation coefficients (R2) for zinc, lead, and iron are all high at 0.99, 0.94 and 0.99, respectfully. As a result, this information can be utilized for purposes of resource modelling.

It is recommended that Darnley Bay maintain continuous and systematic QA/QC procedures for all future drilling programs on all the deposits.

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12 Data Verification

12.1 Cominco Drilling All Pine Point Mines Ltd. (Cominco) core was logged by geologists, or by geology students who were supervised by experienced Pine Point Mines geologists. Logging followed rigorous standard procedures, utilizing pre-printed, standardized logging cards to systematically record run-by-run recoveries, geochemical and full assays, verbose geological logs, and digital codes for geology, alteration, karst and mineralization codes. Logging protocols were provided to geologists, the last- used protocol being issued by Pine Point Mines (1985). Pine Point Mines Ltd.’s (Cominco) stratigraphic logging scheme was as shown on Figure 12.1.

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Figure 12.1: Selected Cominco Lithologic Unit Codes

Source: CAM Report (2007)

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The database for Pine Point drilling was compiled in “GEORES” format by the Pine Point Mines Ltd. (Cominco) exploration staff (GEORES was Cominco’s proprietary geological modelling software). Drill data exists for all holes drilled on the Pine Point Mines property. In the final years of operations the Pine Point Mines data were systematically updated and coded throughout the property according to the latest geological nomenclature developed from Pine Point Mines Ltd.’s (Cominco) geological research group at Pine Point. These data were subsequently acquired by Tamerlane Ventures Inc. and then by Darnley Bay.

The geologists marked the core for sampling, which was then undertaken by trained technicians and sent to the laboratory for preparation and assay. Sample intervals were determined visually and measured to the nearest 0.03 m (0.1 ft). Nearly all the sampled intervals contain +2% Zn+Pb. In the initial phases of each deposit, core was split pneumatically and 50% of the core was preserved. Later in the program, the entire mineralized interval was sent for assay, and no core remains of mineralized intercepts.

The remaining core including non-mineralized intervals was preserved in a large core library called the "Back 40" west of the (former) Pine Point mill. It is stacked in forklift-moveable cradles containing approximately 609.6 m (2,000 ft) of core per cradle. This core was left in an orderly and well-labeled condition, but in the 20+ years since drilling ceased, many of the labels have deteriorated or fallen off, and many of the wooden cradles no longer can be moved intact. This, plus the fact that the entire mineralized core was assayed in most holes, renders the historic core only marginally useful for re-assaying purposes, but is still excellent for geology purposes.

Cominco's work at Pine Point spanned 56 years (1929-1985), of which 20 years (1964-1984) involved intensive mining and exploration. There is sufficient documentation to conclude that the historic Pine Point Mines Ltd. (Cominco) drilling on the deposits can be used to generate mineral Resource estimates in accordance with NI 43-101 guidelines.

12.2 Westmin Drilling Westmin, as the operator for a joint venture of Westmin (Boliden) and DuPont, utilized a competent geological staff and industry-standard methods. However, few details of their specific procedures are known. The available information suggests that core recovery was very high, exceeding 90%. The Westmin geological logging scheme varied from that of Pine Point Mines Ltd. (Cominco) (described above).

The principal lithologic units recorded by Westmin were as shown in Figure 12.2.

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Figure 12.2: Selected Westmin Lithologic Unit Codes

Source: CAM Report (2007)

Westmin's geologic and drill data are presently housed at the C.S. Lord Geological Institute in Yellowknife (Turner, et al., 2002), and are available to the public. A complete set of the available data relating to the Westmin program was obtained by Tamerlane and it after review by a QP, it was determined that Westmin's drilling and sampling were of a quality sufficient to use together with Tamerlane's more recent drill data, for estimation of mineral Resources.

12.3 Data Verification Procedures The data verification procedure used to verify the data used for the calculation of resources for this report was the visual examination of assay sections through the deposits produced from the data

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DARNLEY BAY RESOURCES LTD. PINE POINT PEA using the Vulcan geological modelling program. Chlumsky, Armbrust and Meyer (CAM, 2007) reported that they used the following procedure to verify the data:

 Check for duplicate collars;  Check for twin holes;  Check for statistically anomalous down-hole surveys;  Check for overlapping assays;  Check for 0 length assays;  Review of assay statistics by grade class;  Review of assay statistics by length class;  Checks for holes bottomed in Potentially Economic Material (PEM);  Check for assay values successively the same;  Check for assay spikes;  Check for down-hole contamination by decay analysis; and  Check of total grade thickness by hole.

Because historic assay certificates were not available, CAM (2007) was unable to verify the database against source documents.

The assay procedures and down-hole surveys by Pine Point Mines Ltd. (Cominco) were confirmed by Tamerlane drilling. MASKWA Engineering, Hay River, confirmed UTM coordinates for CP and underground deposit drill holes by field identifiable Pine Point Mines Ltd. (Cominco) collars. Further survey verification should be conducted for feasibility level reports for the CP’S and N-204.

12.4 QP’s Opinion The QP’s have reviewed and verified the data referred to in this report. Hard copies of reports and various files are stored in a warehouse located in Nevada.

As Westmin was an exploration company that invested heavily into the exploration of the Pine Point deposits that consisted of major drilling expenditures, it would have needed to follow standard industry practices in its logging, sampling, and QA/QC procedures. It could be expected that in order to have had sufficient funds to cover its exploration costs, it would have been required to follow established methods and protocols.

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13 Mineral Processing and Metallurgical Testing

Metallurgical test programs were conducted on samples from multiple deposits in the Pine Point District. Heavy media and flotation test work indicated that standard zinc and lead flotation preceded by dense media separation will likely yield excellent recoveries. The results are summarized in the following test reports:

 Continental Metallurgical Services, “Metallurgical Testwork Report”, November 23, 2006 (Continental, 2006) – R-190 Deposit;  SGS Lakefield Research Ltd., Project No. 11574-001 “Sulphide Ore Upgrading Using Gravity Separation”, May 22, 2007 (SGS, 2007a) – R-190 Deposit;  SGS Lakefield Research Ltd., Projection No. 11633-001 “The Recovery of Lead and Zinc from Pine Point Ore Samples Using a Non-cyanide Flotation Method”, August 15, 2007 (SGS, 2007b) – R-190 Deposit;  SGS Lakefield Research Ltd., Project 11790-001 “The Recovery of Lead and Zinc Using a Non- cyanide Flotation Method on the Pine Point Ore Samples”, June 2, 2008 (SGS, 2008) – O-556 / Z-155 Deposit;  G&T Metallurgical Services, KM2984 “Metallurgical Testing on Samples from the Pine Point Project”, August 23, 2011 (G&T, 2011) – R-190 Deposit;  SGS Canada Inc., Project No. 13001-002 “Bulk Concentrate Production from Heavy Liquid Separation on a Sample from the Pine Point N-204 Deposit”, December 26, 2011 (SGS, 2011) – N-204 Deposit; and  G&T Metallurgical Services Ltd., KM3195 “Flotation Testing Dense Media Product N-204 Deposit”, January 13, 2012 (G&T, 2012) – N-204 Deposit.

The following sections are a compilation of three previous technical reports and the author’s own observations. For further reference, please consult the following technical reports:

 Pincock, Allen & Holt, “NI 43-101 Technical Report Update Pine Point Project Northwest Territories, Canada”, July 30, 2008 (Pincock, Allen & Holt, 2008);  MineTech International Ltd., “Technical Report on the R‐190, X‐25, P‐499, O‐556, Z‐155, and G‐ 03 Deposits of the Pine Point Project”, April 2, 2012 (MineTech, 2012); and  Siega, Albert & Paul Gann, “NI 43-101 Summary Technical Report Update of the Pine Point Mine Development Project Northwest Territories, Canada”, March 14, 2014 (Siega & Gann, 2014).

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Historically, in the Pine Point area, the mineralization from many deposits was processed through the same metallurgical plant with only slight variations to the process being required to achieve the targeted recoveries. Therefore, the proposed process plant should be able to treat mineralized rock from all deposits. Historical operating data and R-190 test results were used to predict the recoveries and grades expected for CP.

13.1 Metallurgical Test Work for R-190 and O-556 / Z-155 Initially, an objective of the metallurgical program for the Pine Point Project was to produce a Direct Shipment “Ore” (DSO) product with a +45% combined lead and zinc grade. It was envisioned DSO production would use a simple process flowsheet including crushing to liberate the sulphides from the non-sulphides followed by screening to remove the crushed “Fines”. DMS would then reject coarse liberated non-sulfide gangue material and produce a final, marketable DSO product. This process approach did concentrate the lead and zinc minerals with good recoveries to the DSO product, but did not attain the desired +45% combined lead plus zinc grade (see Continental, 2006).

A second phase of testing then evaluated a finer crush and secondary upgrading by gravity separation and coarse particle flotation. These process options also failed to improve the overall product grade to the desired level (see SGS, 2007a). The metallurgical test program then shifted focus to the development of a suitable flotation process following DMS (see SGS, 2007b and SGS, 2008). The material samples used throughout the metallurgical test program were obtained from diamond drill core. Information concerning the program samples is summarized in Table 13.1.

Table 13.1: Material Samples Used in the Pine Point Metallurgical Test Program

Composite Material Testing Type of Sample Purpose of Sample Name Deposit Laboratory Mineralized drill core increments from 10 drill Amenability to DMS First Continental R-190 holes assembled as high, medium and low process using Heavy liquid Composite Services grade composites separation Remaining mineralized drill core increments Second Continental Optimize pre-DMS crush R-190 from the same holes used in the first Composite Services size for DSO production composite Remainder of second composite plus R-190 Third Flotation reagent scheme R-190 drill increments from holes adjacent to mine SGS Composite development plan area DMS amenability testing. Fourth O-556 Mineralized drill core increments from 2 holes SGS Optimize pre-DMS crush Composite size Fifth Z-155 Mineralized drill core increments SGS DMS amenability testing Composite Sixth O-556 / 300 kg bulk sample of fourth and fifth Flotation flowsheet SGS Composite Z-155 composites combined development Source: Pincock, Allen & Holt (2008)

13.1.1 Heavy Liquid Separation Tests

The objective of the initial Metallurgical Test program was to evaluate the effectiveness of Heavy Liquid Separation (HLS) to remove barren gangue from the ROM material for DSO production or

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Diamond drill core from the R-190 material deposit as shown in Table 13.1 comprised the “First Composites” used for the initial testing. Selected mineralized drill core intervals from seven holes of R-190 were assembled into three composite samples with different head grades as indicated in Table 13.2.

Table 13.2: Calculated Head Grades for the R-190 DMS Composite

Composite Name Pb Grade (%) Zn Grade (%) High Grade 15.5 24.7 Medium Grade 6.2 7.5 Low Grade 4.1 5.1 Source: Continental (2006)

The three composites were each divided into three smaller lots; each lot was individually crushed to either 100% minus 1”, 5/8” or 3/8”. Each crushed sample (nine in total) was then screened to remove the -28 mesh fines that would contaminate the HLS separation media. Finally, each +28 mesh screened sample was subjected to HLS tests at specific gravity cut-points of 2.75, 2.85 and 2.95.

The “Sink” product, the “Float” rejects and the screen fines were weighed and assayed to determine the mass balance for the test. Table 13.3 presents the results for those samples crushed to minus 5/8”, showing only the 2.85 specific gravity cut point.

Table 13.3: DMS Results for R-190

Wt Assay Recovery Composite Sample Product (%) Pb (%) Zn (%) Pb (%) Zn (%) Float 2.85 6.2 0.62 2.22 0.3 0.6 Sink 2.85 85.8 16.72 26.38 92.4 91.5 High Grade -18 mesh Fines 8.0 14.3 24.6 7.3 7.9 Sinks and Fines 93.8 16.51 26.23 99.7 99.4 Float 2.85 38.4 0.51 0.45 3.7 2.3 Sink 2.85 53.1 9.25 12.14 94.7 86.3 Medium Grade -18 mesh Fines 8.5 0.95 10.1 1.6 11.4 Sinks and Fines 61.6 8.1 11.87 96.3 97.7 Float 2.85 62.1 0.37 0.54 5.6 6.5 Sink 2.85 26.9 14.15 15.77 93 82.8 Low Grade -18 mesh Fines 11.0 0.52 5 1.4 10.7 Sinks and Fines 37.9 10.21 12.64 94.4 93.5 Source: Continental (2006)

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The HLS results for the high grade composite show only a minimal upgrade with minor weight rejection to the float product. This result indicates that high grade materials may by-pass the DMS stage and report directly to the grind circuit. The medium and low grade sample results, however, indicate that 40 to 60% weight of the ROM can be rejected with only a small loss of lead and zinc content. Thus, the DMS process can be projected to save significant operating costs by rejecting much of the ROM material prior to grinding and flotation. The low grade results were used as the design basis to predict DMS recovery for the R-190 deposit. The physical separation test program continued at MSRDI utilizing the Second Composite shown in Table 13.1. This testing attempted to optimize the crush size and employ gravity separation or coarse particle flotation on the “fines” fraction to further upgrade the R-190 material to DSO quality. Although the material was upgraded significantly, the goal was not achieved (see SGS, 2007a). 13.1.2 Flotation Test Work Development

Since attempts to produce a direct shipping material product from the Pine Point deposits by gravity mineral separation methods were unsuccessful, flotation became the focus of the continuing metallurgical program. The objective was to develop a process flowsheet that assumed a DMS (or HLS for laboratory simulation) pre-concentration step followed by conventional grinding and selective lead, zinc flotation. However, unlike traditional lead/zinc flotation operations, a reagent scheme that did not utilize sodium cyanide as a depressant for zinc was desired. The flotation test program was conducted at SGS Lakefield, under the direction of Srdjan Bulatovic (see SGS, 2007b). During the first phase of development, a conventional grind/flotation flowsheet was evaluated on the third Metallurgical Composite shown in Table 13.1. This sample was composed of the remaining Second Metallurgical Composite and diamond drill core increments from the R-190 deposit (that were adjacent to the mine plan area). Batch flotation tests were done to determine the grind (particle size) and reagent scheme required to produce marketable lead concentrate and zinc concentrate. Published information on the reagent scheme and flowsheet details used by the 1978 Cominco, Pine Point Mill guided the developing test program. Since the Cominco mill relied heavily on the use of sodium cyanide, a compound with potential environmental impacts that would be hard to mitigate under present-day regulatory requirements, an array of alternative depressants were evaluated. The development testing was followed with a laboratory locked-cycle test (LCT) to confirm the batch flotation test flowsheet, concentrate grades and overall metal recoveries in the concentrates. The mass balance of the locked-cycle test, which demonstrates the very good results obtained from the flotation test program, is presented in Table 13.4. These results were used as the design basis to predict flotation recovery for the R-190 deposit. Table 13.4: Locked Cycle Flotation Test Results for R-190

Wt Assay Recovery Test No. Product (%) Pb (%) Zn (%) Pb (%) Zn (%) Pb Cleaner Concentrate 9.4 60.7 3.35 92.3 2.4 Zn Cleaner Concentrate 20.2 0.91 61.9 3 94.2 5 Zn Combined Tailings 70.4 0.42 0.64 4.7 3.4 Head (Calculated) 100 6.19 13.2 100 100 Source: SGS (2007b)

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13.1.3 Large Scale Heavy Liquid Test Work

The second phase of process development work was performed on samples from two deposits adjacent to the R-190 deposit, the O-556 and the Z-155 deposits. All the available mineralized diamond drill core increments from each deposit were combined into a composite sample for each deposit shown as the fourth and fifth composite in Table 13.1. Small-scale heavy liquid tests were performed on these two composites to assess the amenability of the two materials to DMS. Once it was verified, both deposits were amenable to DMS, the fourth and fifth composites were combined to form the sixth composite as described in Table 13.1. The head grades of the composites are shown in Table 13.5. A complete summary of results can be found SGS (2008).

Table 13.5: Head Grades for O-556 and Z-155

Composite Name Deposit Pb Grade (%) Zn Grade (%) Fe Grade (%) Fourth Composite O-556 0.84 2.47 0.79 Fifth Composite Z-155 2.12 6.59 1.2 Combined O-556 & Z- Sixth Composite 1.48 4.56 1 155 Source: SGS (2008)

The 300 kg sixth composite was crushed to minus ½” and screened on 14 mesh to remove the fines. The plus 14 mesh material was subjected to heavy liquid separation at a specific gravity of 2.9. The products of the HLS test were the -14 mesh fines, “sink” and “float” reject. The fines and sink product were recombined as feed material for flotation testing. The results of the bulk HLS test are shown in Table 13.6. The bulk HLS test confirms that the Pine Point materials are amenable to pre- concentration by HLS. More than 66% of the material was rejected as a light fraction at a specific gravity cut point of 2.9 with very little loss of metal value.

Table 13.6: DMS Test Results for O-556 / Z-155

Wt Assay Recovery Product (%) Pb (%) Zn (%) Pb (%) Zn (%) -1/2 inch +14 mesh 2.9 SG Floats 66.38 0.058 0.21 2.8 3.1 -1/2 inch +14 mesh 2.9 SG Sinks 8.18 10.25 35 60.3 63.5 -14 mesh Fines 25.45 2.02 5.93 37 33.5 Combined Sinks and Fines 33.62 4.02 13 97.2 96.9 Calculated Head 100 1.39 4.51 100 100 Source: SGS (2008)

13.1.4 Flotation Process Verification

Flotation flowsheet and reagent scheme development continued with the HLS pre-concentrated sixth composite sample. The following circuit parameters were examined in this testing:

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 Fineness of grind;  Type of secondary lead collectors;  Various depressant reagent systems; and  Different flowsheet configurations for the lead cleaner circuit.

During the course of the development test work, it was discovered that cavities within some cores of the O-556 material were filled with bituminous oil which had the undesired effect of activating the flotation of fine clay slimes. Considerable work was devoted to developing a depressant scheme to alleviate the negative effect of bitumen-activated slimes. At the conclusion of the development test program three, locked cycle tests were conducted. The main parameter examined in these tests was the lead flowsheet configuration to optimize the lead concentrate grade. The mass balance generated from the final two locked cycle tests is presented in Table 13.7.

Table 13.7: Locked Cycle Flotation Test Results for O-556 / Z-155

Wt Assay Recovery Test No. Product (%) Pb (%) Zn (%) Pb (%) Zn (%) Pb Cleaner Concentrate 4.5 72.27 1.37 90.1 0.5 Zn Cleaner Concentrate 18.7 1.05 64.07 5.4 96.3 22 Zn Combined Tailings 76.7 0.21 0.52 4.4 3.2 Calculated Head 100 3.62 12.47 100 100 Pb Cleaner Concentrate 4.3 73.76 1.62 91.1 0.6 Zn Cleaner Concentrate 19.4 0.94 59.14 5.3 96.6 23 Zn Combined Tailings 76.3 0.16 0.44 3.6 2.8 Calculated Head 100 3.45 11.85 100 100 Pb Cleaner Concentrate 4.39 73.02 1.5 90.7 0.6 Zn Cleaner Concentrate 19.03 0.99 61.61 5.3 96.4 Average 22 & 23 Zn Combined Tailings 76.58 0.18 0.48 4 3 Calculated Head 100 3.54 12.16 100 100 Source: SGS (2008) and Pincock, Allen & Holt (2008)

Table 13.8 and Table 13.9 presents the final reagent scheme developed and quantities of reagent consumed. It should be noted that only a small quantity of frother (MIBC) was used in each flotation stage as required.

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Table 13.8: Finalized Flotation Reagent Scheme Pb Circuit

Reagents (g/t of mill feed) Stage Lime P92* SQ6** Na2S PAX 3418A Grinding 1,000 1,000 250 800 - - Pb Rougher 1 - - - - 6 10 Pb Rougher 2 - - - - 6 4 Pb Rougher 3 - - - - 2 2 Pb 1st Cleaner 100 250 25 - - - Pb 1stCleaner / Scavenger 100 100 50 - - - Pb Scalping - - - - 6 2 Pb 2nd Cleaner 100 100 25 - - - Total 1,300 1,450 350 800 25 18 * P92 is composed of ZnS04 (66%), Na2S205 (17%) and Oxalix Acid (17%) ** SQ6 is composed of Acrboxymethyl Cellulose (CMC) (39%), Suspendol PKK (Cognis) (39%) and Ethylenediamine Tetra Acetic Acid (22%) Source: SGS (2008) and Pincock, Allen & Holt (2008)

Table 13.9: Finalized Flotation Reagent Scheme Zn Circuit

Reagents (grams per tonne of mill feed) Stage Lime CuSO4 3894 SQ6* PAX Conditioner 1 1,200 - - 200 - Conditioner 2 - 1,200 10 - - Conditioner 3 - - - - 20 Zn Rougher 1 - - - - - Zn Rougher 2 - - 2 - 5 Zn Rougher 3 - - 2 - 2 Zn 1st Cleaner - - - 100 - Zn 1st Cleaner / Scavenger - - 2 - 5 Zn 2nd Cleaner - - - 50 - Zn 3rd Cleaner - - - 50 - Total 1,200 1,200 16 400 32 * SQ6 is composed of Acrboxymethyl Cellulose (CMC) (39%), Suspendol PKK (Cognis) (39%) and Ethylenediamine Tetra Acetic Acid (22%) Source: SGS (2008) and Pincock, Allen & Holt (2008)

The lead and zinc concentrates produced by the final two locked cycle tests were analyzed for trace elements that might constitute smelter impurities. The result of that analysis is shown in Table 13.10. The analysis shows the concentrates are of good quality and should not be subject to smelter penalties.

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Table 13.10: Final Concentrate Impurity Analysis

Element Units Pb Concentrate Zn Concentrate Lead % 71.3 0.81 Zinc % 1.81 60 Copper % 0.013 0.038 Iron % 2.24 1.42 Arsenic % <0.001 <0.001 Antimony % <10 <10 Tin % <0.002 0.006 Bismuth g/t <20 <20 Cadmium g/t 39 930 Cobalt g/t <10 <10 Indium g/t <200 <200 Aluminum g/t 530 570 Calcium g/t 6,600 5,500 Magnesium g/t 2,700 2,600 Manganese g/t <20 45 Silica g/t 970 1,700 Titanium g/t 38 28 Gallium g/t 2.1 22 Germanium g/t <4 130 Selenium g/t <10 <10 Carbon % 2.23 0.43 Sulphur % 17.3 30.7 Chlorine g/t 88 554 Fluorine g/t <0.01 <0.01 Mercury g/t <0.3 <0.3 Gold g/t 0.06 0.03 Silver g/t <0.5 <0.5 Source: SGS (2008) and Pincock, Allen & Holt (2008)

13.1.5 Comminution and Liberation Results

One Bond ball mill work index test was performed on a composite of R-190 (see G&T, 2011). A work index of 7.4 kWh/t was recorded, indicating a soft material.

G&T (2011) also performed a Particle Mineral Analysis (PMA) (by means of A QEMSCAN) on flotation feed ground to a sizing of 70 µm P80. At this grind sizing, both galena and sphalerite had estimated two-dimensional liberations of about 70 and 80%, respectively. The unliberated galena

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13.2 Metallurgical Test Work for N-204 In 2011, heavy liquid separation (HLS) test work was conducted on N-204 composites to confirm the deposit’s amenability to DMS pre-concentration (see SGS, 2011). HLS sinks, at an SG cut point of 2.8, were then shipped to G&T metallurgical Services for flotation testing (see G&T, 2012). The following sections summarize the main results used as the design basis.

13.2.1 Sample Composition

Approximately 110 kg of drill core material from the N-204 deposit was delivered to SGS Lakefield on November 24, 2010 under sample receipt number 0317-NOV10. The original shipment contained 11 pails, nine of which were to be used in the HLS test work. The samples were combined, crushed to -13 mm and blended. Approximately 5 kg was removed and further crushed to nominal -10 mesh and a sub-sample removed for head assay. All of the remaining of -13 mm crushed material (approximately 100 kg) was screened at 20 mesh to remove the fines in advance of heavy liquid separation (SGS, 2011). The head assays for the resulting composite sample are shown in Table 13.11.

Table 13.11: N-204 Composite Head Assays

Pb (%) Zn (%) Fe (%) Cu (%) S (%) 0.92 3.67 0.93 <0.001 2.21 Source: SGS (2011)

13.2.2 DMS Test Work Results

A large-scale HLS test was completed on ~100 kg of -13 mm + 20 mesh material at a specific gravity of 2.7. The heavy liquid used was methylene iodide split with acetone to achieve the specific gravity target. Once the separation was complete, small samples were removed from each fraction and prepared for assaying (Pb, Zn, Fe). Between 97% and 98% of the Pb and Zn were recovered to the sinks. It was noted that a large proportion of the mass reported to the sinks resulting in low upgrading of the feed. A decision was made to re-pass the ~80 kg of 2.7 sinks at a higher specific gravity of 2.8. Following the separation, samples were extracted for assaying (SGS, 2011). An overall summary of the results is presented in Table 13.12. The 2.8 SG results were used as the design basis to predict DMS recovery for the N-204 deposit.

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Table 13.12: DMS Test Results for N-204

Wt Assay Recovery Product (%) Pb (%) Zn (%) Pb (%) Zn (%) SG 2.8 Sinks 22.1 3.54 12.1 82.4 83.5 SG 2.8 Floats 54 0.19 0.65 12.2 11 SG 2.7 Floats 21.2 0.1 0.33 2.5 2.2 -20 Mesh Fines 2.7 0.9 4.02 2.9 3.4 Calculated Head 100 0.84 3.2 100 100 Flotation Feed (Average) 24.8 2.9 11.2 85.3 86.9 HLS Rejects 75.2 0.16 0.56 14.7 13.1 Source: SGS (2011)

13.2.3 Flotation Test Work Results

A limited flotation testing program was conducted on Master Composite 3 (HLS sinks). The flotation testing consisted of a series of open circuit batch cleaner tests followed by a single locked cycle test (G&T, 2011). The flotation flowsheet is presented in Figure 13.1 and the results for the locked cycle test are shown in Table 13.13. These results were used as the design basis to predict flotation recovery for the N-204 deposit.

Figure 13.1: Locked Cycle Flotation Flowsheet for N-204

Source: G&T (2011)

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Table 13.13: Locked Cycle Flotation Test Results for N-204

Wt Assay Recovery Test No. Product (%) Pb (%) Zn (%) Pb (%) Zn (%) Pb Feed 100 2.65 10.9 100 100 Pre-float 1.65 5.81 9.2 3.6 1.4 Pb Concentrate 3.98 54.3 10.7 81.6 3.9 6 Zn Concentrate 17.23 1.04 55.7 6.8 88.3 Zn 1st Cleaner Tails 7.5 1.77 7.2 5 5 Zn Rougher Tails 69.65 0.11 0.23 3 1.5 Source: G&T (2011)

13.2.4 DMS Discussion

The metallurgical characteristics of the N-204 deposit differs from the Central and Southwest sections of the trend in that it is fine grained versus the coarse grained nature of the deposits southwest of N-204. The DMS test data from N-204 were recalculated to show the nature of coarse grained mineralogy versus its fine grained nature as it pertains to recovery in the DMS circuit. The results of this review showed that the DMS required capacity to produce 1,800 t of concentrator feed per day dropped from 7,200 to 6,100 t for a coarse grain N-204 based on the heavy media recoverability of coarse grained versus fine grained material. The DMS capacity can be forecasted for the CP material, which is coarse grained, based upon the grade of the deposits.

Figure 13.2 shows the graph of the R-190 to N-204 correlation and the projected tonnages for various deposits in the CP zone.

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Figure 13.2: DMS Feed Rates vs. Feed Grades

Source: Siega & Gann (2014)

13.3 Process Selection The DMS process is a key step in the beneficiation of the Pine Point deposits, as it was shown to be capable of removing 30 to 60% of the waste material while recovering greater than 87% of the Pb and Zn.

An earlier paper by Allen and Walker, which was presented at a Los Angeles AIME meeting in October of 1946, lists nine advantages of heavy media separation:

 Ability to make sharp separations at any predetermined specific gravity ranging from 1.25 to 3.4 and continuously maintains this preselected specific gravity within plus or minus 0.01. Efficient cleaning methods eliminated viscosity problems;  The specific gravity of the separating medium can be changed at any time, and within a few minutes, when necessary to meet changing characteristics of the feed;  Ability to remove sink, or refuse, continuously;  Ability to treat full-size range material without pre-sizing;  Heavy-media plants can be started and shut down without loss of values or operating efficiency;

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 The medium used for separation is relatively low cost and losses incident to the operation are negligible (100 grams to 400 g/t of feed);  Low operating and maintenance cost;  Separate units have large capacity and occupy relatively small space; and  Capital costs are low in comparison with installation of competitive processes of equal capacity” (Allen and Walker, 1949).

Flotation of the DMS sinks product in SGS (2008) was shown to have excellent recoveries of 90.7% Pb and 96.4% Zn. The process will incorporate DMS pre-concentration followed by sequential lead and zinc froth flotation. All float rejection from the DMS (waste rock) will be utilized as construction material on site to build roads, possibly sold as aggregates, or used as backfill and reintroduced back into the mining stopes once the stopes are mined.

The flotation reagent scheme at the plant will not use cyanide due to environmental commitments. An alternative reagent scheme was investigated and cyanide is no longer used as a reagent. The depressant designed to alleviate the harmful effect of pre-active clay slimes was P92 and SQ6. P92 is composed of ZnSO4 (66%), Na2S2O5 (17%), and Oxalic Acid (17%). SQ6 is composed of Carboxymethyl Cellulose (CMC) (39%), Suspendol PKK (Cognis) (39%) and Ethylenediamine Tetra Acetic Acid (22%).

As reviewed by Pincock, Allen & Holt (2008) and Siega & Gann (2014), all deposits in the Pine Point have similar geological and metallurgical characteristics. Historically, in the Pine Point area, mineralization from many deposits was processed through the same metallurgical plant with only slight variations to the process being required to achieve the required recovery rates. There is therefore a high level of confidence that the proposed process plant will be able to treat mineralized rock from all deposits. Pine Point Mines Ltd. (Cominco) mined and concentrated over 64 Mt of material at Pine Point operating through a single concentrator for the 23 years. The deposits of the CP are deposits which are primarily situated between the Pine Point Mines Ltd. (Cominco) mined out deposits.

13.4 Recovery Projections Flotation grade recovery curves for the CP material incorporate data from the SGS test program 11633-001 and historical operating information. Test work performed on R190 was used to predict the CP recoveries and grades based on the geological similarities of the material. The lead and zinc grade recovery curves are shown below in Figure 13.3.

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Figure 13.3: Lead and Zinc Grade Recovery Curves

Lead Grade Recovery Curve 98

96

94 Test 1

92 Test 2 90

88 Test 3

Lead Recovery, % 86

84 Test 4 82 40 50 60 70 80 Historical Operating Data Lead Grade, %

Zinc Grade Recovery Curve 100

95

PDC 90 Test 1

85 Test 2 Test 3 80 Zinc Recovery, % Test 4

75 Op. Data

70 40 50 60 70 Zinc Grade, %

Source: JDS, CP: 2007 SGS Project 11633-001 R-190 and operating data; N-204: 2012 G&T KM3195-06

The recovery projections for each deposit based on the metallurgical results discussed in the above sections and historical operating data are summarized in Table 13.14. These are preliminary estimates and additional test work is planned in the next stage of engineering to confirm these values.

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Table 13.14: Recovery Projections for Each Deposit

Grade Recovery Stage Pb (%) Zn (%) Pb (%) Zn (%) R-190 DMS Stage Recovery 6.91 13.95 94.4 93.5 Flotation Stage Recovery 70.0 61.9 90.0 94.2 Overall Recovery 70.0 61.9 85.0 88.1 LOM Head 2.31 4.71 100 100 N-204 DMS Stage Recovery 2.47 9.32 87.1 88.4 Flotation Stage Recovery 55.7 55.3 82.0 88.2 Overall Recovery 55.7 55.3 71.4 78.0 Calculated Head 0.7 2.6 100 100 Source: JDS (2017)

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14 Mineral Resource Estimates

14.1 National Instrument 43-101 Qualified Resources The Pine Point Barrier Reef, which hosts the Pine Point deposits, plunges gently south-west at a rate of 0.9 m/km. Deposits on the east side of the Buffalo River are considered to be shallow enough to be mined by open pit and 49 deposits have been mined by open pit in this area but a large number of potentially economic open pit deposits remain.

The main open pit deposits that Darnley Bay is focused on at this time are the N-204 located on the South Trend, the W-85 and X-65 located on the North Trend and the Main Trend deposits which include J-68, K-68, M-62/63, M-67, O-53, L-65 and HZ (Hinge Zone). In addition to these there are a number of other deposits that currently remain in the historical resources category but are not included in this current PEA mine plan. These historical resources are the R-190, X25, G-O3, P-499, O-556, and Z-155 deposits.

Confirmatory drilling was conducted at deposit N-204; the Cluster Pit deposits were modeled using Pine Point Mines historic drilling intersections and additional drilling was completed by Tamerlane at the HZ, M-67, W-85 and L-65 deposits.

14.1.1 R-190, X25, G-O3, P-499, O-556, Z-155

The numerical and geostatistical analysis of drill sampling for the deposits R-190, X25, G-O3, P-499, O-556, Z-155 and associated mineral resource estimations, were carried out by Messrs, Craig Horlacher, Stuart Collins and Barton Stone of Pincock, Allen & Holt (PAH), 2008, all qualified persons according to 43-101 regulations. Their work was reviewed and found to be adequate and is accepted by the authors; a summary is presented in Table 14.1 below. The deposits in Table 14.1 are prismatic in type and generally follow the modelling procedures for prismatics as indicated below. PAH found no capping was required, models were completed using a geological boundary corresponding to approximately 1% Pb+Zn, inverse distance, power 2 was used for estimation.

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Table 14.1: Mineral Resource Estimates, Underground Deposits: R-190, X-25, G-03, O-556 & Z-155

Pb + Zn cut-off %Zn Deposit Category Tonnes %Pb (%) Measured 3.50% 647,308 6.1 12.47

R-190 Indicated 3.50% 357,311 3.79 8.27

TOTAL 1,004,619 5.28 10.98 Measured 2.50% 182,841 4.07 7.78

P-499 Indicated 2.50% 708,811 2.32 5.38

TOTAL 891,652 2.68 5.87 Measured 2% 433,860 3.47 3.92

O-556 Indicated 2% 596,398 2.11 3.49

TOTAL 1,030,258 2.68 3.67 Measured 2.50% 1,625,380 2.61 7.21

X-25 Indicated 2.50% 482,933 1.33 5.11

TOTAL 2,108,313 2.32 6.73 Measured 2.50% 426,121 3.54 5.89

Z-155 Indicated 2.50% 349,297 1.74 3.96

TOTAL 775,418 2.73 5.02 Measured 2% 960,804 3.57 6.09

G-03 Indicated 2% 1,019,207 2.52 3.91

TOTAL 1,980,011 3.03 4.97 Measured + TOTAL 7,790,271 3.01 6.16 Indicated Source: PAH (2008)

Cut-off grades are included in the current resource tables since economic pit shells were developed at a zinc and lead price of US$1/lb, in the 2014 report: NI 43-101 Summary Technical Report Update of the Pine Point Mine Development Project, Northwest Territories, Canada March 2014.

The mineral resources listed in Table 14.1 are not included in this current PEA mine plan.

14.1.2 N-204

From MineTech 2012 (Technical Report on the Pine Point N-204 Lead Zinc Deposit), the associated mineral resource estimation was carried out by Roy and Hannon both qualified persons according to NI 43-101 regulations. The current QP, Mr. Siega, has reviewed the MineTech assumptions and calculations, related to the N-204 deposit, and concurs with the findings.

Pit shells based on the Lerch Grossman (LG) pit optimization routine were developed by JDS using the criteria listed in Table 14.3. The N-204 open pit shell provided by JDS was reviewed by Mr. Siega and he finds it reasonable for this application. Results of the analysis are summarized in Table 14.5.

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Table 14.2: Criteria - N204

Parameter Unit Value Revenue, Smelting & Refining

Lead Price US$/lb Pb 0.90 Zinc Price US$/lb Zn 1.05 Exchange Rate C$:US$ 0.75 Royalty (3%) % NSR 3.0% OPEX Estimates

OP Waste Mining Cost C$/t waste mined 2.85 OP Mill Feed Mining Cost C$/t feed mined 2.85 OP Mining Cost C$/t processed 14.25 Processing Cost (DMS and Milling) (N-204) C$/t processed 7.43

Ore haulage from DMS to Mill C$/t processed 1.57

Additional Waste Rock and Tailings Reclamation C$/t processed 0.25 G&A C$/t processed 7.13 Total OPEX (excluding Mining) (N-204) C$/t processed 16.38 Total OPEX (including Mining) (N-204) C$/t processed 30.63 Recovery and Dilution

OP External Mining Dilution % 12% OP Mining Recovery % 95% Metal Recovery (DMS & Flotation)

Lead (N-204) % 71.4% Zinc (N-204) % 78.0% Other

Overall Pit Slope Angles degrees 52 Mill Production Rate tpd 1,800 Smelter Terms and Offsite Costs

Concentrate Grades

Lead concentrate (N-204) % 55.7 Lead concentrate moisture content % 8.00 Zinc concentrate (N-204) % 55.3 Zinc concentrate moisture content % 8.00 Payables

Lead % 95.0 Zinc % 85.0

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Parameter Unit Value Unit deductions

Lead % 3.0 Zinc % 8.0 Smelting & Refining Costs

Lead concentrate smelter cost (TC) $US$/dmt 135.00 Zinc concentrate smelter cost (TC) $US$/dmt 135.00

Freight & Marketing

Lead Concentrate US$/wmt 167 Zinc Concentrate US$/wmt 154 Source: JDS (2017)

Table 14.3: Resource Estimates, N-204

Deposit Tonnes %Pb %Zn N-204 Measured 6,321,000 0.81 3.04 Indicated 3,684,000 0.79 2.88 Total M+I 10,005,000 0.80 2.98 Inferred 3,527,000 0.78 2.89 Source: JDS (2017)

14.2 Methods, Cluster Pit Deposits The Report filed by Siega & Gann March, 27th 2017 audited the work completed by Pine Point Mines Ltd (Cominco) on the Cluster Pit deposits (made up of the North and Main trends) and upgraded the historical resources to NI 43-101 Mineral Resources. Pit shells based on the LG pit optimization routine were developed by JDS using the criteria noted in Table 14.4. The Cluster Pit open pit shells provided by JDS were reviewed by Mr. Siega and he finds them reasonable for this application. Results of the analysis are summarized in Table 14.7.

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Table 14.4: Criteria - Cluster Deposits

Parameter Unit Value Revenue, Smelting & Refining Lead Price US$/lb Pb 0.90 Zinc Price US$/lb Zn 1.05 Exchange Rate C$:US$ 0.75 Royalty (3%) % NSR 3.0% OPEX Estimates OP Waste Mining Cost C$/t waste mined 2.85 OP Mill Feed Mining Cost C$/t feed mined 2.85 OP Mining Cost C$/t processed 14.25 Processing Cost (DMS and Milling) (CP) C$/t processed 9.20

Ore haulage from DMS to Mill C$/t processed 1.57

Additional Waste Rock and Tailings Reclamation C$/t processed 0.25 G&A C$/t processed 7.13 Total OPEX (excluding Mining) (CP) C$/t processed 18.15 Total OPEX (including Mining) (CP) C$/t processed 32.40 Recovery and Dilution

OP External Mining Dilution % 12% OP Mining Recovery % 95% Metal Recovery (DMS & Flotation) Lead (CP) % 85.0% Zinc (CP) % 88.1% Other Overall Pit Slope Angles degrees 52 Mill Production Rate tpd 1,800 Smelter Terms and Offsite Costs Concentrate Grades

Lead concentrate (CP) % 61.0 Lead concentrate moisture content % 8.00 Zinc concentrate (CP) % 62.0 Zinc concentrate moisture content % 8.00 Payables Lead % 95.0 Zinc % 85.0 Unit deductions Lead % 3.0

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Parameter Unit Value Zinc % 8.0 Smelting & Refining Costs Lead concentrate smelter cost (TC) $US$/DMT 135.00 Zinc concentrate smelter cost (TC) $US$/DMT 135.00

Freight & Marketing Lead Concentrate US$/WMT 167 Zinc Concentrate US$/WMT 154 Source: JDS (2017)

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Table 14.5: Resource Estimates, Cluster Pits

Deposit Tonnes %Pb %Zn Hinge Zone (HZ) Measured 825,000 3.23 3.55 Indicated 724,000 2.22 3.61 Total M+I 1,549,000 2.76 3.58 J-68 Measured 301,000 2.27 4.93 Indicated 3,000 0.56 2.09 Total M+I 304,000 2.25 4.90 K-68 Measured 300,000 0.94 2.97 Indicated 890,000 0.76 2.70 Total M+I 1,193,000 0.81 2.77 X-65 Indicated 5,430,000 0.80 2.41 Total M+I 5,430,000 0.80 2.41 Inferred 56,000 0.00 3.67 W-85 Measured 2,703,000 2.34 3.84 Indicated 1,280,000 1.23 2.65 Total M+I 3,983,000 1.98 3.46 M-67 Measured 880,000 0.92 3.33 Indicated 485,000 0.68 3.97 Total M+I 1,365,000 0.84 3.55 M-62/63 Measured 352,000 1.42 2.13 Total M+I 352,000 1.42 2.13 O-53 Indicated 81,000 0.91 2.90 Total M+I 81,000 0.91 2.90 Inferred 121,000 0.83 2.79 L-65 Measured 374,000 0.52 2.22 Indicated 1,204,000 0.76 1.87 Total M+I 1,578,000 0.70 1.95 Source: JDS (2017)

14.2.1 Density

Density was calculated on a block by block basis by the formula:

Density = (0.95)/(0.3509-((0.0025*Pb)+(0.0015*Zn)+(0.0033*Fe)))

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14.2.2 Domaining

Geological boundaries for the underground deposits were domained by developing a 1% Pb+Zn shell boundary. The CP deposits were developed by iterative grade shells. Domains utilized represent greater than 1.0% mineralization (Pb%+Zn%) with contiguity (Maptek™). Mineralized materials outside of domain boundaries were not included.

14.2.3 Compositing

Composites for modeling were one metre broken at the geological/domain boundaries. Composites for N-204 were two metres, boundary broken.

14.2.4 Capping

Basic Statistics for mineralized intersections in composite datasets tagged by Geological code SIGI for CP deposits are indicated below.

Cumulative frequency data on 1 m composite datasets are also indicated below. Caps were applied at the 95th percentile. PAH (2011) & Minetech 2012) found a 25% zinc cap and a 6.5% lead cap appropriate for N-204, their effort is judged as reasonable and not duplicated here.

For R-190, PAH “ran standard statistical check procedures on the database and found few statistically anomalous results for a database for this type of deposit. The analysis indicated the database is internally consistent. Review of the cumulative frequency plots for zinc and lead indicated that there was no need for capping” (PAH 2008).

PAH’s analysis is reasonable; and not re-evaluated here. Furthermore, the QP agrees with PAH’s decision not to apply capping as no significant difference in overall resources occur.

14.2.5 Statistical Analysis for the CP Deposits

Table 14.6 through Table 14.14 show descriptive statistics for the CP deposits. The outliers’ exclusion is efficient at the 95th percentile, i.e. grades higher than the 95th percentile are set at the 95th percentile.

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Table 14.6: Descriptive Statistics for HZ

Tonnage Pb + Zn (%) Mean 1,817,793 Mean 6.66 Standard Error 133,638 Standard Error 0.30 Median 1,635,144 Median 6.71 Mode #N/A Mode #N/A Standard Deviation 681,424 Standard Deviation 1.55 Sample Variance 4.6 E+11 Sample Variance 2.40 Kurtosis -0.44 Kurtosis -1.28 Skewness 0.77 Skewness -0.10 Range 2,295,759 Range 4.88 Minimum 1,047,868 Minimum 4.05 Maximum 3,343,627 Maximum 8.93 Sum 4,7262,612 Sum 173.22 Count 26 Count 26 Confidence Level (95.0%) 275,233 Confidence Level (95.0%) 0.63 Source: Siega & Gann (2017)

Table 14.7: Descriptive Statistics for J-68

Tonnage Pb + Zn (%) Mean 259,673 Mean 9.00 Standard Error 10,878 Standard Error 0.27 Median 251,911 Median 9.01 Mode 275,354 Mode 8.41 Standard Deviation 55,465 Standard Deviation 1.38 Sample Variance 3,076,383,165 Sample Variance 1.90 Kurtosis -0.39 Kurtosis -0.96 Skewness 0.69 Skewness -0.23 Range 192,022 Range 4.63 Minimum 188,778 Minimum 6.45 Maximum 380,800 Maximum 11.08 Sum 6,751,509 Sum 233.95 Count 26 Count 26 Confidence Level (95.0%) 22,403 Confidence Level (95.0%) 0.56 Source: Siega & Gann (2017)

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Table 14.8: Descriptive Statistics for K-68

Tonnage Pb + Zn (%) Mean 1,025,468 Mean 4.41 Standard Error 105,675 Standard Error 0.23 Median 896,121 Median 4.31 Mode #N/A Mode #N/A Standard Deviation 538,840 Standard Deviation 1.15 Sample Variance 290,348,205,265 Sample Variance 1.32 Kurtosis -0.79 Kurtosis -1.19 Skewness 0.63 Skewness 0.21 Range 1,736,040 Range 3.66 Minimum 383,073 Minimum 2.72 Maximum 2,119,113 Maximum 6.38 Sum 26,662,166 Sum 114.57 Count 26 Count 26 Confidence Level (95.0%) 217,642 Confidence Level (95.0%) 0.46 Source: Siega & Gann (2017)

Table 14.9: Descriptive Statistics for X-65

Tonnage Pb + Zn (%) Mean 3,848,664 Mean 4.63 Standard Error 332,475 Standard Error 0.20 Median 3,393,345 Median 4.67 Mode #N/A Mode #N/A Standard Deviation 1,695,298 Standard Deviation 1.01 Sample Variance 2.87 E+12 Sample Variance 1.01 Kurtosis -0.36 Kurtosis -1.15 Skewness 0.81 Skewness -0.11 Range 5,792,072 Range 3.31 Minimum 1,851,875 Minimum 2.91 Maximum 7,643,947 Maximum 6.22 Sum 100,065,262 Sum 120.31 Count 26 Count 26 Confidence Level (95.0%) 684,746 Confidence Level (95.0%) 0.41 Source: Siega & Gann (2017)

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Table 14.10: Descriptive Statistics for W-85

Tonnage Pb + Zn (%) Mean 3,416,305 Mean 6.73 Standard Error 155,363 Standard Error 0.20 Median 3,282,666 Median 6.77 Mode #N/A Mode #N/A Standard Deviation 792,198 Standard Deviation 1.03 Sample Variance 627,576,931,055 Sample Variance 1.06 Kurtosis -1 Kurtosis -1.20 Skewness 0 Skewness -0.05 Range 2,582,817 Range 3.33 Minimum 2,332,767 Minimum 5.04 Maximum 4,915,584 Maximum 8.37 Sum 88,823,927 Sum 174.91 Count 26 Count 26 Confidence Level (95.0%) 319,976 Confidence Level (95.0%) 0.42 Source: Siega & Gann (2017)

Table 14.11: Descriptive Statistics for M-67

Tonnage Pb + Zn (%) Mean 1,323,378 Mean 5.25 Standard Error 87,159 Standard Error 0.19 Median 1,222,415 Median 5.30 Mode #N/A Mode #N/A Standard Deviation 444,426 Standard Deviation 0.99 Sample Variance 1.98 E+11 Sample Variance 0.97 Kurtosis -0.73 Kurtosis -1.10 Skewness 0.61 Skewness -0.08 Range 1,495,242 Range 3.29 Minimum 741,466 Minimum 3.59 Maximum 2,236,708 Maximum 6.88 Sum 34,407,816 Sum 136.42 Count 26 Count 26 Confidence Level (95.0%) 179,507 Confidence Level (95.0%) 0.40

Source: Siega & Gann (2017)

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Table 14.12: Descriptive Statistics for M-62/63

Tonnage Pb + Zn (%) Mean 1,116,839 Mean 3.62 Standard Error 148,075 Standard Error 0.18 Median 841,681 Median 3.68 Mode #N/A Mode #N/A Standard Deviation 755,042 Standard Deviation 0.90 Sample Variance 5.70 E+11 Sample Variance 0.81 Kurtosis 0.20 Kurtosis -1.22 Skewness 1.06 Skewness -0.11 Range 2,643,258 Range 2.94 Minimum 338,270 Minimum 2.09 Maximum 2,981,528 Maximum 5.03 Sum 29,037,808 Sum 94.17 Count 26 Count 26 Confidence Level(95.0%) 304,968 Confidence Level(95.0%) 0.36 Source: Siega & Gann (2017)

Table 14.13: Descriptive Statistics for O-53

Tonnage Pb + Zn (%) Mean 247,335 Mean 4.54 Standard Error 29,674 Standard Error 0.27 Median 189,152 Median 4.57 Mode 144,129 Mode 5.25 Standard Deviation 157,019 Standard Deviation 1.42 Sample Variance 24,655,120,105 Sample Variance 2.02 Kurtosis 0.008 Kurtosis -1.13 Skewness 1.01 Skewness 0.10 Range 541,711 Range 4.69 Minimum 81,151 Minimum 2.31 Maximum 622,862 Maximum 7 Sum 6,925,392 Sum 127.16 Count 28 Count 28 Confidence Level (95.0%) 60,886 Confidence Level (95.0%) 0.55 Source: Siega & Gann (2017)

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Table 14.14: Descriptive Statistics for L-65

Tonnage Pb + Zn (%) Mean 2,594,564 Mean 3.22 Standard Error 337,650 Standard Error 0.13 Median 2,050,582 Median 3.27 Mode #N/A Mode #N/A Standard Deviation 1,721,681 Standard Deviation 0.69 Sample Variance 2,964,186,865,237 Sample Variance 0.47 Kurtosis -0.54 Kurtosis -1.24 Skewness 0.79 Skewness -0.07 Range 5,650,441 Range 2.19 Minimum 658,821 Minimum 2.1 Maximum 6,309,262 Maximum 4.29 Sum 67,458,673 Sum 83.82 Count 26 Count 26 Confidence Level (95.0%) 695,402 Confidence Level(95.0%) 0.28

Source: Siega & Gann (2017)

14.2.6 Variography

Variography was completed on composites for each deposit. Spatial continuity was verified on a deposit basis. Modeled blocks were resource categorization tagged (measured/indicated) according to best fit semi-variograms. A generally accepted practice of 33% measured and 67% indicated (inside the range of correlation) was used. Tamerlane geologists completed significant geo- statistical work confirmed by the QP.

Average categorization ranges for measured and indicated resources for tabular and prismatic deposits are shown below. Sample counts and ellipsoid search dimensions for grade estimation have been verified as reasonable by other consultants (including Sinclair & Giroux; Pincock, Allen & Holt; and MineTech), reviewed, and endorsed.

14.2.7 Tabular Deposits

14.2.7.1 Measured If the number of samples for the block estimate were greater than or equal to six and the average distance from the block center was less than or equal to 45 m, then the block was categorized as measured, the number of holes must have been greater than or equal to two, no limit was placed on the number of samples per drill hole.

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14.2.7.2 Indicated If the number of samples for the block estimate is greater than or equal to four and the average distance from the block center is less than or equal to 90 m, then the block is categorized as indicated, the number of holes must be greater than or equal to two, no limit is placed on the number of samples per drill hole.

Estimation was via Vulcan (©Maptek) inverse distance power 2, Ellipsoid search, Ellipsoid major axis orientation is parallel to mineralization strike.

14.2.8 Prismatic type Deposits

14.2.8.1 Measured If the number of samples for the block estimate was greater than or equal to 6 and the average distance from the block center to the samples was less than or equal to 15 m then the block was categorized as measured, the number of holes must have been greater than or equal to two, a limit of two samples per drill holes was placed on the estimation.

14.2.8.2 Indicated If the number of samples for the block estimate was greater than or equal to four and the average distance from the block center to the samples was less than or equal to 30 m, then the block was categorized as indicated, the number of holes must have been greater than or equal to two, a limit of two samples per drill holes was placed on the estimation.

Estimation was via Vulcan (©Maptek) inverse distance power 2, Ellipsoid search. Samples for both deposit types were composites, broken at geological boundaries, reviewed and endorsed by PAH.

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For Deposit W-85, a double pass estimation was run to more closely represent the high grade zones.

First pass:

Second Pass:

For W-85, resource tags were as indicated above for prismatics.

Confirmation of Pine Point Mines Ltd. (Cominco) drill holes has been carried out in four of the nine CP deposits. The author believes that the Main Trend mineralization and the North Trend mineralization are basically continuous deposits whose origin and deposition occurred in Devonian time on a continuous event basis. Verification drilling completed on four deposits confirmed geology, mineralization types and grades as established by Pine Point Mines Ltd. (Cominco). Additional discussion regarding site geology is in Sections 7 and 8. In total, including the underground deposits, nine individual deposits have been tested with confirmation holes. In most cases confirmation drilling correlated with proximal historical drill intercepts. Resources calculated via block modelling, were typically more conservative than Cominco Historical estimates for the deposits considered herein.

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15 Mineral Reserve Estimates

This PEA report does not state a Mineral Reserve.

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16 Mining Methods

16.1 Open Pit Mine Design Open Pit mine design was conducted using a combination of software packages, including Hexagon MineSight™ and Datamine NPV Scheduler.

The resultant 10 open pit shapes for the Project will contain 29.5 Mt of mineralized material with an average lead grade of 1.08% and average zinc grade of 2.93%. The total waste tonnage will be 117.2 Mt for a strip ratio of 4.0:1. The majority of the pits are also referred to as the Cluster pits (CP).

Table 16.1 summarizes the combined tonnages and grades contained within the planned ultimate pit shells for the 10 deposits (using the incremental cut-off value of $18.15/t for the Cluster pits and $16.38/t for the N-204 deposit).

Table 16.1: Pit Shell Summary

Description Unit Value Total Process Feed Material Mt 29.5 Total Diluted Lead Grade % 1.08 Total Diluted Zinc Grade % 2.93 Total Contained Lead Mlbs 703 Total Contained Zinc Mlbs 1,910 Total Waste Mt 117.2 Total Material Mt 146.7 Overall Strip Ratio w:o 4.0 Note: *diluted tonnes and grade @ 12% external dilution and 95% mining recovery Source: JDS (2017)

16.2 Optimization 16.2.1 Input Parameters

The various 3D Mineral Resource block models for the Project, were used as the basis for deriving the economic shell limits for the Pine Point Project.

Estimates were made for metal prices, mining dilution, process recovery, off-site costs and royalties. Mining, processing, and general administration operating cost estimates (OPEX) were also calculated based on calculated processing throughput and, along with geotechnical parameters, formed the basis for OP optimization. The OP mining costs were estimated for both plant feed material and waste mining, where variations in haulage profiles and equipment selection were taken into account in the cost estimate.

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The block value is based on the NSR derived from using dense media separation plants followed by traditional grinding and flotation to upgrade mineralized material for the shipment of lead & zinc concentrates. Table 16.2 summarizes the NSR inputs and optimization parameters.

Table 16.2: Input Parameters OP Optimization

Parameter Unit Value Revenue, Smelting & Refining Lead Price US$/lb Pb 0.90 Zinc Price US$/lb Zn 1.05 Exchange Rate US$:C$ 0.75 Royalty (3%) % NSR 3.0% OPEX Estimates OP Waste Mining Cost C$/t waste mined 2.85 OP Mill Feed Mining Cost C$/t feed mined 2.85 Strip Ratio (estimated) W:O 4.0 OP Mining Cost C$/t processed 14.25 Processing Cost (DMS and Milling) (CP) C$/t processed 9.20 Processing Cost (DMS and Milling) (N-204) C$/t processed 7.43 Haulage from DMS to Mill C$/t processed 1.57 Additional Waste Rock and Tailings Reclamation C$/t processed 0.25 G&A C$/t processed 7.13 Total OPEX (excluding Mining) (CP) C$/t processed 18.15 Total OPEX (excluding Mining) (N-204) C$/t processed 16.38 Recovery and Dilution OP External Mining Dilution % 12 OP Mining Recovery % 95 Metal Recovery (DMS & Flotation) Lead (CP) % 85.0 Zinc (CP) % 88.1 Lead (N-204) % 71.4 Zinc (N-204) % 78.0 Other Overall Pit Slope Angles degrees 52 Mill Production Rate t/d 1,800 Smelter Terms and Offsite Costs Concentrate Grades %

Lead concentrate (CP) % 61.0 Lead concentrate (N-204) % 55.7 Lead concentrate moisture content % 8.00 Zinc concentrate (CP) % 62.0 Zinc concentrate (N-204) % 55.3 Zinc concentrate moisture content % 8.00 Payables

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Parameter Unit Value Lead % 95.0 Zinc % 85.0 Unit deductions Lead % 3.0 Zinc % 8.0 Smelting & Refining Costs Lead concentrate smelter cost (TC) $US$/dmt 135.00 Zinc concentrate smelter cost (TC) $US$/dmt 135.00 Freight & Marketing Lead Concentrate US$/wmt 167 Zinc Concentrate US$/wmt 154 *The values in this table vary slightly from those used in the economic model as parameters were further refined in the economic model. The differences are not considered material to the pit shape definition. Source: JDS (2017)

The mineral inventory block models for the deposits were then used with OP optimization software (NPVS) that incorporates the Lerchs-Grossman (LG) algorithm to determine optimal mining shells. This evaluation included the aforementioned parameters. The economic shell limits for some of the deposits also include Inferred Mineral Resources.

Inferred Mineral Resources are considered too speculative geologically to have the economic considerations applied to them to be categorized as Mineral Reserves, and there is no certainty that the Inferred Resources would be upgraded to a higher resource category.

16.2.1.1 Cut-off Grade Table 16.2 summarizes the parameters used to determine the incremental (or mill) cut-off grade (COG) (based on NSR). The incremental (or mill) COG incorporates all OPEX except mining and incorporates dilution. This incremental cut-off is applied to material contained within an economic pit shell where the decision to mine a given block was determined by the NPVS optimization.

This mill cut-off was applied to all of the estimates that follow. For the Cluster pits the NSR cut-off is $18.13/t milled, while for the N-204 deposit it is $16.38/t milled.

16.2.2 Optimization Results

A series of optimized shells were generated for the deposits based on varying revenue factors. The results were analyzed with shells chosen as the basis for ultimate limits and preliminary pit stage selection.

The LG algorithm produces both “best case” (i.e. mine out shell 1, the smallest shell, and then mine out each subsequent shell from the top down, before starting the next shell) and “worst case” (mine each bench completely to final limits before starting next bench) scenarios. These two scenarios provide a bracket for the range of possible outcomes.

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To better determine the optimum shells, and to gain a better understanding of the various deposits, the shells were analyzed in a preliminary schedule. The schedule assumed a maximum processing rate of 5.5 kt/d DMS feed for the CP pits and 7.3 kt/d for N-204. No stockpiles were used in the analysis and no CAPEX was added. Ultimate shells were chosen based on a review of mineralized rock and waste tonnages, incremental strip ratios and impact on the NPV.

Based on the analysis of the shells and preliminary schedule, the maximum NPV shell was chosen for all deposits other than N-204 (Revenue Factor (RF) 0.84) and X65 (RF 0.90).

Table 16.3 summarizes the tonnages and grades contained within the planned ultimate pit shells for the 10 deposits (using the incremental cut-off value of $18.15/t for the Cluster pits and $16.38/t for the N-204 deposit)

Table 16.3: Pit Shell Summary

Diluted Feed Strip Deposit Feed Pb Zn Pb Zn Waste Total Ratio (kt) (%) (%) (M Lb) (M Lb) (kt) (kt) (t:t) J68 304 2.25 4.9 15 33 385 1.3 689 HZ 1,549 2.76 3.58 94 122 6,271 4 7,820 W85 3,983 1.98 3.46 174 304 13,633 3.4 17,617 X65 5,485 0.79 2.42 95 293 11,154 2 16,639 O53 202 0.86 2.83 4 13 921 4.6 1,123 N-204 13,531 0.8 2.96 238 882 54,087 4 67,619 M67 1,365 0.84 3.55 25 107 9,783 7.2 11,148 K68 1,193 0.81 2.77 21 73 7,431 6.2 8,623 L65 1,578 0.7 1.95 25 68 9,884 6.3 11,463 M62/63 352 1.42 2.13 11 16 3,628 10.3 3,980 Total 29,544 1.08 2.93 703 1,910 117,178 4 146,722 Source: JDS (2017)

The general location of the various deposits is illustrated in Figure 16.1 along with existing roads and previously disturbed areas. Figure 16.2 through to Figure 16.11 illustrate the individual pit shapes.

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Figure 16.1: Overview of Deposit Locations

Source: JDS (2017)

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Figure 16.2: J68 Pit Shell

Source: JDS (2017)

Figure 16.3: Hinge Zone Pit Shell

Source: JDS (2017)

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Figure 16.4: W85 Pit Shell

Source: JDS (2017)

Figure 16.5: X65 Pit Shell

Source: JDS (2017)

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Figure 16.6: N-204 Pit Shell

Source: JDS (2017)

Figure 16.7: M67 Pit Shell

Source: JDS (2017)

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Figure 16.8: O53 Pit Shell

Source: JDS (2017)

Figure 16.9: K68 Pit Shell

Source: JDS (2017)

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Figure 16.10: L65 Pit Shell

Source: JDS (2017)

Figure 16.11: M62/63 Pit Shell

Source: JDS (2017)

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16.2.3 Pit Stages

The resultant pit shapes for Pine Point were further analyzed and optimizations were conducted in order to better define the possible stage shapes within the ultimate pit limits. The ultimate pit shapes for a number of the deposits were divided into smaller stages. This provided flexibility in the mining sequence to maximize the grade in the early years, reduce the pre-stripping requirements, and to maintain the process facility at full production capacity. The pit tonnages, grades, and contained metal of the stages for all deposits are summarized in Table 16.4 and Figure 16.12.

Table 16.4: Pit Stage Summary

Diluted Feed Waste Strip Total Deposit (kt) Pb (%) Zn (%) Pb (M Lb) Zn (M Lb) (kt) Ratio (kt) J68 304 2.25 4.9 15 33 385 1.3 689 HZ 1,549 2.76 3.58 94 122 6,271 4 7,820 W85A 1,618 2.42 3.06 86 109 5,125 3.2 6,743 W85B 2,365 1.68 3.73 88 194 8,508 3.6 10,873 X65A 3,282 0.75 2.51 54 182 3,624 1.1 6,906 X65B 1,099 0.92 2.28 22 55 1,517 1.4 2,616 X65C 1,104 0.78 2.3 19 56 6,014 5.4 7,117 N-204A 5,789 0.73 2.88 93 368 18,890 3.3 24,679 N-204B 3,627 0.89 3.04 72 243 18,683 5.2 22,310 N-204C 1,363 0.75 2.95 22 89 5,597 4.1 6,960 N-204D 2,752 0.83 3 51 182 10,918 4 13,669 O53 202 0.86 2.83 4 13 921 4.6 1,123 M67 1,365 0.84 3.55 25 107 9,783 7.2 11,148 K68 1,193 0.81 2.77 21 73 7,431 6.2 8,623 L65A 1,048 0.5 2.12 11 49 6,146 5.9 7,194 L65B 530 1.11 1.61 13 19 3,739 7 4,269 M62/63 352 1.42 2.13 11 16 3,628 10.3 3,980 Total OP 29,544 1.08 2.93 702 1,910 117,178 4 146,722 Source: JDS (2017)

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Figure 16.12: Pit Stages

Darnley Bay - Pine Point PEA Pit Stage Summary 30,000 10.50

25,000 8.50

20,000 6.50

15,000 4.50 Tonnage (kt)

10,000 2.50 Mined Grade (%)

5,000 0.50

0 -1.50

Pit Stage

Diluted Feed Waste Pb (%) Zn (%) Strip

Source: JDS (2017)

16.3 Mine Production Schedule The production schedule for the Pine Point deposits incorporates the various pits and pit stages mentioned above. The majority of the feed tonnes are comprised of Measured and Indicated resources. Only 13% of the total feed tonnes are based on Inferred Resources.

Inferred Mineral Resources are considered too speculative geologically to have the economic considerations applied to them to be categorized as Mineral Reserves, and there is no certainty that the Inferred Resources would be upgraded to a higher resource category.

The mining schedule is based on developing the open pit deposits with mineralized material fed initially to DMS plants followed by traditional grinding and flotation to upgrade mineralized material into lead & zinc concentrates. Mining rates will range from 4,000 to 6,800 t/d of mineralized material for processing over the life of the mine.

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Mineral processing will include a combination of portable and fixed crushers and dense media separation plants located near each deposit, which will concentrate the mineralized material, and a pre-concentrate will then be trucked to a central 1,800 t/d milling facility, where it will be subject to standard grinding and flotation methods to produce a final concentrate.

The DMS recovery and hence required throughput varies by deposit and is based on the combined lead and zinc grades, with the higher grades having less float rejects.

The current life-of-mine (LOM) plan focuses on achieving consistent processing feed production rates, mining of higher value material early in the schedule, balancing grade and strip ratios, while trying to maximize NPV. Consideration was also given to waste rock and tailings management, with tailings to be conveyed from the centralized milling facility and deposited within historic and proposed open pits.

The open pits will be mined sequentially, with up to three stages and/or pits active at any one time. Mining is planned to commence at the cluster pits with N-204 commencing in Year 3.

The average mining rate over the 13-year LOM is planned to be 31,000 t/d, with a maximum of 40,000 t/d during Year 3.

Table 16.5 is a summary of total material movement by year for the LOM production schedule (both as totals, as well as by each stage) as well as the proposed mill schedule. Figure 16.13 and Figure 16.14 further illustrate the Pine Point LOM schedule.

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Table 16.5: LOM Production Schedule

Year Item Units Total -1 1 2 3 4 5 6 7 8 9 10 11 12 13 J68 DMS Feed (kt) 304 304 Waste (kt) 385 385 Total (kt) 689 689 Mill Feed (kt) 107 107 Pb Grade (%) 2.25 2.25 Zn Grade (%) 4.9 4.9 Strip Ratio (w:o) 1.3 1.3 HZ DMS Feed (kt) 1,549 1,549 Waste (kt) 6,271 6,271 Total (kt) 7,820 7,820 Mill Feed (kt) 523 523 Pb Grade (%) 2.76 2.76 Zn Grade (%) 3.58 3.58 Strip Ratio (w:o) 4 4 W85A DMS Feed (kt) 1,618 83 1,536 Waste (kt) 5,125 2,871 2,254 Total (kt) 6,743 2,953 3,790 Mill Feed (kt) 524 27 497 Pb Grade (%) 2.42 4.39 2.32 Zn Grade (%) 3.06 1.76 3.13 Strip Ratio (w:o) 3.2 34.7 1.5 W85B DMS Feed (kt) 2,365 620 1,745 Waste (kt) 8,508 7,893 616 Total (kt) 10,873 8,513 2,360 Mill Feed (kt) 608 159 449 Pb Grade (%) 1.68 1.93 1.59 Zn Grade (%) 3.73 3.74 3.73 Strip Ratio (w:o) 3.6 12.7 0.4 X65A DMS Feed (kt) 3,282 518 1,717 357 691 Waste (kt) 3,624 885 1,560 1,079 100 Total (kt) 6,906 1,403 3,277 1,435 791 Mill Feed (kt) 965 152 505 105 203 Pb Grade (%) 0.75 0.84 0.83 0.63 0.55 Zn Grade (%) 2.51 2.56 2.51 2.35 2.56 Strip Ratio (w:o) 1.1 1.7 0.9 3 0.1 X65B DMS Feed (kt) 1,099 1,099 Waste (kt) 1,517 1,517 Total (kt) 2,616 2,616 Mill Feed (kt) 323 323 Pb Grade (%) 0.92 0.92 Zn Grade (%) 2.28 2.28 Strip Ratio (w:o) 1.4 1.4 X65C DMS Feed (kt) 1,104 293 810

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Year Item Units Total -1 1 2 3 4 5 6 7 8 9 10 11 12 13 Waste (kt) 6,014 5,603 411 Total (kt) 7,117 5,896 1,221 Mill Feed (kt) 325 86 238 Pb Grade (%) 0.78 0.57 0.86 Zn Grade (%) 2.3 1.97 2.42 Strip Ratio (w:o) 5.4 19.1 0.5 N-204A DMS Feed (kt) 5,789 218 590 2,150 1,712 1,119 Waste (kt) 18,890 10,566 6,435 1,814 30 45 Total (kt) 24,679 10,784 7,025 3,964 1,743 1,164 Mill Feed (kt) 1,489 56 152 553 440 288 Pb Grade (%) 0.73 0.33 0.52 0.9 0.68 0.67 Zn Grade (%) 2.88 1.27 1.73 3.72 2.59 2.63 Strip Ratio (w:o) 3.3 48.4 10.9 0.8 0 0 N-204B DMS Feed (kt) 3,627 53 178 2,219 1,177 Waste (kt) 18,683 4,827 5,468 5,288 3,013 87 Total (kt) 22,310 4,827 5,521 5,466 5,232 1,264 Mill Feed (kt) 933 14 46 571 303 Pb Grade (%) 0.89 0.31 0.52 0.96 0.86 Zn Grade (%) 3.04 1.59 2.13 3.28 2.8 Strip Ratio (w:o) 5.2 103.4 29.7 1.4 0.1 N-204C DMS Feed (kt) 1,363 1,131 232 Waste (kt) 5,597 5,598 -1 Total (kt) 6,960 6,729 232 Mill Feed (kt) 351 291 60 Pb Grade (%) 0.75 0.71 0.91 Zn Grade (%) 2.95 2.75 3.97 Strip Ratio (w:o) 4.1 4.9 0 N-204D DMS Feed (kt) 2,752 1,725 1,027 Waste (kt) 10,918 9,725 1,193 Total (kt) 13,669 11,450 2,220 Mill Feed (kt) 708 444 264 Pb Grade (%) 0.83 0.76 0.96 Zn Grade (%) 3 2.72 3.47 Strip Ratio (w:o) 4 5.6 1.2 M67 DMS Feed (kt) 1,365 375 991 Waste (kt) 9,783 9,531 252 Total (kt) 11,148 9,906 1,242 Mill Feed (kt) 422 116 306 Pb Grade (%) 0.84 0.5 0.96 Zn Grade (%) 3.55 2.27 4.04 Strip Ratio (w:o) 7.2 25.4 0.3 O53 DMS Feed (kt) 202 202 Waste (kt) 921 921 Total (kt) 1,123 1,123 Mill Feed (kt) 61 61 Pb Grade (%) 0.86 0.86

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Year Item Units Total -1 1 2 3 4 5 6 7 8 9 10 11 12 13 Zn Grade (%) 2.83 2.83 Strip Ratio (w:o) 4.6 4.6 K68 DMS Feed (kt) 1,193 1,193 Waste (kt) 7,431 3,782 3,648 Total (kt) 8,623 3,782 4,841 Mill Feed (kt) 356 356 Pb Grade (%) 0.81 0.81 Zn Grade (%) 2.77 2.77 Strip Ratio (w:o) 6.2 3.1 L65A DMS Feed (kt) 1,048 841 207 Waste (kt) 6,146 6,150 -5 Total (kt) 7,194 6,992 202 Mill Feed (kt) 301 242 59 Pb Grade (%) 0.5 0.53 0.38 Zn Grade (%) 2.12 2.16 1.94 Strip Ratio (w:o) 5.9 7.3 0 L65B DMS Feed (kt) 530 530 Waste (kt) 3,739 1,996 1,743 Total (kt) 4,269 1,996 2,273 Mill Feed (kt) 152 152 Pb Grade (%) 1.11 1.11 Zn Grade (%) 1.61 1.61 Strip Ratio (w:o) 7 3.3 M62/63 DMS Feed (kt) 352 352 Waste (kt) 3,628 3,628 Total (kt) 3,980 3,980 Mill Feed (kt) 105 105 Pb Grade (%) 1.42 1.42 Zn Grade (%) 2.13 2.13 Strip Ratio (w:o) 10.3 10.3 TOTAL Open Pit DMS Feed (kt) 29,544 304 1,632 2,156 2,481 2,307 2,507 2,456 2,396 2,512 2,363 2,324 2,267 2,462 1,378 Waste (kt) 117,178 385 9,142 10,147 12,066 7,995 7,720 5,598 6,850 8,615 10,029 10,553 11,794 11,463 4,821 Total (kt) 146,722 689 10,774 12,303 14,547 10,302 10,227 8,054 9,246 11,128 12,392 12,877 14,060 13,925 6,199 Mill Feed (kt) 8,252 107 550 657 657 656 658 657 657 657 657 658 658 655 369 Pb Grade (%) 1.08 2.25 2.84 2.21 1.32 0.75 0.86 0.64 0.77 0.91 0.8 0.83 0.71 0.81 1.07 Zn Grade (%) 2.93 4.9 3.49 3.3 3.27 2.31 3.53 2.56 2.43 3.13 2.59 3.3 2.67 2.41 3.13 Strip Ratio (w:o) 4 1.3 5.6 4.7 4.9 3.5 3.1 2.3 2.9 3.4 4.2 4.5 5.2 4.7 3.5 Mill Feed Mill Feed (kt) 8,252 - 657 657 657 657 657 657 657 657 657 657 657 657 368 Pb Metal Contained (Mlbs) 646 111 99 68 35 42 31 37 44 38 39 33 39 29 Zn Metal Contained (Mlbs) 1,741 148 147 167 109 173 124 116 154 122 155 124 117 85 Pb Mill Head Grade (%) 3.55 7.65 6.83 4.71 2.45 2.89 2.12 2.57 3.06 2.61 2.69 2.3 2.7 3.6 Zn Mill Head Grade (%) 9.57 10.22 10.13 11.52 7.52 11.97 8.6 8.04 10.62 8.44 10.69 8.54 8.1 10.45 Source: JDS (2017)

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Figure 16.13: LOM Schedule

Pine Point PEA - LOM Schedule

16,000 6.00

14,000 5.00 Thousands 12,000

4.00 10,000

8,000 3.00

6,000 Material Tonnage 2.00

4,000

1.00 2,000 Mined Grade (pb%, Zn and %) Strip Ratio (t:t)

0 0.00 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 Year

OP DMS Feed OP Waste Avg Mined Pb Grade Avg mined Zn grade Strip Ratio

Source: JDS (2017)

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Figure 16.14: Mineralized Material by Pit Phase

Pine Point PEA - Mined DMS Feed Tonnage by Phase

3,000 8,000 2,500 7,000 6,000 2,000 5,000 1,500 4,000 1,000 3,000 2,000 500 1,000 0 0 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 Mined DMS Feed Tonnage (kt) Tonnage Feed DMS Mined HZ 0 1,549 0 0 0 0 0 0 0 0 0 0 0 0 J68 304 0 0 0 0 0 0 0 0 0 0 0 0 0 W85A 0 83 1,536 0 0 0 0 0 0 0 0 0 0 0

W85B 0 0 620 1,745 0 0 0 0 0 0 0 0 0 0 (tonnes/day) Feed DMS Mined Total X65A 0 0 0 518 1,717 357 691 0 0 0 0 0 0 0 X65B 0 0 0 0 0 0 0 1,099 0 0 0 0 0 0 X65C 0 0 0 0 0 0 0 0 293 810 0 0 0 0 N204A 0 0 0 218 590 2,150 1,712 1,119 0 0 0 0 0 0 M67 0 0 0 0 0 0 0 0 0 375 991 0 0 0 O53 0 0 0 0 0 0 0 0 0 0 202 0 0 0 K68 0 0 0 0 0 0 0 0 0 0 0 1,193 0 0 N204B 0 0 0 0 0 0 53 178 2,219 1,177 0 0 0 0 N204C 0 0 0 0 0 0 0 0 0 0 1,131 232 0 0 L65B 0 0 0 0 0 0 0 0 0 0 0 0 530 0 L65A 0 0 0 0 0 0 0 0 0 0 0 841 207 0 M62/63 0 0 0 0 0 0 0 0 0 0 0 0 0 352 N204D 0 0 0 0 0 0 0 0 0 0 0 0 1,725 1,027 Total DMS Feed/day 834 4,472 5,906 6,798 6,320 6,868 6,728 6,564 6,883 6,473 6,366 6,210 6,745 3,776 Year

Source: JDS (2017)

16.4 Mine Equipment The PEA envisions the use of conventional open pit mining methods. Mining rates will range from 4,000 to 6,800 t/d of mineralized material and, on average, a total of 32,000 t/d of total material mined over the life of the mine.

The primary, all diesel, mining fleet will consist of 90-tonne haul trucks, 8.0 m3 front shovels, 8.0 m3 front end loaders and 150 mm diameter drills. The ancillary mining fleet will consist of track dozers, graders, wheel dozers and water trucks. Table 16.6 summarizes the annual open pit mobile equipment requirements for the Project.

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Table 16.6: Mobile Mine Equipment Fleet (average number of units)

Equipment Type Units Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Fleet Requirements (Production Equipment) Drill (89-152 mm) # 1 3 3 3 3 3 3 3 3 3 3 3 3 2 Shovel (8.0 m3, Hydraulic) # 1 2 2 2 2 2 2 2 2 2 2 2 2 2 FEL (8.0 m3) # 1 1 2 2 2 2 2 2 2 2 2 2 2 1 Haul Truck (90 t) # 4 6 8 8 8 8 8 8 8 8 8 8 8 5 Track Dozer (D275 class) # 2 3 3 4 4 4 4 4 4 4 4 4 4 3 Wheel Dozer (824 class) # - 1 1 1 1 1 1 1 1 1 1 1 1 1 Grader (14H class) # 1 3 3 3 3 3 3 3 3 3 3 3 3 2 Water Truck (55 m3) # 1 2 2 2 2 2 2 2 2 2 2 2 2 1 Source: JDS (2017)

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16.5 Mine Personnel The mine personnel requirements include mine operations equipment operators, mine maintenance, supervision and technical services. Table 16.7 summarizes annual total manpower estimates, where:

 Mine General – front line supervision, superintendents;  Drill & Blast – drillers, blasters, blaster helpers;  Load & Haul – shovel/loader operators, haul truck drivers, ancillary equipment operators, labourers;  Maintenance – heavy equipment mechanics, electricians and welders, maintenance supervisors, tireman, labourers; and  Technical Services - mine engineering staff including superintendents, engineers, geologists, surveyors and clerks.

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Table 16.7: Total Mine Personnel Requirements

Position UNITS Y-1 Y1 Y2 Y3 Y4 Y5 Y6 Y7 Y8 Y9 Y10 Y11 Y12 Y13 Mine Operations Personnel Mine General # 5 5 5 5 5 5 5 5 5 5 5 5 5 5 Drill & Blast # 10 14 18 18 14 14 14 14 18 18 18 18 18 14 Load & Haul # 27 64 75 82 72 72 72 72 72 77 77 81 78 54 Maintenance # 30 34 38 42 34 34 34 34 38 38 38 42 38 30 Technical Services # 15 15 15 15 15 15 15 15 15 15 15 15 15 15 Source: JDS (2017)

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17 Process Description/Recovery Methods

The Pine Point Project has identified eight deposits of similar composition, referred to as the Cluster Pits (CPs), in addition to those identified in the MineTech NI 43-101 Technical Report of April 2, 2012 and the MineTech NI 43-101 Technical Report of March 23, 2012. As reported in the MineTech NI 43-101 Technical Report of April 2, 2012, mineralization occurs as coarse galena and sphalerite in a matrix of limestone, dolomite and shale. The recent metallurgical test program demonstrates that standard zinc and lead flotation techniques preceded by dense media separation (DMS) will yield zinc and lead recoveries of over 90% and concentrate grades of over 60%. These flotation test results are in line with historical operational data of zinc and lead recoveries and concentrate grades produced over a 23-year period by Pine Point Mines Ltd. (Cominco) at Pine Point.

Two similar crushing plants and DMS plants are proposed to treat the mineralized material from the respective mining operations at the cluster pits and N-204 pit. These plants will be located near the respective mining operations.

The mining schedule dictates the requirement for two crushing/DMS facilities, although there will be periods where only one facility will be operational at any one time.

A central grinding and flotation plant, located at the cluster pit area, will process the DMS concentrate products

The lead and zinc grades of mined material from the cluster pits (CP) delivered to the DMS plant is expected to average at 2% lead and 5% zinc. The grade of the CP DMS concentrate product is expected to average approximately 7% lead and 14% zinc.

The process plant will consist of secondary and tertiary crushing; a grinding circuit consisting of a ball mill in closed-circuit with cyclones, sequential lead and zinc flotation circuits each incorporating three stages of cleaning, concentrate dewatering circuits using thickeners and pressure filters, concentrate storage and load-out facilities, and tailings disposal. Existing and new mined out open pits will be used for tailings disposal.

17.1 Introduction The process design criteria and flowsheets have been developed based on both the recent metallurgical test work results detailed in Section 13, and historical operating data. The test data indicates that the Pine Point mineralized material can be treated using conventional mineral processing techniques, applying conventional differential flotation techniques for the recovery of lead and zinc concentrates.

The two crushing plants and DMS plants will each process between 5,700 and 7,300 t/d of material from the cluster pits (CP) and N-204 pit, respectively. Each facility will include three stages of crushing and a DMS plant.

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The concentrator plant will consist of a grinding mill, lead rougher flotation and three stages of lead cleaner flotation and zinc rougher flotation, and three stages of zinc cleaner flotation. Both concentrates will be dewatered in concentrate thickeners and pressure filters.

The LOM average metal recoveries are expected to be in the range of 71 to 85% for lead at a grade of 56 to 70% lead. Zinc recoveries are expected to be in the range 78 to 88% with concentrate grades of 55 to 62%.

The crushing circuits will operate on average 18 hours a day. The process plant will operate 24 hours per day, 365 days per year at an availability of 92%.

The processing facilities will consist of the following unit operations:

 Two plants - 3-stage crushing and screening;  Two DMS plants;  Materials handling and Primary grinding;  Lead flotation using conventional and column cells;  Zinc flotation using conventional and column cells;  Lead concentrate dewatering, filtration and truck load-out facility;  Zinc concentrate dewatering, filtration and truck load-out facility; and  Tailings disposal.

17.2 Plant Design Criteria The process plant design criteria used for this PEA is provided in Table 17.1 with the process flowsheet illustrated in Figure 17.1.

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Table 17.1: Process Design Criteria

Design Criteria Description Units Source (CP/N-204) t/d 1,800 Plant Throughput Siega/Gann 2014, t/a 657,000 Crusher Availability % 75 JDS Crusher/DMS Throughput t/d 5,700 to 7,300 JDS Mill/Flotation Availability % 92 JDS Mill Throughput t/h 82 Siega /Gann 2014 Physical Characteristics BWi kWh/t 7.4 G&T KM2984 2011

Primary Grind Size P80 µm 75 Siega /Gann 2014 Concentrate Grind Size (N-204 P80 µm 22 Test Work Regrind) Head Grade to Plant % Pb 6.9/2.5 Test Work % Zn 14.0/9.3 Lead % 85.0/71.4 Test Work Overall Recovery Zinc % 88.1/78.0 Test Work Lead % 70.0/55.7 Test Work Concentrate Grade Zinc % 61.9/55.3 Test Work

Lead Circuit Residence time Roughers minutes 6/8 Test Work

Cleaner minutes 11/11 Test Work Zinc Circuit Residence time Roughers mins 6/6 Test Work Cleaners mins 15/15 Test Work Lead % 8.9/3.6 JDS Concentrate Mass Pull Zinc % 21.2/14.9 JDS Settling Rate Concentrates m3/m2h 0.4 2008 - SGS Project 11790-001 Zinc Thickener Underflow Density Target % w/w 60 Tailings Concentrate Moisture Target % 8 Source: JDS, CP: 2007 SGS Project 11633-001 R-190 and operating data; N-204: 2012 G&T KM3195-06

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17.3 Process Plant Description 17.3.1 Crushing and DMS

Two mobile crushing and DMS plants will operate to process the mined material from CP and N-204 pits, respectively. The crushing plants include a jaw crusher for primary crushing followed by secondary and tertiary cone crushers. The N-204 and CP crushing circuit will process up to 7,300 t/d through a 160 kW 42”x48” jaw crusher, a 220 kW secondary cone crusher and a 300 kW tertiary cone crusher. The N-204 crushing circuit has an additional screen to handle the higher tonnage. Run-of-mine mineralized rock will be crushed and screened to a final product size P80 of 6.5 mm (¼ inch).

The 6.5 mm material will be conveyed to the DMS plant. The material will be screened with a bottom deck screen aperture of 1.4 mm (14 mesh) to eliminate the fines fraction from the DMS process. The fines (minus 14 mesh) and the DMS sinks product will be trucked to the process plant. The DMS float material (waste / low grade) will be trucked to the waste dumps or used as aggregate for road construction.

The DMS plant includes screening, DMS cyclones, ferrosilicon recovery, magnetic separator and ancillary equipment for materials handling. Ferrosilicon will be used as the media in the DMS plant at an operating density between 2.8 and 2.9 g/cm3.

The DMS flowsheet is provided in Figure 17.2.

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Figure 17.2: DMS Flowsheet

Source: Vendor Supplied (2017)

17.3.2 Grinding

DMS sinks and fines will be trucked to plant and dumped into a hopper and conveyed to a surge bin that feeds the grinding circuit at a rate of 1,800 t/d. The grinding circuit consists of a 3.7 m diameter by 5.5 m long, 933 kW ball mill operating in closed circuit with a cyclone cluster. Cyclone underflow will be recycled to the mill for further size reduction and cyclone overflow, with a target grind size of P80 75 micron, will flow to either pre-float circuit, N-204 material, or to the lead conditioning tank.

17.3.3 Flotation

Preliminary metallurgical test work has identified the presence of bituminous-activated slimes in the N-204 material.

The flotation circuit has thus been designed to include a pre-float stage to remove this material prior to the lead flotation stage for N-204 material

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When processing the N-204 material, the mill cyclone overflow process stream will be directed to the pre-float stage to scalp the bituminous-activated slimes. When processing the CP material, the mill cyclone overflow product will flow directly to the lead conditioning tank. The N-204 pre-float concentrate will report to final tailings and the cell tailings will report to the lead conditioning tank.

Test work indicates N-204 material will require regrind of the rougher concentrates to a target P80 of 22 µm.

17.3.3.1 Lead Circuit The lead flotation circuit consists of conventional tank cells and column flotation cells. The rougher- scavenger circuit includes 6 to 10 m3 flotation cells. The rougher concentrate will be pumped to the cleaner circuit, and the tailings will feed the zinc flotation circuit. The rougher concentrate will be pumped to two 9 m3 flotation columns operating in parallel. Cleaner column tailings will be pumped to 2 to 3 m3 cleaner scavenger cells. The cleaner scavenger concentrate will be recycled to the column cells and the tailings will be delivered to the rougher circuit. Flotation column concentrate gravitates to the 6 m diameter lead thickener. The thickener overflow will be used in the process as spray water within the lead cleaner circuit. Thickened lead concentrate will be pumped to a pressure filter for further dewatering to approximately 8% moisture for loading into trucks for transporting to the railhead at Hay River. 17.3.3.2 Zinc Circuit The tailings of the lead circuit will report to the zinc conditioning tank. The slurry will be conditioned with lime, to pH 10-11, and various flotation reagents before being fed to the zinc rougher and zinc scavenger flotation cells. Tailings from the zinc rougher circuit will be collected in the final tailings pump box and then pumped to various disposal sites. The zinc rougher concentrate will feed the zinc first cleaner flotation cells and the first cleaner scavenger tailings will be combined with the rougher tailings in the final tailings pump box. The first cleaner scavenger concentrate will be pumped back to the rougher conditioning tank. The first cleaner concentrate will feed four 9 m3 second cleaner columns operating in parallel. Second cleaner column tailings will report to the first cleaner circuit. Column (final) concentrate will flow to the 8 m diameter zinc concentrate thickener. Zinc thickener overflow will be used as spray water in the zinc cleaner circuit. Zinc concentrate thickener underflow process stream will be pumped to a pressure filter for further dewatering to approximately 8% moisture for loading into trucks for transport to the railhead at Hay River. 17.3.4 Reagents

Reagents added to the flotation circuits are prepared and distributed from the reagent handling facility. This area includes various mixing and storage tank units. All reagent areas will be bermed with sump pumps, which transfer spills to the final tailings pump box, with the exception of the flocculant. Flocculant spillage will be returned to the storage tank. The reagents will be mixed, stored and then delivered through individual supply loops with dosage controlled by flow meters and manual control valves. The storage tanks have been sized for a minimum of 1-day storage capacity. The reagents will be delivered to the mine site as either in powder form or as solutions. The reagents and consumptions are listed in Table 17.2.

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Table 17.2: Reagents and Consumptions

Consumption, t/y Lime 1,183 Flocculant 26 Copper Sulphate 56 PAX 22 MIBC 2

ZnSO4 404

Na2S 104 Oxalic Acid 104 SQ6 296 CMC 116 PPK 116 EDTA 65 Aerophine 8 Thionocarbamate - Danafloat 262 4 Source: JDS (2017)

17.3.5 Plant Air

The primary consumers of compressed air are the primary crushing plant, cleaner flotation columns and the lead and zinc concentrate filter. In addition, minor users of compressed air are dust collection/suppression, samplers, mill gear lubrication systems and air hose stations located throughout the plant.

Blowers will be used to supply air to the flotation circuits.

17.3.6 Assay Laboratory

The assay laboratory will consist of a sample preparation/metallurgical module and a wet laboratory module. The two containers will be located outside the process plant.

The laboratory will be performing test work for the mine, the mill, and the environmental group. On- site analyses might include Pb, Zn, and Fe. Samples may also be analyzed for C, SiO2, S, and SO4. The concentrates will likely be tested for Pb, Zn, As, Sb, Fe, and Cd and possibly also SiO2 and C. The high grade concentrates will be titrated for Pb and Zn.

Analysis of samples for environmental monitoring requirements might be conducted at the on-site laboratory. However, it is anticipated that some of the analyses will be sent to third party laboratories.

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17.4 Process Personnel The process personnel requirements include mine operations equipment operators, mine maintenance, supervision and technical services. Table 17.3 summarizes annual total manpower estimates, where:

 Mine Management – mill manager, administration and front line supervision;  Mill Operations – operators and labourers;  Maintenance – millwrights, electricians and welders, maintenance supervisors and planners; and  Technical Services – metallurgists, assayers and lab technicians. Table 17.3: Total Process Personnel

Position Number of Employees Mill Management 6 Mill Operations 58 Mill Maintenance 22 Technical Services 12 Total - Processing Staff 98 Source: JDS (2017)

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18 Project Infrastructure and Services

18.1 Summary The Project envisions the upgrading or construction of the following key infrastructure items:  Primary power supply from the NWT Power grid;  Secondary and back-up power supply from LNG generators;  Tailings management facilities in existing and mined out open pits;  Waste rock storage areas (WRSA) in existing open pits or external stockpiles;  Pit dewatering wells and reinjection wells;  2 x Crushing and dense media separation circuits;  Lead/Zinc process facility with grinding, flotation, thickening and dewatering;  Assay laboratory;  Administration, mine dry and first aid facilities;  Mine maintenance facilities;  Plant warehouse facilities;  Bulk fuel storage and distribution; and  Temporary worker housing complex.

18.2 General Site Arrangement An overall site plan for the Project area is shown in Figure 18.1

18.3 Site Access The Pine Point property is easily accessible via NWT Highway 6, a 365 day a year maintained paved highway. Use of existing roads and historic exploration cut-lines will be maximized to the extent feasible in order to minimize the need for ground disturbance. Roads will be constructed with waste rock from the Cluster Pits or N-204. Inert waste rock will be used for road construction, good to excellent main haul roads currently exist in the Cluster Pit area. Short access roads will be required from these haul roads to access the each cluster pit and the N-204 deposit.

Effective Date: April 18, 2017 18-1

620,000 630,000 640,000 650,000 660,000

6,760,000 6,760,000

6,750,000 6,750,000

6,740,000 6,740,000 620,000 630,000 640,000 650,000 660,000

LEGEND: NOTES: DARNLEY BAY RESOURCES LIMITED 1 0.5 0 1 2 3 4 5 km 1. BASE MAP: BING ONLINE AND PROJECT IMAGERY. HISTORIC PIT PINE POINT PROJECT 2. COORDINATE GRID IS IN METRES. SUBSTATION COORDINATE SYSTEM: NAD 1983 UTM ZONE 11N.

DEPOSIT 3. THIS FIGURE IS PRODUCED AT A NOMINAL SCALE OF 1:110,000 SITE GENERAL ARRANGEMENT FOR 11x17 (TABLOID) PAPER. ACTUAL SCALE MAY DIFFER PLANT SITE ACCORDING TO CHANGES IN PRINTER SETTINGS OR PRINTED PAPER SIZE. P/A NO. REF NO. 4. HISTORICAL SITE LOCATIONS REFERENCED FROM VA101-60/6 VA17-00752 KK GLS 0 23MAY'17 ISSUED WITH TRANSMITTAL KK STEVENSON INTERNATIONAL GROUNDWATER CONSULTANTS LTD's REV REV DATE DESCRIPTION DRAWN REVIEWED REPORT OF THE HYDROGEOLOGY OF R190 MINERALIZED REGION (1983). SAVED: M:\1\01\00060\06\A\GIS\Figs\VA17-00752\Fig18-1_SiteGeneralArrangement_r0.mxd; May 23, 2017 10:27 AM; M:\1\01\00060\06\A\GIS\Figs\VA17-00752\Fig18-1_SiteGeneralArrangement_r0.mxd;kkrauszova 23, May 2017 10:27 SAVED: DESIGNED FIGURE 18-1 0 DARNLEY BAY RESOURCES LTD. PINE POINT PEA

18.4 Power Northwest Territories Power Corporation (NTPC) supplies power to South Slave region of the Northwest Territories via the Taltson Dam. The Taltson facility was brought online in 1965 to supply power to the original Pine Point Mines Ltd. (Cominco) facilities at Pine Point, which closed down in 1987, and also to the town of Fort Smith, NWT. The facility now supplies power to the towns of Hay River, Fort Smith, Fort Resolution and Enterprise. The dam complex currently has an 18.5 MW capacity, of which 3 MW is currently unused and available for use during the winter months and 12 MW unused and available in the summer months.

Primary power to the Cluster Pit area will be supplied by a 72 kV transmission line that originates at the existing Pine Point Substation. This line runs west from the substation and parallels Highway 6 past the CP deposits on its way to Hay River.

Primary power to the N-204 deposit will be via the existing 12.5 kV transmission line that originates at the Pine Point Substation and travels east along Highway 6 to Fort Resolution, NT.

Additional power required at the Cluster Pit area during the winter months will be provided by five LNG reciprocating engine generator sets (gensets) in a N+2 (3+2) arrangement. Each generator will be prime rated for 1,450 kW running at 1,200 rpm and generating power of 4,160 V. The peak gross power is 4.35 MW (three operating gensets @ 100%; 3 x 1.45 MW). The average estimated electrical loads is 6.3 MW with the average annual power consumption estimated to be approximately 46,400,000 kWh/a.

An onsite substation near the process plant at the Cluster Pit area will step-down power from 72 kV to 4,160 V. A substation near the DMS plant at the N-204 deposit will step-down the power from 12.5 kV to 4,160V. The mine power distribution will be at 4160 V. Feeders, both cable and overhead line, will run out from the main substation and feed substations located on the mine site. Portable 4160V-600V substations will transform voltages down to utilization levels for motor loads for the various pieces of equipment.

18.5 Ancillary Facilities 18.5.1 Truck Shop & Warehouse Building

Truck shop complexes will be located at both the Cluster Pit area and the N-204 deposit. Each shop will consist of an 800 m2 structural steel, pre-engineered building designed to accommodate facilities for repair and maintenance of mining equipment and light vehicles. A covered cold storage warehouse will be provided at the cluster pit area with a 320 m2 pre-engineered fabric building.

18.5.2 Mine Dry and Office Complex

The main office complex will be located at the Cluster Pit area with a smaller satellite office at the N- 204 deposit area. Each location will have a similar size mine dry facilities. The mine dry and office complexes will be constructed from modular units manufactured off-site and in compliance with highway transportation size restrictions. Modules will rest on wood cribbing. The complexes will

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DARNLEY BAY RESOURCES LTD. PINE POINT PEA comply with all building and fire code requirements and be provided with sprinklers throughout. Arctic corridors will be provided to connect the mine dries and office complexes.

The mine dry facilities will service construction and operations staff during the life of the Project. Each facility will be capable of servicing 100 workers during shift change and contain the following:

 Male and female lockers; and  Showers and washroom facilities with separate male and female sections.

The main office facility at the cluster pit area will contain the following items:  Private offices;  Main boardroom; and  Mine operations line-up area.

The office facility at the N-204 area will contain the following items:  Mine operations offices; and  Mine operations line-up area.

18.5.3 Camp

The camp will be located in close proximity to the process facilities at the Cluster Pits. The camp will comprise single-occupancy rooms with central washrooms. It will be used during the construction phase and throughout the operations phase. The camp is design to accommodate approximately 105 management and support staff.

The kitchen / dining / recreation complex will include the following:

 Kitchen complete with cooking, preparation and baking areas, dry food storage and walk-in freezer / cooler. The kitchen will be provided with appropriate specialized fire detection and suppression systems;  Dining room with serving and lunch preparation areas;  First aid room;  Mudroom complete with coat and boot racks, benches and male-female washrooms;  Housekeeping facilities;  Reception desk and lobby; and  Recreation area.

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The camp will be constructed from modular units manufactured off-site in compliance with highway transportation size restrictions. Camp modules will rest on wood cribbing. The camp will comply with all building and fire code requirements and be provided with sprinklers throughout. Arctic corridors will connect the main camp complex and dormitory wings.

18.5.4 Fuel Storage

On-site fuel storage is designed with a one-week supply capacity. One - 100,000 L double walled enviro-tank will be installed at each of the Cluster Pit area and N-204 area. Each tank will be set within a lined containment berm. Fuel dispensing equipment for mining, plant services, and freight vehicles will be located adjacent to the fuel tank bund and the fueling area will drain into the bund.

Fuel will be transported to site with 50,000 L tanker trucks on an as-required basis.

18.5.5 Explosives Storage

Explosive storage for the Pine Point Project will consists of the following three main components:

 Bulk ammonium nitrate (AN) storage;  Bulk emulsion storage; and  Explosive storage magazines.

Bulk ammonium nitrate (AN) prill will be shipped to site with 40-tonne B-train trucks and stored in 80-tonne silos, one located at the CP area and one at the N-204 area. Bulk emulsion will be shipped to site in 20-tonne tanker trucks and stored in 60-tonne silos, one located at the CP area and one at the N-204 area.

Packaged explosives and explosive detonators will be stored in approved explosive magazines located on separate pads. There will be two packaged explosives magazines each with the capacity to store 20,000 kg of explosives.

The design of all storage facilities will meet government regulations and will be located according to required separation distances as regulated by the Explosives Regulatory Division (ERD) of Natural Resources Canada (NRC).

The final location of the explosives storage site would be determined as part of future prefeasibility or feasibility studies.

18.6 Mine Water Management 18.6.1 Hydrogeological Summary

Groundwater conditions at the Pine Point Mine are reasonably well documented owing much to the historic open pit dewatering during mining operations. The understanding of the site hydrogeology is developed from summaries on the groundwater flow regime by Geologic Testing Consultants (GTC;

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DARNLEY BAY RESOURCES LTD. PINE POINT PEA

1983) and Stevenson International Groundwater Consultant Ltd. (Stevenson; 1984), and summaries on the historic dewatering efforts by Vogwill (1976, 1981) and Durston (1979).

18.6.1.1 Regional Hydrogeology On a regional scale, groundwater flow originates in topographic highs (recharge areas) such as the Caribou Mountains located 200 km south of the Pine Point area as shown on Figure 18.2. The Caribou Mountains rise approximately 600 m above the surrounding land surface and groundwater flow is more or less radially distributed from the mountains. Groundwater flow in a northerly direction from the mountains towards the north discharges primarily to the Hay River valley to the northwest, Great Slave Lake to the north, and the Little Buffalo River and Slave River valleys to the northeast.

The area surrounding Great Slave Lake represents a lowland and is considered a major groundwater discharge area. Groundwater discharge is evident as sulphurous springs and as areas of high specific conductance in rivers.

Effective Date: April 18, 2017 18-6

500,000 600,000 700,000

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PINE POINT

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88

697 300 20 10 0 20 30040 60 80 100 km 500

500,000 300 600,000 700,000 600 300 600 LEGEND: DARNLEY BAY RESOURCES LIMITED REGIONAL GROUND WATER FLOW DIRECTIONS PROJECT LOCATION PINE POINT PROJECT RIVER/STREAM NOTES: ROAD LAKE 1. BASE MAP: TOPORAMA, NTS. PINE POINT AREA LIMITED-USE ROAD PROTECTED AREA 2. COORDINATE GRID IS IN METRES. REGIONAL PLAN HYDROGEOLOGY COORDINATE SYSTEM: NAD 1983 UTM ZONE 11N. CONTOUR 100 M 3. THIS FIGURE IS PRODUCED AT A NOMINAL SCALE OF 1:1,000,000 P/A NO. REF NO. CONTOUR 20 M FOR 11x17 (TABLOID) PAPER. ACTUAL SCALE MAY DIFFER VA101-60/6 VA17-895 0 26MAY'17 ISSUED WITH TRANSMITTAL KK KK CAS ACCORDING TO CHANGES IN PRINTER SETTINGS OR REV PRINTED PAPER SIZE. REV DATE DESCRIPTION DESIGNED DRAWN REVIEWED 0

SAVED: M:\1\01\00060\06\A\GIS\Figs\VA17-00895 Regional Hydrogeology\Fig18-2_RegionalHydrogeology.mxd; May 26, 2017 9:27 AM; kkrauszovaAM; 9:27 Hydrogeology\Fig18-2_RegionalHydrogeology.mxd;26, May 2017 Regional M:\1\01\00060\06\A\GIS\Figs\VA17-00895 SAVED: FIGURE 18.2 DARNLEY BAY RESOURCES LTD. PINE POINT PEA

18.6.1.2 Site Hydrogeology Local groundwater recharge to the bedrock aquifer at the Pine Point site is likely to be variable and largely controlled by the overburden geology. High rates of recharge are expected in areas where sinkholes are present, whereas recharge generally will be limited by the presence of till overburden. Recharge to groundwater is estimated at 10% of the 34 cm annual precipitation (Stevenson, 1984).

The most productive aquifer at Pine Point is the dolomitized Presqu’ile barrier complex within the Pine Point Group. The dolomitized Presqu’ile unit comprises a major aquifer and consists of highly porous, well-fractured dolomite with paleo-karst cavity networks. Skall (1975) describes the dolomitized reef unit as approximately 65 m thick along the Main Hinge and 20 m thick along the North Hinge. Groundwater within the saturated bedrock is expected to flow along solution channels, bedding planes and fractured zones. According to Stevenson (1984), the major Presqu’ile aquifer is laterally confined along hinge zones by the Buffalo River shales to the north and the Muskeg evaporites to the south. Overlying clay till overburden and Watt Mountain argillaceous limestones of generally low permeability act to confine the aquifer on top. Fine-grained dolomites of the lower Pine Point Group and the Keg River Member underlie the coarsely-crystalline Presqu’ile dolomite and comprise a minor aquifer thought to consist primarily of fracture permeability. The Chinchaca Formation evaporites underlie the minor aquifers and form an effective lower barrier to the bedrock aquifer. A simplified stratigraphy relating geology to hydrogeology (for the A70 Pit area in the northwest portion of the Pine Point area) developed by Vogwill (1976) is shown on Figure 18.3.

A shallow water table is present in the glacial till overburden at depths between 7 to 30 m across the site (Vogwill, 1981). The till unit has a relatively low permeability and negligible communication between the upper and lower water bearing zones was historically observed. As a result, groundwater in the overburden is not expected to significantly influence dewatering requirements.

Properties of the aquifers have been extensively documented as a result of the historic dewatering activity. The permeability of the Presqu’ile aquifer is very high with transmissivities in the order of 1x10-2 m2/s (GTC, 1983). Historic aquifer testing has shown that the bedrock permeability can be anisotropic, particularly near the hinge zones. The hinge-lines are lines of weakness in the bedrock where preferential karstification and coarse-crystalline dolomitization are concentrated and hydraulic continuity is more predominant along the trend of the hinge-lines (GTC, 1983). The recognition of the anisotropic permeability should permit a more efficient placement of dewatering wells. Similarly, the close proximity of the W-85 and X-65 deposits to the northern barrier of the Presqu’ile aquifer may limit groundwater inflow from the north into the developed open pits.

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Figure 18.3: Geologic and Hydrogeologic Stratigraphy of the Presqu’ile Barrier Reef Complex at the A70 Pit Area

Source: KP (2017)

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Based on work completed by Stevenson (1984), the groundwater hydraulic gradient in the Pine Point area is generally northwards towards Great Slave Lake while to the south of the Pine Point area, the gradient trends from west to east (Figure 18.4). Interpretation of the groundwater contours in relation to the topography indicates that the groundwater level is up to 30 m below ground along the northeastward trending ridge in the east-central part of the site. In the northwest portion of the site, the piezometric surface is higher than the ground surface. High piezometric levels have resulted in groundwater discharge as springs along the incised Buffalo River channel and other small tributary channels in the area, as well as conditions of leaky artesian flow from the bedrock units (GTC, 1983).

Although the Presqu’ile aquifer has a high permeability, the flow through it is thought to be slow due to the low gradient in the Pine Point area. Based on preliminary calculations taking into account the documented transmissivity values, gradients and geological sections, the flow velocity is estimated to be less than 1 m/day. Due to the high porosity, the storativity of the aquifer is quite high. It is estimated that about 1 billion cubic meters of water was removed during mining activities from 1968 to 1984. According to Stevenson (1984), this included water stored within the aquifer (16%), recharge from local precipitation (76%), with the remainder from the regional groundwater flow.

Effective Date: April 18, 2017 18-10

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Great Slave Lake N204

6,760,000 6,760,000

P24 X15 X53 K32 Y54 K35 X59 Z53 O32 W17 L35 N34 N32 Z61 Z57 A55 N33 O33 O30 X65 L37 P31 6,750,000 66-17 I46 N37 P32 6,750,000 J44 M40 X68 Z-64 N36 N35 AIR K48 N45 N42 O37 STRIP O42 A70 64-47 K53 P41 V-90 W85 O44 O45 K62 K57 I65 O52 HZ Pine Point K66 O53 K60 J68 L65 M62/63 J69 M-64 K68 M67 T-58 K-77 M57 65-15 66-12 S-65 R-61 R-57

6,740,000 N-81 6,740,000

G-03 6 H-02 V-46 W-19 ZI-55 X25

O-556 R190 POPLAR P-499 LAKE

6,730,000 6,730,000 600,000 610,000 620,000 630,000 640,000 650,000 660,000

LEGEND: NOTES: DARNLEY BAY RESOURCES LIMITED 2 1 0 2 4 6 8 10 km HISTORIC OPEN PIT/DEPOSIT GROUNDWATER CONTOUR (SEE NOTE 4) 1. BASE MAP: NTS. PINE POINT PROJECT PLANT SITE GROUNDWATER FLOW DIRECTION 2. COORDINATE GRID IS IN METRES. RIVER/STREAM/DRAINAGE COORDINATE SYSTEM: NAD 1983 UTM ZONE 11N. PROPOSED DEPOSIT PINE POINT AREA WATER 3. THIS FIGURE IS PRODUCED AT A NOMINAL SCALE OF 1:175,000 6 TERRITORIAL HIGHWAY FOR 11x17 (TABLOID) PAPER. ACTUAL SCALE MAY DIFFER GROUNDWATER GRADIENT PLAN ROAD ACCORDING TO CHANGES IN PRINTER SETTINGS OR PRINTED PAPER SIZE. RAIL BED (TRACK REMOVED) P/A NO. REF NO. CONTOUR 20 M 4. HISTORICAL SITE LOCATIONS REFERENCED FROM VA101-60/6 VA17-00887 KK CAS 0 23MAY'17 ISSUED WITH TRANSMITTAL KK STEVENSON INTERNATIONAL GROUNDWATER CONSULTANTS LTD's REV REV DATE DESCRIPTION DRAWN REVIEWED REPORT OF THE HYDROGEOLOGY OF R190 MINERALIZED REGION (1983). SAVED: M:\1\01\00060\06\A\GIS\Figs\VA17-00887\Fig18-4_GroundwaterGradientPlan_r0.mxd;5:38 23, PM;May 2017 kkrauszova SAVED: DESIGNED FIGURE 18-4 0 DARNLEY BAY RESOURCES LTD. PINE POINT PEA

18.6.2 Water Management Plan

Site water management at the Pine Point Mine has historically been an important aspect of mine operations, which will also be the case for the proposed mine development plan. The substantial historic site-specific data set related to pit dewatering and pumping tests provided a unique opportunity to calibrate an appropriate pit dewatering and management system for the Project.

Well-documented summaries of the challenges and successes of past dewatering methodologies are provided in Vogwill (1976) and Durston (1979). A summary of dewatering efforts compiled by Durston (1979) indicates that an average of 50 to 60 wells were operated per year in five or six pits in the years preceding 1979, and average well discharge had reached a reported 60 million US gallons per day. Historic experience indicates most Pine Point wells efficiently yielded 1,000 to 1,200 US gpm and an optimum pump size was in the range of 100 to 125 hp. Approximately 80% of the wells yielded 800 US gpm or higher. Dewatering wells were advanced using 14¾” diameter drill holes through the bedrock with no casing or screen installed. Dewatering created regional drawdowns oriented along an east-west direction, which resulted in additional dewatering at adjacent pits (Vogwill, 1976).

For the PEA mine plan, water that will be managed includes:

 Groundwater o Pit Dewatering (as described above) – perimeter wells / in-pit sumps;  Surface Water o Waste Rock Storage Areas (WRSAs) – contact water; and o Pit and WRSA upstream catchments – non-contact water.

The concept developed to manage pit dewatering and Waste Rock Storage Area (WRSA) contact water involves disposal in existing open pits or through well injection. An alternate method of water management that will be considered during optimization studies during the next level of design includes overland disposal of pit dewatering and contact water. Water will be conveyed by gravity from each of the proposed pits to the final disposal locations through surface ditches, many of which are already in place from the previous mine operations. Table 18.1 identifies possible methods for water disposal for each of the proposed deposits.

Pit dewatering and WRSA contact water from deposits J68, HZ, M67, O53, K68, L65A, L65B and M62/63 could be conveyed by ditches and disposed of in the historic A70 pit, which is located downgradient of these deposits. Pit dewatering and WRSA contact water from the N-204, W85 and X65 deposits could be conveyed by ditches to downgradient areas where groundwater well injection fields will be developed.

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Table 18.1: Water Management Summary

Estimated Historic Pit Groundwater Well Injection Field YEAR PIT Number of Dewatering Wells A70 W85 / X65 N-204 -1 J68 9 1 HZ 20 1 – 2 W85A 16 2 – 3 W85B 13 3 – 6 X65A 15 7 X65B 7 8 – 9 X65C 18 3 – 7 N-204A 8 5 – 9 N-204B 8 10 – 11 N-204C 8 12 – 13 N-204D 4 9 – 10 M67 12 10 O53 9 10 – 11 K68 21 11 – 12 L65A 11 11 – 12 L65B 14 13 M62/63 19 Source: KP (2017)

Similar to pit dewatering techniques utilized during past mine operations at Pine Point, pit dewatering will primarily occur through the installation of perimeter dewatering wells. Surface ditches around each pit will be constructed for two purposes:

 Convey pit dewatering water to their final disposal location, and  Divert upstream catchment surface water runoff from entering the pit, thereby reducing the volume of dewatering water.

The depth of the pit dewatering wells will vary depending on a number of factors, including the local groundwater regime in the area of each deposit, as well as the depth of each pit. However, based on past site experience, the wells will be drilled to approximately 60 m below the bottom elevation of each pit. As a result, the average depth of perimeter pit dewatering wells for each of the proposed pits is approximately 130 m below ground surface (BGS). The dewatering wells will be drilled with 15” diameter holes, with pumping capacity to dewater at a rate of 230 m3/hr (1,000 US gpm). Table 18.1 also presents the estimated number of wells for each of the proposed pits.

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18.7 Waste Rock Management Waste rock and reject from the Dense Media Separation (DMS reject) will be managed in the WRSAs. The majority of the waste rock and DMS reject will be back-filled into historic or proposed open pits (Pit Backfill WRSAs). The remaining waste rock and DMS reject will be placed in external stockpiles (External WRSAs) adjacent to the nearest proposed pits. The waste rock and DMS reject allocation for the Project is summarized in Table 18.2 with the capacity of each facility shown in Table 18.3.

The overall final slope of the External WRSAs will be established at 2.5H:1V to facilitate reclamation. Stability assessments will be completed as part of future studies.

Table 18.2: Waste Rock and DMS Reject Storage Summary

EXTERNAL PIT BACKFILL WRSA WRSA

PIT (kt) YEAR WASTE 204A 204A 204B J69 K68 K57 K68 O52 M64 - - - L65A X65A X65A X65B W85A N N N W85A/B WASTE TYPE WR 385 -1 J68 DMS 197 WR 6,271 1 HZ DMS 1,026 WR 5,125 1 – 2 W85A DMS 1,094 WR 1,765 2 – 3 W85B DMS 1,757 WR 3,624 3 – 6 X65A DMS 2,317 WR 1,517 7 X65B DMS 776 WR 6,014 8 – 9 X65C DMS 779 WR 18,890 3 – 7 N-204A DMS 4,301 WR 18,683 5 – 9 N-204B DMS 2,694 WR 5,597 10 – 11 N-204C DMS 1,013 WR 10,918 12 – 13 N-204D DMS 2,044 9 – 10 M67 WR 9,783

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EXTERNAL PIT BACKFILL WRSA WRSA

PIT (kt) YEAR WASTE 204A 204A 204B J69 K68 K57 K68 O52 M64 - - - L65A X65A X65A X65B W85A N N N W85A/B WASTE TYPE

DMS 943 WR 921 10 O53 DMS 141 WR 7,431 10 – 11 K68 DMS 837 WR 6,146 11 – 12 L65A DMS 747 WR 3,739 11 – 12 L65B DMS 378 WR 3,628 13 M62/63 DMS 247 Note: WR = Waste Rock; DMS = DMS Reject Source: KP (2017)

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Table 18.3: Capacity of Waste Rock Storage Facilities

WRSA Waste (kt) INPITJ69 10,168 INPITK57 6,271 INPITM52 921 INPITM64 3,628 Sub-total Historic In-pit 20,988 INPITK68 6,146 INPITL65A 3,748 INPITN-204A 18,683 INPITN-204B 16,514 INPITW85A 1,836 INPITX65A 1,517 INPITX65B 6,014 Sub-total Proposed In-pit 54,458 K68WRSA 7,431 N-204AWRSA 18,890 W85AWRSA 5,125 W85BWRSA 6,672 X65AWRSA 3,635 Sub-total External WRSA 41,753 Grand Total 117,198 Source: JDS (2017)

18.8 Tailings Management Facility Given there are many historic pits at the Pine Point area from past mine operations, as well as more than twenty new open pits within the ten deposits for the proposed mine development plan, there are many opportunities to manage tailings in pits, negating the need to construct a separate surface tailings management facility. Tailings slurry will be generated at a solids content of 25% (by weight) with an assumed dry-density of 1.3 t/m3 and a daily throughput of 1,800 tonnes per day. Pumping will be required to convey the tailings slurry through a pipeline along a generally flat gradient for distances up to 2.5 km from the process plant.

Three Pit Backfill Tailings Management Facilities (PBTMFs) have been identified to manage the tailings for the life of mine for the proposed project. Table 18.4 presents the PBTMF designation, the pit name and the approximate timing for when each pit will be actively receiving tailings for disposal.

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Table 18.4: Pit Backfill Tailings Management Facilities

Year Designation Historic Or Proposed Pit 1 – 2 PBTMF 1 K62 (historic) 3 – 9 PBTMF 2 HZ (proposed) 10 – 13 PBTMF 3 M67 (proposed) Source: KP (2017)

The historic K62 pit will be used for tailings disposal for the first two years of mine operations, with the designation as PBTMF1. Starting in Year 3 and continuing for seven years, the new HZ pits will be used to manage tailings, with the designation as PBTMF2. Finally, starting in the first few months of Year 10 and continuing until the end of the mine life, tailings will be managed at PBTMF3 within the new M67 pit, once mining ceases from this pit. Figure 18.5 presents the locations for the three proposed PBTMFs.

Figure 18.5: Pit Backfill Tailings Management Facilities

Source: KP (2017)

Reclaim water will be sourced from groundwater wells in the vicinity of the process plant. Thus, there will be no reclaim pumping and pipeline system from the PBTMFs to the process plant. Given the high groundwater flow rates that exist at the Pine Point site, past experience has shown that all pits will fill with water over a short timeframe once pit dewatering ceases. For the proposed pits, once the pits have filled with water and/or tailings (in the case of the PBTMFs), no discharge to the surface environment is expected, similar to the current conditions of the historic pits, as noted during the March 2017 site visit.

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19 Market Studies and Contracts

19.1 Market Studies No preliminary market studies on the potential concentrate sales from the Pine Point Project were completed. The terms selected for the economic analysis are considered to be in line with the current market conditions. The indicative terms were reviewed and found to be acceptable by QP Garett Macdonald, P.Eng. The concentrate will be trucked to the Hay River railway facility approximately 65 km west, and from there it will be shipped to markets. The study recommends that as the Project advances towards development, a detailed marketing report and logistics study should be undertaken to ensure the accuracy of the terms. Table 19.1 outlines the terms used in the economic analysis.

Table 19.1: Concentrate Terms

Assumptions & Inputs Unit Value Lead Concentrate Metal Recovery to Pb Con (CP) % 85.0 Metal Recovery to Pb Con (N-204) % 71.4 Average Pb Recovery to Concentrate % 80.4 Average Pb Concentrate Grade Produced % 65.0 Minimum Deduction % Pb/t 3.0 Pb Payable % 95.0 Pb Treatment Charge C$/dmt conc. 200.00 Pb Concentrate Transportation Cost C$/wmt 167.00 Zinc Concentrate Metal Recovery to Zn Con (CP) % 88.1 Metal Recovery to Zn Con (N-204) % 78.0 Average Zn Recovery to Concentrate % 83.4 Average Zn Concentrate Grade Produced % 58.9 Minimum Deduction % Zn/t 8.0 Zn Payable % 85.0 Zn Treatment Charge C$/dmt conc 200.00 Zn Concentrate Transportation Cost C$/wmt 167.00 Source: JDS (2017)

19.2 Contracts No contractual arrangements for concentrate trucking, port usage, shipping, or treatment exist at this time. Furthermore, no contractual arrangements have been made for the lead concentrate or zinc concentrate at this time.

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19.3 Royalties The Pine Point Project is subject to a royalty payment of 3.0% NSR based on the sale of minerals payable to Karst Investments LLC. The Project is not subject to any other royalties, back-in rights, payments, or other agreements or encumbrances other than the territorial royalty (calculated as a tax but called a royalty).

This financial commitment is included in the cash flow. Total third party royalties for the Project amount to $62.5 M over the LOM.

19.4 Metal Prices The precious metal markets are highly liquid and benefit from terminal markets around the world (London, New York, Tokyo, and Hong Kong). Historical lead and zinc prices are shown in Figure 19.1 and Figure 19.2. Historical exchange rate trends are plotted in Figure 19.3.

Figure 19.1: Historical Lead Price

Source: London Metals Exchange (2017)

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Figure 19.2: Historical Zinc Price

Source: London Metals Exchange (2017)

Figure 19.3: Historical US$:C$ F/X Rates

Source: Bank of Canada (2017)

The lead price and exchange rate selected was based on the 6-month trailing average as at March 2017. The zinc price selected was based on the 12-month trailing average as at March 2017. These parameters are in line with recently released comparable Technical Reports.

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A sensitivity analysis on metal prices and exchange rates was completed as part of the overall economic analysis. The results of this are discussed in Section 23. Table 19.2 outlines the metal prices used in the economic analysis.

It must be noted that metal prices are highly variable and are driven by complex market forces and are extremely difficult to predict.

Table 19.2: Metal Price and Exchange Rate

Parameter Unit Value Lead Price US$/lb 1.00 Zinc Price US$/oz 1.10 Exchange Rate US$:C$ 0.75 Source: JDS (2017)

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20 Environmental Studies, Permitting and Social or Community Impact

20.1 Environment The PEA shows the Pine Point Project has key features which will function to minimize environmental impacts, namely:

 No tailings pond, no tailings dam: the tailings management concept that is developed for the Pine Point Project is based on “in-pits” disposal (refer to Section 18.7- Tailings Management); as the Project does not include a tailings pond facility with dams, this approach thus eliminates the geotechnical risks associated with long term operation of tailings dams, while also eliminating the need for long term treatment and discharge of tailings pond water;  Minimal surface discharge of mine dewatering water: most water from the dewatering of the open pits will be returned, by gravity in most instances, to adjacent historical open pits or recently mined pits (refer to Section 18.6 – Mine Water Management);  No discharge of liquid effluents from the process plant: the process plant will operate under a zero discharge operating plan whereby water will recycled/reused from a basin, available or constructed (refer to Section 17.);  Geochemistry of waste rock and DMS rejects: Previous available results from geochemistry tests by Tamerlane where reviewed by the consultant “pHase Geochemistry” as part of the PEA. Phase Geochemistry concluded that the available results (static ABA) bring evidence that the waste rock samples tested based on the Tamerlane mine plan at the time were non-PAG. However, pHase Geochemistry concluded that the representativeness of the samples tested at this point will need further evaluation at the Feasibility Study level as a function of for the proposed mine plan under the PEA  Small footprint: The management concept for the Waste Rock and the DMS rejects is to minimize the surface storage footprint.

20.2 Baseline Studies The Pine Point Project benefits from considerable previous baseline work and studies (published and unpublished) by Cominco, Tamerlane, Avalon, Government Agencies, and others. As part of the PEA, a review was completed of published and unpublished recent studies that have been conducted across the Pine Point District to document: archaeological resources, aquatic resources (including water quality, fisheries, wetlands, and habitat), air quality, terrestrial resources (including upland habitat, vegetation, and wildlife), hydrology and hydrogeology. Traditional Knowledge and Socioeconomic studies have also been completed.

The above baseline work will provide the Project with solid support for permitting.

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20.3 Closure The mine Closure Plan to be developed at the FS step will follow the “MVLWB/AANDC Guidelines for the closure and reclamation of advanced mineral exploration and mine sites in the Northwest Territories (2013)”. At the level of the PEA, an order of magnitude closure cost has been allocated for the demolition of the process plant building and associated facilities, removal of equipment and safe disposal of demolition debris, structural steel, etc.

20.4 Permitting Development of open pit mines at Pine Point will require that Darnley Bay obtain a new Land Use Permit and a new Water License from the Mackenzie Valley Land and Water Board (MVLWB). This permitting is subject to review and approval by the MVLWB and the federal department of Aboriginal Affairs and Northern Development Canada (AANDC). It is likely that the MVLWB will require that the Project complete an Environmental (Impact) Assessment process with the Mackenzie Valley Environmental Impact Review Board (MVEIRB) before they make any decision related to the land use permit and water license applications. Both the MVLWB and the MVEIRB have public consultation and comment periods built into their processes. It is typical for it to take two to three years (or more) for a new applicant to obtain the necessary environmental permits for a mining Project in the NWT. It is possible that permits for the Pine Point deposits described in this report can be obtained in less than two years due to the following advantages:

 Much of the new development will use existing infrastructure (haul roads, quarry pits, etc.) which minimizes impacts;  Most, if not all, of the background data collection is completed;  Some of the issues of concern to the Metis and First Nations have been addressed in this current mine plan; and  Some of the processing steps to be advanced with this Project have already been pre- screened. A variety of additional permits and licenses will be required from local, territorial and federal agencies once the Project is financed and final construction plans are complete. These vary from business license, to explosive magazine license, to electrical permit, to health and safety permits among others.

20.5 Social Engagement The Mackenzie Valley Land and Water Board requires a formal public engagement process be observed as part of the land use and water license application processes. Darnley Bay has started its engagement activities in the Pine Point District. Initial information about the Project has been shared with First Nations (Metis, Akaitcho and DehCho Councils and Nations) as well as with municipal governments. Active engagement will continue throughout the project planning process as well as over the life of the entire Project.

The land that Darnley Bay has leased is Crown Land; it does not belong to an Aboriginal group. There are three Aboriginal groups with (unsettled) land claims in the Pine Point area: the Katlodeeche First Nation, the Deninu Kue First Nation, and the Northwest Territory Metis Nation.

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Darnley Bay has committed to entering into negotiations for Impact Benefit Agreements (IBA) with each Nation for future phases of the mining activities at Pine Point. IBAs are developed through a negotiation process that typically lasts 6 – 12 months. The negotiations can occur before, during or after the permitting process; it is typical for the negotiations to be completed sometime during the permitting process at which time the First Nations will then publicly support the Project.

The most common terms in IBAs are cash and/or stock payments; employment and/or contracting commitments; and education and training opportunities for Band/Nation members. Unemployment rates are high in the Hamlet of Fort Resolution (where most of the population is Aboriginal) and on the Katlodeeche FN’s Hay River Reserve.

Residents of these communities expressed support for more and better job opportunities at the public hearings held for Tamerlane’s R190 pilot project. Darnley Bay has proactively involved local residents and businesses (including Aboriginal-owned enterprises) in jobs and other economic opportunities in the on-going exploration program. It is expected that Darnley Bay will be able to build upon these connections in future phases of the Project, including mine construction, operations, and closure.

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21 Capital Cost Estimate

21.1 Capital Cost Summary Estimated project capital costs total $271 M, consisting of the following distinct stages:

 Initial Capital Costs – includes all pre-production costs to develop the cluster pit resources to a 1,800 t/d operation. Initial capital costs total $154 M and are expended over a 24-month pre- production construction and commissioning period;  Sustaining Capital Costs – includes all costs related to the acquisition, replacement, or major overhaul of assets during the mine life required to sustain operations. This includes the DMS plant and infrastructure required to develop the N-204 resource. Sustaining capital costs total $100 M and are expended in operating Year 1 through 13; and  Closure Costs – includes all costs related to the closure, reclamation, and ongoing monitoring of the mine post operations. Closure costs total $17 M, and are primarily incurred in Year 14, with costs extending into Year 18 for ongoing monitoring activities.

The capital cost estimate was compiled using a combination of quotations, database costs, and database factors.

Table 21.1 presents the capital estimate summary for initial, sustaining, and closure capital costs in Q1 2017 Canadian dollars with no escalation.

Table 21.1: Capital Cost Summary

Sustaining/ Pre-Production Total Capital Costs Closure (C$M) (C$M) (C$M) Open Pit Mining 17.1 14.4 31.5 Mine Water Management 2.7 17.0 19.8 Site Development 3.6 2.7 6.3 Processing 52.7 29.0 81.6 Infrastructure 24.8 15.8 40.6 Indirect & Owner’s Costs 32.8 9.9 42.7 Closure - 17.5 17.5 Subtotal 133.7 106.3 240.1 Contingency @15% 20.1 11.2 31.2 Total Capital Costs 153.8 117.5 271.3 Source: JDS (2017)

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21.2 Capital Cost Profile All capital costs for the Project have been distributed against the development schedule in order to support the economic cash flow model. Figure 21.1 presents an annual LOM capital cost profile (excluding closure years).

Figure 21.1: Capital Cost Profile (Excluding Closure Years)

120 300 109 250 252 253 254 239 242 243 100 232 235 238 250

189 80 178 184 200 154 60 150 45 42 40 100

Period Capital Cost (C$ M) 24

45 Cumulative Capital Costs (C$ M) 20 50 7 6 5 4 3 0 3 1 2 2 0 - - -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 Operating Year

Source: JDS (2017)

21.3 Key Estimate Assumptions The following key assumptions were made during development of the capital cost estimates:

 Open pit mine development activities will be performed by an Owner-operated team;  The primary mining fleet will be leased at the onset of the Project. A down payment of 15% of the purchase price of the equipment has been included in pre-production or sustaining capital. Monthly payments have been considered an operating cost; and  All surface construction (civil, structural, architectural, mechanical, piping, electrical, and instrumentation) will be performed by contractors under the management of an Engineering, Procurement & Construction Management (EPCM) contractor.

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21.4 Key Estimate Parameters The following key parameters apply to the capital cost estimates:

 Estimate Class: The capital cost estimates are considered Class 4 estimates (-20%/+30%). The overall project engineering definition is estimated to be 10%;  Estimate Base Date: The base date of the estimate is April 18, 2017. No escalation has been applied to the capital cost estimate for costs occurring in the future;  Units of Measure: The International System of Units (SI) is used throughout the capital estimate; and  Currency: All capital costs are expressed in Canadian Dollars (C$). Portions of the estimate were estimated in US Dollars (US$) and converted to Canadian Dollars at a rate of US$0.75:C$1.00.

21.5 Basis of Estimate 21.5.1 Labour Rates

The majority of installation costs within the estimate have been factored based on mechanical equipment costs. Where applicable within the estimate, an average all-in contractor crew labour-rate of $95/h has been applied, based on buildups from other recent and similar studies.

Operational labour rates were built up from first principles. Base rates are based on prevailing wages in the area, and legal premiums and benefits were built up to create all-in rates. Operational labour rates and staffing levels are described further within Section 22.

21.5.2 Fuel & Energy Supply

Where applicable, a delivered fuel price of $0.88/L has been used throughout the estimate. Power energy supply has been estimated at $0.11/kWh for grid power and $0.14/kWh for LNG generated power.

21.5.3 Mine Capital Cost Estimate

Capital cost estimates are based on a combination of budgetary quotes from equipment suppliers, in-house cost databases and from similar mines in Northern Canada. Table 21.2 summarizes the mining capital cost estimate.

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Table 21.2: Mining Capital Costs

Pre-Production Sustaining Total Capital Costs (C$M) (C$M) (C$M)

Pre-Stripping 4.8 - 4.8 Mining Equipment 12.3 14.4 26.7 Mine Dewatering 2.7 17 19.7 Total Mining Capital Cost 19.8 31.4 51.2 Source: JDS (2017)

21.5.3.1 Pre-Stripping Pre-stripping costs include all labour and consumables related to pre-production waste stripping from the J68 pit. Costs were assembled from first principles using the open pit mining schedule as the basis. Database unit costs were applied to labour, equipment, and material requirements.

Stripping of all other cluster pits and the N-204 deposit are included as an operating cost within the mine cost model.

21.5.3.2 Mining Equipment Open pit mining equipment quantities were determined through buildup of mine plan quantities and associated equipment utilization requirements. Database unit prices were applied to the required quantities. The mining equipment is assumed to be leased over a period of 60 months at a lease rate of 4.6%. A 15% initial down payment is included in the pre-production and sustaining capital costs. Monthly lease payments for the equipment are included as an operating cost at an annual percentage rate (APR) of 4.6%.

21.5.3.3 Mine Dewatering Mine dewatering costs include the following:

 Drilling of perimeter wells around each pit;  Supply and installation of perimeter well pumps;  Surface dewatering ditches and ponds;  External pit pump stations and pipelines; and  Injection well fields, where applicable.

21.5.4 Site Development & Earthworks

Site development costs are generally based on high level material take-offs and database unit pricing.

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21.5.5 Processing Cost Estimate

The process plant capital costs include all of the direct costs to construct the 1,800 t/d flotation process plant, the 5,700 t/d DMS plant for the cluster pit area and the 7,300 t/d DMS plant for the N- 204 area.

The process plant capital cost estimate was assembled form a combination of supplier quotations and database allowances. Table 21.3 presents a summary basis of estimate for the various commodity types within the process plant estimate.

Table 21.3: Process Plant Basis of Estimate

Commodity Estimate Basis Equipment Budget quotations were solicited from qualified Major Equipment suppliers for the major equipment identified in the flow sheets and equipment register. In-house data (firm and budgetary quotations from Minor Equipment recent projects) was used for minor or low value equipment. Installation (Labour & Materials) High level take-off quantities were developed from general arrangement drawings and database Concrete concrete quantity ratios per facility area (m3/m2). Database unit rates were applied to the take-off quantities. Internal Structural Steel Factored based on mechanical equipment costs. Database unit costs ($/m2) applied to areas determined from the general arrangement drawings. Process Plant Building Pre-engineering steel buildings were assumed for the process area building. Lump sum allowances included for modular control and lunch rooms. Mechanical Equipment Installation Factored based on mechanical equipment costs. Piping Factored based on mechanical equipment costs. Electrical & Instrumentation Factored based on mechanical equipment costs. Source: JDS (2017)

21.5.6 On-Site Infrastructure

The on-site infrastructure estimate at the Pine Point Project includes an assay laboratory, permanent accommodation complex, administration complex, on-site power distribution, water and waste handling infrastructure, the surface mobile support fleet, and information technology (IT) and communications systems. Table 21.4 presents a summary basis of estimate for the various commodity types within the process plant estimate.

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Table 21.4: Infrastructure Basis of Estimate

Component Estimate Basis Factored database costs, based on the Accommodations, Administration Complex, and Assay accommodations, dry, and office requirements Laboratory determined for the operations and construction phases. Database unit costs ($/m2) applied to areas determined from the general arrangement drawings. Pre-engineered Maintenance Facilities steel buildings were assumed for the maintenance buildings. Site utilities include on-site power distribution, secondary and emergency power generation, a chlorinator water Site Utilities treatment plant and incinerator. Lump sum allowances have been applied to these facilities based on experience at similar operations. Surface equipment fleet requirements are determined based on material movement requirements and experience at similar operations, and considering site conditions specific to the Project. No equipment Surface Mobile Equipment replacements are anticipated for the surface equipment fleet due to the short mine life and relatively low utilization of equipment. Database unit pricing has been applied to the surface equipment fleet quantities. Lump sum allowance based on other recently Bulk Fuel Storage & Distribution constructed facilities. Lump sum allowances based on experience at similar Explosives Management Facilities operations. Lump sum allowances based on experience at similar IT & Communications operations. Source: JDS (2017)

21.6 Indirect Cost Estimate Indirect costs are those that are not directly accountable to a specific cost object. Table 21.5 presents the basis of estimate for each of the indirect cost categories. The majority of indirect costs in the estimate are factors or allowances based on recently completed definitive estimates for similar projects.

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Table 21.5: Indirect Cost Basis of Estimate

Commodity Basis Time based cost allowance for general construction site services (temporary power, heating & hoarding, contractor Construction Support Services support, etc.) applied against the surface construction schedule Allowance for construction offices and ablution facilities Temporary Facilities & Utilities Allowance for diesel construction power Factored allowance (1.0%) of direct costs for contractor Contractor Indirect Costs facilities and auxiliary expenses Factored allowance (10.0% of direct equipment costs and Logistics & Freight 5% for all direct material costs) for all freight and logistics. Factored allowance (2.0%) for spare parts Start-up and Commissioning Factored allowance (1.0%) for the provision of vendor services for commissioning support Factored allowance for Engineering and Procurement Civil 5% of total direct site development costs

EPCM Process 7% of total direct process costs Infrastructure 5% of total direct infrastructure costs Factored allowance (7%) of total direct costs (excluding mining) for Construction Management. Source: JDS (2017)

21.7 Owners Cost Estimate Owner’s costs are items that are included within the operating costs during production. These items are included in the initial capital costs during the construction phase and capitalized. The cost elements described below are described in more detail within Section 22.

 Pre-production processing: costs of the Owner’s processing labour, power, and consumables incurred before declaration of commercial production; and  Pre-production general & administration: costs of the Owner’s labour and expenses (camp and catering, safety, finance, security, purchasing, support labour, maintenance, equipment usage, management, etc.) incurred prior to commercial production.

21.8 Closure Cost Estimate Closure costs have been estimated based on the typical closure, reclamation, and monitoring activities for an open pit mine. Typical activities include:

 Removal of all surface infrastructure and buildings;  Access and haul road closure;  Power transmission line and substation removal;  Re-vegetation and seeding; and

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 Ongoing site monitoring. A total lump sum closure cost of $17.5 M has been used for the estimate, based on factored costs from similar projects.

21.9 Contingency An overall contingency of 15% was applied to the LOM capital costs of the Project. The LOM Project contingency amounts to $31.2 M.

21.10 Capital Estimate Exclusions The following items have been excluded from the capital cost estimate:

 Working capital (included in the financial model);  Financing costs;  Currency fluctuations;  Lost time due to severe weather conditions beyond those expected in the region;  Lost time due to force majeure;  Additional costs for accelerated or decelerated deliveries of equipment, materials or services resultant from a change in project schedule;  Warehouse inventories, other than those supplied in initial fills, capital spares, or commissioning spares;  Any project sunk costs (studies, exploration programs, etc.);  Costs arising from the establishment of Impact Benefit Agreements;  Territorial or Federal sales tax;  Closure bonding; and  Escalation cost.

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22 Operating Cost Estimate

The OPEX estimate is based on a combination of experience, reference projects, budgetary quotes and factors as appropriate with a preliminary study.

The operating cost estimate in this study includes the costs to mine, process the mineralized material to produce a lead concentrate and a zinc concentrate, and general and administrative expenses (G&A). These items total the mine operating costs and are summarized in Table 22.1. The target accuracy of the operating cost is -25/+30%.

The operating cost estimate is broken into three major sections:

 Open Pit Mining;  Processing; and  General & Administrative.

The total operating unit cost is estimated to be $37.64/t processed (or $38.03/t processed excluding pre-production tonnes). Total LOM and unit operating cost estimates are summarized in Table 22.1. Figure 22.1 illustrates the operating cost distribution.

Operating costs are expressed in Canadian dollars. No allowance for inflation has been applied. This estimate is based on leasing the mine production equipment.

Table 22.1: Breakdown of Estimated Operating Costs

LOM Operating Costs C$/t milled* (C$M) Mining** 21.32 630 Processing 11.49 339 G&A 4.83 143 Total 37.64 1,112 *includes pre-production tonnes milled. **Average LOM mining cost amounts to $4.29/t mined (including pre-production tonnes) at a 4:1 strip ratio. Source: JDS (2017)

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Figure 22.1: Operating Cost Distribution

Source: JDS (2017)

The main operating cost component assumptions are shown in Table 22.2. Table 22.2: Main OPEX Assumptions

Item Unit Value

Electrical power cost (NWT Power) $/kWh 0.11 Electrical power cost (LNG generators) $/kWh 0.14 Average power consumption MW 6.3 Overall power consumption (all facilities) kWh/t processed 20.4 Diesel cost (delivered) $/litre 0.88 LOM average manpower (including contractors, excluding corporate) employees 321 Source: JDS (2017)

22.1 Operations Labour This section provides an overview of total workforce and the methods used to compile the labour rates.

Table 22.3 summarizes the total planned workforce during project operations. The peak workforce is 371 persons in Year 5.

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Table 22.3: Annual Manpower Numbers by Discipline

Year Year Year Year Discipline Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 10 11 12 13 Mining 132 152 146 144 154 156 156 146 146 146 146 152 120 Process 78 78 98 98 98 98 98 98 98 98 98 98 98 G&A 69 69 69 69 69 69 69 69 69 69 69 69 69 Construction 50 - 50 - 50 ------Total - Manpower 329 299 363 311 371 323 323 313 313 313 313 319 287 Source: JDS (2017)

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Labour is a significant portion of annual operating cost. Labour rates include base wage and allowances for overtime, insurance, tax, and benefits. Labour burdens were assembled using first principles. The following items are included in the burdened labour rates:

 Canada Pension Plan at 4.95%;  Worker’s Compensation (WSCC) at 1.62%;  Employment Insurance at 2.28%; and  Company Benefits at $5,000/year.

22.2 Basis of Estimate 22.2.1 Mine Operating Cost Estimate

Costs for OP mining activities for the Pine Point Project, assumed to be undertaken by an Owner- operated fleet, were built up from first principles, as well as JDS experience of similar-sized OP operations and local conditions. OP mining costs for both mineralized and waste material take into account variations in haulage profiles and equipment selection. Equipment efficiency was estimated based on Pine Point conditions (e.g. haul routes for each phase). Local labour rates and diesel fuel pricing estimates were utilized for estimation purposes. The OP mining costs encompass pit and dump operations, road maintenance, and mine supervision, technical services and mine dewatering costs.

The average mine OPEX for the LOM plan was estimated to be $4.29/t material mined or $21.32/t plant feed, for pit and dump operations, road maintenance, mine supervision, technical services and pit dewatering.

No contingency was added.

Table 22.4 provides a breakdown of the mine operating costs by category.

Table 22.4: Mine Operating Costs by Category

Cost Category $/t processed* $/t mined LOM (M$) Labour 7.14 1.43 211 Fuel 3.59 0.72 106 Operating & Maintenance Consumables 6.91 1.39 204 Tools, Supplies & Contract Services 0.47 0.10 14 Lease Payments 2.13 0.43 63 Dewatering 1.08 0.22 32 Total Mine Costs 21.32 4.29 630 *includes pre-production tonnes milled. Source: JDS (2017)

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22.2.2 Processing Operating Cost Estimate

Process operating costs were developed using labour rates based on operating mines in the area and sufficient personnel to operate the process plant, factored maintenance cost, budget quotes for consumables and a factored power requirement. Process operating costs are summarized below in Table 22.5. Costs are subdivided into operating categories.

Table 22.5: Processing Operating Cost by Category

Category $/t processed* LOM ($M) Labour 4.71 139 Equipment Maintenance & Consumables (Reagents, Media, Liners 4.37 129 and other Wear Parts) Power & Fuel 1.15 34 Mill Feed Haulage 1.26 37 Grand Total by Activity 11.49 339 *includes pre-production tonnes milled. Source: JDS (2017)

Process labour includes burden for salaried and hourly employees to account for in-country benefits, training, production bonus and potential ex-patriot benefits and costs.

Equipment maintenance was calculated by applying a factor of 4% to major process equipment cost. Costs for media were determined using engineering calculations based on mill power draw, abrasion index and vendor quotes for media as a cost per tonne. Reagent requirements from recent test work and budget quotes from vendors were used to calculate the cost of reagents. Mill liners and wear parts for major equipment were based on vendor recommended requirements and quotes. Mill feed haulage from the DMS plants to the process plant was calculated based on cycle times assuming a contractor operating the loading and haulage equipment.

Power costs were calculated from the total installed power assuming $0.11/kWh.

22.2.3 General and Administration Operating Cost Estimate

The general and administration costs include all off-site and on-site activities including personnel transportation, camp catering and cleaning, surface support equipment, water management, environmental monitoring, facilities maintenance, insurance, and all associated labour. The G&A operating cost is estimated to be $4.83/t processed (including pre-production tonnes milled) and can be attributed to five categories:

 Labour;  Camp & travel;  Infrastructure operation & maintenance;  Site support equipment; and  Other G&A items.

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G&A operating costs are summarized below in Table 22.6.

Table 22.6: Summary of G&A Costs

Cost Category $/t processed* LOM (C$M) Labour 2.20 65 Camp & Travel 1.08 32 Infrastructure Operation & Maintenance 0.24 7 Site Support Equipment 0.37 11 Other G&A Items 0.94 28 Total G&A Costs 4.83 143 *includes pre-production tonnes milled. Source: JDS (2017)

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23 Economic Analysis

23.1 Summary of Results An economic model was developed to estimate annual cash flows and sensitivities of the Project. Pre-tax estimates of project values were prepared for comparative purposes, while after-tax estimates were developed and are likely to approximate the true investment value. It must be noted, however, that tax estimates involve many complex variables that can only be accurately calculated during operations and, as such, the after-tax results are only approximations.

Univariate sensitivity analyses were performed for variations in metal prices, head grades, operating costs, capital costs, and discount rates to determine their relative importance as project value drivers.

This technical report contains forward-looking information regarding projected mine production rates, construction schedules and forecasts of resulting cash flows as part of this study. The mill head grades are based on sufficient sampling that is reasonably expected to be representative of the realized grades from actual mining operations. Factors such as the ability to obtain permits to construct and operate a mine, or to obtain major equipment or skilled labour on a timely basis, to achieve the assumed mine production rates at the assumed grades, may cause actual results to differ materially from those presented in this economic analysis.

The estimates of capital and operating costs have been developed specifically for this project and are summarized in Section 21 and Section 22 of this report (presented in 2017 dollars). The economic analysis has been run with no inflation (constant dollar basis).

It must be noted that this PEA is preliminary in nature and includes the use of 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 results of the preliminary economic assessment will be realized.

23.2 Key Assumptions The model excludes all pre-development and sunk costs up to the start of detailed engineering (i.e. exploration and resource definition costs, engineering fieldwork and studies costs, environmental baseline studies costs, financing costs, etc.).

Table 23.1 outlines the metal prices and exchange rate assumptions used in the economic analysis. The lead price and exchange rate selected was based on the 6-month trailing average as at March 2017. The zinc price selected was based on the 12-month trailing average as at March 2017. These parameters are in line with recently released comparable Technical Reports.

The reader is cautioned that the metal prices and exchange rates used in this study are only estimates based on recent historical performance and there is absolutely no guarantee that they will

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Table 23.1: Metal Price & Exchange Rates used in Economic Analysis

Parameter Unit Value Lead Price US$/lb 1.00 Zinc Price US$/oz 1.10 Exchange Rate US$:C$ 0.75 Source: JDS (2017)

Other economic factors include the following:

 Discount rate of 8% (sensitivities using other discount rates have been calculated);  Closure cost of $17.5 M;  Nominal 2017 dollars;  Revenues, costs, taxes are calculated for each period in which they occur rather than actual outgoing/incoming payment;  Working capital calculated as three months of operating costs (mining, processing, and G&A) in Year 1;  Results are presented on 100% ownership; and  No management fees or financing costs (equity fund-raising was assumed).

23.3 Mine Production The summary of the mine plan and contained metal is outlined in Table 23.2.

Table 23.2: LOM Plan Summary

Parameter Unit Value Mine Life Years 13 PEA Mineral Resources Mt 29.5 Waste Mined Mt 117.2 Total Mined Mt 146.7 Mined Lead Grade % 1.08 Mined Zinc Grade % 2.93 Contained Lead Mlbs 702.7 Contained Zinc Mlbs 1,910.4 Source: JDS (2017)

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23.4 Processing & Concentrate Terms Mine revenue is derived from the sale of zinc concentrate and lead concentrate into the international marketplace. No contractual arrangements for refining exist at this time. Details regarding the terms used for the economic analysis can be found in the market studies (Section 19) of this report.

Table 23.3 outlines the recoveries, payable terms, treatment charges and transportation costs used in the economic analysis.

Table 23.3: Concentrate Terms

Assumptions & Inputs Unit Value Lead Concentrate Metal Recovery to Pb Con (CP) % 85.0 Metal Recovery to Pb Con (N-204) % 71.4 Average Pb Recovery to Concentrate % 80.4 Average Pb Concentrate Grade Produced % 65 Minimum Deduction % Pb/t 3.0 Pb Payable % 95.0 Pb Treatment Charge C$/dmt conc. 200.00 Pb Concentrate Transportation Cost C$/wmt 167.00 Zinc Concentrate Metal Recovery to Zn Con (CP) % 88.1 Metal Recovery to Zn Con (N-204) % 78.0 Average Zn Recovery to Concentrate % 83.4 Average Zn Concentrate Grade Produced % 58.9 Minimum Deduction % Zn/t 8.0 Zn Payable % 85.0 Zn Treatment Charge C$/dmt conc 200.00 Zn Concentrate Transportation Cost C$/wmt 167.00 Source: JDS (2017)

Figure 23.1 and Figure 23.2 show breakdowns of the payable lead and zinc recovered during the mine life. A total of 536.5 Mlbs of lead, and 1,354.9 Mlbs of zinc are projected to be produced during the mine life. Zinc accounts for about 73% of project revenues, and lead for about 27%, this is illustrated in Figure 23.3.

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Figure 23.1: Payable Lead Production by Year

Source: JDS (2017)

Figure 23.2: Payable Zinc Production by Year

Source: JDS (2017)

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Figure 23.3: Revenue Distribution

Source: JDS (2017)

23.5 Taxes The Project has been evaluated on an after-tax basis in order to provide a more indicative, but still approximate, value of the potential project economics. A tax model was prepared by a tax consultant with applicable NWT mineral tax regime experience. Current tax pools were used in the analysis. The tax model contains the following assumptions:

 15% federal income tax rate; and  11.5% NWT tax rate. Total taxes for the Project amount to $214 M.

23.6 Royalties A 3.0% NSR royalty is payable to Karst as described in Section 19. Total third party royalties for the Project amount to $62.5 M over the LOM. More details on the structure of this royalty can be found in Section 19.

23.7 Results At this preliminary stage, the Project has a pre-tax IRR of 48% and a net present value using an 8% discount rate (NPV8%) of $341 M using the metal prices described in Section 19.

Figure 23.4 shows the projected after-tax cash flows, and Table 23.4 summarizes the economic results of the Pine Point Project.

The pre-tax break-even zinc price for the Project is approximately US$0.76/lb, based on the LOM plan presented herein, a lead price of US$1.00/lb and an FX rate of 0.75 US$:C$.

Effective Date: April 18, 2017 23-11

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

Figure 23.4: Annual After-Tax Cash Flow

Source: JDS (2017)

Effective Date: April 18, 2017 23-12

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

Table 23.4: Summary of Results

Parameter Unit Value Pre-Tax Operating Income C$M 908 Pb (Straight Cash Cost) US$/lb 2.51 Pb Cash Cost (Net of By-Products) US$/lb (0.27) Zn (Straight Cash Cost) US$/lb 0.99 Zn Cash Cost (Net of By-Products) US$/lb 0.60 Capital Costs Pre-Production Capital C$M 134 Pre-Production Contingency C$M 20 Total Pre-Production Capital C$M 154 Sustaining & Closure Capital C$M 106 Sustaining & Closure Contingency C$M 11 Total Sustaining & Closure Capital C$M 118 Total Capital Costs Incl. Contingency C$M 271 Cash Flows Working Capital C$M 20 LOM C$M 637 Pre-Tax Cash Flow C$M/a 49 Taxes LOM C$M 214 LOM C$M 423 After-Tax Cash Flow C$M/a 33 Economic Results

Pre-Tax NPV8% C$M 341 Pre-Tax IRR % 48 Pre-Tax Payback Years 1.4

After-Tax NPV8% C$M 211 After-Tax IRR % 35 After-Tax Payback Years 1.8 Source: JDS (2017)

23.8 Sensitivities A univariate sensitivity analysis was performed to examine which factors most affect the project economics when acting independently of all other cost and revenue factors. Each variable evaluated was tested using the same percentage range of variation, from -40% to +40%, although some variables may actually experience significantly larger or smaller percentage fluctuations over the LOM. For instance, the metal prices were evaluated at a +/- 10% range to the base case, while the mill feed grade and all other variables remained constant. This may not be truly representative of market scenarios, as metal prices may not fluctuate in a similar trend. The variables examined in

Effective Date: April 18, 2017 23-13

DARNLEY BAY RESOURCES LTD. PINE POINT PEA this analysis are those commonly considered in similar studies – their selection for examination does not reflect any particular uncertainty.

Notwithstanding the above noted limitations to the sensitivity analysis, which are common to studies of this sort, the analysis revealed that the Project is most sensitive to exchange rate, followed by recovery, metal prices, mill feed grade, and operating costs. The Project showed the least sensitivity to capital costs. Table 23.5 and Figure 23.5 show the results of the sensitivity tests.

A sensitivity analysis was also performed on discount rate. The results of this test are demonstrated in Table 23.6 and Figure 23.6.

The economic cash flow model for the Project is illustrated in Figure 23.7.

Table 23.5: After-Tax Sensitivity Results on NPV @ 8% Discount Rate

Parameter -40% -30% -20% -10% Base +10% +20% +30% +40% Price (271) (123) 6 111 211 310 410 508 607 CAD:US FX 870 635 459 321 211 120 41 (30) (104) Mill Feed Grade (135) (31) 54 133 211 287 364 441 517 Recovery (334) (163) (11) 108 211 310 410 508 607 OPEX 377 335 294 252 211 169 127 84 39 CAPEX 298 276 254 233 211 189 167 145 123 Source: JDS (2017)

Effective Date: April 18, 2017 23-14

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

Figure 23.5: Sensitivity, Post-Tax NPV @ 8% Discount Rate

Source: JDS (2017)

Table 23.6: Discount Rate Sensitivity

Discount Rate Pre-Tax NPV After-Tax NPV (%) (C$M) (C$M)

0 637 423 5 428 273 8 341 211 10 294 177 12 254 149 15 205 114 20 143 70 Source: JDS (2017)

Effective Date: April 18, 2017 23-15

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

Figure 23.6: Post-Tax NPV Sensitivity to Discount Rate

Source: JDS (2017)

Effective Date: April 18, 2017 23-16

Pre-Production Production Life of Mine Parameters Source Unit Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Year 11 Year 12 Year 13 Year 14 Year 15 Year 16 Year 17 Year 18 Year 19 Year 20 Total Total Total KEY INPUTS Metal Prices Pb input US$/lb 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 1.00 Zn input US$/lb 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 Exchange Rate input C$:US$ 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75 0.75

OPEN PIT MINING SCHEDULE Open Pit Mine Quantities Mineralized Material link (M1) tonnes 304,398 29,239,797 29,544,195 - 304,398 1,632,238 2,155,867 2,481,220 2,306,704 2,506,894 2,455,669 2,395,953 2,512,326 2,362,566 2,323,532 2,266,642 2,461,826 1,378,362 ------Waste Mined link (M1) tonnes 384,891 116,792,770 117,177,661 - 384,891 9,141,609 10,147,037 12,066,248 7,995,251 7,719,947 5,597,991 6,849,945 8,615,296 10,029,044 10,553,092 11,793,591 11,462,787 4,820,934 ------Total Material Mined calc tonnes 689,289 146,032,567 146,721,856 - 689,289 10,773,847 12,302,904 14,547,468 10,301,956 10,226,840 8,053,659 9,245,898 11,127,622 12,391,609 12,876,624 14,060,232 13,924,612 6,199,296 ------Average Strip Ratio calc w/o 1.3 4.0 4.0 - 1.3 5.6 4.7 4.9 3.5 3.1 2.3 2.9 3.4 4.2 4.5 5.2 4.7 3.5 ------Open Pit Contained Metals Contained Lead link (M1) lb 15,109,000 687,601,400 702,710,400 - 15,109,000 102,204,596 104,820,064 72,429,745 38,035,757 47,667,693 34,607,390 40,809,178 50,493,073 41,791,976 42,713,441 35,637,602 43,738,266 32,652,620 ------Contained Zinc link (M1) lb 32,887,200 1,877,473,600 1,910,360,800 - 32,887,200 125,407,064 156,935,016 178,785,865 117,645,992 195,034,208 138,585,180 128,474,964 173,216,606 134,710,006 169,256,239 133,400,933 131,038,849 94,982,679 ------Open Pit Head Grades Mined Pb Grade calc % 2.25% 1.07% 1.08% - 2.25% 2.84% 2.21% 1.32% 0.75% 0.86% 0.64% 0.77% 0.91% 0.80% 0.83% 0.71% 0.81% 1.07% ------Mined Zn Grade calc % 4.90% 2.91% 2.93% - 4.90% 3.49% 3.30% 3.27% 2.31% 3.53% 2.56% 2.43% 3.13% 2.59% 3.30% 2.67% 2.41% 3.13% ------Average Head Grade Value calc $CAD/tonne $ 224.64 $ 125.53 $ 126.55 $ - $ 224.64 $ 196.17 $ 171.59 $ 144.60 $ 96.79 $ 139.46 $ 101.56 $ 101.36 $ 127.92 $ 107.21 $ 131.35 $ 107.28 $ 101.76 $ 132.65 $ - $ - $ - $ - $ - $ - $ -

UNDERGROUND MINING SCHEDULE Underground Mine Quantities Mineralized Material link (M1) tonnes ------Underground Contained Metals Contained Lead link (M1) lb ------Contained Zinc link (M1) lb ------Underground Head Grades Mined Pb Grade calc % ------Mined Zn Grade calc % ------Average Head Grade Value calc $CAD/tonne $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ - $ -

TOTAL MINE PRODUCTION Total Mine Quantities Mineralized Material calc tonnes 304,398 29,239,797 29,544,195 - 304,398 1,632,238 2,155,867 2,481,220 2,306,704 2,506,894 2,455,669 2,395,953 2,512,326 2,362,566 2,323,532 2,266,642 2,461,826 1,378,362 ------Total Contained Metals Contained Lead calc lb 15,109,000 687,601,400 702,710,400 - 15,109,000 102,204,596 104,820,064 72,429,745 38,035,757 47,667,693 34,607,390 40,809,178 50,493,073 41,791,976 42,713,441 35,637,602 43,738,266 32,652,620 ------Contained Zinc calc lb 32,887,200 1,877,473,600 1,910,360,800 - 32,887,200 125,407,064 156,935,016 178,785,865 117,645,992 195,034,208 138,585,180 128,474,964 173,216,606 134,710,006 169,256,239 133,400,933 131,038,849 94,982,679 ------Average Mined Head Grades Mined Pb Grade calc % 2.25% 1.07% 1.08% - 2.25% 2.84% 2.21% 1.32% 0.75% 0.86% 0.64% 0.77% 0.91% 0.80% 0.83% 0.71% 0.81% 1.07% ------Mined Zn Grade calc % 4.90% 2.91% 2.93% - 4.90% 3.49% 3.30% 3.27% 2.31% 3.53% 2.56% 2.43% 3.13% 2.59% 3.30% 2.67% 2.41% 3.13% ------Average Head Grade Value calc $CAD/tonne $ 224.64 $ 125.53 $ 126.55 $ - $ 224.64 $ 196.17 $ 171.59 $ 144.60 $ 96.79 $ 139.46 $ 101.56 $ 101.36 $ 127.92 $ 107.21 $ 131.35 $ 107.28 $ 101.76 $ 132.65 $ - $ - $ - $ - $ - $ - $ -

PROCESSING SCHEDULE Crushing/DMS Plant #1 (CP) Total Feed link (M1) tonnes 304,398 15,708,430 16,012,828 - 304,398 1,632,238 2,155,867 2,263,102 1,716,894 356,582 690,506 1,099,304 293,278 1,185,329 1,192,541 2,034,143 737,098 351,548 ------Total Output link (M1) tonnes 106,973 4,665,632 4,772,605 - 106,973 550,083 656,908 601,059 504,804 104,843 203,024 323,219 86,230 354,172 366,795 597,885 211,810 104,801 ------Operating Days input days 90 4,380 4,470 - 90 365 365 365 365 365 365 365 365 365 - 365 - 365 365 ------DMS Plant Throughput calc tpd 3,380 3,590 3,580 - 3,380 4,470 5,910 6,200 4,700 980 1,890 3,010 800 3,250 - 5,570 - 960 ------Crushing/DMS Plant #2 (UG) Total Feed link (M1) tonnes ------Total Output link (M1) tonnes ------Operating Days input days - 3,285 3,285 - - 365 365 365 365 365 365 365 365 365 ------DMS Plant Throughput calc tpd ------Crushing/DMS Plant #3 (N204) Total Feed link (M1) tonnes - 13,531,367 13,531,367 - - - - 218,119 589,811 2,150,311 1,765,163 1,296,649 2,219,048 1,177,237 1,130,991 232,499 1,724,727 1,026,814 ------Total Output link (M1) tonnes - 3,479,494 3,479,494 - - - - 56,088 151,666 552,937 453,899 333,424 570,612 302,718 290,826 59,785 443,501 264,038 ------Operating Days input days - 3,650 3,650 ------365 365 365 365 365 365 365 365 365 365 ------DMS Plant Throughput calc tpd - 3,710 3,710 ------5,890 4,840 3,550 6,080 3,230 3,100 640 4,730 2,810 ------Total Crushing/DMS Total Feed calc tonnes 304,398 29,239,797 29,544,195 - 304,398 1,632,238 2,155,867 2,481,220 2,306,704 2,506,894 2,455,669 2,395,953 2,512,326 2,362,566 2,323,532 2,266,642 2,461,826 1,378,362 ------Total Output clac tonnes 106,973 8,145,126 8,252,099 - 106,973 550,083 656,908 657,146 656,469 657,780 656,923 656,643 656,842 656,890 657,621 657,670 655,311 368,838 ------Operating Days clac days 90 5,110 5,200 - 90 365 365 365 365 365 365 365 365 365 365 365 365 365 365 ------Total DMS Throughput calc tpd 3,380 5,720 5,680 - 3,380 4,470 5,910 6,800 6,320 6,870 6,730 6,560 6,880 6,470 6,370 6,210 6,740 3,780 ------Flotation Plant Total DMS Concentrate Processed link (M1) tonnes - 8,252,099 8,252,099 - - 657,000 657,000 657,000 657,000 657,000 657,000 657,000 657,000 657,000 657,000 657,000 657,000 368,099 ------Operating Days input days - 4,745 4,745 - - 365 365 365 365 365 365 365 365 365 365 365 365 365 ------Flotation Plant Throughput calc tpd - 1,740 1,740 - - 1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,010 ------

Pb CONCENTRATE PRODUCTION & REVENUES Production Metal Recovery to Pb Con (CP; R190; X25) input % Pb - 85.0% 85.0% - - 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% ------Metal Recovery to Pb Con (N-204) input % Pb - 71.4% 71.4% - - 71.4% 71.4% 71.4% 71.4% 71.4% 71.4% 71.4% 71.4% 71.4% 71.4% 71.4% 71.4% 71.4% ------Average Pb Recovery to Concentrate calc % Pb - 80.3% 80.4% - - 85.0% 85.0% 84.7% 82.6% 72.8% 74.7% 78.8% 72.4% 77.7% 79.3% 83.2% 76.0% 76.0% ------calc Pb tonnes - 256,251 256,251 - - 45,231 40,414 27,828 14,251 15,743 11,727 14,590 16,579 14,736 15,369 13,452 15,076 11,253 ------Metal in Pb Concentrate calc Pb lbs - 564,930,536 564,930,536 - - 99,716,557 89,097,054 61,350,135 31,418,169 34,707,565 25,853,647 32,164,719 36,550,167 32,486,734 33,883,175 29,657,034 33,237,010 24,808,570 ------Pb Concentrate Grade (CP; R190; X25) input % - 70.0% 70.0% - - 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% 70.0% ------Pb Concentrate Grade (N-204) input % - 55.7% 55.7% - - 55.7% 55.7% 55.7% 55.7% 55.7% 55.7% 55.7% 55.7% 55.7% 55.7% 55.7% 55.7% 55.7% ------Average Pb Concentrate Grade Produced calc % - 65.0% 65.0% - - 70.0% 70.0% 69.7% 67.4% 57.1% 59.0% 63.3% 56.7% 62.2% 63.8% 68.0% 60.4% 60.3% ------calc dmt - 394,347 394,347 - - 64,616 57,735 39,943 21,156 27,565 19,865 23,046 29,247 23,703 24,073 19,772 24,979 18,649 ------Pb Concentrate Produced calc wmt - 428,638 428,638 - - 70,235 62,755 43,416 22,995 29,962 21,592 25,050 31,790 25,764 26,166 21,491 27,151 20,271 ------Moisture Content input % - 8.0% 8.0% - - 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% ------Payables/Revenues Minimum Deduction input %Pb/tonne - 3.0% 3.0% - - 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% ------Pb Payable input % Payable - 95.0% 95.0% - - 95.0% 95.0% 95.0% 95.0% 95.0% 95.0% 95.0% 95.0% 95.0% 95.0% 95.0% 95.0% 95.0% ------Pb Payable based on Min. deduction calc % - 62.0% 62.0% - - 67.0% 67.0% 66.7% 64.4% 54.1% 56.0% 60.3% 53.7% 59.2% 60.8% 65.0% 57.4% 57.3% ------Pb Payable based on % calc % - 61.7% 61.7% - - 66.5% 66.5% 66.2% 64.0% 54.3% 56.1% 60.1% 53.9% 59.1% 60.7% 64.6% 57.3% 57.3% ------calc tonnes - 243,340 243,340 - - 42,970 38,393 26,437 13,539 14,916 11,131 13,860 15,702 13,999 14,601 12,780 14,322 10,690 ------Payable Pb in Pb Concentrate calc lbs - 536,468,354 536,468,354 - - 94,730,729 84,642,202 58,282,628 29,847,261 32,884,485 24,539,839 30,556,483 34,615,831 30,862,397 32,189,016 28,174,182 31,575,160 23,568,141 ------Revenues Pb in Pb Concentrate calc $CAD - 715,291,139 715,291,139 - - 126,307,638 112,856,269 77,710,171 39,796,348 43,845,981 32,719,785 40,741,977 46,154,441 41,149,863 42,918,689 37,565,576 42,100,213 31,424,189 ------Treatment & Refining Charges input C$/dmt conc - 200.00 200.00 - - 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 ------Pb Treatment Charge calc CAD$ - (78,869,474) (78,869,474) - - (12,923,181) (11,546,903) (7,988,507) (4,231,115) (5,512,955) (3,972,930) (4,609,251) (5,849,396) (4,740,658) (4,814,523) (3,954,435) (4,995,816) (3,729,803) ------input C$/wmt - 167.00 167.00 - - 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 ------Pb Concentrate Transport Cost calc $CAD - (71,582,621) (71,582,621) - - (11,729,191) (10,480,069) (7,250,439) (3,840,197) (5,003,606) (3,605,866) (4,183,397) (5,308,962) (4,302,663) (4,369,703) (3,589,080) (4,534,246) (3,385,202) ------Pb Concentrate NSR (Pre-Royalties) calc $CAD - 564,839,044 564,839,044 - - 101,655,266 90,829,297 62,471,224 31,725,036 33,329,419 25,140,989 31,949,329 34,996,083 32,106,542 33,734,463 30,022,061 32,570,151 24,309,183 ------Pre-Production Production Life of Mine Parameters Source Unit Year -2Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Year 11 Year 12 Year 13 Year 14 Year 15 Year 16 Year 17 Year 18 Year 19 Year 20 Total Total Total Zn CONCENTRATE PRODUCTION & REVENUES Production Metal Recovery to Zn Con (CP; R190; X25) input % Pb - 88.1% 88.1% - - 88.1% 88.1% 88.1% 88.1% 88.1% 88.1% 88.1% 88.1% 88.1% 88.1% 88.1% 88.1% 88.1% ------Metal Recovery to Zn Con (N-204) input % Pb - 78.0% 78.0% - - 78.0% 78.0% 78.0% 78.0% 78.0% 78.0% 78.0% 78.0% 78.0% 78.0% 78.0% 78.0% 78.0% ------Average Zn Recovery to Concentrate calc % Pb - 83.4% 83.4% - - 88.1% 88.1% 87.8% 86.2% 79.0% 80.8% 82.3% 78.7% 82.6% 84.0% 86.6% 80.1% 79.8% ------calc Pb tonnes - 723,010 723,010 - - 63,257 62,714 71,166 45,981 69,849 50,814 47,985 61,870 50,501 64,502 52,376 47,632 34,361 ------Metal in Zn Concentrate calc Pb lbs - 1,593,948,147 1,593,948,147 - - 139,457,247 138,259,749 156,892,835 101,370,261 153,990,200 112,025,322 105,787,965 136,398,662 111,334,693 142,201,232 115,468,472 105,010,116 75,751,394 ------Zn Concentrate Grade (CP; R190; X25) input % - 61.9% 61.9% - - 61.9% 61.9% 61.9% 61.9% 61.9% 61.9% 61.9% 61.9% 61.9% 61.9% 61.9% 61.9% 61.9% ------Zn Concentrate Grade (N-204) input % - 55.3% 55.3% - - 55.3% 55.3% 55.3% 55.3% 55.3% 55.3% 55.3% 55.3% 55.3% 55.3% 55.3% 55.3% 55.3% ------Average Zn Concentrate Grade Produced calc % - 59.9% 58.9% - - 61.9% 61.9% 61.7% 60.6% 55.9% 57.2% 58.2% 55.8% 58.4% 59.2% 60.9% 56.7% 56.5% ------calc dmt - 1,228,195 1,228,195 - - 102,193 101,315 115,387 75,820 124,888 88,891 82,517 110,897 86,546 108,873 86,004 83,998 60,865 ------Zn Concentrate Produced calc wmt - 1,334,995 1,334,995 - - 111,079 110,125 125,420 82,413 135,748 96,621 89,693 120,540 94,071 118,340 93,483 91,303 66,157 ------Moisture Content input % - 8.0% 8.0% - - 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% ------Payables/Revenues Minimum Deduction input %Pb/tonne - 8.0% 8.0% - - 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% 8.0% ------Zn Payable input % Payable - 85.0% 85.0% - - 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% 85.0% ------Zn Payable based on Min. deduction calc % - 51.9% 50.9% - - 53.9% 53.9% 53.7% 52.6% 47.9% 49.2% 50.2% 47.8% 50.4% 51.2% 52.9% 48.7% 48.5% ------Zn Payable based on % calc % - 50.9% 50.0% - - 52.6% 52.6% 52.4% 51.5% 47.5% 48.6% 49.4% 47.4% 49.6% 50.4% 51.8% 48.2% 48.0% ------calc tonnes - 614,559 614,559 - - 53,769 53,307 60,491 39,084 59,372 43,192 40,787 52,590 42,926 54,827 44,520 40,487 29,207 ------Payable Zn in Zn Concentrate calc lbs - 1,354,855,925 1,354,855,925 - - 118,538,660 117,520,787 133,358,910 86,164,722 130,891,670 95,221,523 89,919,770 115,938,862 94,634,489 120,871,047 98,148,201 89,258,599 64,388,685 ------Revenues Zn in Zn Concentrate calc $CAD - 1,987,122,024 1,987,122,024 - - 173,856,701 172,363,821 195,593,068 126,374,926 191,974,449 139,658,234 131,882,329 170,043,665 138,797,251 177,277,536 143,950,695 130,912,611 94,436,738 ------Treatment & Refining Charges input C$/dmt conc - 200.00 200.00 - - 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 200.00 ------Zn Treatment Charge calc CAD$ - (245,639,040) (245,639,040) - - (20,438,576) (20,263,073) (23,077,316) (15,164,054) (24,977,698) (17,778,256) (16,503,484) (22,179,363) (17,309,153) (21,774,629) (17,200,796) (16,799,682) (12,172,960) ------input C$/wmt - 167.00 167.00 - - 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 167.00 ------Zn Concentrate Transport Cost calc $CAD - (222,944,129) (222,944,129) - - (18,550,229) (18,390,941) (20,945,173) (13,763,027) (22,669,976) (16,135,700) (14,978,705) (20,130,183) (15,709,938) (19,762,843) (15,611,592) (15,247,537) (11,048,284) ------Zn Concentrate NSR (Pre-Royalties) calc $CAD - 1,518,538,855 1,518,538,855 - - 134,867,896 133,709,806 151,570,578 97,447,845 144,326,775 105,744,279 100,400,140 127,734,120 105,778,160 135,740,064 111,138,306 98,865,392 71,215,493 ------

NET SMELTER RETURN Pb & Zn NSR (Pre-Royalty) calc $CAD - 2,083,377,899 2,083,377,899 - - 236,523,162 224,539,104 214,041,803 129,172,882 177,656,195 130,885,267 132,349,469 162,730,203 137,884,702 169,474,527 141,160,367 131,435,543 95,524,676 ------link % NSR - 3.0% 3.0% - - 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% 3.0% ------Karst Royalty calc $CAD - (62,501,337) (62,501,337) - - (7,095,695) (6,736,173) (6,421,254) (3,875,186) (5,329,686) (3,926,558) (3,970,484) (4,881,906) (4,136,541) (5,084,236) (4,234,811) (3,943,066) (2,865,740) ------Pb & Zn NSR (Post-Royalty) calc $CAD - 2,020,876,562 2,020,876,562 - - 229,427,467 217,802,931 207,620,549 125,297,695 172,326,509 126,958,709 128,378,985 157,848,297 133,748,161 164,390,291 136,925,556 127,492,477 92,658,936 ------

OPERATING COSTS input C$/t mined - 4.31 4.29 - - 4.52 4.49 4.21 4.97 4.61 5.13 4.42 3.99 3.94 3.99 3.87 3.76 5.30 ------Open Pit Mining calc CAD$ - (629,897,884) (629,897,884) - - (48,651,629) (55,262,192) (61,212,469) (51,160,174) (47,121,007) (41,299,182) (40,884,337) (44,454,504) (48,820,863) (51,437,026) (54,347,426) (52,397,477) (32,849,599) ------input C$/t milled ------Underground Mining calc CAD$ ------input C$/t milled - 11.61 11.49 - - 13.52 11.14 11.02 11.79 11.18 11.33 11.51 11.17 11.56 11.58 11.14 10.94 15.18 ------Processing calc CAD$ - (339,453,623) (339,453,623) - - (22,067,984) (24,007,395) (27,336,821) (27,199,278) (28,038,922) (27,828,246) (27,576,892) (28,069,488) (27,318,961) (26,906,503) (25,239,344) (26,934,867) (20,928,922) ------input C$/t milled - 4.88 4.83 - - 4.83 4.83 4.83 4.83 4.83 4.83 4.83 4.83 4.83 4.83 4.83 4.83 4.83 ------General & Administration calc CAD$ - (142,770,086) (142,770,086) - - (11,005,075) (10,931,942) (11,036,985) (10,956,875) (11,012,985) (10,966,364) (10,995,575) (10,998,180) (10,986,319) (10,986,319) (10,984,319) (10,954,575) (10,954,575) ------calc C$/t milled - 38.03 37.64 - - 50.07 41.84 40.14 38.72 34.37 32.62 33.16 33.24 36.88 38.45 39.96 36.67 46.96 ------Total OPEX calc CAD$ - (1,112,121,592) (1,112,121,592) - - (81,724,687) (90,201,529) (99,586,275) (89,316,327) (86,172,914) (80,093,792) (79,456,804) (83,522,172) (87,126,143) (89,329,848) (90,571,089) (90,286,918) (64,733,095) ------

OPERATING INCOME calc CAD$ - 908,754,970 908,754,970 - - 147,702,779 127,601,402 108,034,274 35,981,368 86,153,595 46,864,917 48,922,182 74,326,125 46,622,018 75,060,443 46,354,467 37,205,559 27,925,841 ------Gross Pre-Tax Operating Income calc C$/t milled - 31.08 30.76 - - 90.49 59.19 43.54 15.60 34.37 19.08 20.42 29.58 19.73 32.30 20.45 15.11 20.26 ------

CAPITAL COSTS Open Pit Mining input CAD$ (17,134,359) (14,403,559) (31,537,918) - (17,134,359) (4,736,449) (962,560) (210,235) - (420,470) (210,235) - (3,023,983) (374,698) (4,464,930) ------Underground Mining input CAD$ ------Mine Water Cutoff Systems input CAD$ ------Mine Water Management input CAD$ (2,718,937) (17,040,770) (19,759,707) - (2,718,937) (7,074,370) (720,850) (2,254,590) - (601,020) (971,460) (336,420) - (669,030) (1,324,210) (1,280,210) (1,438,170) (370,440) ------Site Development input CAD$ (3,610,000) (2,705,000) (6,315,000) (2,610,000) (1,000,000) (630,000) - (630,000) - (315,000) (815,000) - - - (315,000) ------Crushing & Concentration input CAD$ (20,226,393) (23,028,993) (43,255,386) (5,056,598) (15,169,795) (300,000) - - (21,978,993) - (150,000) - - (150,000) (150,000) (300,000) ------Flotation Plant input CAD$ (32,436,085) (5,944,770) (38,380,855) (8,109,021) (24,327,064) - - - (5,944,770) ------Infrastructure input CAD$ (24,846,246) (15,751,000) (40,597,246) (13,110,909) (11,735,337) (7,500,000) (3,450,000) (80,000) (4,291,000) - (240,000) - - - (190,000) ------Project Indirect Costs input CAD$ (17,078,321) (9,918,129) (26,996,449) (6,655,677) (10,422,643) (1,189,278) (486,715) (1,444,913) (4,544,759) (1,926,575) (169,998) - - (21,162) (92,405) (42,323) ------Owners Costs input CAD$ (15,714,547) - (15,714,547) (3,614,690) (12,099,857) ------Closure input CAD$ - (17,500,000) (17,500,000) ------(7,500,000) (2,500,000) (2,500,000) (2,500,000) (2,500,000) Total - Project Costs (Pre-Contingency) calc CAD$ (133,764,887) (106,292,221) (240,057,108) (39,156,895) (94,607,992) (21,430,098) (5,620,125) (4,619,737) (36,759,522) (3,263,065) (2,556,693) (336,420) (3,023,983) (1,214,890) (6,536,545) (1,622,533) (1,438,170) (370,440) (7,500,000) (2,500,000) (2,500,000) (2,500,000) (2,500,000) - - Contingency calc CAD$ (20,064,733) (11,158,299) (31,223,032) (5,873,534) (14,191,199) (2,504,047) (698,635) (661,425) (5,513,928) (426,389) (351,969) (50,463) - (126,029) (310,742) (243,380) (215,726) (55,566) ------Total - Project Costs (Incl. Contingency) calc CAD$ (153,829,620) (117,450,520) (271,280,140) (45,030,429) (108,799,190) (23,934,145) (6,318,760) (5,281,163) (42,273,450) (3,689,455) (2,908,661) (386,883) (3,023,983) (1,340,919) (6,847,287) (1,865,913) (1,653,896) (426,006) (7,500,000) (2,500,000) (2,500,000) (2,500,000) (2,500,000) - - Total CAPEX calc CAD$ (153,829,620) (117,450,520) (271,280,140) (45,030,429) (108,799,190) (23,934,145) (6,318,760) (5,281,163) (42,273,450) (3,689,455) (2,908,661) (386,883) (3,023,983) (1,340,919) (6,847,287) (1,865,913) (1,653,896) (426,006) (7,500,000) (2,500,000) (2,500,000) (2,500,000) (2,500,000) - -

WORKING CAPITAL Working Capital calc CAD$ (20,431,000) 20,431,000 - - (20,431,000) ------20,431,000 ------

CASH FLOW Pre-Tax Net Pre-Tax Cash Flow calc CAD$ (174,260,620) 811,735,450 637,474,830 (45,030,429) (129,230,190) 123,768,635 121,282,642 102,753,111 (6,292,082) 82,464,140 43,956,256 48,535,299 71,302,142 45,281,100 68,213,156 44,488,554 35,551,663 47,930,835 (7,500,000) (2,500,000) (2,500,000) (2,500,000) (2,500,000) - - Cumulative Pre-Tax Cash Flow calc CAD$ (174,260,620) 637,474,830 637,474,830 (45,030,429) (174,260,620) (50,491,985) 70,790,656 173,543,767 167,251,685 249,715,825 293,672,082 342,207,380 413,509,522 458,790,622 527,003,778 571,492,332 607,043,995 654,974,830 647,474,830 644,974,830 642,474,830 639,974,830 637,474,830 637,474,830 637,474,830 Post-Tax Income Taxes calc CAD$ - (214,302,006) (214,302,006) - - (16,358,802) (34,532,856) (31,920,513) (3,719,305) (22,305,820) (10,368,885) (12,431,957) (16,492,400) (14,353,656) (24,256,547) (14,065,325) (10,942,039) (7,933,256) 4,044,557 668,263 666,534 Net Post-Tax Cash Flow calc CAD$ (174,260,620) 597,433,444 423,172,824 (45,030,429) (129,230,190) 107,409,833 86,749,785 70,832,597 (10,011,387) 60,158,320 33,587,372 36,103,342 54,809,742 30,927,444 43,956,609 30,423,229 24,609,624 39,997,579 (3,455,443) (1,831,737) (1,833,466) (2,500,000) (2,500,000) - - Cumulative Post-Tax Cash Flow calc CAD$ (174,260,620) 423,172,824 423,172,824 (45,030,429) (174,260,620) (66,850,787) 19,898,998 90,731,596 80,720,209 140,878,529 174,465,901 210,569,243 265,378,985 296,306,429 340,263,038 370,686,267 395,295,891 435,293,469 431,838,026 430,006,289 428,172,824 425,672,824 423,172,824 423,172,824 423,172,824

ECONOMIC RESULTS Pre-Tax Pre-Tax IRR calc % 47.8% Pre-Tax Payback calc year 1.4 1.0 0.4 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Pre-Tax NPV @ 8% calc CAD$ 340,764,000 Pre-Tax NPV @ 0% calc CAD$ 637,475,000 Post-Tax Post-Tax IRR calc % 34.5% Post-Tax Payback calc year 1.8 1.0 0.8 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Post-Tax NPV @ 8% calc CAD$ 210,547,000 Post-Tax NPV @ 0% calc CAD$ 423,173,000 DARNLEY BAY RESOURCES LTD. PINE POINT PEA

24 Adjacent Properties

There are no adjacent operational mining properties that would lead to a better understanding of this property.

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DARNLEY BAY RESOURCES LTD. PINE POINT PEA

25 Other Relevant Data and Information

There is no other relevant data or information related to the scope of this report.

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26 Interpretations and Conclusions

It is the conclusion of the QP’s that the PEA summarized in this technical report contains adequate detail and information to support the positive economic result herein contained. The PEA proposes the use of industry standard equipment and operating practices. To date, the QP’s are not aware of any fatal flaws for the Project.

Using the assumptions highlighted in this report, the Pine Point Project offers sufficient economic potential to be advanced to the next stage of study (Pre-Feasibility or Feasibility Study).

26.1 Risks As with most mining projects, many risks could affect the economic outcome of the Project. Most of these risks are external and largely beyond the control of the project proponents. They can be difficult to anticipate and mitigate, although, in many instances, some risk reduction can be achieved. Table 26.1 identifies what are currently deemed the most important internal project risks, potential impacts, and possible mitigation approaches, excluding those external circumstances that are generally applicable to all mining projects (e.g., changes in metal prices, exchange rates, smelter terms, transport costs, investment capital availability, government regulations, local and regional project support, etc.).

Factors such as the ability to obtain permits to construct and operate a mine, or to obtain major equipment or skilled labour on a timely basis, to achieve the assumed mine production rates at the assumed grades, may cause actual results to differ materially from those presented in this economic analysis.

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Table 26.1: Main Project Risks

Risk Explanation/Potential Impact Possible Risk Mitigation Dewatering of the open pits was an important component of the historic mining at Pine Point. Continued collection and analysis of Pit dewatering is a significant cost to the Ground Water Quantity data relating to ground and surface operations and underestimating the ground water needs to be continued. water quantity could affect the project economics. Conversion of historical resources All Mineral Resource estimates carry some into Measured and Indicated Mineral Resource risk. Mineral Resources that are not Mineral Resources, through confirmation Estimate Reserves do not have demonstrated economic drilling. viability.

Changes to metallurgical assumptions could lead to reduced metal recovery, increased Additional sampling and test work processing OPEX costs, and/or changes to the Metallurgical Recoveries optimization to be conducted as part processing circuit design. If LOM metal of ongoing studies. recovery is lower than assumed, the project economics would be negatively impacted. The ability to achieve the estimated CAPEX and OPEX costs are important elements of Further cost estimation accuracy with project success. the next level of study, as well as the active investigation of potential cost CAPEX and OPEX If OPEX increases, then the mining cut-off reduction measures would assist in grade would increase and, all else being equal, the support of reasonable cost the open pit recovered tonnage would reduce estimates. yielding fewer tonnes in the PEA mine plan. Continued development of close relationship with the local The ability to secure all of the permits to build communities. Development of and operate the project is of paramount relationships with government along Permit Acquisition importance. Failure to secure the necessary with a thorough Environmental and permits could stop or delay the project. Social Impact Assessment that gives appropriate consideration to the environment and local people. The project development could be delayed for a number of reasons and could impact project economics. If an aggressive schedule is to be Development Schedule followed, early engagement with all

stakeholders is important. A change in schedule would alter the project economics. The ability to attract and retain competent, experienced professionals is a key success factor for the project, particularly due to the The early search for professionals, Ability to Attract and remote nature of the project. as well as maintaining competitive Retain Experienced salaries, flexible work schedules and

Professionals benefits all help to identify, attract High turnover or the lack of appropriate and retain critical people, technical and management staff at the project could result in difficulties meeting project goals. Source: JDS 2017

Effective Date: April 18, 2017 26-2

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

26.2 Opportunities 26.2.1 Addition of Underground Resources

In addition to the open pit mineral resources identified in the PEA there is potential opportunity to include underground mineralization in the mine plan. Two deposits in particular, R190 (1.0 Mt (Measured & Indicated) grading 5.28% Pb and 10.98% Zn) and X25 (2.1 Mt (Indicated) grading 2.32% Pb and 6.73% Zn), represent an opportunity to enhance the project’s economics by supplying higher grade feed to the processing plant. Further evaluation of these underground deposits is recommended at the feasibility study stage.

26.2.2 Other Potential Opportunities

Other potential opportunities to improve project economics include:

 Conversion of historical resources into Measured and Indicated Resources, through confirmation drilling. Historically there are 46 undeveloped deposits on the Pine Point Project and of these, 10 are included in the PEA. Darnley Bay intends to evaluate the remaining deposits and, where considered appropriate, initiate confirmation drilling programs in an effort to add additional deposits to the mine plan prior to the Pre-Feasibility or Feasibility Study;  Explore for additional deposits. Although more than 100 distinct deposits have been outlined historically in the Pine Point area, there are several portions of the mineralized trends that remain under-explored and Darnley Bay intends to begin exploring this spring to identify additional deposits;  Additional metallurgical testing planned for 2017 could increase either recoveries, concentrate grades, or both;  Additional surface and groundwater studies to determine ways to lower the costs of water handling;  The PEA envisages the use of electric power sourced from a combination of local grid power and natural gas generators on the property. The Company intends to initiate a dialogue with the Northwest Territories Power Corporation to discuss the potential of obtaining more or all of Pine Point’s power needs from the grid through upgrades to the Taltson River hydroelectric dam, which would reduce capital and operating costs significantly; and  Evaluate the potential to purchase used mining and/or processing equipment to reduce the capital cost of the Project.

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DARNLEY BAY RESOURCES LTD. PINE POINT PEA

27 Recommendations

In the opinion of JDS, the financial analysis of this PEA demonstrates that the Pine Point Project has positive economics and warrants consideration for advancement to the feasibility level engineering by Darnley Bay. This more advanced study will further detail:

 A Mineral Reserve estimate;  Processing engineering based on the PEA flowsheets;  Project scheduling;  Capital and operating cost estimation; and  Economic results.

The study will improve the confidence in the Project design and execution and will result in an improved accuracy of project economics.

It is estimated that a FS and its supporting work programs would cost approximately $2.6 M. A breakdown of the key components of the next study phase is summarized in Table 27.1.

Table 27.1: Feasibility Study Cost Estimate

Item Cost ($) Hydrogeology/Geochemistry/Geotechnical Studies 700,000 Processing and Metallurgy 150,000 Feasibility Study 1,750,000 TOTAL 2,600,000 Source: JDS (2017)

Effective Date: April 18, 2017 27-1

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

28 References

Advanced Terra Testing, Inc., Testing report, Includes lab tests for (bulk) density, (particle) specific gravity, moisture content, porosity, and compressive strength on core samples from drill holes W-85- TV02, W85-TV03, R-190-TV05, R-190-TV07 and R190-TV10. 2006. AeroQuest Limited, Report on a helicopter-borne AeroTEM II electromagnetic and magnetic survey, Consulting report prepared for Tamerlane Ventures Inc., January 2005. Allen, C.F. and Walker, G.B., “Beneficiation of Industrial Minerals by Heavy-media Separation” AIME, January 1949 Mining Transactions. AMC Mining Consultants Canada “Pine Point Geotechnical Review Phase 2 R190 Deposit”. 2012. AMC Report #: AMC711018. Burns, R. F. and M.M. Kent (2001). “The Great Slave Reef Lead-Zinc Deposits, Pine Point, N.W.T.” – Private Report for Kent Burns Group. Calver, B. and D.J.M. Farnsworth. 1969. Open Pit Dewatering at Pine Point Mines. Canadian Mining and Metallurgical Bulletin (CIM). December, 1969. pp. 1290-1298. Campbell, Neil, “Stratigraphy and Structure of Pine Point Area, NWT”, Structural Geology of Canadian Ore Deposits, Volume II, Sixth Commonwealth Mining and Metallurgical Congress, 1957. Chlumsky, Armbrust & Meyer, LLC. (CAM). 2007. NI 43-101 Technical Report, Pine Point Project, Northwest Territories, Canada. Prepared for Tamerlane Ventures, Inc. Prepared by: F. Barnard, S. Milne and R.L. Sandefur. CMS, Pine Point Project, Metallurgical Report. Consulting report prepared by Confidential Metallurgical Services (Mississauga, ON) for Tamerlane Ventures, 2 September 2006. Cominco Ltd. 1984, Evaluation of Great Slave Reef Project. Internal report by R.C. Armstrong of Cominco subsidiary Pine Point Mines Ltd. dated 5 June 1984. Continental Metallurgical Services, “Metallurgical Testwork Report”, November 23, 2006 (Continental, 2006) – R-190 Deposit; Durston, K.J., 1979. Open Pit Dewatering at the Pine Point Operations of Cominco Limited. First International Mine Drainage Symposium, Denver, Colorado, May 20-23, 1979. Gann, 2017, Ni 43-101 Summary Technical Report Update of the Pine Point Mine Development Project, Northwest Territories, Canada Giroux, 1982a, Global Ore Reserve Estimates - O556 deposit: consulting report prepared October 1982 for Westmin Resources Ltd. by Sinclair, A.J., and Giroux, G.H, Approx. 50 pages. Giroux, 1982b, Global Ore Reserve Estimates - P499 deposit: consulting report dated November 1982 for Westmin Resources Ltd. by Sinclair, A.J., and Giroux, G.H., approx. 60 pages Giroux, 1982c, Global Ore Reserve Estimates - X25 deposit: consulting report dated November 1982 for Westmin Resources Ltd. by Sinclair, A.J., and Giroux, G.H., Approx. 75 pages

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Giroux, G.H. and McCartney, I., Report on the Great Slave Reef Lead-Zinc Deposits, Pine Point, NWT, for Tamerlane Ventures, November 2001, amended May 13, 2004. Giroux, G.H. and I. McCartney, Report on the Great Slave Reef lead-zinc deposits, Pine Point, NWT. Consulting report prepared for Kent Burns Group LLC November 2001, 37 pages. G&T Metallurgical Services, KM2984 “Metallurgical Testing on Samples from the Pine Point Project”, August 23, 2011 (G&T, 2011) – R-190 Deposit; G&T Metallurgical Services Ltd., KM3195 “Flotation Testing Dense Media Product N-204 Deposit”, January 13, 2012 (G&T, 2012) – N-204 Deposit. GTC Geologic Testing Consultants, 1983. Hydrogeologic Evaluation of the Pine Point – Great Slave Lake Region. Report to National Hydrology Research Institute, Environment Canada. Hannigan, P. 2007. “Metallogeny of the Pine Point Mississippi Valley-Type Zinc-Lead District, Southern Northwest Territories”, Mineral Deposits of Canada, ed. W.D. Goodfellow, pp. 609 to 632. Leach, D.L., J.B. Viets, N. Foley-Ayuso and D.P. Klein. Date Unknown. Mississippi Valley-Type Pb- Zn Deposits. Available at: pubs.usgs.gov/of/1995/ofr-95-0831/CHAP30.pdf Site Accessed: 19 August 2013. MineTech International Ltd. Technical Report on the Pine Point N-204 Lead-Zinc Deposit Latitude N 60° 57’, Longitude W 114° 02’ Near Fort Resolution, Northwest Territories, Canada. Prepared for: Tamerlane Ventures Inc. Authors: Wm. Douglas Roy, M.A.Sc., P. Eng., Patrick J. Hannon, M.A.Sc., P. Eng., Ian M. Flint, PhD., P. Eng. Effective Date of Report: March 23, 2012. MineTech International Ltd. Technical Report on the Pine Point Project Deposits: R-190, P-499, O- 556, X-25, Z-155, G-03 Near Pine Point, Northwest Territories, Canada NTS 85B/(10, 11, 14, 15, 16) N 60° 44’, W 115° 07’. Prepared for Tamerlane Ventures Inc. Authors: Patrick J. Hannon, M.A.Sc., P.Eng., William Douglas Roy, M.A.Sc., P.Eng., Ian M. Flint, PhD., P.Eng., Bradford Ernst, B.Eng., P.Eng., Edward Thornton, P. Eng. Effective Date of Report: April 2, 2012. Northwest Territories and Nunavut Mining Regulations (C.R.C., c. 1516), Schedule 1, accessed from http://laws-lois.justice.gc.ca/PDF/C.R.C.,_c._1516.pdf on July 22, 2013. PAH (Pincock, Allen & Holt). 2008. NI 43-101 Technical Report Update, Pine Point Project Northwest Territories, Canada. Prepared for Tamerlane Ventures Inc. July 30, 2008. Authors: Stuart E. Collins, P.E., J. Ross Conner, P.Geo., Craig F. Horlacher, Don M. Larsen, Ph.D., P.E., Barton G. Stone, C.P.G. PAH. 2011. (DRAFT) Technical Report for Resources Pine Point N204 Deposit, Northwest Territories, Canada. DRAFT Report Prepared for Tamerlane Ventures Inc. February 18, 2011. Prepared by: Craig F. Horlacher, MAusIMM. PHase Geochemistry, 2017. Review of Geochemical Data for the Pine Point Project. Prepared for Knight Piésold. Pincock, Allen & Holt, “NI 43-101 Technical Report Update Pine Point Project Northwest Territories, Canada”, July 30, 2008 (Pincock, Allen & Holt, 2008); Rhodes, D., et al. 1975. Pine Point Ore bodies and Their Relationship to the Stratigraphy, Structure, Dolomitization, and Karstification of the Middle Devonian Barrier Complex. Economic Geology. Vol. 79, pp. 991-1055.

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Rhodes, D., et al., 1984. Pine Point ore bodies and their relationship to the stratigraphy, structure, dolomitization, and karstification of the Middle Devonian Barrier Complex. Economic Geology. 79(5): 991-1055. SGS Lakefield Research Ltd., Project No. 11574-001 “Sulphide Ore Upgrading Using Gravity Separation”, May 22, 2007 (SGS, 2007a) – R-190 Deposit; SGS Lakefield Research Ltd., Projection No. 11633-001 “The Recovery of Lead and Zinc from Pine Point Ore Samples Using a Non-cyanide Flotation Method”, August 15, 2007 (SGS, 2007b) – R-190 Deposit; SGS Lakefield Research Ltd., Project 11790-001 “The Recovery of Lead and Zinc Using a Non- cyanide Flotation Method on the Pine Point Ore Samples”, June 2, 2008 (SGS, 2008) – O-556 / Z- 155 Deposit; SGS Canada Inc., Project No. 13001-002 “Bulk Concentrate Production from Heavy Liquid Separation on a Sample from the Pine Point N-204 Deposit”, December 26, 2011 (SGS, 2011) – N- 204 Deposit; and Siega, Albert & Paul Gann, “NI 43-101 Summary Technical Report Update of the Pine Point Mine Development Project Northwest Territories, Canada”, March 14, 2014 (Siega & Gann, 2014). Sinclair, A.J. and G.H. Giroux. 1982a Global Ore Reserve Estimates - O556 deposit. Consulting report prepared October 1982 for Westmin Resources Ltd. 50 pages. Sinclair, A.J. and G.H. Giroux. 1982b. Global Ore Reserve Estimates - P499 deposit. Consulting report dated November 1982 for Westmin Resources Ltd. 60 pages. Sinclair, A.J. and G.H. Giroux. 1982c.. Global Ore Reserve Estimates - X25 deposit. Consulting report dated November 1982 for Westmin Resources Ltd. 75 pages. Skall, H. 1975. The Paleoenvironment of the Pine Point Lead-Zinc District. Economic Geology. 70(1): 22-47. Stevenson International Groundwater Consultants Ltd., 1984. A Study of the Great Slave Reef Pine Point Mines Aquifer, Based on Analyses of Selected Pine Point Mines Pumping Test Data. Prepared for: Mr. A.W. Randall, Westmin Resources Limited. Turner W.A., Pierce K.L., and Cairns K.A., 2002. Great Slave Reef (GSR) Project Drill Hole Database. A compilation of the drill hole locations, drill logs, and associated geochemical data for the Great Slave Reef Joint Venture Project; Interior Platform, Northwest Territories, Canada (NTS 85B11 to 14). NWT Open Report 2002-001, C.S. Lord Northern Geoscience Centre, Yellowknife, Northwest Territories. U.S. Geological Survey, 2013, Mineral commodity summaries 2013: U.S. Geological Survey, Technical Report. 198 pp. Vogwill, R.I.J. 1976. Some Practical Aspects of Open-Pit Dewatering at Pine Point. Bull. Can. Inst. Min. Met. 69(768): 76-88. Vogwill, R.I.J., 1981. R-190 Zone Aquifer Test Analysis and Preliminary Dewatering Design. Prepared for: Western Mines Ltd. Great Slave Reef Project.

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Westmin Resources Ltd. 1981. Great Slave Reef Project - Summary report of 1981 drilling program. Internal report by A.W. Randall. May 1981. 20 pages plus maps of O-555 and O-556. Westmin Resources Ltd., 1983a, Special Studies Group: internal geological report by J.L. Hardy dated April 1983, Approx. 100 pages. Westmin Resources Ltd., 1983b, Great Slave Reef - R190, Review of mine planning: internal report by C.T. Penney dated Nov 1983, 18 pages plus appendices. Westmin Resources Ltd., 1983c, Great Slave Reef Project, Review of Mine Planning - Phase I: internal report by C.T. Penney dated 8 Dec 1983, Approx. 25 pages. Westmin Resources Ltd., 1984, Great Slave Reef - Exploration meeting; internal report by A.W. Randall dated November 20-22, 1984, Approx 30 pages. Westmin Resources Ltd., (undated, 1980's) Great Slave Reef Project, West Reef, list of diamond drill holes: internal report, handwritten and typed, Approx 60 pages.

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DARNLEY BAY RESOURCES LTD. PINE POINT PEA

29 Units of Measure, Abbreviations and Acronyms

Symbol/Abbreviation description ' minute (plane angle) " second (plane angle) or inches ° degree °C degrees Celsius 3D three-dimensions A ampere a annum (year) ac acre Acfm actual cubic feet per minute amsl above mean sea level AN ammonium nitrate ARD acid rock drainage Au gold B billion BD bulk density BRS B-reef substrate Bt billion tonnes BTU British thermal unit bya billion years ago C$ dollar (Canadian) Ca calcium CAPEX Capital cost estimate cfm cubic feet per minute CHP combined heat and power plant CIM Canadian institute of mining and metallurgy cm centimetre cm2 square centimetre cm3 cubic centimetre COG Cut-off grade cP centipoise CP Cluster Pit Cr chromium Cu copper d day d/a days per year (annum) d/wk days per week Darnley Bay Darnley Bay Resources Ltd. dB decibel dBa decibel adjusted DMS dense media separation dmt dry metric tonne DSO Direct shipment ore DWT dead weight tonnes EA environmental assessment EIS environmental impact statement

Effective Date: April 18, 2017 29-5

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

Symbol/Abbreviation description EM Electromagnetic ERD explosives regulatory division Fe Iron FEL front-end loader Fm Formation FOC fisheries and oceans Canada FS Feasibility study ft foot ft2 square foot ft3 cubic foot ft3/s cubic feet per second g gram G&A general and administrative g/cm3 grams per cubic metre g/L grams per litre g/t grams per tonne Ga billion years gal gallon (us) GJ gigajoule GPa gigapascal gpm gallons per minute (us) GSC geological survey of Canada GSR The Great Slave Reef GW gigawatt h hour h/a hours per year h/d hours per day h/wk hours per week ha hectare (10,000 m2) ha hectare HG high grade HLS Heavy liquid separation hp horsepower HQ drill core diameter of 63.5 mm Hz hertz HZ Hinge Zone in inch in2 square inch in3 cubic inch INAC Indigenous and Northern Affairs Canada IP Induced polarization IRR internal rate of return JDS JDS Energy & Mining Inc. K hydraulic conductivity k kilo (thousand) kg kilogram kg/h kilograms per hour

Effective Date: April 18, 2017 29-6

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

Symbol/Abbreviation description kg/m2 kilograms per square metre kg/m3 kilograms per cubic metre km kilometre km/h kilometres per hour km2 square kilometre KP Knight Piesold kPa kilopascal kt kilotonne kV kilovolt kVA kilovolt-ampere kW kilowatt kWh kilowatt hour kWh/a kilowatt hours per year kWh/t kilowatt hours per tonne L litre L/min litres per minute L/s litres per second LG low grade LNG Liquid natural gas LOM life of mine m metre M million M$ Million dollars m/min metres per minute m/s metres per second m2 square metre m3 cubic metre m3/h cubic metres per hour m3/s cubic metres per second Ma million years MAAT mean annual air temperature MAE mean annual evaporation MAGT mean annual ground temperature mamsl metres above mean sea level MAP mean annual precipitation masl metres above mean sea level Mb/s megabytes per second mbgs metres below ground surface Mbm3 million bank cubic metres Mbm3/a million bank cubic metres per annum mbs metres below surface mbsl metres below sea level Mct million carats mg milligram mg/L milligrams per litre min minute (time) mL millilitre

Effective Date: April 18, 2017 29-7

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

Symbol/Abbreviation description mm millimetre Mm3 million cubic metres MMER metal mining effluent regulations MMSIM metamorphosed massive sulphide indicator minerals mo month MPa megapascal Mt million metric tonnes MVA megavolt-ampere MVT Mississippi Valley-Type MW megawatt NAD North American datum NG normal grade Ni nickel NI 43-101 national instrument 43-101 Nm3/h normal cubic metres per hour NPV Net present value NQ drill core diameter of 47.6 mm NRC natural resources Canada NRS net smelter return NTPC Northwest Territories Power Corp. NWT Northwest Territories OP open pit OPEX Operating cost estimate OSA overall slope angles oz troy ounce P.Geo. professional geoscientist Pa Pascal PAG potentially acid generating Pb lead PEA preliminary economic assessment PEM Potential economic material PFS preliminary feasibility study PMA Particle mineral analysis PMF probable maximum flood ppb parts per billion PPHC Pine Point Holding Corp. ppm parts per million psi pounds per square inch QA/QC quality assurance/quality control QP qualified person RC reverse circulation RF Revenue factor RMR rock mass rating ROM run of mine rpm revolutions per minute RQD rock quality designation s second (time) S.G. specific gravity

Effective Date: April 18, 2017 29-8

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

Symbol/Abbreviation description Scfm standard cubic feet per minute SEDEX sedimentary exhalative SFD size frequency distribution SG specific gravity SI International system of units (metric) SRM Standard reference material st/kg stones per kilogram st/t stones per metric tonne t tonne (1,000 kg) (metric ton) t metric tonne t/a tonnes per year t/d tonnes per day t/h tonnes per hour Tamerlane Tamerlane Ventures Inc. TCR total core recovery TFFE target for further exploration TMF tailings management facility tph tonnes per hour ts/hm3 tonnes seconds per hour metre cubed US united states US$ dollar (American) UTM universal transverse mercator V volt VMS volcanic massive sulphide w/w weight/weight Westmin Westmin Resources Ltd. wk week wmt wet metric tonne WRSA waste rock storage area XRF X-ray fluorescence μm microns μm Micrometer Zn zinc

Scientific Notation Number Equivalent 1.0E+00 1 1.0E+01 10 1.0E+02 100 1.0E+03 1,000 1.0E+04 10,000 1.0E+05 100,000 1.0E+06 1,000,000 1.0E+07 10,000,000 1.0E+09 1,000,000,000 1.0E+10 10,000,000,000

Effective Date: April 18, 2017 29-9

DARNLEY BAY RESOURCES LTD. PINE POINT PEA

Rock Type Description

CaSO42H2O Gypsum

CaSO4 Anhydrite NaCl Sodium salt dl Dolomite py Pyrite Sph Sphalerite gn galena

Effective Date: April 18, 2017 29-10

Appendix A – QP Certificates PARTNERS IN JDS Energy & Mining Inc. ACHIEVING Suite 900 – 999 West Hastings Street MAXIMUM Vancouver, BC V6C 2W2 RESOURCE t 604.558.6300 DEVELOPMENT VALUE jdsmining.ca

CERTIFICATE OF AUTHOR

I, Garett Macdonald, P. Eng., do hereby certify that:

1. This certificate applies to the Technical Report entitled “NI 43-101 Preliminary Economic Assessment Technical Report on the Pine Point Zinc Project, Northwest Territories, Canada”, with an effective date of April 18, 2017, (the “Technical Report”) prepared for Darnley Bay Resources Limited;

2. I am currently employed as Vice President Project Development with JDS Energy & Mining Inc. with an office at Suite 3670, 130 King Street West, Toronto, ON M5X 1E2;

3. I am a graduate of Laurentian University with a B.Eng. in Mining Engineering, 1996. I have practiced my profession continuously since 1996;

I have worked in technical, operations and management positions at mines in Canada. I have been an independent consultant for over one year and have managed preliminary economic assessments, pre- feasibility studies, feasibility studies and technical due diligence reviews.

I am a Registered Professional Mining Engineer in Ontario (#90475344)

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. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of NI 43-101;

4. I have not visited the Pine Point project.

5. I am responsible for Sections 1 to 3, 18 (except 18.6 and 18.8), 19 to 27 of this Technical Report;

6. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of the NI 43-101;

7. I have had no prior involvement with the property that is the subject of this Technical Report

8. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, this Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading;

9. I have read NI 43-101, and the Technical Report has been prepared in accordance with NI 43-101 and Form 43-101F1.

Effective Date: April 18, 2017 Signing Date: June 1, 2017

(original signed and sealed) “Garett Macdonald, P.Eng.”

Garett Macdonald, P. Eng.

VANCOUVER | TORONTO | KELOWNA | WHITEHORSE | YELLOWKNIFE | TUCSON | HERMOSILLO

PARTNERS IN JDS Energy & Mining Inc. ACHIEVING Suite 900 – 999 West Hastings Street MAXIMUM Vancouver, BC V6C 2W2 RESOURCE t 604.558.6300 DEVELOPMENT VALUE jdsmining.ca

CERTIFICATE OF AUTHOR

I, Kelly McLeod, P. Eng., do hereby certify that:

1. This certificate applies to the Technical Report entitled “NI 43-101 Preliminary Economic Assessment Technical Report on the Pine Point Zinc Project, Northwest Territories, Canada”, with an effective date of April 18, 2017, (the “Technical Report”) prepared for Darnley Bay Resources Limited;

2. I am currently employed as a Senior Engineer, Metallurgy, with JDS Energy & Mining Inc. with an office at Suite 900 – 999 West Hastings Street, Vancouver, British Columbia, V6C 2W2;

3. I am a Professional Metallurgical Engineer registered with the APEGBC, P.Eng. #15868. I am a graduate of McMaster University with a Bachelors of Engineering, Metallurgy, 1984. I have practiced my profession intermittently since 1984 and have worked for the last 10 years consulting in the mining industry in metallurgy and process design engineering;

4. 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. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of NI 43-101;

5. I have not visited the Pine Point project.

6. I am responsible for Sections 13 and 17 of this Technical Report;

7. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of the NI 43-101;

8. I have had no prior involvement with the property that is the subject of this Technical Report

9. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, this Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading;

10. I have read NI 43-101, and the Technical Report has been prepared in accordance with NI 43-101 and Form 43-101F1.

Effective Date: April 18, 2017 Signing Date: June 1, 2017

(original signed and sealed) “Kelly McLeod, P.Eng.”

Kelly McLeod, P. Eng.

VANCOUVER | TORONTO | KELOWNA | WHITEHORSE | YELLOWKNIFE | TUCSON | HERMOSILLO

PARTNERS IN JDS Energy & Mining Inc. ACHIEVING Suite 900 – 999 West Hastings Street MAXIMUM Vancouver, BC V6C 2W2 RESOURCE t 604.558.6300 DEVELOPMENT VALUE jdsmining.ca

CERTIFICATE OF AUTHOR

I, Dino Pilotto, P. Eng., do hereby certify that:

1. This certificate applies to the Technical Report entitled “NI 43-101 Preliminary Economic Assessment Technical Report on the Pine Point Zinc Project, Northwest Territories, Canada”, with an effective date of April 18, 2017, (the “Technical Report”) prepared for Darnley Bay Resources Limited;

2. I am currently employed as Engineering Manager with JDS Energy & Mining Inc. with an office at Suite 900-999 West Hastings Street, Vancouver, BC, V6C 2W2;

3. I am a Professional Mining Engineer (P.Eng. #2527) registered with the Association of Professional Engineers Yukon. I am also a registered Professional Mining Engineer in British Columbia, Alberta, Northwest Territories and Nunavut. I am a graduate of the University of British Columbia with a B.Sc. in Mining and Mineral Process Engineering (1987). I have practiced my profession continuously since June 1987. I have been involved with mining operations, mine engineering and consulting covering a variety of commodities at locations in North America, South America, Africa, and Eastern Europe;

4. 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;

5. I have visited the Pine Point project on March 3, 2017.

6. I am responsible for Sections 15 and 16 of this Technical Report;

7. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of the NI 43-101;

8. I have had no prior involvement with the property that is the subject of this Technical Report

9. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, this Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading;

10. I have read NI 43-101, and the Technical Report has been prepared in accordance with NI 43-101 and Form 43-101F1.

Effective Date: April 18, 2017 Signing Date: June 1, 2017

(original signed and sealed) “Dino Pillotto, P.Eng.”

Dino Pilotto, P. Eng.

VANCOUVER | TORONTO | KELOWNA | WHITEHORSE | YELLOWKNIFE | TUCSON | HERMOSILLO

www.knig h tpiesold .com

CERTIFICATE OF AUTHOR

I, Kenneth Embree, P. Eng., do hereby certify that:

1. This certificate applies to the Technical Report entitled “NI 43-101 Preliminary Economic Assessment Technical Report on the Pine Point Zinc Project, Northwest Territories, Canada”, with an effective date of April 18, 2017, (the “Technical Report”) prepared for Darnley Bay Resources Limited;

2. I am employed as Managing Principal of Knight Piésold Ltd. with an office at Suite 1400 - 750 West Pender Street, Vancouver, British Columbia, V6C 2T8.

3. I am a Professional Engineer in good standing with the Association of Professional Engineers and Geoscientists of British Columbia in the area of geological engineering (No. 17439). I am also registered as a Professional Engineer in Ontario (No. 100040332).

4. I am a graduate of the University of Saskatchewan with a B.Sc. in Geological Engineering (1986). I have practiced my profession continuously since 1986. My experience includes tailings and waste and water management for mine developments in Canada, the US and South America.

5. I have read the definition of "qualified person" set out in National Instrument 43-101 (NI 43-101) and certify that by reason of my education, 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 visited the Pine Point Project on March 3, 2017.

7. I am responsible for Sections 18.6 and 18.8 of this Technical Report.

8. I am independent of the Issuer and related companies applying all of the tests in Section 1.5 of NI 43-101.

9. I have had no prior involvement with the property that is the subject of this Technical Report.

10. As of the effective date of this Technical Report, to the best of my knowledge, information and belief, this Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

11. I have read NI 43-101, and the Technical Report has been prepared in accordance with NI 43-101 and Form 43-101F1.

Effective Date: April 18, 2017 Signing Date: June 1, 2017

(original signed and sealed) “Ken Embree, P.Eng.”

Ken Embree, P. Eng.

Knight Piésold Ltd. | Suite 1400 – 750 West Pender St, Vancouver, BC Canada V6C 2T8 | p. +1.604.685.0543 f. +1.604.685.0147 CERTIFICATE OF QUALIFICATIONS

I, Albert Siega, do hereby certify that:

1. I am a mining engineer, located at #102, 15466, 16th Ave, Whiterock, British Columbia, Canada 2. I am a graduate of the University of Albert with a B.Sc. degree in Mining Engineering (1986) and an MBA in finanace (1992). 3. I am a Competent Person. I am regisitered as a Professional Engineer in British Columbia, Canada: License #34957 & NWT L3635 4. I have worked in the mineral resource industry for over twenty years. My expereince for the purpose of this technical report is twenty years of mine engineering with : Areva, Syncrude, Athabasca Gold, Hunter Dickinson, Century Mining, SilverCorp and Golder Associates. 5. I have read the definition of “Qualified Person” set out in National Instrument 43- 101 (NI-43-101), and certify that I am qualified. 6. I am Independent of the issuer and in compliance with section 1.5 of NI-43-101. I worked for Tamerlane Ventures Ltd. during various periods over the past 10 years. I own no securities in this issuer Darnley Bay Resources Ltd., or any related companies. 7. I have read NI43-101, Form 43-101-F1 and the Companion Policy 43-1021CP, this technical report has been prepared in compliance with that instument, form and policy. 8. My prior involvement with the property included block modeling, mine design and economic analysis of deposits in the project area. 9. I conducted a personal inspection of the property on September 5, 2013. 10. I am responsible for Sections 1-7, and 14. I conducted the block modeling used to derive the Resources Estimates. 11. I have read the Technical Report. To the best of my knowledge, information and belief, it contains all the information required to be disclosed, such that the report is not misleading; it fairly and accurately represents the information that I am reponsible for. 12. I consent to the filing of this Technical Report with any stock excahange or regulatory authority and publication by them, including electronic publication except where publication would violate exchange rules.

June 1, 2017

Orignal signed

______

Albert Daniel Siega, P.Eng CERTIFICATE OF QUALIFICATIONS

I, Paul Gann, do herby certify that:

1. I am a practicing geologist residing at #204 -55 Blackberry Dr. New Westminster, B.C. 2. This certificate applies to the report titled: NI 43-101 Summary Technical Report, Pine Point Mine Development Project, Northwest Territories, Canada, written for Darnley Bay Resources Inc. 3. I graduated from the University of Calgary in 1982 with a BSc in Geology. I have been a practicing geologist for over 33 years. I have been involved in the exploration for base and precious metals, lead & zinc SEDEX deposits, REE, diamonds, tin, tungsten and platinum/palladium ultramafic deposits. My work has been conducted at sites across Canada as well as the USA. 4. I am a Competent Person under the professional geologist registration requirements of British Columbia. I am a Professional Geoscientist (P.Geo.) registered with the Association of Professional Engineers and Geoscientists of British Columbia, member #30164 and have been a member in good standing since 2006. I am also a member in good standing of Association of Professional Geologists of Ontario member #1848 5. Seasonal weather conditions have prevented me from accessing the property and obtaining beneficial information from it. Therefore, I have not visited the property in discussion in this report. 6. I am responsible the sections 4.0 - 12.0. 7. I am an independent issuer as applied in section 1.5 of the Nl 43-101.I have never been employed by the issuer in the past. I own no securities in the issuer, Darnley Bay Resources Inc. or any related companies. 8. I have read the definition of "Qualified Person" set out in National Instrument 43-101 (NI 43-101) and certify that I am qualified. 9. I have read NI 43-101, Form 43-101-F1and the Companion Policy 43-101CP, including the June,2011 amendments; this technical report has been prepared in compliance with that instrument, form and policy. 10. On this date of Certificate, I have read the Technical Report. To the best of my knowledge, information and belief, it contains all the scientific and technical information that is required to be disclosed to make the technical report not misleading. It fairly and accurately represents the information that I am responsible for. 11. I consent to the filling of this Technical report with any stock exchange or regulatory authority and any publication by them, including electronic publication in the public company files or websites by the public, except where this would violate exchange rules.

June 1, 2017

Orignal signed

______

Paul Gann PGeo