EX-99.1 2 exhibit99-1.htm EXHIBIT 99-1 NovaGold Resources Inc.: Exhibit 99.1 - Filed by newsfilecorp.com

About IMPORTANT NOTICE
 

This report was prepared as a National Instrument 43-101 Technical Report for NovaGold Resources Inc. (NovaGold) by AMEC Americas Limited (AMEC). The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in AMEC's services, 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. This report is intended for use by NovaGold subject to terms and conditions of its contract with AMEC. Except for the purposes legislated under Canadian provincial securities law, any other uses of this report by any third party is at that party's sole risk.



 
DONLIN GOLD PROJECT
ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

C O N T E N T S

1.0  SUMMARY 1-1 
  1.1  Principal Outcomes 1-1 
  1.2  Location, Climate, and Access 1-2 
  1.3  Agreements 1-2 
  1.4  Mineral Tenure 1-4 
  1.5  Surface Rights 1-4 
  1.6  Royalties 1-4 
  1.7  Environment, Permitting and Socio-Economics 1-4 
  1.8  Geology and Mineralization 1-5 
  1.9  Exploration 1-6 
  1.10  Exploration Potential 1-7 
  1.11  Drilling 1-7 
  1.12  Sample Analysis and Security 1-8 
  1.13  Data Verification 1-9 
  1.14  Metallurgical Testwork 1-9 
  1.15  Mineral Resource Estimate 1-13 
  1.16  Mineral Reserve Estimate 1-15 
  1.17  Proposed Mine Plan 1-17 
  1.18  Process Design 1-18 
  1.19  Planned Project Infrastructure 1-20 
  1.20  Markets 1-21 
  1.21  Capital Costs 1-21 
  1.22  Operating Costs 1-22 
  1.23  Financial Analysis 1-22 
  1.24  Preliminary Development Schedule 1-23 
  1.25  Conclusions 1-23 
  1.26  Recommendations 1-24 
2.0  INTRODUCTION 2-1 
  2.1  Terms of Reference 2-1 
  2.2  Qualified Persons 2-4 
  2.3  Site Visits and Scope of Personal Inspections 2-4 
  2.4  Effective Dates 2-5 
  2.5  Previous Technical Reports 2-5 
  2.6  Information Sources and References 2-6 
3.0  RELIANCE ON OTHER EXPERTS 3-1 
  3.1  Mineral Tenure 3-1 
  3.2  Surface Rights 3-2 
  3.3  Agreements 3-2 
  3.4  Royalties 3-3 
  3.5  Marketing 3-3 
  3.6  Taxation 3-3 
4.0  PROPERTY DESCRIPTION AND LOCATION 4-1 

     
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ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

  4.1  Location 4-1 
  4.2  Project Ownership History 4-1 
  4.3  Lease Rights 4-4 
  4.4  Mineral Tenure 4-5 
  4.5  Surface Rights 4-1 
  4.6  Royalties and Encumbrances 4-2 
  4.7  Permits 4-2 
  4.8  Environmental Liabilities 4-3 
  4.9  Social License 4-3 
  4.10  Significant Risk Factors 4-3 
  4.11  Comments on Section 4 4-3 
5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY 5-1 
  5.1  Accessibility 5-1 
  5.2  Climate 5-1 
  5.3  Local Resources and Infrastructure 5-2 
  5.4  Physiography 5-2 
  5.5  Sufficiency of Surface Rights 5-2 
  5.6  Comments on Section 5 5-2 
6.0  HISTORY 6-1 
7.0  GEOLOGICAL SETTING AND MINERALIZATION 7-1 
  7.1  Regional Geology 7-1 
  7.2  Project Geology 7-2 
    7.2.1  Lithologies 7-2 
    7.2.2  Structure 7-3 
  7.3  Deposit Setting 7-3 
  7.4  Paragenesis 7-4 
  7.5  Deposit Geology 7-5 
    7.5.1  Sedimentary Rocks 7-5 
    7.5.2  Igneous Rocks 7-6 
    7.5.3  Structure 7-7 
  7.6  Deposits 7-9 
  7.7  Mineralization 7-10 
    7.7.1  Vein and Disseminated Mineralization 7-10 
  7.8  Alteration 7-12 
  7.9  Minor Elements 7-13 
  7.10  Comments on Section 7 7-13 
8.0  DEPOSIT TYPES 8-1 
  8.1  Comments on Section 8 8-1 
9.0  EXPLORATION 9-1 
  9.1  Grids and Surveys 9-1 
  9.2  Geological Mapping 9-1 
  9.3  Geochemical Sampling 9-1 
  9.4  Geophysics 9-1 

     
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NI 43-101 TECHNICAL REPORT
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  9.5  Pits and Trenches 9-4 
  9.6  Petrology, Mineralogy, and Research Studies 9-4 
  9.7  Geotechnical and Hydrological Studies 9-4 
  9.8  Metallurgical Studies 9-4 
  9.9  Exploration Potential 9-5 
    9.9.1  Far Side 9-5 
    9.9.2  Duqum 9-5 
    9.9.3  Snow/Quartz 9-7 
    9.9.4  Dome 9-7 
    9.9.5  Ophir 9-8 
  9.10  Comments on Section 9 9-8 
10.0  DRILLING 10-1 
  10.1  Drill Methods 10-1 
  10.2  Geological Logging 10-6 
  10.3  Recovery 10-7 
  10.4  Collar Surveys 10-7 
  10.5  Down-hole Surveys 10-7 
  10.6  Geotechnical and Hydrological Drilling 10-8 
  10.7  Metallurgical Drilling 10-8 
  10.8  Condemnation Drilling 10-8 
  10.9  Drill Orientations 10-10 
  10.10  Twin Drilling 10-11 
  10.11  Drilled Width versus True Thickness 10-11 
  10.12  Summary of Drill Intercepts 10-11 
  10.13  Comments on Section 10 10-11 
11.0  SAMPLE PREPARATION, ANALYSES, AND SECURITY 11-1 
  11.1  Sampling Methods 11-1 
  11.2  Metallurgical Sampling 11-1 
  11.3  Density/Specific Gravity Determinations 11-1 
  11.4  Analytical and Test Laboratories 11-3 
  11.5  Sample Preparation and Analysis 11-3 
  11.6  Quality Assurance and Quality Control 11-5 
    11.6.1  1995–2002 QA/QC Protocol 11-5 
    11.6.2  2005–2006 QA/QC Protocol 11-6 
    11.6.3  2007–2010 QA/QC Protocol 11-6 
    11.6.4  Standard Reference Materials 11-6 
    11.6.5  Blank Materials 11-7 
  11.7  Databases 11-7 
  11.8  Sample Security 11-8 
  11.9  Comments on Section 11 11-9 
12.0  DATA VERIFICATION 12-1 
    12.1.1  AMEC (2002) 12-1 
    12.1.2  NovaGold (2005) 12-1 
    12.1.3  NovaGold (2008) 12-1 
  12.2  AMEC (2011) 12-2 

     
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ALASKA, USA
NI 43-101 TECHNICAL REPORT
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  12.3  Comments on Section 12 12-2 
13.0  MINERAL PROCESSING AND METALLURGICAL TESTING 13-1 
  13.1  Metallurgical Testwork 13-1 
    13.1.1  Domains 13-1 
    13.1.2  Gold Deportment 13-4 
    13.1.3  Mercury, Chlorine, Carbonates and Organic Carbon Deportment 13-4 
    13.1.4  Samples 13-5 
    13.1.5  Comminution 13-5 
    13.1.6  Flotation 13-18 
    13.1.7  Pressure Oxidation 13-25 
    13.1.8  Neutralization 13-34 
    13.1.9  Carbon-in-Leach (CIL) 13-42 
    13.1.10  Thickening and Counter-Current Decantation (CCD) Washing 13-47 
    13.1.11  Environmental Testwork 13-48 
  13.2  Recovery Estimates 13-49 
    13.2.1  Flotation 13-50 
    13.2.2  Pressure Oxidation 13-55 
    13.2.3  Overall Plant Gold Recovery 13-57 
  13.3  Metallurgical Variability 13-57 
  13.4  Deleterious Elements 13-57 
  13.5  Comments on Section 13 13-57 
14.0  MINERAL RESOURCE ESTIMATES 14-1 
  14.1  Key Assumptions/Basis of Estimate 14-1 
  14.2  Geological Models 14-1 
  14.3  Exploratory Data Analysis 14-2 
  14.4  Density Assignment 14-4 
  14.5  Grade Capping/Outlier Restrictions 14-4 
    14.5.1  Gold, Sulphur, Arsenic, Mercury, and Antimony Grade Caps 14-4 
    14.5.2  Neutralization Potential Grade Caps 14-5 
  14.6  Composites 14-5 
    14.6.1  Gold, Sulphur, Arsenic, Mercury, and Antimony Composites 14-5 
    14.6.2  Neutralization Potential Composites 14-6 
  14.7  Gold and Sulphur Indicator Models 14-6 
    14.7.1  Overburden 14-7 
  14.8  Variography Performed in Support of PAG Model 14-8 
  14.9  Estimation/Interpolation Methods 14-8 
    14.9.1  Gold 14-8 
    14.9.2  Sulphur 14-9 
    14.9.3  Arsenic, Mercury and Antimony 14-9 
    14.9.4  Calcium, Magnesium and Carbon Di-oxide 14-10 
    14.9.5  Neutralization Potential 14-10 
    14.9.6  Classification of Waste Rock Management Categories 14-10 
  14.10  Block Model Validation 14-11 
  14.11  Dilution 14-11 
  14.12  Classification of Mineral Resources 14-12 

     
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ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

  14.13  Reasonable Prospects of Economic Extraction 14-12 
    14.13.1  NSR Calculations for Marginal Cut-off Application 14-12 
  14.14  AMEC Review 14-14 
  14.15  Mineral Resource Statement 14-17 
  14.16  Comments on Section 14 14-18 
15.0  MINERAL RESERVE ESTIMATES 15-1 
  15.1  Key Assumptions/Basis of Estimate 15-1 
  15.2  Dilution and Mining Losses 15-1 
  15.3  Conversion Factors from Mineral Resources to Mineral Reserves 15-4 
    15.3.1  Mining Costs 15-4 
    15.3.2  Processing Costs 15-5 
    15.3.3  Recovery 15-6 
    15.3.4  Overhead Costs 15-6 
    15.3.5  Refining, Freight, and Royalties 15-6 
    15.3.6  Metal Prices 15-6 
    15.3.7  Pit Slopes 15-6 
    15.3.8  Sensitivity of Optimized Pit 15-7 
  15.4  Mineral Reserves Statement 15-8 
  15.5  Comments on Section 15 15-9 
16.0  MINING METHODS 16-1 
  16.1  Throughput Considerations 16-1 
  16.2  Pit Design 16-1 
  16.3  Geotechnical Considerations 16-1 
    16.3.1  Rock Mass Model 16-1 
    16.3.2  Open Pit Slope Design 16-2 
    16.3.3  Recommended Design Parameters 16-3 
  16.4  Pit Phases 16-3 
  16.5  Haul Roads 16-5 
  16.6  Production Schedule 16-6 
    16.6.1  Planned Production Schedule 16-6 
    16.6.2  Pit–Phase Mining Rates 16-8 
    16.6.3  Mill Feed Plan 16-9 
  16.7  Ore Stockpiles 16-11 
  16.8  Waste Rock Scheduling and NAG/PAG Management 16-13 
    16.8.1  Overburden Scheduling and Concurrent Reclamation 16-14 
  16.9  Water Management and Treatment 16-14 
  16.10  Ore Control 16-15 
  16.11  Blasting and Explosives 16-16 
  16.12  Mining Equipment 16-16 
    16.12.1  Drilling 16-16 
    16.12.2  Loading 16-17 
    16.12.3  Hauling 16-17 
    16.12.4  Secondary Fleet 16-17 
    16.12.5  Support Equipment 16-17 
    16.12.6  Maintenance Considerations 16-22 

     
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ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

    16.12.7  Health and Safety Considerations 16-22 
    16.12.8  Communications Considerations 16-23 
  16.13  Consumables 16-23 
  16.14  Work Schedule 16-23 
  16.15  Comments on Section 16 16-24 
17.0  RECOVERY METHODS 17-1 
  17.1  Plant Design 17-1 
    17.1.1  General 17-1 
    17.1.2  Crushing and Coarse Ore Stockpile 17-1 
    17.1.3  Grinding and Pebble Crushing 17-3 
    17.1.4  Flotation 17-4 
    17.1.5  Thickening, Concentrate Storage, Acidulation, and CCD Washing   17-5 
    17.1.6  Autoclave Plant 17-6 
    17.1.7  CCD POX Thickening and Washing 17-8 
    17.1.8  Flotation Tailings (FT) Neutralization 17-8 
    17.1.9  Solids CIL Neutralization 17-9 
    17.1.10  Carbon-in-Leach Cyanidation Circuit 17-9 
    17.1.11  Cyanide Destruction System 17-10 
    17.1.12  Carbon Elution, Electrowinning, Reactivation, and Gold Refining  17-10 
    17.1.13   Mercury Abatement Systems 17-11 
    17.1.14  Reagent Preparation 17-13 
  17.2  Process Services 17-14 
    17.2.1  Air 17-14 
    17.2.2  Plant Water Distribution 17-15 
  17.3  Process Ventilation 17-16 
  17.4  Control System 17-17 
  17.5  Laboratories 17-18 
  17.6  Mill Feed Schedule 17-18 
  17.7 Comments on Section 17 17-19 
18.0  PROJECT INFRASTRUCTURE 18-1 
  18.1  Access and Logistics 18-1 
    18.1.1  Port-to-Mine Access Road 18-1 
    18.1.2  Road Construction 18-1 
    18.1.3  Airstrip 18-2 
    18.1.4  Cargoes 18-3 
    18.1.5  Fuel 18-3 
  18.2  Site Facilities 18-4 
    18.2.1  Site Investigations 18-4 
    18.2.2  Plant Site Design Considerations 18-5 
    18.2.3  Plant Site Facilities 18-7 
  18.3  Camps and Accommodation 18-9 
  18.4  Waste Storage Facilities 18-9 
    18.4.1  Location 18-9 
    18.4.2  Acid-base Accounting 18-9 
    18.4.3  Construction Plan 18-11 
  18.5  Tailings Storage Facilities 18-12 

     
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ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

    18.5.1  Design Considerations 18-14 
  18.6  Water Management 18-16 
    18.6.1  Water Balance 18-16 
    18.6.2  Construction Water Management Strategy 18-17 
    18.6.3  Operations Water Management Strategy 18-21 
    18.6.4  Closure Water Management Strategy 18-24 
    18.6.5  Potable Water 18-26 
    18.6.6  Fire Water 18-26 
  18.7  Bethel Marine Terminal 18-26 
  18.8  Kuskokwim River Dock Site 18-27 
  18.9  Power and Electrical 18-29 
  18.10  Gas Pipeline 18-30 
  18.11  Fuel 18-31 
    18.11.1  Diesel 18-31 
    18.11.2  Natural Gas 18-31 
  18.12  Comment on Section 18 18-31 
19.0  MARKET STUDIES AND CONTRACTS 19-1 
  19.1  Marketing Partnership Agreement 19-1 
  19.2 Gold Marketing 19-1 
  19.3  Comments on Section 19 19-1 
20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT 20-1 
  20.1  Baseline Studies 20-1 
  20.2  Environmental Issues 20-1 
  20.3  Closure Plan 20-1 
    20.3.1  Water Treatment Plant 20-4 
    20.3.2  Tailings Storage Facility 20-5 
    20.3.3  Waste Rock Facility 20-5 
    20.3.4  Roads and Airstrip 20-6 
    20.3.5  Foundations and Buildings 20-6 
    20.3.6  Waste Disposal 20-6 
    20.3.7  Port Facilities, Access Road, Airstrip, and Personnel Camp 20-6 
    20.3.8  Mobile Equipment 20-7 
    20.3.9  Trust Fund 20-7 
    20.3.10  Closure Cost Estimate 20-7 
  20.4  Permitting 20-8 
    20.4.1  Exploration Stage Permitting 20-9 
    20.4.2  Pre-Application Phase 20-9 
    20.4.3  The NEPA Process and Permit Applications 20-11 
    20.4.4  Laws, Regulations, and Permit Requirements 20-12 
  20.5  Considerations of Social and Community Impacts 20-12 
    20.5.1  Stakeholders 20-17 
    20.5.2  Community Development and Sustainability 20-19 
  20.6  Comments on Section 20 20-20 
21.0  CAPITAL AND OPERATING COSTS 21-1 
  21.1  Capital Cost Estimates 21-1 

     
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DONLIN GOLD PROJECT
ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

    21.1.1  Basis of Estimate 21-1 
    21.1.2  Contingency 21-2 
    21.1.3  First Fill 21-3 
    21.1.4  Sustaining Capital 21-3 
    21.1.5  Capital Cost Summary 21-4 
  21.2  Operating Cost Estimates 21-6 
    21.2.1  Basis of Estimate 21-6 
    21.2.2  Mine Operating Costs 21-6 
    21.2.3  Process Operating Costs 21-6 
    21.2.4  General and Administrative Operating Costs 21-8 
    21.2.5  Operating Cost Summary 21-8 
  21.3  Comments on Section 21 21-10 
22.0  ECONOMIC ANALYSIS 22-1 
  22.1  Valuation Methodology 22-1 
  22.2  Financial Model Parameters 22-2 
    22.2.1  Production Forecast 22-2 
    22.2.2  Metallurgical Recoveries 22-2 
    22.2.3  Smelting and Refining Terms 22-2 
    22.2.4  Metal Prices 22-3 
    22.2.5  Capital Costs 22-3 
    22.2.6  Operating Costs 22-3 
    22.2.7  Royalties 22-3 
    22.2.8  Working Capital 22-3 
    22.2.9  Taxes 22-4 
    22.2.10  Closure Costs and Salvage Value 22-4 
    22.2.11   Financing 22-4 
    22.2.12   Inflation 22-5 
  22.3  Financial Results 22-5 
  22.4  Sensitivity Analysis 22-5 
  22.5  Comment on Section 22 22-10 
23.0  ADJACENT PROPERTIES 23-1 
24.0  OTHER RELEVANT DATA AND INFORMATION 24-1 
  24.1  Preliminary Development Schedule 24-1 
  24.2  Project Opportunities 24-3 
25.0  INTERPRETATION AND CONCLUSIONS 25-1 
  25.1  Agreements, Mineral Tenure, Surface Rights, and Royalties 25-1 
  25.2  Geology and Mineralization 25-2 
  25.3  Exploration, Drilling, and Data Analysis 25-2 
  25.4  Metallurgical Testwork 25-3 
  25.5  Mineral Resource and Mineral Reserve Estimation 25-4 
  25.6  Mine Plan 25-5 
  25.7  Process Design 25-6 
  25.8  Infrastructure Considerations 25-8 
  25.9  Markets and Contracts 25-10 

     
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DONLIN GOLD PROJECT
ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

  25.10  Environmental, Social Issues and Permitting 25-10 
  25.11  Capital and Operating Cost Estimates 25-12 
  25.12  Financial Analysis 25-13 
  25.13  Preliminary Development Schedule 25-13 
  25.14  Conclusions 25-13 
26.0  RECOMMENDATIONS 26-1 
  26.1  Geology and Modeling 26-1 
  26.2  Data 26-1 
  26.3  Geotechnical 26-2 
  26.4  Pit Slope Dewatering 26-2 
  26.5  Mine Plan/ROM Stockpiling 26-3 
  26.6  Process, Metallurgy, and Water Treatment 26-3 
  26.7  River Surveys 26-4 
  26.8  Third-Party Logistics Service Providers 26-4 
  26.9  Permitting, Environment, and Social 26-4 
27.0  REFERENCES 27-1 
  27.1  Bibliography 27-1 

T A B L E S

Table 1-1: Donlin Gold Project Financial Summary 1-3
Table 1-2: Mineral Resources Summary Table, (Inclusive of Mineral Reserves) Effective Date 11 July 2011, Gordon Seibel, SME Registered Member 1-16
Table 1-3: Proven and Probable Mineral Reserves, Effective Date 11 July 2011, K.Hanson, P.E 1-17
Table 1-4: Summary of Key Financial Evaluation Metrics (Base Case is highlighted) 1-24
Table 2-1: Consulting Firms or Entities Contributing to FSU2 2-3
Table 2-2: QPs, Areas of Report Responsibility, and Site Visits 2-5
Table 7-1: Donlin Gold Project Stratigraphy 7-6
Table 7-2: Donlin Gold Project Intrusive Rocks 7-7
Table 7-3: Vein Stages 7-11
Table 9-1: Work History Summary for Donlin Gold Project 9-2
Table 9-2: Far Side 9-7
Table 9-3: Duqum 9-7
Table 9-4: Snow/Quartz 9-8
Table 9-5: Dome 9-8
Table 10-1: RC and Core Drill Summary Table 10-2
Table 10-2: Drill Hole Intercept Summary Table 10-12
Table 11-1: Specific Gravity Values by Rock Type 11-3
Table 11-2: Specific Gravity Values by Grouped Rock Type 11-3
Table 13-1: Intrusive and Sedimentary Lithologies of the Donlin Gold Project 13-2
Table 13-2: Major Geological Domains of the Donlin Gold Project 13-2
Table 13-3: Vein Types 13-3
Table 13-4: Typical Sulphide and Metals Mineralization in the Donlin Ores 13-3
Table 13-5: Grinding Testwork Results from Hazen Research 13-7

     
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ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

Table 13-6:  Summary of Grindability Testing 13-7
Table 13-7:  Comparison of Average Results from 2006 and 2007 Test Programs 13-10
Table 13-8:  Adjustments Made to the 2006 Test Program Data 13-10
Table 13-9:  Orebody Estimation of Crushing Index (Ci) 13-14
Table 13-10: Orebody Estimation of SAG Power Index (SPI) 13-14
Table 13-11: Orebody Estimation of Bond Ball Work Index (BWI) 13-14
Table 13-12:  Productivity Improvement Assumptions for the FSU 13-19
Table 13-13:   CIL Results from Pilot Flotation Tails 13-45
Table 13-14: Summary of Average Flotation Recovery in Variability Testwork Program, by Geological Domain 13-53
Table 13-15: Summary of Flotation Recovery in Variability Testwork Program by Geological Domain and Adjusted to MCF2 Pilot Result 13-55
Table 14-1:  Summary of Capping Grades for Major Rock Types 14-5
Table 14-2:  Summary of Capping Values for Neutralization Potential, with COV and GT Lost 14-6
Table 14-3:  Donlin Gold Project Mineral Resource Classification Methodology 14-12
Table 14-4:  Assumptions used for Calculation of NSR Values for Mineral Resources 14-13
Table 14-5:  Mill Recoveries used in Calculation of NSR for Mineral Resources 14-13
Table 14-6: Mineral Resources Summary Table, (Inclusive of Mineral Reserves) Effective Date 11 July 2011, Gordon Seibel, SME Registered Member 14-18
Table 15-1:  Assumptions used for Calculation of NSR Values for Mineral Reserves 15-2
Table 15-2:  Net Model Adjustments (within pit design) 15-5
Table 15-3:  Pit Optimization Process Recoveries 15-7
Table 15-4:  Pit Optimization Slopes 15-8
Table 15-5:  Proven and Probable Mineral Reserves, Effective Date 11 July 2011, K.Hanson, P.E.  15-9
Table 16-1:  Summary Projected Mine Production Plan by Year 16-10
Table 16-2:  Annual Required Drill Fleet 16-20
Table 16-3:  Annual Shovel Fleet Required 16-20
Table 16-4:  Mine Support Equipment 16-21
Table 16-5:  Mine Auxiliary Equipment 16-21
Table 16-6:  Mine Equipment Requirements 16-22
Table 17-1:  Projected Process Schedule and Recoveries 17-19
Table 20-1:  Environmental Baseline Studies (1995 to 2010) 20-2
Table 20-2:  Key Environmental Issues 20-3
Table 20-3: Estimated Reclamation Costs 2012–2264 20-8
Table 20-4:  Key Pre-Application Regulatory Meetings 20-11
Table 20-5:  Potential Federal Agency Permits and Authorizations 20-13
Table 20-6:  Potential State Agency Permits and Authorizations 20-14
Table 21-1:  Capital Cost Contingency 21-4
Table 21-2:  Summary of Capital Costs by Discipline 21-5
Table 21-3:  Summary of Capital Costs by Major Area 21-5
Table 21-4:  LOM Direct Process Operating Costs ($000) 21-7
Table 21-5:  Summary of G&A Cost Estimate by Cost Centre ($000) 21-9
Table 21-6:  LOM Operating Costs ($000) 21-9
Table 21-7:  Annual Operating Costs ($000) 21-10
Table 22-1:  Summary of Key Evaluation Metrics (Base Case is highlighted) 22-6
Table 22-2:  Cashflow Analysis 22-7
Table 22-3:  Base Case Project Sensitivity to Gold Price (Base Case is highlighted) 22-9

     
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ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

Table 22-4: Project Sensitivity to Operating Cost (Base Case is highlighted) 22-9 
Table 22-5: Project Sensitivity to Capital Cost (Base Case is highlighted) 22-10 
Table 22-6: Project Sensitivity to Oil Price (Base Case is highlighted) 22-10 
Table 22-7: Project Sensitivity to LNG Price (Base Case is highlighted) 22-10 

F I G U R E S

Figure 2-1:  Regional Project Setting 2-2
Figure 2-2:  Local Project Setting 2-2
Figure 4-1:  Proposed Pit Location in Relation to Calista Lease Boundary 4-2
Figure 4-2:  Key Deposit and Prospect Areas 4-3
Figure 4-3:  Donlin Gold Project Land Status Map 4-6
Figure 7-1:  Regional Geology of Central Kuskokwim Area 7-2
Figure 7-2:  Interpreted Property-Scale Igneous Rocks 7-4
Figure 7-3:  Interpreted Surface Geology of Resource Area 7-6
Figure 7-4:  100 m Bench Level Geology 7-8
Figure 7-5:  Lewis Area Section 7-8
Figure 7-6:  ACMA Area Section 7-9
Figure 7-7:  100 m Bench Level Gold Distribution (>1 g/t Au grade blocks) 7-11
Figure 9-1:  Regional Magnetic Image Showing Magnetic Low Intensity Zone 9-6
Figure 9-2:  Gold-in-Soils Compilation Plan 9-6
Figure 10-1:  Project Drill Hole Location Plan 10-4
Figure 10-2:  Resource Area Drill Holes 10-5
Figure 10-3:  Proposed Facility Sites (FSU1 Layout) and Drill Hole Locations 10-9
Figure 10-4:  Example Drill Cross-Section ACMA 10-15
Figure 10-5:  Vertical Cross Section Through ACMA and Lewis Block Model, Looking 315° 10-16
Figure 10-6:  Example Drill Cross-Section, Lewis 10-17
Figure 10-7:  Vertical Cross Section Through Lewis Block Model, Looking 45° 10-18
Figure 13-1:  Donlin Gold Project Geological Domains 13-3
Figure 13-2:  Test P80 vs. Measured BWI Results on Blend Composite Sample 13-12
Figure 13-3: Illustration of MCF2 Generic Flowsheet 13-15
Figure 13-4: Plant Utilization Ramp-up Schedule for Donlin Feasibility Compared to Other Available Commissioned Sites 13-17
Figure 13-5: Plant Throughput Ramp-up Schedule for Donlin Feasibility Compared to Other Available Sites 13-17
Figure 13-6:  Comparison of SGS Lakefield Dec 2006 Key Bench and Pilot-Plant Results 13-22
Figure 13-7:  Sulphide Oxidation Pressure Oxidation Kinetics at 220°C 13-27
Figure 13-8:  Gold Recovery Profiles from Pressure Oxidation at 220°C 13-27
Figure 13-9:  2007 Phase 1 Neutralization Pilot pH Profiles 13-39
Figure 13-10: Lime Demand Test Results of 2007 Phase 1 Pilot Samples, Plotted against Initial pH 13-39
Figure 13-11:  2007 Phase 2 Neutralization Pilot pH Profiles 13-41
Figure 13-12: Lime Demand Test Results of 2007 Phase 2 Pilot Samples, Plotted against Initial pH 13-41
Figure 13-13: Plot of Neutralization Variability Testing Lime Demand Results at 6 Hours’ Residence Time 13-43

     
Project No.: 166549
December 2011
TOC xi



DONLIN GOLD PROJECT
ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

Figure 13-14:  MCF2 Pilot-Plant Campaign Survey Results 13-52 
Figure 13-15:  Flotation Recovery Trend throughout Mine Life 13-55 
Figure 14-1: Donlin Geology and Mineral Domains 14-3 
Figure 16-1:  ACMA Phases in Plan at 94 m Elevation 16-4 
Figure 16-2:  Lewis Phases in Plan at 178 m Elevation 16-5 
Figure 16-3:  End-of Mine Plan Layout of Open Pit 16-6 
Figure 16-4: Proposed Mining Rate per Pit Phase 16-11 
Figure 16-5:  Projected Ore and Waste Production Schedule 16-11 
Figure 16-6:  Proposed Locations, Ore Stockpiles 16-12 
Figure 16-7: Donlin Ore Stockpile Projected Capacity by Year 16-12 
Figure 16-8:  Waste Dump Locations 16-14 
Figure 16-9:  Mine Equipment Consumables Distribution 16-24 
Figure 17-1:  Donlin Gold Project Process Plant Block Flow Diagram 17-2 
Figure 17-2:  Planned Gold Production by Year 17-19 
Figure 18-1:  Basic Route of Mine Access Road 18-2 
Figure 18-2:  Plant Site Layout 18-6 
Figure 18-3:  Plan of Ultimate Waste Rock Facility 18-10 
Figure 18-4:  Location of Tailings Storage Facility 18-13 
Figure 18-5:  Construction Water Management Layout 18-18 
Figure 18-6:  Operations Water Management Layout 18-22 
Figure 18-7:  Closure Water Management Layout 18-25 
Figure 21-1:  LOM Mine Production Operating Costs 21-7 
Figure 22-1:  After Tax LOM Total Cash Flow Sensitivity Spider Graph 22-9 

A P P E N D I C E S
Appendix A: Process List

     
Project No.: 166549
December 2011
TOC xii



DONLIN GOLD PROJECT
ALASKA, USA
NI 43-101 TECHNICAL REPORT
ON SECOND UPDATED FEASIBILITY STUDY

1.0

SUMMARY

NovaGold Resources Inc. (NovaGold) requested AMEC Americas Limited (AMEC) to prepare a summary report (the Report) on the results of the second updated feasibility study (FSU2) for the Donlin Gold Project (the Project) in Alaska, USA.

The Project is a 50:50 partnership between NovaGold Resources Alaska, Inc, a wholly-owned subsidiary of NovaGold) and Barrick Gold U.S. Inc, (a wholly-owned subsidiary of Barrick). The partners use an operating company, Donlin Gold LLC (Donlin Gold) to manage the Project. For the purposes of this Report, Donlin Gold is used as a synonym for the partnership. Prior to July 2011, Donlin Gold was known as Donlin Creek LLC (DCLLC).

NovaGold is using the Report in support of a press release dated 5 December, 2011, entitled "NovaGold Passes Key Milestone On Path to Becoming Premier North American Gold Producer; Completes Positive Feasibility Study On Donlin Gold Project Natural Gas Pipeline's Economic Benefits Confirmed Capex Estimate Declines From Previous Guidance Project Ready to Advance to Permitting", and a press release dated 12 January 2012 entitled ""NovaGold Files Donlin Gold Feasibility Study Technical Report".

1.1

Principal Outcomes


 
  •  
  • Proven and Probable Mineral Reserves estimated for approximately 34 Moz contained gold:

           
     
  •  
  • Proven Mineral Reserves: 7.7 Mt at 2.32 g/t Au (0.6 Moz contained gold)

     
  •  
  • Probable Mineral Reserves: 497 Mt at 2.08 g/t Au (33.3 Moz contained gold)

           
     
  •  
  • 25 year operating mine life

           
     
  •  
  • 27 year process life at 53,500 t/d throughput (includes two years of stockpile processing at the end of the operating mine life)

           
     
  •  
  • Average annual gold production:

           
     
  •  
  • 1.1 Moz over the projected life of mine

     
  •  
  • 1.5 Moz over the first full 5 years

     
  •  
  • 1.4 Moz over the first full 10 years

           
     
  •  
  • Predicted total cash costs:

           
     
  •  
  • $585/oz 1 Au over the life of mine

     
  •  
  • $409/oz over the first full 5 years

     
  •  
  • $452/oz over the first full 10 years

           
     
  •  
  • Net present after-tax cash flow (net present value (NPV) 5%)

           
     
  •  
  • At $1,000/oz gold price negative 1,342 million
     
  •  
  • At $1,200/oz gold price (Base Case) $547 million

     
  •  
  • At $1,700/oz gold price (Alternative Case 1) $4,581 million

    ______________________
    1 All dollar figures quoted in this summary are in US dollars

         
    Project No.: 166549
    December 2011
    Page 1-1



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

     
  •  
  • At $2,000/oz gold price (Alternative Case 2) $6,722 million

     
  •  
  • At $2,500 oz gold price $10,243 million

         
     
  •  
  • Average annual cash flow for first full five years of production2

         
     
  •  
  • At $1,000/oz gold price $673 million

     
  •  
  • At $1,200/oz gold price $950 million

     
  •  
  • At $1,700/oz gold price $1,500 million

     
  •  
  • At $2,000/oz gold price $1,783 million

     
  •  
  • At $2,500 oz gold price $2,184 million

           
     
  •  
  • Increase in contained gold ounces of approximately 4.7 Moz in Proven and Probable Mineral Reserve over the previous Proven and Probable Mineral Reserve estimate of 31 December 2008.

    Table 1-1 summarizes the key physical, technical, and financial parameters and the results of the FSU2 report.

    1.2

    Location, Climate, and Access

       

    The Donlin deposits are situated approximately 280 miles (450 km) west of Anchorage and 155 miles (250 km) northeast of Bethel up the Kuskokwim River. The closest village is the community of Crooked Creek, approximately 12 miles (20 km) to the south, on the Kuskokwim River. There is no road or rail access to the site. All access to the Project site for personnel and supplies is by air.

       

    The nearest roads are in the Anchorage area. Access to Bethel and Aniak, the regional centres, is limited to river travel by boat or barge in the summer and air travel year-round. The Kuskokwim River is a regional transportation route and is serviced by commercial barge lines.

       

    The area has a relatively dry interior continental climate with typically about 20 inches (500 mm) of total annual precipitation.

       
    1.3

    Agreements

       

    On December 1, 2007, NovaGold entered into a limited liability company agreement with Barrick that provided for the conversion of the Donlin Gold Project into a new limited liability company, the Donlin Creek LLC, which is jointly owned by NovaGold and Barrick on a 50/50 basis. In July 2011, the Board of Donlin Creek LLC voted to change the name of the company to Donlin Gold LLC.

    __________________________
    2 Total revenues minus total operating costs and royalties before interest, taxes, depreciation and amortization.

         
    Project No.: 166549
    December 2011
    Page 1-2



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

      Table 1-1: Donlin Gold Project Financial Summary

    Item Unit LOM $/oz  $/t milled  $/t mined
    Total Mined Mt 3,260
    Ore Tonnes Treated Mt 505
    Gold Grade g/t 2.09
    Gold Contained Moz 33.849
    Gold Recovery % 89.8
    Gold Recovered Moz 30.401
    Gold Payable Moz 30.371
    Gold Price $/oz 1,200
    Gold Gross Revenue $M 36,481 1,200 72.27 11.19
    OP Mining $M 8,200 270 16.24 2.52
    Processing $M 7,808 257 15.47 2.39
    G&A + Land Payments $M 3,068 101 6.08 0.94
    Payable Metal Deduction - Gold $M 36 1 0.07 0.01
    Doré TC+RC+Freight+Insure $M 31 1 0.06 0.01
    Direct Operating Costs + Metal Charges $M 19,144 630 37.92 5.87
    IFRS Total Capitalized Opex $M (1,386) (46) (2.75) (0.43)
    Stockpile Inventory Adjustment - Opex $M
    Total Operating Costs $M 17,758 584 35.18 5.45
    Depreciation $M 9,846 324 19.50 3.02
    Total Costs Before Taxes $M 27,604 908 54.68 8.47
    Cash Taxes $M 2,741 90 5.43 0.84
    Total Costs Including Taxes $M 30,345 998 60.11 9.31
    EBITDA $M 18,581 611 36.81 5.70
    Excluded from Cash Costs:          
    Community & Social Development Costs $M 141
    Project Development / Start-up Expenses $M 2
    Funding of Closure “Trust Fund” $M 274
    Note: EBITDA = earnings before interest, taxes, depreciation, and amortization    

    The Donlin exploration and mining lease currently includes a total of 72 sections in the vicinity of the deposit and additional partial sections associated with the Project infrastructure leased from Calista Corporation, an Alaska Native Corporation that holds the subsurface (mineral) estate for Native-owned lands in the region. Following a renegotiation in March 2010, the lease runs through April 2031 with provisions to extend beyond that time. Title to all of these sections has been conveyed to Calista by the Federal Government. Calista owns the surface estate on 27 of these 72 sections.

    A separate Surface Use Agreement with The Kuskokwim Corporation (TKC), an Alaska Native Village Corporation that owns the majority of the private surface estate in the area, grants non-exclusive surface use rights to Donlin Gold on at least 34 sections overlying the mineral deposit, with provisions allowing for adjusting that area in conjunction with adjustments to the subsurface included in the Calista lease. The term of the Surface Use Agreement runs through 5 June 2015 with provisions to extend beyond that time so long as mining, processing, or marketing operations are continuing and the Calista lease remains in effect.

    The Lyman family owns a small (13 acre) private parcel in the vicinity of the deposit and holds a placer mining lease from Calista that covers approximately four sections.

         
    Project No.: 166549
    December 2011
    Page 1-3



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    1.4

    Mineral Tenure

       

    Donlin Gold has 49,261 acres (20,081 hectares) leased from Calista as mineral rights. In addition, Donlin Gold holds 242 Alaska State mining claims comprising 31,740 acres (12,845 hectares), bringing the total land package to 81,361 acres (32,926 hectares). Of these claims, three are on State-selected lands and a total of 158 are tentatively approved from conveyance from Federal to State-owned, pending survey. None of the claims held by Donlin Gold have been surveyed.

       
    1.5

    Surface Rights

       

    Donlin Gold, through native lease agreements, holds a significant portion of the surface rights that will be required to support mining operations in the proposed mining area. Negotiations with TKC will be required for surface rights for additional lands supporting mining and access infrastructure. The currently identified Mineral Resources and the bulk of the proposed primary infrastructure (mill and waste rock facilities) are located on the leased lands.

       

    Other lands required for offsite infrastructure, such as those required for the Jungjuk port site, road to the port site, gas pipeline, and tailings storage facility in Anaconda Creek, are categorized as Native, State of Alaska conveyed, or Bureau of Land Management (BLM or Federal) lands.

       

    Rights-of-way will be required from the State and BLM for the road and pipeline alignments where they cross state and federal lands, respectively. Discussions regarding the extension and expansion of the TKC Surface Use Agreement and the disposition of the Lyman family land parcel and lease are ongoing.

       
    1.6

    Royalties

       

    A net proceeds royalty is payable to Calista of equal to 8% of the net proceeds realized by Donlin Gold at the Project after deducting certain capital and operating expenses (including an overhead charge, actual interest expenses incurred on borrowed funds and a 10% per annum deemed interest rate on investments not made with borrowed funds). Part of this royalty is paid as advance, pre-set, minimum royalty payments.

       

    There are currently no Government royalty obligations.

       
    1.7

    Environment, Permitting and Socio-Economics

       

    There has been a focused effort for at least 15 years to collect comprehensive environmental baseline data and lay the groundwork with local and regulatory stakeholders for the successful permitting and development of a large-scale mining operation at Donlin. Baseline data collected has included studies covering wetland delineation, water quality, fish and aquatic habitats, air quality, wildlife habitats, cultural resources and heritage, subsistence, traditional knowledge, socio-economics, health, mercury data, overburden, ore and waste rock characterization studies, noise, visual aesthetics, and river and land use.


         
    Project No.: 166549
    December 2011
    Page 1-4



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    The National Environmental Policy Act (NEPA) process and formal permit applications will require the preparation of an environmental impact statement (EIS). Upon completion of the NEPA process, a Record of Decision (ROD) will be prepared that approves the preferred alternative for the Project, describes the conditions of the approval, and explains the basis for the decision. The State permitting process typically is not finalized until the NEPA process is completed.

       

    Key environmental issues from stakeholders and regulatory authorities are likely to include mercury and cyanide management and water usage and management.

       

    Donlin Gold and Barrick have maintained all of the necessary permits for exploration and camp facilities. Project development will require appropriate permits from both State and Federal regulatory authorities, and operational and construction permitting is likely to require at least 80 separate permits. Each Federal and State permit will have compliance stipulations requiring review and possibly negotiation by the applicant and appropriate agency. The comprehensive permitting process will determine the exact number of management plans required to address all aspects of the Project to ensure compliance with environmental design and permit criteria.

       

    A preliminary closure plan has been prepared, which includes both concurrent reclamation during mining activities, and post-mining rehabilitation and monitoring. A modified version of the Barrick Reclamation Cost Estimator (BRCE) was used to develop the reclamation and closure cost estimate of $131.3 million. This amount is included in a Trust Fund for Reclamation, Closure costs and Post-Closure Obligations model prepared to determine the funding required to generate sufficient cash flow to cover the following costs: spillway construction from Anaconda creek to Crevice Creek; capital to construct a water treatment plant; perpetual water treatment; long-term monitoring; and associated facility and access maintenance. The total amount to cover reclamation / closure costs and post-reclamation and closure maintenance is estimated at $273.7 million, paid annually at $8.6 million over 32 years, including the construction period and 27-year life-of-mine.

       
    1.8

    Geology and Mineralization

       

    The Donlin mineralization model is a high-level, reduced intrusion-related vein system. The Lewis–ACMA part of the district is a low sulphidation, reduced intrusion related, epizonal system with both vein and disseminated mineral zones.

       

    The Donlin gold deposits lie in the central Kuskokwim basin of southwestern Alaska, which contains a back-arc continental margin basin fill assemblage of the Upper Cretaceous Kuskokwim Group, and Late Cretaceous volcano-plutonic complexes. The Project area is underlain by a 5 mile (8.5 km) long x 1.5 mile (2.5 km) wide granite porphyry dike and sill swarm hosted by lithic sandstone, siltstone, and shale of the Kuskokwim Group.


         
    Project No.: 166549
    December 2011
    Page 1-5



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    The deposits are hosted primarily in igneous rocks and are associated with an extensive Upper Cretaceous gold–arsenic–antimony–mercury hydrothermal system. The northeast, elongated, roughly 5,000 ft (1.5 km) wide x 10,000 ft (3 km) long cluster of gold deposits has an aggregate vertical range that exceeds 3,100 ft (945 m). These areas consist of the ACMA and 400 Zone, Aurora and Akivik mineralized areas (grouped as ACMA) and the Lewis, South Lewis, Vortex, Rochelieu and Queen mineralized areas (grouped as Lewis).

       

    Gold occurs primarily in sulphide and quartz–carbonate–sulphide vein networks in igneous rocks and, to a much lesser extent, in sedimentary rocks. Broad disseminated sulphide zones formed in igneous rocks where vein zones are closely spaced. Sub-microscopic gold, contained primarily in arsenopyrite and secondarily in pyrite and marcasite, is associated with illite–kaolinite–carbonate–graphite-altered host rocks.

       

    In the opinion of the QPs, knowledge of the deposit settings, lithologies, and structural and alteration controls on mineralization is sufficient to support Mineral Resource and Mineral Reserve estimation. The mineralization style and setting of the Project deposit is also sufficiently well understood to support Mineral Resource and Mineral Reserve estimation.

       
    1.9

    Exploration

       

    Placer gold was first discovered at Snow Gulch, a tributary of Donlin Creek, in 1909. Early stage exploration in the modern era was performed by Resource Associates of Alaska (1974–1975), Western Gold Exploration and Mining Co. LP (WestGold) during 1988–1989 and Teck Exploration Ltd. (Teck) in 1993. Exploration included geological mapping, trenching, rock and soil sampling, an airborne magnetic and VLF survey, ground magnetic surveys, and initial Mineral Resource estimates.

       

    The majority of the work completed on the Project has been primarily undertaken, in chronological order, by Placer Dome (1995 to 2000, and again from 2002 to 2005), NovaGold (2001 to 2002), Barrick (2006) and from 2007 to date by Donlin Gold.

       

    Activities have included construction of infrastructure to support exploration activities, reconnaissance and geological mapping; aerial photography; rock chip and soil sampling; trenching; max-min (EM) geophysical surveys; airborne geophysical surveys; RC and core drilling for resource infill, geotechnical, engineering, condemnation, waste rock, calcium carbonate exploration and metallurgical purposes; environmental baseline studies; community consultations; detailed metallurgical test work; geotechnical and hydrogeological studies; sampling of prospective calcium carbonate source areas; exploration and auger drilling program for sand and gravel sources; a series of Mineral Resource and Mineral Reserve estimates; and initial mining and engineering studies. This work culminated in a feasibility study in 2007, and updates to this study in 2009 and 2011.


         
    Project No.: 166549
    December 2011
    Page 1-6



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    In the opinion of the QPs, the exploration programs completed to date are appropriate to the style of the deposits and prospects within the Project. The exploration and research work supports the genetic and affinity interpretations.

       
    1.10

    Exploration Potential

       

    The Project retains exploration potential. The Akivik and East ACMA areas have good potential for lateral extensions of mineralization to the northwest and southeast of the FSU2 pit footprint. In addition, known gold mineralization is likely to extend at depth at the base of the designed pit, and in some areas immediately adjacent the planned pit floor, has been intersected by current drilling. Several drilled prospects and other exploration targets along the 3.7 mile (6 km) igneous trend north of the resource area remain under- explored, for example the Snow and Dome prospects.

       
    1.11

    Drilling

       

    Approximately 1,834 exploration and development diamond core (90%) and reverse circulation (RC) (10%) drill holes, totalling 1,337,321 ft (407,720 m), were completed from 1988 through 2010. Approximately 50% of the core and 40% of the holes were drilled during 2006–2007. All but about 20% (district exploration, carbonate resource, facilities condemnation, hydrology, infrastructure engineering) of this drilling was utilized for the current resource model. Supporting the FSU2 model are a total of 1,396 core (89%) and RC (11%) holes totalling 1,114,324 ft (339,733 m), and 282 trenches totalling 70,344 ft (21,441 m).

       

    Core sizes used on the Project include: NQ3 (45.1 mm core diameter), NQ (47.6 mm), HQ3 (61.2 mm), HQ (63.5 mm), and PQ (85 mm). Since 2002, core drills have been used exclusively for all resource delineation, and RC drilling was relegated to condemnation and hydrology studies.

       

    Standard logging and sampling conventions were used to capture information from the drill core and, where applicable, RC chips. Data captured included lithology, mineralization, alteration (visual), structural and geotechnical, with provision for geologists to add comments on the core if required.

       

    A survey of nearly 200,000 core recovery records in the database revealed an overall length-weighted average core recovery of 95%. Average recovery increases from 80 to 95% from 0 to 40 m and then ranges from 95 to 100% below 40 m where overburden and surface weathering effects are generally absent.

       

    Collar survey methods to 2001 included Brunton compass and hip chain, a Motorola GPS system and conventional theodolite survey methods. From 2002, an Ashtech Promark2 GPS post-processed system consisting of a base unit and up to two roving units has been employed.


         
    Project No.: 166549
    December 2011
    Page 1-7



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    The Sperry Sun single-shot camera method was used through 2000 for directional surveys to determine down-hole deviation. Reflex EZ Shot instrumentation was introduced in 2001. Approximately 60% of the core holes drilled within the resource model area were oriented to collect structural information for geotechnical and geological studies. Core orientation methods included clay impression, EZ Mark, and Reflex ACT instrument.

         

    The quantity and quality of the lithological, geotechnical, and collar and down-hole survey data collected in the exploration and delineation drill programs are sufficient to support Mineral Resource and Mineral Reserve estimation in the opinion of the QPs.

         

    Core is digitally photographed and split in half with an electric rock saw equipped with water-cooled diamond saw blades. Drill holes are sampled from the top of bedrock to the end of the hole. The maximum sample length in zones consisting of intrusive rocks or that contain appreciable sulphide/arsenic minerals is 6.6 ft (2 m), whereas sample lengths in sedimentary rock zones that lack appreciable sulphide/arsenic minerals can be 9.8 ft (3 m). A minimum of three additional 6.6 ft (2 m) sample intervals are placed before and after each intrusive rock or mineralized zone.

         

    Specific gravity data were collected primarily in 2006 by Barrick staff, using the wax immersion, water displacement method. The weighted average of all SG data points was 2.69.

         

    In the opinion of the QPs, sampling methods are acceptable, meet industry-standard practice, and are acceptable for Mineral Resource and Mineral Reserve estimation.

    Quality assurance and quality control (QA/QC) programs have been in place since 1995, and consist of the insertion of blank, standard reference material (SRM) and duplicate samples.

         
    1.12

    Sample Analysis and Security

         

    The primary laboratory for all assaying has been ALS Chemex in Vancouver, BC. During the exploration programs, ALS Chemex held accreditations typical for the time, including, at various times, ISO9001:2000 and ISO 9002, and from 2005, ISO/IEC 17025 accreditations.

         

    Most core samples from 2005 through 2008 were crushed at the Donlin camp sample preparation facility and pulverized at the ALS Chemex Vancouver laboratory facility. Samples of 2006 core split in Anchorage were shipped to an ALS Chemex preparation laboratory for crushing and pulverizing. Crushing requirements have been to 70% minus 10 mesh (2 mm) at the Donlin facility, and subsequently to better than 85% passing minus 200 mesh or 75 µm at ALS Chemex.

         

    A 1 oz (30 g) subsample of the pulp was assayed by ALS Chemex using fire assay-atomic absorption spectroscopy (AAS). Before 2007, the primary gold assay method was Au- AA23, which had an analytical range of 0.005 to 10 g/t Au. The Au-AA25 gold assay method was initiated in 2007 and had an analytical range of 0.01 to 100 g/t Au. Samples that exceeded the analytical limit for a given method were re-assayed by fire-assay and gravimetric finish or “ore grade” fire-assay AAS.


         
    Project No.: 166549
    December 2011
    Page 1-8



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    ALS Chemex determined the sulphur content of each sample according to the Leco method. The Leco method was also used to analyze samples flagged for acid base accounting (ABA) for carbon content as well as to determine neutralization potential (NP) and acid potential (AP) according to the industry-standard ALS Chemex ABA procedure.

       

    Most trace and major element data for drill holes located within the resource model boundary were acquired prior to the 2005 program by various laboratories using industry-standard acid digestions followed by atomic absorption (AA) or inductively coupled plasma (ICP) instrumental determinations. Subsequent multi-element trace analyses were performed at ALS (Chemex) using aqua regia or four-acid digestions followed by ICP ± mass spectrometry.

       

    Sample security measures practiced included moving of core form the drill site to the core shack at the end of each drill shift, and tracking of sample shipments using industry- standard procedures. Donlin Gold is of the opinion that core storage is secure because Donlin is a remote camp and access is strictly controlled.

       

    In the opinion of the QPs, the quality of the gold and sulphur analytical data are sufficiently reliable to support Mineral Resource and Mineral Reserve estimation without limitations on Mineral Resource confidence categories and sample preparation, analysis, and security are generally performed in accordance with exploration best practices and industry standards.

       
    1.13

    Data Verification

       

    A number of data verification programs and audits have been performed over the Project history, primarily in support of compilation of technical reports on the Project and in support of mining studies. Checks were performed in 2002 (AMEC), 2005 and 2008 (NovaGold), and AMEC (2011).

       

    In the opinion of the QPs, the data verification programs undertaken on the data collected from the Project adequately support the geological interpretations, the analytical and database quality, and therefore support the use of the data in Mineral Resource and Mineral Reserve estimation.

       
    1.14

    Metallurgical Testwork

       

    Testwork completed by SGS-Lakefield Research, Hazen Research, and G&T Metallurgical Services (G&T) under Barrick’s supervision has shown that the Donlin ore requires pre-treatment prior to cyanidation to recover the gold. Process development work has determined that pressure oxidation is the preferred method of pre-treatment. Extensive testwork on composites has shown that acceptable gold recoveries can be produced through a combination of flotation pre-concentration, POX, and CIL cyanidation.

       

    Air flotation using the MCF2 flowsheet provides an estimated life-of-mine (LOM) average of 93.0% recovery, with CIL recoveries after POX at approximately 96.6% for an estimated combined plant total gold recovery of 89.8% . The concentrate pull will vary from 15% to 17% and that will result in a concentrate grade of 13.0 to 12.7 g/t Au.


         
    Project No.: 166549
    December 2011
    Page 1-9



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
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    Process selection is supported by extensive testwork. Placer Dome undertook an initial phase of testwork from 1995 to 1999 to define the basic process. Early on, it became apparent that direct cyanidation or CIL of ore or flotation concentrate returned very low recoveries. Pre-treatment by oxidation was considered necessary.

    Placer Dome testwork included grinding, gravity concentration, flotation, POX, cyanidation, and neutralization. Subsequently from 2002 to 2005, Placer Dome also explored HPGR comminution, arsenopyrite/pyrite separation, nitrogen aerated flotation, and oxidation both by bio-oxidation and pressure autoclave.

    At the end of 2005, another round of work began with some testing at G&T, but this was interrupted by the acquisition of Placer Dome by Barrick Gold, which subsequently assumed management of remaining testwork.

    Major programs at the bench-scale level were initiated in 2006 to test grinding, flotation, POX, and neutralization. In addition to bench-scale work, major pilot-plant runs were performed in flotation, POX, and neutralization at the Barrick Technology Centre, SGS-Lakefield, G&T, and Hazen Research (Golden, U.S.A.). Both bench-level and pilot-plant scale testwork were conducted to develop process parameters and expand engineering information.

    The key testing results and recommendations considered in the FSU2 are summarized as follows:

    Mineralogy

    • Sulphur occurs primarily as pyrite and arsenopyrite. Marcasite is an additional minor sulphide present in the ore. Pyrite contains only a minor portion of the gold, while arsenopyrite is the main gold carrier, with gold in solid solution (sub-microscopic) form. In particular, it is the finest arsenopyrite that has the highest grade of gold. The proportion of pyrite to arsenopyrite ranges from 4 to 2:1, with 3:1 being typical.

    • Mercury in the ore at ~2 ppm average is primarily hosted by pyrite in solid solution (sub-microscopic form). No mercury minerals have been observed.

    • Arsenic in the ore at ~2,800 ppm average is primarily hosted by arsenopyrite. However, arsenic also occurs as native arsenic and realgar.

    • Antimony in the ore at ~80 to 90 ppm average is primarily hosted by stibnite, but also occurs at trace levels hosted by tetrahedrite.

    • Chloride in the ore at ~20 to 25 ppm average is primarily hosted by muscovite, but is also carried to a lesser degree by apatite.

    • Carbonate in the ore at 2.4% to 2.5% (analysis specified as CO2). The most common carbonate within the Donlin ores is ferroan dolomite (impure dolomite containing varying quantities of iron) followed by ankerite. Calcite and siderite are present but not common.

         
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    Direct Leach / CIL

    • The whole ore is refractory to direct and CIL cyanidation processing, with very low recoveries (<15%) from either leaching methodology. High gold recovery is achieved by destruction of the sulphidic host matrix of the gold.

    Crushing / Grinding

    • The ores are considered moderately hard, with an average Ball Work Index (BWI) of 15 kWh/t and an average Minnovex SAG Power Index (SPI) of 87.5 minutes.

    • The ores are amenable to SAG milling with reasonable operating efficiencies.

    • Ore hardness is controlled significantly by rock lithology.

    Flotation

    • Flotation gold recoveries are highest from intrusive ores (94.7% to 97.5%), lower from the sedimentary ores (89.7% to 91.3%), and problematic for partially geologically oxidized ores (average 75.7%).

    • All the testwork showed a very close relationship between arsenic and gold recovery, indicating the presence of gold in close combination with that element.

    • An MCF2 style milling, chemical addition, and flotation duplicated (mill/chemical/float, mill/chemical/float) flowsheet provides a recovery increase of 1.8% to a 7% sulphur flotation concentrate and is economically favoured for Donlin.

    • Using the MCF2 flowsheet, it is possible to concentrate the gold-bearing sulphides into a 7% sulphur concentrate recovering an overall average 93.0% of the gold (including 10% oxide ore in blend) into 15% of the plant feed mass. Required flotation residence time and reagent dosages are relatively high compared to other iron sulphide flotation processes.

    • The partially oxidized (altered) ores, which are predominantly near surface, perform poorly through flotation, with an average flotation recovery of 75.7%. Initial testing using sulphidizing reagents to promote flotation recovery improvement have been unsuccessful on these ores.

    • CIL leaching of the flotation tailings does not yield economically justifiable recovery of gold.

    Pressure Oxidation (POX)

    • POX allows for 96.6% recovery of the gold in CIL following oxidation.

         
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    • A concentrate pre-acidification circuit (to dissolve carbonates) and subsequent CCD wash of the pre-acidified concentrate (using uncontaminated waters) was indicated and has been incorporated as part of the process flowsheet.

    • As a precautionary measure, a mercury recovery system will be incorporated on the autoclave gas products prior to emission to the atmosphere.

    Neutralization

    • The presence of carbonates in the flotation tailings allows for autoclave acid solution neutralization, thus decreasing the overall lime requirements. The carbonate content in the ore is an estimated average 224% of the stoichiometric content of the total sulphur in the ore.

    CIL / Gold Recovery

    • Carbon in Leach (CIL) processing of washed autoclave product provides for optimized gold recoveries of 96% to 97%, requiring relatively low amounts of cyanide.

    • Lime consumption of the CIL feed can be minimized by operating the CIL circuit at a pH of ~9.0 to minimize lime consumption by precipitation of magnesium hydroxide. The Donlin ore contains a naturally high content of magnesium (6,500 to 6,600 ppm average), which is liberated from the ore through reaction with acid, both in the autoclave and neutralization circuits. Since this testwork program was completed, an alternative to operating at low pH in CIL has been identified whereby soluble magnesium is removed prior to CIL to enable operation at conventional pH.

    • Reagent addition is minimized by high-efficiency washing of the autoclave product to 98% or greater washing efficiency.

    • Pilot testing of the CIL circuit on autoclaved blended concentrate demonstrated that CIL recovery is not sensitive to carbon gold loadings. However, the carbon elution circuit has been designed to allow for low carbon loadings in the event that preg-robbing concentrates are encountered.

    • Mercury leached into solution from the autoclave product by cyanide, which is not adsorbed onto carbon, is controlled to low levels within the recirculating process water streams by precipitation as a sulphide using a Cherokee UNR reagent.

    • Current design incorporates mercury gas/vapour recovery systems on the carbon regeneration kiln, electrowinning, retort system, and smelting furnace off-gases.

    Environmental Considerations

    • The high temperature and pressure oxidation process is considered best practice for generation of stable arsenic compounds suitable for long-term disposal in a tailings storage facility. Sufficient iron content is present in the Donlin ores to provide the recommended minimum stoichiometric ratio of 4:1 iron to arsenic.

         
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    • A portion of the arsenic in the Donlin ores is water soluble and liberates into solution within the operation of the grinding and flotation circuits alone.

    • The tailings decant water from the Donlin process plant will likely contain elevated levels (above current aquatic life or drinking water standard) of As, Hg, Mn, Mo, Se, and Sb. The tailings water could also be elevated in sulphates (greater than 10 g/L), particularly due to the presence of magnesium, which increases solubility level of sulphate in solution.

    1.15

    Mineral Resource Estimate

    The cut-off date for information used in the geologic model and resource model (termed the DC-9 model by Donlin Gold) was 1 November 2009.

    The mineral estimate was prepared by Mr. Chris Valorose of Barrick and audited by AMEC. Three-dimensional solids for the geological model were constructed from polygons resulting from geologic interpretation of cross-section and level plans. Nine mineral and geological domains were assigned to the database. Geotechnical domain zone codes were input into the resource model, as required for the Lerchs–Grossmann (LG) pit optimization, using domain solids provided by BGC Engineering Inc (BGC) on 27 June 2008. A waste rock management category (WRMC) model was coded to identify overburden from the other WRMC codes.

    Two specific gravity values were used: 2.65 for intrusive rocks, and 2.71 for sedimentary units.

    Raw assay data were grouped by rock type, and capping values for gold were determined for each major rock type. Gold assays were capped above above 30 g/t Au. Values for neutralization potential (NP) were also capped. Total sulphur, arsenic, mercury, and antimony assays were not capped.

    Composites were created down each hole at 20 ft (6 m) intervals. The composites were not broken at intrusive or sedimentary rock contact boundaries. Indicator semi-variograms generated at 0.25 g/t Au for the 6 m composites were fitted with a spherical model. Ranges of 98.4 ft (30 m) and 147.6 ft (45 m) were observed at 80% and 90% of the total sill variance.

    A gold indicator model was used to estimate gold, arsenic, antimony, and mercury grades based on gold composite data. A separate sulphur indicator model was used to estimate sulphur.

    Gold grades were estimated into the block model using an inverse distance to the third power methodology for two populations:

    • Internal to the mineralized envelope, defined as blocks with indicator values greater than or equal to 50%

         
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    • External to the mineralized envelope, defined as blocks with indicator values less than 50%.

    Interpolation of grade into the blocks was broken into five passes based upon increasing search distances, out to a maximum of 125 m. Gold grades were estimated separately for intrusive rocks, shales, and greywackes, and further sub-divided based upon whether blocks were internal or external to the mineralized envelope.

    Sulphur grades were estimated using the same methods and parameters as for the gold grade estimation. Arsenic, mercury, and antimony grades were estimated using methods and parameters similar to those for the gold grade estimation. Values for CO2, calcium, and magnesium were estimated into the block model based on an ordinary kriging method within nine estimation domains. Neutralization potential (NP) was estimated into the block model for use in the classification of waste rock.

    The block model grades were validated visually against drill holes and composites in section and plan view. A nearest-neighbour block model was also generated. Grade profile plots were generated for the 6 m x 6 m x 6 m Measured and Indicated resource model as a further validation check. No estimation biases were noted from the validation reviews.

    Dilution and selectivity of mineralized material were determined using a Barrick in-house program referred to as SMUman. The extent of the classified material that might have reasonable expectation for economic extraction was assessed by applying a LG pit outline using Whittle® software to the Mineral Resources. A net sales return (NSR) value per tonne was then coded into each block of the resource model. For those blocks with a resource classification of Measured or Indicated, the NSR per tonne value was calculated with the following equations:

    General:

    NSR = [Au grade] * [Recovery] * [Price of Gold less Refining and Royalty Costs] - [Processing Costs+ General and Administrative Costs + Rehandling Costs] US$/tonne

    For Mineral Resources, the figures were:

    NSR = [Au grade] * [Recovery] * [US$1200 – (1.85 + (( US$1200 – 1.85) * 0.045))] - [(10.65 + (2.1874 * S%)) + 2.29 + 0.20] US$/tonne

    For Mineral Reserves the figures were:

    NSR = [Au grade] * [Recovery] * [US$975 – (1.78 + ((US$975 – 1.78) * 0.045))] - [(10.65 + (2.1874 * S%)) + 2.27 + 0.19] US$/tonne

    The NSR cut-off for Mineral Resource reporting purposes was $0.001/t milled, which represents the net sales return marginal cut-off strategy.

         
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    Mineral Resources take into account geologic, mining, processing and economic constraints, and have been confined within appropriate LG pit shells, and therefore are classified in accordance with the 2010 CIM Definition Standards for Mineral Resources and Mineral Reserves. The Qualified Person for the Mineral Resource estimate is Gordon Seibel, SME Registered Member, an employee of AMEC. Mineral Resources are reported in Table 1-2 at a commodity price of $1,200/oz gold, have an effective date of 11 July 2011, and are inclusive of Mineral Reserves.

    Factors which may affect the Mineral Resource estimate include the commodity price; changes to the assumptions used to generate the NSR cut-off; changes to the 0.25 g/t Au threshold used to define the indicator mineralized domains, changes in interpretations of fault geometry; changes to the search orientations used for grade estimation in the ACMA area, results of a review of the Measured classification criteria; and changes to the assumptions used to generate the LG pit constraining the estimate, in particular slope design assumptions.

    1.16

    Mineral Reserve Estimate

    Mineral Reserves were optimized for all Measured and Indicated blocks assuming a gold selling price of $975/oz.

    The ore considered for processing in the optimization was based on a marginal cut-off grade that varied from block to block. Material was considered to be ore if the revenue of the block exceeded the processing and G&A cost. The revenue was based on net gold price after refining charges and royalties had been deducted. The processing cost was a function of the sulphur content of the material being processed.

    Dilution was considered for bulk mineable (12 m bench height) and selective mineable (6 m bench height) scenarios. All blocks classified as Inferred were set to waste in the selective mining plan. In the bulk mineable plan, the entire 12 m block was assigned the highest confidence category of the sub-blocks in the plan. The grade of all Inferred blocks was set to zero at the start of the process. Therefore the combined grade of the 12 m block is derived from the Measured or Indicated metal grades only.

    Pit shell generation was constrained in the northwestern part of the ACMA mining area, to prevent it from encroaching on Crooked Creek, which is a salmon-bearing stream, but was not constrained by any infrastructure considerations.

    Geotechnical domains, design sectors, slope angles, and associated assumptions were provided by BGC. Mine design has incorporated geotechnical and hydrogeological considerations.

         
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    NI 43-101 TECHNICAL REPORT
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      Table 1-2: Mineral Resources Summary Table, (Inclusive of Mineral Reserves)
        Effective Date 11 July 2011,
        Gordon Seibel, SME Registered Member

    Category Tonnage Au Contained Au S
      (kt) (g/t) (koz) (%)
    Measured 7,731 2.52 626 1.15
    Indicated 533,607 2.24 38,380 1.08
    Total Measured and Indicated 541,337 2.24 39,007 1.08
    Inferred 92,216 2.02 5,993 1.08

    Notes to Accompany Mineral Resources Table

      1.

    Mineral Resources are inclusive of Mineral Reserves. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability

      2.

    Mineral Resources are contained within a conceptual Measured, Indicated and Inferred optimized pit shell using the following assumptions: gold price of US$1,200/oz; variable process cost based on 2.1874 * (sulphur grade) + 10.6485; administration cost of US$2.29/t; refining, freight & marketing (selling costs) of US$1.85/oz recovered; stockpile rehandle costs of 0.20/t processed assuming that 45% of mill feed is rehandled; variable royalty rate, based on royalty of 4.5% * (Au price – selling cost)

      3.

    Mineral resources have been estimated using a constant net sales return (NSR) cut-off of US$0.001/t milled. The NSR was calculated using the formula: NSR = Au grade * Recovery * (1,200 - (1.85 + (1,200 - 1.85) * 0.045)) (10.65 + 2.1874 * (S%) + 2.29 + 0.2) and reported in US$/tonne., Assuming an average recovery of 89.54% and an average S% grade of 1.07%, the marginal gold cutoff grade would be approximately 0.46 g/t, or the gold grade that would equate to a $0.001 NSR cutoff at these same values

      4.

    Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grade and contained metal content

      5.

    Tonnage and grade measurements are in metric units. Contained gold ounces are reported as troy ounces.

    The base mining cost (before incremental mining cost with depth) is $1.51/st ($1.668/t), the average processing cost is $13.06/st ($14.39/t), and the G&A cost is $2.06/st ($2.27/t) . These costs are considered reasonable.

    Recoveries for non-oxide ores are quoted as a constant for each rock type, whereas recoveries for oxide ores vary with sulphur grade. Recoveries range from 88.6% in shale to 94.2% in the Akivik zone.

    Mineral Reserves have been modified from Mineral Resources by taking into account geologic, mining, processing, and economic parameters and therefore are classified in accordance with the 2010 CIM Definition Standards for Mineral Resources and Mineral Reserves. The Qualified Person for the Mineral Reserve estimate is Kirk Hanson, P.E., an AMEC employee. Mineral Reserves are reported at a gold price of $975/oz gold, and have an effective date of 11 July 2011.

    Mineral Reserves are summarized in Table 1-3.

    Factors which may affect assumptions used in estimating Mineral Reserves include the commodity price; unrecognized structural complications in areas with relatively low drill hole density that could introduce unfavourable pit slope stability conditions; changes in interpretation of the fault orientiations, in particular the Vortex and Lo Faults; changes in orientations of the bedding or ash layer orientations which may necessitate flatter slope angles than currently assumed; in-pit and pit wall water management if water inflows are higher than predicted; and the likelihood of obtaining required permits and social licenses to construct the gas pipeline and operate the planned mine.

         
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      Table 1-3: Proven and Probable Mineral Reserves, Effective Date 11 July 2011,
        K.Hanson, P.E.

      Tonnage Au Contained Au S
    Category (kt) (g/t) (koz) (%)
    Proven 7,683 2.32 573 1.12
    Probable 497,128 2.08 33,276 1.06
    Total Proven and Probable 504,811 2.09 33,849 1.06

      Notes to Accompany Mineral Reserves Table
    1.

    Mineral Reserves are contained within Measured and Indicated pit designs, and supported by a mine plan, featuring variable throughput rates, stockpiling and cut-off optimization. The pit designs and mine plan were optimized on diluted grades using the following economic and technical parameters: Metal price for gold of US$975/oz; reference mining cost of $1.67/t incremented $0.0031/t/m with depth from the 220 m elevation (equates to an average mining cost of $2.14/t), variable processing cost based on the formula 2.1874 x (S%) + 10.65 for each $/t processed; general and administrative cost of US$2.27/t processed; stockpile rehandle costs of 0.19/t processed assuming that 45% of mill feed is rehandled; variable recoveries by rocktype, ranging from 86.66% in shale to 94.17% in intrusive rocks in the Akivik domain; refining and freight charges of US$1.78/oz gold; royalty considerations of 4.5%; and variable pit slope angles, ranging from 23º to 43º.

    2.

    Mineral Reserves are reported using an optimized net sales return (NSR) value based on the following equation: NSR = Au grade * Recovery * (US$975 - (1.78 + ($US975 - 1.78) * 0.045)) (10.65 + 2.1874 * (S%) + 2.27 + 0.19) and reported in US$/tonne. Assuming an average recovery of 89.54% and an average S% grade of 1.07%, the marginal gold cutoff grade would be approximately 0.57 g/t, or the gold grade that would equate to a $0.001 NSR cutoff at these same values.

      3.

    The life of mine strip ratio is 5.48. The assumed life-of-mine throughput rate is 53.5 kt/d

    4.

    Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grade and contained metal content

      5.

    Tonnage and grade measurements are in metric units. Contained gold ounces are reported as troy ounces.


    1.17

    Proposed Mine Plan

    The preferred development is for a 55 kst/d (50 kt/d) process facility with on-site power; and a mine capacity of 485 kst/d (440 kt/d) with an elevated cut-off policy applied in the initial part of the mine life. The processing rate was upgraded to 58 kst/d (53.5 kt/d) during the FSU2 design phase to take into account processing design constraints and rationalization of the proposed pressure oxidization circuit to be installed.

    The ACMA ultimate pit has been divided into nine phases, the Lewis pit into six phases. The initial phases of the two pits are independent, but they partially merge later in the mine life, forming a final single pit. The mine design, complete with haulage access, includes 556,459 kst (504,811 kt) of ore containing 33,849 koz (1,052,815 kg) of in-situ gold and has a strip ratio of 5.48. The mine design is considered appropriate to the quantity of Measured and Indicated Mineral Resources estimated for the Project. AMEC notes that the engineered pit design includes approximately 5% less ore tonnage and 7% fewer Au ounces than the pit optimization shell it was based on. This is at the upper end of the generally accepted limit of a 10% reduction in tonnes. As such, there is a risk that the engineered pit design contains less ore than optimum.

    Mineable pit phases were designed based on optimized nested pit shell guidance, gold grade, strip ratio, access, and backfilling of the ACMA phases. Ramps in final walls have a design width of 131 ft (40 m) and a gradient of 10%. A nominal minimum mining width of 492 ft (150 m) was used for phase design.

         
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    NI 43-101 TECHNICAL REPORT
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    Dates in this paragraph are for illustrative purposes only, as no Project permits and approvals have been received, and Project development and construction has not been approved by the respective Boards of Donlin Gold, NovaGold and Barrick. Preproduction has been defined as starting in April 2018 and finishing at the end of December 2018, when the main orebody is exposed. Mill production starts in July 2019. The operating mine life is estimated to be 25 years based on a nominal processing rate of 59,000 stpd. The schedule incorporates long-term and short-term ore stockpiles. The long-term stockpile will hold all ore produced at the mine in excess of plant feed, separated into three sections according to sulphur grade for blending purposes.

    The maximum long-term stockpile volume is 104.8 Mt at the end of 2031. This includes 18.5 Mt of high sulphur-grade material, 31.9 Mt of medium sulphur-grade material, and 53.9 Mt of low sulphur-grade material.

    The short-term stockpile was established to cope with daily variations in plant capacity and to accommodate fluctuations in the average daily mill feed; this stockpile was assumed to have an average 45% annual re-handle.

    After plant ramp-up, mill feed averages 52.7 kt/d and reaches a maximum of 54.4 kt/d in 2030 (Year 12). Contained gold in the mill feed averages approximately 1.3 Moz per year, while gold production averages 1.6 Moz per year for the first five years, with a maximum of 1.731 Moz in 2024 (Year 6).

    In the opinion of Donlin Gold, the proposed plant feed supports that the amount of sulphur in the feed can be controlled through a blending strategy combining ore feed directly from the mine and from stockpiles.

    A total of 2,460 Mst (2,232 Mt) of waste will be stored in a single ex-pit waste rock facility, in the American Creek Valley, east of the pit area. Another 466 Mst (423 Mt) of waste rock will be stored in the ACMA backfill dump and 18.7 Mst (17 Mt) of overburden in the overburden stockpiles for reclamation use. The remaining 114 Mst (103 Mt) is used as construction material, of which 99 Mst (90 Mt) is for tailings dam wall construction. Backfilling will commence in 2035 (Year 18) and continue until the end of mine life. In addition, 103 Mt of waste rock will be used for construction purposes, and 16.6 Mt of overburden will be stored in overburden stockpiles for reclamation purposes.

    Surface ditches, a contact water pond (CWP) immediately upstream of the pit, plus diversion systems further upstream, will control surface waters in the pit and waste dump areas. Dewatering systems consisting of perimeter and in-pit vertical dewatering wells, horizontal drains, and in-pit sump pumps will be required to manage groundwater.

    1.18

    Process Design

    Run-of-mine (ROM) ore at 59,000 stpd (53,500 t/d) from the Donlin deposits will be crushed in a gyratory crusher followed by a semi-autogenous grinding (SAG) mill and two-stage ball milling, addition of chemicals, and a flotation circuit (MCF2). The primary ball milling circuit will produce a P80 particle size of 120 to 150 µm as feed to the primary rougher flotation section. The secondary ball milling circuit will produce a P80 particle size of 50 µm as feed to the secondary rougher flotation section.

         
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    Gold-bearing sulphides, recovered by flotation, generate a concentrate containing 7% sulphur. The concentrate is refractory and will be treated in a pressure oxidation circuit prior to cyanidation. Overall gold recovery from flotation, pressure oxidation and cyanidation is estimated to be in the order of 89.83% . Excess acid from the autoclave circuit will be neutralized with flotation tailings and slaked lime. Tailings from the process will be impounded in a zero-discharge tailings storage facility; water reclaimed from here will be re-used in the process plant.

    Mineralogical studies have shown that the gold is not visible. Testwork analysis indicates a high level of association of gold with arsenopyrite. Other sulphides such as pyrite and marcasite are also present, with reduced tenors of gold. Organic carbon, a potential preg robber, is present in the sedimentary ore. It is also present at lower levels in the intrusive ores, believed to be in the form of well-ordered graphite. This form of organic carbon is possibly less likely to preg-rob.

    The average Bond work index for the ore is in the range of 15 kWh/t. Flotation work has shown that kinetics are initially rapid, but to achieve high recoveries, a combined primary and secondary rougher residence time over 100 minutes, together with a high reagent loading in the system, is required. Clay-like minerals will affect slurry viscosity and settling. Slurry density in the underflow will be less than 50% solids for the concentrate thickeners.

    Partially geologically oxidized (altered) ore in the deposit, up to 7% of the mill feed, is the key non-performing ore type in the flotation circuit. Degradation of the sulphide ore via oxidation in the stockpile will also affect the flotation recovery, applied as 5% recovery loss within flotation on all ores stockpiled for longer than one year.

    Pressure oxidation (POX) has been shown to be successful in releasing the valuable constituents, under certain conditions. To optimize oxidation conditions, the water systems design has been modified to use the highest-quality water in the oxidation circuit. The autoclave design incorporates variable level control to provide better control over operating residence time.

    The oxidation circuit discharge will be washed to reduce lime load in carbon-in-leach (CIL); the washed solution will be neutralized by the use of high-carbonate flotation tails to further reduce plant lime consumption prior to tailings disposal.

    Gold recovery by CIL has proven successful in treating Donlin ores and is estimated to be 96.6% . Rheological investigation and CIL testing results have determined that a relatively low CIL feed density of 35% solids should be adopted. In addition, to control lime usage, the CIL circuit will be operated at a pH of approximately 11.0.

         
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    Given the plan to use stockpiles to manage the ore blend into the process from the perspective of gold, sulphur, carbonate, and hardness, allowances were made for ore aging or stockpile degradation for the life-of-mine feed. Ore oxidized through weathering will have a slower flotation response than fresh rock. In general, ore at Donlin does not contain highly reactive sulphide species, and testwork has shown no statistical deviation over a one-year period. While data from a longer timeframe are not presently available, the testwork results for oxidized material show some degradation. Consequently, there is no effect on recovery for material stockpiled for less than one year (sulphide “fresh” material), and a recovery deduction of 5% has been applied to gold and sulphur recoveries for sulphide material stockpiled for longer than one year.

    Alternative flowsheets to flotation-POX-CIL were considered, including whole ore pressure oxidation, roasting a flotation concentrate, and bio-oxidation (BIOX). None of these proved to be a viable economic alternative to the flotation-POX-CIL route.

    1.19 Planned Project Infrastructure

    The Project will require construction of significant infrastructure to support the planned producing facilities. Key infrastructure will include:

    • Access road, 27 miles (44 km) long, from the mine site to the planned Kuskokwim River dock site at Jungjuk

    • Airstrip to support DHC Dash 8 and the Hercules C-130 aircraft

    • Barge cargo terminal at Jungjuk

    • Marine cargo terminal at Bethel

    • Two open pit mines

    • Process plant site in the Anaconda valley

    • Primary crusher area on a ridge on the south side of American Creek

    • Fuel storage compound adjacent the process plant site

    • Mining and road fleet truckshops in association with the primary crusher area

    • Contact water management dams and a freshwater storage reservoir

    • Water management pumping systems

    • Power plant, located adjacent the process plant

    • Tailings storage facility (TSF) in the Anaconda Creek basin

    • Waste rock storage facility (WRF) in the American Creek valley

    • Gas pipeline

    • Construction (2,560 people) and permanent accommodation (maximum 638 people) camps.

         
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    In general, the design and construction of the mine site infrastructure will be relatively straightforward, although the scope of the work is extensive, especially in terms of the water systems. In addition, the Project involves several development sites considerable distances apart, incurring high infrastructure costs to provide interconnecting roads, pipelines, services, and utilities. The decision to use material from the plant site excavation as a borrow source for constructing the tailings dams is an effective way to reduce the site preparation costs.

       

    The construction schedule for the initial phase of the TSF and the Lower CWD is aggressive, with a great deal of work to be completed in a short duration. Weather delays could affect completion on schedule.

       
    1.20

    Markets

       

    NovaGold will be able to market gold produced from the Donlin Project. Sales contracts that could be negotiated would be expected to be within industry norms. However, the majority of production would be expected to be spot marketed.

       
    1.21

    Capital Costs

       

    The capital cost estimate was developed in accordance with Association for the Advancement of Cost Engineering (AACE) Class 3 requirements, consisting of semi- detailed unit costs and assembly line items. The level of accuracy for the estimate is - 15% +30% of estimated final costs, per AACE Class 3 definition. All costs are expressed in second quarter (Q2) 2011 US dollars. No allowances are included for escalation, interest during construction, taxes, or duties.

       

    The total estimated capital cost to design and build the Donlin Project described in this Report is $6,679 million, including an Owner-provided mining fleet and self-performed pre-development. Included in the estimate are:

    • Direct capital costs: $4,009 million (includes gas pipeline direct cost of $758.1 M)

    • Owner's costs: $414 million

    • Other indirect costs: 1,271 million

    • Contingency: $984 million.

    Sustaining capital costs total $1,504 million. Significant areas include $649 million to replace and supplement mobile mining and support equipment and $631 million for periodic tailings storage facility capacity expansions.

    AMEC notes the following in relation to the proposed natural gas pipeline. The direct costs of the pipeline are estimated at $758.1 M, with indirect costs of an additional $75.7 M ($38.7 M engineering procurement, $32.5 M construction costs and, $4.4 M Owners’ costs, primarily for land), totalling $829.4 M, excluding contingency. When contingency is included, the pipeline costs are estimated to total $973 M.

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    1.22

    Operating Costs

       

    Operating cost estimates have been assembled by area and component, based on estimated staffing levels, consumables, and expenditures, according to the mine plan and process design. Operating costs have been prepared in second quarter (Q2) 2011 U.S. dollars with no allowances for escalation, sales tax, import duties, or contingency.

       

    The estimated life-of-mine operating costs are $5.42/t mined or $34.99/t milled, or $581/oz.

       
    1.23

    Financial Analysis

       

    The results of the economic analysis represent forward-looking information that are subject to a number of known and unknown risks, uncertainties, and other factors that may cause actual results to differ materially from those presented here. Forward-looking information includes Mineral Resource and Mineral Reserve estimates, commodity prices and exchange rates, the proposed mine production plan, projected recovery rates, uncertainties and risks regarding the estimated capital and operating costs, uncertainties and risks regarding the cost estimates and completion schedule for the proposed Project infrastructure, in particular the proposed barging program, and the need to obtain permits and governmental approvals.

       

    The overall economic viability of the Project has been assessed using both undiscounted and discounted cash flow techniques. Undiscounted techniques include total net cash flow, payback period (measured from start of production), earnings before interest, taxes, depreciation, and amortization (EBITDA), and cash costs.

       

    Discounted values are calculated using a 5% discount rate and a discrete, end-of-year convention relative to reference dates of 1 January 2012 (FSU2) and 1 January 2014 (this Report, and Project Base Case). A period of approximately 3.5 years for permitting, starting 1 January 2012, is included prior to start of construction.

       

    The economic evaluation of the Donlin Project was prepared by Donlin Gold and is based upon:

    • Capital cost and sustaining capital cost estimates prepared by AMEC, BGC, and Hatch

    • Owner’s capital costs prepared by Donlin Gold

    • Reclamation and closure costs prepared by SRK

    • Post-closure obligations prepared by Donlin Gold

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY
    • Funding requirements for the reclamation, closure, and post-closure obligations endowment prepared by Donlin Gold

    • Mine schedule prepared by Barrick

    • Mineral Resource estimate prepared by Donlin Gold

    • Mine equipment costs based on quotes received from equipment suppliers

    • Estimated mine, process plant, and general and administration operating costs prepared by Donlin Gold, AMEC, Barrick, and Hatch, based on budget quotations, first principles, and/or costs at operating mines similar to that proposed for Donlin such as Barrick’s Goldstrike operation

    • An allowance for supply inventory and working capital (including doré transportation, in-process inventory, and payment delays); these values sum to zero over the life of the mine.

    • Financial analysis of the Base Case (discount rate of 5%) showed the after-tax Project NPV to be $547 M and the internal rate of return (IRR) to be 6.0% (Table 1-4). The cumulative, undiscounted, after-tax cash flow value for the Project is $6,197 M and the after-tax payback period is 9.2 years.

    • Sensitivity analyses were performed on the Project on a range of -20% to +20% on gold price, operating costs, and capital costs. For purposes of the sensitivity analysis, variations in the gold grade were assumed to mirror variations in the gold price.

    • The Project is particularly sensitive to changes in the gold price. The Project requires a gold price of approximately $902/oz to break even on a cash flow basis and a gold price of approximately $1,141/oz to achieve an IRR of 5%. It is less sensitive to variations in operating cost and capital cost.

    1.24

    Preliminary Development Schedule

       

    A preliminary Project development schedule has been generated. The schedule includes consideration of early work requirements, the environmental permitting process, EPCM and construction activities.

       
    1.25

    Conclusions

       

    AMEC considers that the scientific and technical information available on the Project can support proceeding with additional data collection, trade-off and engineering work and preparation of more detailed studies. However, the decision to proceed with a mining operation on the Project is at the discretion of Donlin Gold, NovaGold and Barrick.


         
    Project No.: 166549
    December 2011
    Page 1-23



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

      Table 1-4: Summary of Key Financial Evaluation Metrics (Base Case is highlighted)

    Item Unit Value
    Total Mined Mt 3,260
    Ore Tonnes Treated Mt 505
    Strip Ratio W/O 5.46
    Gold Recovered Moz 30.401
    Gold Recovery % 89.8
    Gold Price $/oz 1,200
    Total Operating Costs $/oz 584
    Total Costs Before Taxes $/oz 908
    Total Costs Including Taxes $/oz 998
    EBITDA $M 18,581
    Total Cash Flow* $M 6,197
    Jan 2012 NPV @ 5%** $M 337
    Jan 2012 IRR % 5.6
    Jan 2014 NPV @ 5%** $M 547
    Jan 2014 IRR % 6.0
    Payback Period Years 9.2
    Operation Life Years 27.0
    Initial Capital $M 6,679
    Total LOM Capital $M 8,184

       Note: EBITDA = Earnings before interest, taxes, depreciation, and amortization
        * Cash flow excludes sunk costs
    ** Reference dates for DCF metrics are 1 January 2012 and 1 January 2014. The DCF metrics for 1 January 2014 treat funds expended before that date as sunk.
    During 2012 and 2013, Donlin Gold intends to complete basic engineering and commence detailed engineering, in tandem with, and in the case of detailed engineering, subject to, progress achieved on the Environmental Impact Statement and associated permitting process. Aggregate expenditures in these years are expected to be approximately $172 million, which if excluded from the discounted cash flow analysis would result in an increased project NPV5 and IRR from 2014 onwards of $210 million and 0.4%, respectively.

    1.26

    Recommendations

       

    Donlin Gold has completed a feasibility study and two updates on the study. A decision to proceed with any mine development plans would be made by the joint venture partners.

    As a consequence, AMEC's recommendations are restricted to activities that would support permitting and detailed engineering studies. These activities are envisaged as a single phase of work, with no item or area dependent on results of another. The estimated total cost of the proposed work is in the range of $135,000 to $200,000.

         
    Project No.: 166549
    December 2011
    Page 1-24



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    2.0

    INTRODUCTION

       

    NovaGold Resources Inc. (NovaGold) requested AMEC Americas Limited (AMEC) to prepare a summary report (the Report) on the results of the second updated feasibility study (FSU2) for the Donlin Gold Project (the Project) in Alaska, USA (Figure 2-1 and Figure 2-2).

       

    The Project is a 50:50 partnership between NovaGold Resources Alaska, Inc, (a wholly-owned subsidiary of NovaGold) and Barrick Gold U.S. Inc, (a wholly-owned subsidiary of Barrick). The partners use an operating company, Donlin Gold LLC (Donlin Gold) to manage the Project. For the purposes of this Report, "Donlin Gold" is used as a synonym for the partnership. Prior to July 2011, Donlin Gold was known as Donlin Creek LLC (DCLLC).

       

    NovaGold is using the Report in support of a press release dated 5 December, 2011, entitled "NovaGold Passes Key Milestone On Path to Becoming Premier North American Gold Producer; Completes Positive Feasibility Study On Donlin Gold Project Natural Gas Pipeline's Economic Benefits Confirmed Capex Estimate Declines From Previous Guidance Project Ready to Advance to Permitting", and a press release dated 12 January 2012 entitled "NovaGold Files Donlin Gold Feasibility Study Technical Report".

       
    2.1

    Terms of Reference

       

    The second updated feasibility study was completed in October, 2011, and was a compendium of different studies by a number of companies, as indicated in Table 2-1.

       

    AMEC used the information completed by these contributors to support information in the current Report. AMEC’s QPs performed or commissioned independent due diligence reviews on the information supplied by Donlin Gold and made adjustments to the results of the FSU2 report based on the outcome of those reviews.

       

    AMEC notes that the FSU2 project description has changed materially in some areas from the first updated feasibility study of February 2009 (FSU1).

       

    The Report uses Canadian English. Unless otherwise specified in the text, monetary amounts are in US dollars and units are metric.


         
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    December 2011
    Page 2-1



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    Figure 2-1: Regional Project Setting


    Note: Figure courtesy Donlin Gold

    Figure 2-2: Local Project Setting


    Note: Figure courtesy Donlin Gold

         
    Project No.: 166549
    December 2011
    Page 2-2



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

      Table 2-1: Consulting Firms or Entities Contributing to FSU2
    Consulting Firm or Entity Area of Responsibility in FSU2 Report Document
    AMEC Americas Ltd

    Overall study compilation, design of port and process facilities (excluding the autoclave and autoclave ancillary facilities), flowsheet, development of logistics program; equipment pricing, excluding equipment associated with the autoclave, oxygen plant, and mining, quantity estimation for major civil and structural components, capital cost estimates for off-site facilities, on-site facilities, and process facilities, excluding the mine, autoclave, autoclave support facilities, and oxygen plant, operating cost estimates for process, transport, and administration, excluding mining, development of Project plan and schedule

    Donlin Gold and Barrick

    Geologic modelling; resource and reserve estimation, specification and management of metallurgical testwork program; bench and pilot testing facilities for pressure oxidation and neutralization; specification and management of environmental and socioeconomic baseline studies, including impact analysis; permitting requirements; reclamation planning; baseline environmental data; process (excluding EPCM requirements) and mining engineering and preproduction costs; financial evaluation; mine planning; capital cost estimates for the mine; operating cost estimates for the mine; sustaining capital cost estimates for the mine

    BGC Engineering Inc.

    Geotechnical engineering to support the mine pits, waste rock facility, plant site, and tailings storage facility; site water management; mine waste rock management; design of the tailings storage facility and waste rock facility foundations; pit dewatering plans for the mine

    CH2MHill

    Routing and geotechnical studies for the selected alignment of the natural gas pipeline; pipeline design and engineering; construction execution planning and scheduling; capital and operating cost estimates for the natural gas pipeline

    HATCH Ltd.

    Flowsheet development of autoclave process; design of autoclave and autoclave ancillary facilities; equipment pricing for autoclave and autoclave ancillary facilities; quantity estimation for autoclave and autoclave ancillary facilities; capital cost estimate for autoclave and autoclave ancillary facilities; operating cost estimate for autoclave and autoclave ancillary facilities; logistics plan for delivery of autoclave

    Lorax Environmental Services Ltd.

    Water quality modelling for the mine pit lake

    SRK

    Acid rock drainage (ARD) and metal leaching (ML) assessment; closure cost estimate

    Rowland Engineering Consultants

    Geotechnical investigations to support port site, airstrip, and material borrow sources; geotechnical engineering for access roads between port site, airstrip, and plant site


         
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    December 2011
    Page 2-3



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    2.2

    Qualified Persons

    The following people served as the Qualified Persons (QPs) as defined in National Instrument 43-101, Standards of Disclosure for Mineral Projects, and in compliance with Form 43-101F1:

    • Gordon Seibel, R.M. SME, Principal Geologist, AMEC Reno

    • Kirk Hanson, P.E., Technical Director, Open Pits North America, AMEC Reno

    • Tony Lipiec., P.Eng., Manager, Process Engineering, AMEC Vancouver.

    2.3

    Site Visits and Scope of Personal Inspections

    The QPs conducted site visits to the Project as shown in Table 2-2.

    Mr Seibel completed a data verification site visit to the Project on 1 October 2008. During the visit, core logs were compared to the core, lithologies in the resource model were compared to the lithologies in the surface outcrops, and core logging, and sampling protocols were reviewed. Handling of the core and sample preparation, however, could not be observed directly as no drilling or sample preparation was being performed during the site visit.

    During the October 1, 2008, site visit, Mr Hanson undertook a high-level review of the Project geology, inspected drill core, viewed the Project topography, inspected proposed pit and waste dump locations, and the locations of existing and proposed infrastructure, including road cuts and borrow pits.

    In addition to these visits, other AMEC personnel have visited site during preparation of the FSU2 report, and have provided input to the AMEC QPs in the areas of their expertise in support of this Report.

    AMEC considers that although completed in 2008, the site visits are still current. Since the date of the last technical report filed on the Project, Donlin Gold has completed an additional 62 drill holes (25,000 m) out of 1,740 holes (370,000 m). This drilling is not considered to comprise a material change to the Project. Changes to the Mineral Resource and Mineral Reserve statements in Section 15 of this Report are primarily driven by the increases to the gold price used in estimation.

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

      Table 2-2: QPs, Areas of Report Responsibility, and Site Visits
    Qualified Person                              Site Visits Report Sections of Responsibility
        (or Shared Responsibility)
    Tony Lipiec No site visit Sections 1, 2, 3, 4, 5, 6, 13, 14.5.2, 14.6.2, 14.8, 14.9.4, 14.9.5, 14.9.6, 17, 18, 19, 20, 21, 22, 23, 24, 25, 26, and 27
    Gordon Seibel 1 October 2008 Sections 7, 8, 9, 10, 11, 12, 14 (excepting 14.5.2, 14.6.2, 14.8, 14.9.4, 14.9.5 and 14.9.6), and those portions of the Summary, Interpretations and Conclusions and Recommendations that pertain to those Sections
    Kirk Hanson 1 October 2008 Sections 15 and 16, and those portions of the Summary, Interpretations and Conclusions and Recommendations that pertain to those Sections.

    2.4

    Effective Dates

    The Report has a number of effective dates, as follows:

    • Effective date of the assay database that supports estimation: 1 November 2009

    • Effective date of the Mineral Resources: 11 July 2011

    • Effective date of the Mineral Reserves: 11 July 2011

    • Effective date of the tenure and surface rights data: 7 October 2011

    • Effective date of the financial analysis that supports the updated feasibility study: 18 November 2011.

    The overall effective date of the Report, based on the supply of the financial data, is 18 November 2011.

       

    There has been no material change to the scientific and technical information on the Project between the effective date of the Report, and the signature date.

       
    2.5

    Previous Technical Reports

       

    NovaGold has previously filed the following technical reports on the Project:

       

    Francis, K., 2008: Donlin Creek Project, NI 43-101 Technical Report, Southwest Alaska, U.S.: unpublished technical report prepared for NovaGold Resources Inc., effective date 5 February 2008.


         
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    December 2011
    Page 2-5



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    Dodd, S., Francis, K. and Doerksen, G., 2006: Preliminary Assessment Donlin Creek Gold Project Alaska, USA, unpublished NI43-101F1 Technical Report to NovaGoldResources Inc. by SRK Consulting (US), Inc., effective date September 20, 2006

    Dodd, S., 2006: Donlin Creek Project 43-101 Technical Report, unpublished NI43-101F1 Technical Report to NovaGold Resources Inc. by NovaGold Resources Inc., effective date January 19, 2006

    Juras, S. and Hodgson, S., 2002: Technical Report, Preliminary Assessment, Donlin Creek Project, Alaska, unpublished NI43-101F1 Technical Report to NovaGold Resources Inc. by MRDI, report date March 2002.

    Juras, S., 2002: Technical Report, Donlin Creek Project, Alaska, unpublished NI43-101F1 Technical Report to NovaGold Resources Inc. by MRDI, effective date 24 January, 2002.

    2.6

    Information Sources and References

    The primary data source for this Report is the 2011 Feasibility Study Update 2, entitled:

    AMEC Americas Ltd., 2011: Donlin Creek Gold Project, Alaska, Feasibility Study Update 2, Effective Date 7 October 2011: unpublished feasibility study update prepared by AMEC for Donlin Creek LLC, dated 9 October 2011.

    Reports and documents listed in the Section 3, Reliance on Other Experts and Section 27, References sections of this Report were also used to support preparation of the Report. Additional information was sought from NovaGold, Barrick, and Donlin Gold personnel where required.

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    3.0

    RELIANCE ON OTHER EXPERTS

       

    The QPs have relied upon the following other expert reports, which provided information regarding mineral rights, surface rights, property agreements, and marketing sections of this Report as noted below.

       
    3.1

    Mineral Tenure

       

    The QPs have not reviewed the mineral tenure, nor independently verified the legal status, ownership of the Project area, underlying property agreements or permits. AMEC has fully relied upon, and disclaims responsibility for, information derived from Donlin Gold experts and experts retained by Donlin Gold for this information through the following documents:

    • Sellers, T.M., 2009: Legal Opion on Mineral Title: unpublished confidential legal opinion prepared by Reeves Amodio LLC for Donlin Creek LLC, and addressed to James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 20 March 2009

    • Manzer, D.S., 2011: Donlin Gold Project – Administrative Status of State of Alaska Mining Claims, Effective Date October 5, 2011: unpublished title search prepared by Alaska Land Status Inc. for Perkins Coie LLP and addressed to Robert Maynard of Perkins Coie LLP and James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 6 October 2011

    • Manzer, D.S., 2011: Donlin Gold LLC’s Donlin Gold Project, Updated Abstract of Record Title, Effective September 2, 2011: unpublished title search prepared by Alaska Land Status Inc. for Perkins Coie LLP and addressed to Robert Maynard of Perkins Coie LLP and James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 7 October 2011

    • Maynard, R.M., 2011: Update of March 20, 2009 Reeves Amodio LLC Title Opinion Letter for Donlin Crcek Project Real Property Interests: unpublished confidential legal opinion prepared by Perkins Coie LLP for Donlin Creek LLC, and addressed to James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 10 October 2011.

    This information is used in Section 4.3 of the Report and was also used to support considerations of reasonable prospects of economic extraction and declaration of Mineral Resources in Section 14.3 and 14.4, and for consideration of appropriate modifying factors for declaration of Mineral Reserves in Section 15.3.

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    3.2

    Surface Rights

    The QPs have fully relied upon and disclaim responsibility for information supplied by Donlin Gold staff and experts retained by Donlin Gold for information relating to the status of the current surface rights as follows:

    • Sellers, T.M., 2009: Legal Opion on Mineral Title: unpublished confidential legal opinion prepared by Reeves Amodio LLC for Donlin Creek LLC, and addressed to James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 20 March 2009

    • Manzer, D.S., 2011: Donlin Gold Project – Administrative Status of State of Alaska Mining Claims, Effective Date October 5, 2011: unpublished title search prepared by Alaska Land Status Inc. for Perkins Coie LLP and addressed to Robert Maynard of Perkins Coie LLP and James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 6 October 2011

    • Manzer, D.S., 2011: Donlin Gold LLC’s Donlin Gold Project, Updated Abstract of Record Title, Effective September 2, 2011: unpublished title search prepared by Alaska Land Status Inc. for Perkins Coie LLP and addressed to Robert Maynard of Perkins Coie LLP and James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 7 October 2011

    • Maynard, R.M., 2011: Update of March 20, 2009 Reeves Amodio LLC Title Opinion Letter for Donlin Crcek Project Real Property Interests: unpublished confidential legal opinion prepared by Perkins Coie LLP for Donlin Creek LLC, and addressed to James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 10 October 2011.

    This information is used in Section 4.5 of the report and for consideration of appropriate modifying factors for declaration of Mineral Reserves in Section 15.3.

    3.3

    Agreements

    The QPs have fully relied upon and disclaim responsibility for information supplied by Donlin Gold staff and experts retained by Donlin Gold or NovaGold for information relating to the status of the current Property Agreements as follows:

    • Sellers, T.M., 2009: Legal Opion on Mineral Title: unpublished confidential legal opinion prepared by Reeves Amodio LLC for Donlin Creek LLC, and addressed to James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 20 March 2009

    • Manzer, D.S., 2011: Donlin Gold Project – Administrative Status of State of Alaska Mining Claims, Effective Date October 5, 2011: unpublished title search prepared by Alaska Land Status Inc. for Perkins Coie LLP and addressed to Robert Maynard of Perkins Coie LLP and James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 6 October 2011

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY
    • Manzer, D.S., 2011: Donlin Gold LLC’s Donlin Gold Project, Updated Abstract of Record Title, Effective September 2, 2011: unpublished title search prepared by Alaska Land Status Inc. for Perkins Coie LLP and addressed to Robert Maynard of Perkins Coie LLP and James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 7 October 2011

    • Maynard, R.M., 2011: Update of March 20, 2009 Reeves Amodio LLC Title Opinion Letter for Donlin Crcek Project Real Property Interests: unpublished confidential legal opinion prepared by Perkins Coie LLP for Donlin Creek LLC, and addressed to James Fueg, Feasibility Study Manager for Donlin Gold LLC, dated 10 October 2011.

    This information is used in Section 4.4 of the Report and was also used to support considerations of reasonable prospects of economic extraction and declaration of Mineral Resources in Section 14.3 and 14.4, and for consideration of appropriate modifying factors for declaration of Mineral Reserves in Section 15.3.

       
    3.4

    Royalties

       

    The QPs have fully relied upon and disclaim responsibility for information supplied by NovaGold staff and experts retained by NovaGold for information relating to the status of the current royalties payable as follows:

       
  •  
  • Francis, K., 2012: Confirmation Letter Regarding Royalties, Marketing and Taxation Pool: unpublished letter from Kevin Francis, Vice President Resources, Novagold to Scott Mackin, AMEC Project Manager Donlin Gold Project, 6 January 2012
       

    This information is used in Section 4.7 of the report and was also used to support considerations of reasonable prospects of economic extraction and declaration of Mineral Resources in Section 14.3 and 14.4, and for consideration of appropriate modifying factors for declaration of Mineral Reserves in Section 15.3.

       
    3.5

    Marketing

       

    The QPs have fully relied upon and disclaim responsibility for information supplied by NovaGold for information relating to the status of the potential Project marketing regime as follows:

       
  •  
  • Francis, K., 2012: Confirmation Letter Regarding Royalties, Marketing and Taxation Pool: unpublished letter from Kevin Francis, Vice President Resources, Novagold to Scott Mackin, AMEC Project Manager Donlin Gold Project, 6 January 2012

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    This information is used in Section 19, and was used to support considerations of reasonable prospects of economic extraction and declaration of Mineral Resources in Section 14.3 and 14.4, declaration of Mineral Reserves in Section 15.3, and the cashflow analysis in Section 22.

    3.6

    Taxation

       

    The QPs have fully relied upon and disclaim responsibility for information supplied by NovaGold for information relating to the status of the potential taxation pool available to NovaGold through the following:

       
  •  
  • Francis, K., 2012: Confirmation Letter Regarding Royalties, Marketing and Taxation Pool: unpublished letter from Kevin Francis, Vice President Resources, Novagold to Scott Mackin, AMEC Project Manager Donlin Gold Project, 6 January 2012
       

    This information is used in Section 22.5 of the Report.


         
    Project No.: 166549
    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    4.0

    PROPERTY DESCRIPTION AND LOCATION

       
    4.1

    Location

       

    The Donlin deposits are situated approximately 280 miles (450 km) west of Anchorage and 155 miles (250 km) northeast of Bethel up the Kuskokwim River. The closest village is the community of Crooked Creek, approximately 12 miles (20 km) to the south, on the Kuskokwim River.

       

    The resource areas are within T. 23 N., R. 49. W., Seward Meridian, Kuskokwim and Mt. McKinley Recording Districts, Crook Creek Mining District, Iditarod A-5 USGS 1:63,360 topography map. The mineralization is centred on approximately 540222.50 east and 6878534.36 north, using the NAD 83 datum.

       

    These areas consist of the ACMA and 400 Zone, Aurora and Akivik mineralized areas (grouped as ACMA) and the Lewis, South Lewis, Vortex, Rochelieu and Queen mineralized areas (grouped as Lewis). The final proposed pit outline for the combined ACMA and Lewis area in relation to the Calista Lease Boundar y (refer to Section 4.3) is shown as Figure 4-1. Key deposit and prospect areas are shown in Figure 4-2.

       
    4.2

    Project Ownership History

       

    Calista Corporation, an Alaska Native Corporation, has held the mineral rights to the Project since 1974.

       

    Placer Dome acquired a 20-year lease from Calista effective 1 May, 1995. The lease agreement contains a provision that extends the lease period beyond 20 years as long as mining or processing operations continue in good faith or good faith efforts are being made to place a mine on the property into production.

       

    On 13 November, 2002, NovaGold Resources Alaska, Inc., a wholly-owned subsidiary of NovaGold Resources Inc., earned a 70% interest in the Project by expending US$10 million on exploration and development of the Project. Placer Dome retained an option to buy back into the Project.

       

    On February 11, 2003, Placer Dome exercised its back-in right and assumed management of the continued development of the Project. In January 2006, Barrick acquired Placer Dome and assumed Placer Dome’s joint venture responsibilities with regard to Donlin.


         
    Project No.: 166549
    December 2011
    Page 4-1



    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    Figure 4-1: Proposed Pit Location in Relation to Calista Lease Boundary


    Note: Figure courtesy Donlin Gold

         
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    Figure 4-2: Key Deposit and Prospect Areas

    Note: Figure courtesy NovaGold, figure dated September 2011. Coloured portions indicate different generations of intrusive dikes.

    On December 1, 2007, NovaGold entered into a limited liability company agreement with Barrick (the Donlin Creek LLC Agreement) that provided for the conversion of the Project into a new limited liability company, the Donlin Creek LLC, which is jointly owned by NovaGold and Barrick on a 50/50 basis. In July 2011, the Board of Donlin Creek LLC voted to change the name of the company to Donlin Gold LLC.

    As part of the Donlin Creek LLC Agreement, NovaGold has agreed to reimburse Barrick over time for approximately US$64.3 million, representing 50 percent of Barrick’s approximately US$128.6 million expenditures at the Project from April 1, 2006 to November 30, 2007. NovaGold reimbursed US$12.7 million of this amount following the effective date of the agreement by paying US$12.7 million of Barrick’s share of Project development costs. The remaining approximately US$51.6 million will bear interest and be paid out of future mine production cash flow. Funding is currently shared by both parties on a 50/50 basis.

         
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    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
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    Lyman Resources has existing placer mining leases covering approximately four square miles within the Donlin lease area. The Lyman family also has title to approximately 13 acres of surface estate within the Snow Gulch area. This lease area lies immediately to the north of the current open pit shell outline but in the opinion of Donlin Gold, should not result in any significant conflicts with the pit shell or envisioned infrastructure layout. The Calista Exploration and Lode Mining Lease grants priority to extraction of the lode mineralization in the event of a conflict of use between lode and placer mining operations, provided that a two-year notice period is provided to Lyman Resources. Negotiations regarding the future of the Lyman holdings are ongoing.

       
    4.3

    Lease Rights

       

    The Donlin exploration and mining lease currently includes a total of 72 sections in the vicinity of the deposit, and additional partial sections associated with the Project infrastructure, leased from Calista Corporation, an Alaska Native Corporation that holds the subsurface (mineral) estate for Native-owned lands in the region. Title to 20,081 hectares (49,261 acres) of the leased lands has been conveyed to Calista by the Federal Government. Calista owns the surface estate on a portion of these lands.

       

    In March 2010, DCLLC renegotiated its lease with Calista. Amendments to the renegotiated lease provide for:

    • The lease of certain additional lands that may be required for the development of the property

    • An extension of the term of the lease to April 30, 2031 and automatically year to year thereafter, so long as either mining or processing operations are carried out on or with respect to the property in good faith on a continuous basis in such year, or Donlin Gold pays to Calista an advanced minimum royalty of US$3.0 million (subject to adjustment for increases in the Consumer Price Index) for such year

    • The elimination of Calista’s option to acquire a 5% to 15% participating operating interest in the Project and replacement with the payment to Calista of a net proceeds royalty equal to 8% of the net proceeds realized by Donlin Gold at the Project after deducting certain capital and operating expenses (including an overhead charge, actual interest expenses incurred on borrowed funds and a 10% per annum deemed interest rate on investments not made with borrowed funds)

    • An increase in the advanced minimum royalties payable to Calista under the lease to US$0.5 million for the year ending April 30, 2010, increasing on an annual basis thereafter until reaching US$1.0 million for each of the years 2015 to 2024 inclusive and US$2.0 million for each of the years 2025 to 2030 inclusive. All advance minimum royalties paid to Calista continue to be recoverable as a credit against Calista's existing net smelter royalty under the lease agreement, which remains unchanged.

         
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    A separate Surface Use Agreement with The Kuskokwim Corporation (TKC), an Alaska Native Village Corporation that owns the majority of the private surface estate in the area, grants non-exclusive surface use rights to Donlin Gold on at least 34 sections overlying the mineral deposit, with provisions allowing for adjusting that area in conjunction with adjustments to the subsurface included in the Calista lease. The term of the Surface Use Agreement runs through 5 June 2015, with provisions to extend beyond that time so long as mining, processing, or marketing operations are continuing and the Calista lease remains in effect.

    The Lyman family owns a small (13 acre) private parcel in the vicinity of the deposit and holds a placer mining lease from Calista that covers approximately four sections.

    Figure 4-3 shows the land status in the Project area.

    4.4

    Mineral Tenure

    In addition to the 49,261 acres (20,081 hectares) leased from Calista, Donlin Gold holds 242 Alaska State mining claims comprising 31,740 acres (12,845 hectares), bringing the total land package to 81,361 acres (32,926 hectares).

    Most of the rights (surface and subsurface) are governed by conditions defined by the Alaska Native Claims Settlement Act (ANCSA). Section 12(a) of ANCSA entitled each village corporation to select surface estate land from an area proximal to the village in an amount established by its population size. Calista receives conveyance of the subsurface when the surface estate in those lands is conveyed to the village corporation. Section 12(b) of ANCSA allocated a smaller entitlement to the regional corporations with the requirement they reallocate it to their villages as they choose. Calista receives subsurface estate when its villages receive 12(b) lands. Calista reallocated its 12(b) entitlement in 1999, based on a formula that was based on original village corporation enrolments.

         
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    Figure 4-3: Donlin Gold Project Land Status Map

    Note: Figure courtesy Donlin Gold

         
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    The Donlin exploration and mineral lease currently includes lands leased from Calista, which holds the subsurface (mineral) estate for Native-owned lands in the region. The leased land is believed to contain 20,081 ha. Title to all lands has been conveyed to Calista by the Federal Government. Calista owns the surface estate on a portion of these lands.

    A separate Surface Use Agreement with TKC, which owns the surface estate of the remaining lands, grants non-exclusive surface use rights to Donlin Gold. All of these lands have been conveyed to TKC by the Federal Government. Donlin Gold has entered into negotiations with the TKC in regards to the Surface Use Agreement, as that agreement expires in 2015.

    Donlin Gold holds 242 unpatented State mineral claims totalling 12,845 ha, primarily surrounding the leased land in the Kuskokwim and Mount McKinley recording districts. Of these claims:

    • Three are on State-selected lands

    • A total of 158 are tentatively approved from conveyance from Federal to State- owned, pending survey.

    None of the claims held by Donlin Gold have been surveyed.

    All claims are either 64.8 ha or 16.2 ha in size.

    4.5

    Surface Rights

    Donlin Gold, through native lease agreements, holds a significant portion of the surface rights that will be required to support mining operations in the proposed mining area. Negotiations with TKC will be required for surface rights for additional lands supporting mining and access infrastructure.

    Currently, Donlin Gold operates under the Mining Lease with the Calista Corporation and the Surface Use Agreement with TKC. The terms of the TKC Surface Use agreement include the following:

    • Annual aggregate surface use fee of $50,000

    • Once exclusive-use lands are identified, payment of an annual exclusive-use fee of 10% of the fair market value of the property

    • Or, at TKC’s request, purchase the exclusive-use property.

         
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    Drilling operations for the Project are covered under the Alaska Placer Mining process and related permits.

    Other lands required for off-site infrastructure, such as the gas pipeline, port site, and access road, are categorized as Native, State of Alaska conveyed, or BLM (Federal) lands.

    4.6

    Royalties and Encumbrances

    The terms for the Calista Exploration and Lode Mining Lease include the following:

    • Annual advance minimum royalty increasing to $650,000 in 2014

    • Annual advance minimum royalty of $1 million from 2015 to 2024

    • Annual advance minimum royalty of $2 million from 2025 to 2030

    • Net smelter return of 1.5% for the earlier of the first five years following commencement of commercial production or until payback

    • Conversion to 4.5% after the earlier of five years or payback

    • Net proceeds royalty of 8% of the net proceeds realized by Donlin Gold commencing with the first quarter in which net proceeds are first realized

    • Calista shareholder hire preference and Calista 5% bidding preference on competitive contracts for all work on the property leased from Calista.

    Currently there are no Government royalty obligations, and no other royalties are payable on the Project.

    4.7

    Permits

    Donlin Gold advised AMEC that Barrick has maintained all of the necessary permits for exploration and camp facilities. These permits are active at the Alaska Department of Natural Resources (hard rock exploration, temporary water use), the Corp of Engineers (individual 404 and nationwide 26), Alaska State Department of Conservation (wastewater, drinking water, food handling), the Alaska Department of Fish and Game (title 16 – fish), the Environmental Protection Agency (NPDES) and the Federal Aviation Administration (airport).

    Permits required to support Project development are discussed in Section 20.

         
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    4.8

    Environmental Liabilities

       

    Environmental studies, closure plans and costs, and environmental liabilities and issues are discussed in Section 20.

       
    4.9

    Social License

       

    The potential social and community impact assessments of the Project are discussed in Section 20.

       
    4.10

    Significant Risk Factors

       

    Based on Project design concepts developed through the prefeasibility and feasibility engineering work and on the results of four years of community interaction, several key environmental issues of concern have been identified for the Project. These are discussed in detail in Section 20.

       
    4.11

    Comments on Section 4

       

    In the opinion of the QPs, the following conclusions are appropriate:

    • Information from Donlin Gold and legal experts supports that the mining tenure held is valid and is sufficient to support declaration of Mineral Resources and Mineral Reserves. Mineral tenures have not been surveyed. Appropriate claim payments have been made as required, and required expenditure commitments had been met by Donlin Gold.

    • Information from Donlin Gold indicates that surface rights are held by either TKC or Calista. This information supports that Donlin Gold has agreements with both parties as follows. Calista owns the surface estate on 27 of the 72 sections that make up the Project. TKC has granted Donlin Gold non-exclusive surface use rights to Donlin Gold on at least 34 sections overlying the mineral deposit, with provisions allowing for adjusting that area in conjunction with adjustments to the subsurface included in the Calista lease. The Lyman family owns a small (13 acre) private parcel in the vicinity of the deposit and holds a placer mining lease from Calista that covers approximately four sections. The currently identified resource and the bulk of the primary infrastructure (mill and waste rock facilities) are located on the leased lands. Additional lands required for the Jungjuk port site, road to the port site, gas pipeline, and tailings storage facility in Anaconda Creek are located on a combination of Native, State of Alaska, and Federal (Bureau of Land Management, BLM) lands. Rights-of-way will be required from the State and BLM for the road and pipeline alignments where they cross state and federal lands, respectively. Discussions regarding the extension and expansion of the TKC Surface Use Agreement and the disposition of the Lyman family land parcel and lease are ongoing.

         
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  •  
  • Information from Donlin Gold indicates that royalty payments are associated with the Calista lease as follows:

           
     
  •  
  • annual advance minimum royalty increasing to $650,000 in 2014

     
  •  
  • annual advance minimum royalty of $1 million from 2015 to 2024

     
  •  
  • annual advance minimum royalty of $2 million from 2025 to 2030

     
  •  
  • net smelter return of 1.5% for the earlier of the first five years following commencement of production or until payback

     
  •  
  • conversion to 4.5% after the earlier of five years or payback

     
  •  
  • net proceeds royalty of 8% of the net proceeds realized by Donlin Gold commencing with the first quarter in which net proceeds are first realized.

           
     
  •  
  • There are no Government royalty obligations

           
     
  •  
  • Information from Donlin Gold indicates that a payment of an annual Aggregate Surface Use Fee of $50,000 is required to TKC, and either, once exclusive-use lands are identified, payment of an annual exclusive-use fee of 10% of the fair market value of the property, or purchase of the exclusive-use lands

           
     
  •  
  • Exploration to date has been conducted in accordance with Alaskan regulatory requirements

           
     
  •  
  • Additional permits will be required for Project development.

           
     
  •  
  • Key areas identified by stakeholders (refer to Section 20) as areas of social and environmental concern include mercury and cyanide abatement, waste rock and water treatment and management, transportation of goods and materials to and from the Project site, and flora and fauna management. Donlin Gold is of the opinion that these issues have been, or can be, addressed and mitigated through a combination of good baseline data collection, diligent engineering and Project design, and thorough public consultation.


         
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    5.0

    ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY

         
    5.1

    Accessibility

       

    The Donlin deposits are approximately 280 miles (450 km) west of Anchorage and 155 miles (250 km) northeast of Bethel up the Kuskokwim River. The closest village is the community of Crooked Creek, approximately 12 miles (20 km) to the south, on the Kuskokwim River. Bethel, approximately 20 miles (30 km) upstream from the mouth of the Kuskokwim River, is the regional centre for the Yukon-Kuskokwim Delta area of Southwest Alaska. The town of Aniak, also on the Kuskokwim River and about 50 air miles (80 km) southwest of the Project site, is the regional centre for the Upper Kuskokwim Valley.

       

    There is no road or rail access to the site. The nearest roads are in the Anchorage area. Access to Bethel and Aniak, the regional centres, is limited to river travel by boat or barge in the summer and air travel year-round. The Kuskokwim River is a regional transportation route and is serviced by commercial barge lines.

       

    All access to the Project site for personnel and supplies is by air. The Project has an all-season, soft-sided camp with facilities to house up to 150 people. An adjacent 5,000 ft (1,500 m) long airstrip is capable of handling aircraft as large as C-130 Hercules (42,000 lb or 19,050 kg), allowing efficient shipment of personnel, some heavy equipment, and supplies. The Project can be serviced directly by charter air facilities out of both Anchorage and Aniak.

       
    5.2

    Climate

       

    The area has a relatively dry interior continental climate with typically about 20 inches (500 mm) of total annual precipitation. Summer temperatures are relatively warm and may exceed 83°F (30°C). Minimum temperatures may fall to well below -45°F (-42°C) during the cold winter months.

       

    Exploration is possible year round, though snow levels in winter and wet conditions in late autumn and in spring can make travel within the Project area difficult. It is expected that mining operations will be able to be conducted year-round.


         
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    5.3

    Local Resources and Infrastructure

       

    Local resources necessary for the exploration and possible future development and operation of the Project are located in Bethel. Some resources would likely have to be brought in from the Anchorage area.

       

    Alaska and the adjacent Canadian Province of British Columbia have a long mining history and a large resource of equipment and skilled personnel.

       

    The Project is currently isolated from power and other public infrastructure. The exploration camp has a capacity of 160 persons. Power is provided by diesel generators.

       

    Infrastructure assumptions and the proposed infrastructure layout for the Project are discussed in Section 18 of the Report.

       
    5.4

    Physiography

       

    The Project area is one of low topographic relief on the western flank of the Kuskokwim Mountains. Elevations range from 500 to 2,100 ft (150 to 640 m). Ridges are well rounded and easily accessible by all-terrain vehicle.

       

    Hillsides are forested with black spruce, tamarack, alder, birch, and larch. Soft muskeg and discontinuous permafrost are common in poorly drained areas at lower elevations and along north-facing slopes.

       
    5.5

    Sufficiency of Surface Rights

       

    In regard to future mining operations, sufficient space is available to site the various facilities, including personnel housing, stockpiles, tailing storage facility, waste rock storage facilities and processing plants.

       
    5.6

    Comments on Section 5

       

    In the opinion of the QPs:

    • The existing and planned infrastructure, availability of staff, the existing power, water, and communications facilities, the design and budget for such facilities, and the methods whereby goods could be transported to any proposed mine, and any planned modifications or supporting studies are reasonably well-established, or the requirements to establish such, are reasonably well understood by Donlin Gold, NovaGold and Barrick, and can support the declaration of Mineral Resources and Mineral Reserves.

         
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    • There is sufficient area within the Project to host an open pit mining operation, including the proposed open pits, mine and plant infrastructure, waste rock and tailings storage facilities.

    • It is a reasonable expectation that any additional surface rights to support Project development and operations can be obtained.

    • It is expected that any future mining operations will be able to be conducted year- round.

         
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    6.0

    HISTORY

    Placer gold was first discovered at Snow Gulch, a tributary of Donlin Creek, in 1909. Intermittent small-scale placer gold production has continued to the present. Resource Associates of Alaska (RAA) carried out a regional evaluation for Calista in 1974 to 1975. This work included a soil grid and three bulldozer trenches in the Snow area immediately north of the current resource area. Calista followed up with prospecting activities between 1984 and 1986, and completed minor auger drilling in 1987.

    The first substantial exploration drill program was carried out by Western Gold Exploration and Mining Co. LP (WestGold) in 1988 and 1989. WestGold completed geological mapping, trenching, rock and soil sampling, an airborne magnetic and VLF survey, and ground magnetic surveys. WestGold also tested biogeochemical sampling and ground penetrating radar with positive results. Based on this information, WestGold performed an initial Mineral Resource estimate.

    Teck Exploration Ltd. (Teck) carried out a limited trenching and soil sampling program in the Lewis area in late 1993, and updated the Mineral Resource estimate.

    Placer Dome US (Placer Dome) explored the property from 1995 to 2000. Placer Dome constructed an exploration camp and airstrip, undertook reconnaissance and geological mapping, aerial photography, completed rock chip and soil sampling, trenching, max-min (EM) geophysical surveys, airborne geophysical surveys, RC and core drilling, carried out detailed metallurgical test work, and prepared a series of Mineral Resource estimates and initial mining and engineering studies.

    Placer Dome formed the Donlin Creek joint venture (DCJV) with NovaGold Resources, Inc. as operator in 2001. During the period of the DCJV, NovaGold undertook trenching, core and geotechnical drilling, updated Mineral Resource estimates, and completed a Preliminary Assessment.

    Placer Dome reassumed management of the Project as operator in late 2002. From 2002 to 2005, work comprised additional core drilling, condemnation, geotechnical, and water drilling, geotechnical and hydrogeological studies, geological mapping and sampling of prospective calcium carbonate source areas, exploration and auger drilling program for sand and gravel resources, and updated Mineral Resource estimates.

    Barrick Gold (Barrick) acquired the Placer Dome interest in the DCJV through a merger with Placer Dome in early 2006. Work completed in the period 2006-2007 included core drilling for resource infill, geotechnical, engineering, condemnation, waste rock, calcium carbonate exploration and metallurgical purposes, and updated Mineral Resource estimates.

         
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    The DCJV partners formed DCLLC in late 2007, with the subsequent name change to Donlin Gold occurring in 2011.

    To 2011, work completed has consisted of soil and stream sediment sampling, core drilling for resource infill, geotechnical, engineering, condemnation, waste rock, and metallurgical purposes, and estimation of Mineral Resources and Mineral Reserves.

    An initial feasibility study was completed on the Project in 2007, and updated in 2009. A second update was performed in 2011, and is the subject of this Report.

         
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    7.0

    GEOLOGICAL SETTING AND MINERALIZATION

       
    7.1

    Regional Geology

    The Donlin gold deposits lie in the central Kuskokwim basin of southwestern Alaska, and is a northeast-trending basin that subsided between a series of amalgamated terranes. Rocktypes within the basin include Mesozoic marine volcanic rocks, Palaeozoic clastic and carbonate rocks, and Proterozoic metamorphic rocks.

    The Kuskokwim basin is predominately underlain by the Upper Cretaceous Kuskokwim Group, a back-arc continental margin basin fill assemblage that formed in response to a change in the angle of convergence between the Kula oceanic plate and the Cretaceous North American continental margin. Sediments primarily consist of a coarse- to fine-grained turbidite comprising sandstone, siltstone, and shale with minor conglomerate.

    Late Cretaceous and Early Tertiary volcano-plutonic complexes intrude and overlie the Kuskokwim Group sedimentary rocks. Volcanic components of these complexes consist of intermediate tuffs and flows. Subaerial volcanic tuffs, flows, and domes are regionally extensive and dominantly andesitic, locally include dacite, rhyolite, and basalt. Associated plutons are calc-alkaline in composition, ranging from monzonite to granodiorite. Felsic to intermediate hypabyssal granite to granodiorite porphyry dikes, sills, and plugs are also widely distributed and often intruded into northeast-striking extensional faults. Volumetrically minor Upper Cretaceous intermediate to mafic intrusive bodies are also common.

    The centre of the Kuskokwim basin lies between two continental-scale, dextral slip-fault zones: the Denali-Farewell Fault system to the south and the Iditarod-Nixon Fork Fault system to the north. Fold-and-thrust-style deformation formed the earliest structures in response to subduction-related compression shortly after deposition of the Kuskokwim sediments. Eastward-trending folds and thrust faults are common in the central Kuskokwim basin, including the Donlin Creek area. Younger north–northeast-trending folds are dominant near the Iditarod-Nixon Fork Fault and Denali-Farewell Fault but also formed throughout the region in response to basin-scale dextral movement. Most of the folds predate emplacement of the volcano-plutonic complexes. Pre-, syn-, and post-(?) intrusion, northeast-striking normal and oblique slip faults formed during subsequent late compressional and extensional events and focused intrusive igneous rocks and hydrothermal systems across the basin.

    A regional geological plan is included as Figure 7-1.

         
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    Figure 7-1: Regional Geology of Central Kuskokwim Area

    Figure courtesy Donlin Gold.

    7.2

    Project Geology

       

    Because outcrop is limited and of generally poor quality, property-scale geology is largely interpreted from trenches, drill holes, aeromagnetic surveys, and soil geochemistry.

       

    The Project area is underlain by a 5 mile (8.5 km) long x 1.5 mile (2.5 km) wide granite porphyry dike and sill swarm hosted by lithic sandstone, siltstone, and shale of the Kuskokwim Group.

       
    7.2.1

    Lithologies

       

    The oldest igneous rocks at Donlin Creek are intermediate to mafic dikes and sills. They are not abundant but occur widely throughout the property as generally thin and discontinuous bodies. The younger and much more voluminous granite porphyry intrusive rocks vary from a few feet to 200 ft (60 m) wide and occur as west–northwest- trending sills in the southern resource area and north–northeast-trending dikes farther north. The granite porphyry dikes and sills all have similar mineralogy, and the porphyry texture indicates relatively shallow emplacement. Although these rocks belong to the regionally important granite porphyry igneous event, geologists working on the Project classify them into five textural varieties of rhyodacite. These units are chemically similar, temporally and spatially related, and probably reflect textural variations of related intrusive events.


         
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    Figure 7-2 illustrates the interpreted property-scale distribution of igneous rocks, including the mineral resource area between the Queen deposit area on the northeast and the airstrip on the southwest.

    7.2.2

    Structure

    The Project is located in a structurally complex area about 15 miles (25 km) southeast of the Iditarod–Nixon Fault (refer to Figure 7-1). Sedimentary bedding generally strikes northwest and dips 10° to 50° to the southwest. Overall, sedimentary structure in the northern resource area is monoclinal, while sedimentary rocks in the southern resource area display open eastward-trending folds. East–southeast-trending and plunging folds or monoclinal warps are the oldest recognized structures and are associated with north-vergent thrust faults. Thrust faults are generally southwest-dipping, parallel to the bedding plane, and account for imbrication of the sedimentary rocks and locally moderate to steep southwest and northeast dips. Younger, low-amplitude north–northeast-trending folds crop out in the airstrip exposures along American Creek and are recorded on historic trench geology maps. Lack of cleavage or other evidence of dynamic recrystallization suggests that folds and thrust faults formed at relatively shallow depths.

    7.3

    Deposit Setting

    A northeast, elongated, roughly 5,000 ft (1.5 km) wide x 10,000 ft (3 km) long cluster of gold deposits has an aggregate vertical range that exceeds 3,100 ft (945 m). The deposits are hosted primarily in igneous rocks, and are associated with an extensive Upper Cretaceous gold–arsenic–antimony–mercury hydrothermal system. Gold occurs primarily in sulphide and quartz–carbonate–sulphide vein networks in igneous rocks and, to a much lesser extent, in sedimentary rocks. Broad disseminated sulphide zones formed in igneous rocks where vein zones are closely spaced. Submicroscopic gold, contained primarily in arsenopyrite and secondarily in pyrite and marcasite, is associated with illite–kaolinite–carbonate–graphite-altered host rocks.

         
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    Figure 7-2: Interpreted Property-Scale Igneous Rocks

    Note: RDA = Aphanitic Porphyry; RDX = Crowded Porphyry; RDXL = Lath-Rich Porphyry; and RDXB = Blue Porphyry. Figure courtesy Donlin Gold.

    7.4 Paragenesis

    Fluid inclusion studies and field and drill hole observations define three distinct styles of gold mineralization that are locally telescoped and cross-cut one another. The earliest is a porphyry-style stockwork vein system at the Dome prospect.

    Dome is located within the same dike-and-sill swarm that hosts the ACMA–Lewis resource, but the Kuskokwim sedimentary rocks are thermally metamorphosed to a siliceous hornfels. Quartz veins have a Au–Ag–Cu–Zn–Bi ± Te trace metal signature (Ebert et al., 2003c; Drexler, 2010) with up to 3% arsenopyrite–pyrite–chalcopyrite–pyrrhotite ± Fe-rich sphalerite and trace amounts of electrum, native bismuth, and bismuth tellurides and selenides. Veins cut both the hornfels and porphyry dikes.

    ACMA–Lewis-style mineralization post-dates the Dome veins and consists of sparse Au–Ag–As–Sb–Hg ± W (Ebert et al., 2003c; Drexler, 2010), trace metal-bearing quartz-Fe-dolomite veins with <3% auriferous arsenopyrite-pyrite ± stibnite ± late realgar, native arsenic, and graphite. Veins and related disseminated sulphide zones are primarily hosted in illite-carbonate-kaolinite-altered rhyodacite dikes and sills but also occur in Kuskokwim Group sedimentary rocks near igneous contacts.

         
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    Variations between Dome and ACMA–Lewis vein habits, vein mineralogy, wall rock alteration, geochemical signatures, stable isotope variations (Drexler, 2010), and fluid inclusion chemistry (Ebert et al., 2003c) indicate that hydrothermal fluids were sourced at depth northeast of the Dome prospect, precipitated the base metal assemblage at Dome from metals sequestered in the vapour phase, and then migrated southwestward to the more distal ACMA–Lewis environment, where gold-bearing minerals were precipitated due to mixing with meteoric waters and boiling.

    The last event consists of gold-bearing quartz–stibnite veins up to 3 ft (1 m) thick with variable carbonate, pyrite, and arsenopyrite found mainly around the margins of Dome and partially overlapping ACMA–Lewis. Quartz–stibnite veins also contain anomalous Au–As–Cu–Zn–Bi and have fluid chemistry and temperatures intermediate between Dome and ACMA–Lewis (Ebert et al., 2003). In the opinion of Donlin Gold, these veins do not contain significant gold mineralization.

    7.5

    Deposit Geology

    Most of the detailed trench, road cut, and outcrop maps have not yet been compiled into a geological “fact map” in the resource area. The surface geology illustration in Figure 7-3 is a projection of the 3D geological model of intrusive rock units and faults shown in a perspective view of an orthophoto-draped digital elevation model (DEM) image.

    7.5.1

    Sedimentary Rocks

    Informal sedimentary stratigraphy in the immediate deposit area is shown in Table 7-1. The approximate thicknesses of each unit are from the southern, or ACMA, resource area.

    The stratigraphy in the deposit area consists of basin margin clastic rocks (MacNeil, 2009) dominated by greywacke (lithic sandstone) units with complex transition zones of interbedded siltstone, shale, and greywacke. Marker beds are not yet recognized, so absolute stratigraphic breaks are difficult to identify. Greywacke is dominant in the northern part of the resource area (Lewis, Queen, Rochelieu, Akivik), whereas shale–siltstone-rich units are common in the southern part (South Lewis, ACMA).

         
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    Figure 7-3: Interpreted Surface Geology of Resource Area

    Note: Oblique view looking northeastward showing igneous rock units, faults, drill holes, and Mineral Reserve (DC9) pit outline. Figure courtesy Donlin Gold

    Table 7-1: Donlin Gold Project Stratigraphy

        Apparent Thickness
    Assigned Nomenclature Principal Rock Type (ft) (m)
    Upper Greywacke greywacke 328+ 100+
    Upper Siltstone siltstone/shale 164 50
    Main Greywacke greywacke 262 80
    Main Shale shale/siltstone up to 459 (with sills) up to 140 (with sills)
    Basal Greywacke greywacke 656 200+

    Note: After Piekenbrock and Petsel (2003)

    7.5.2

    Igneous Rocks

    The mafic igneous rocks and the five textural varieties of rhyodacite recognized in the Donlin deposits were also shown in Figure 7-3. Table 7-2 lists the intrusive rocks, also from oldest to youngest.

         
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    Table 7-2: Donlin Gold Project Intrusive Rocks

     Name Code Age Rock Types
    Mafic
    Dikes/Sills
    MD oldest

    Intermediate to mafic dikes and sills; locally host high-grade gold; generally less than 10 ft thick. In the transition area between Akivik and ACMA, mafic sills are extremely abundant within the Lower Greywacke, immediately below the Main Shale

    Fine-
    Grained
    Porphyry
    RDF -

    Earliest rhyodacite intrusions recognized. grey, typically fine-grained, felsic porphyries. RDF intrusives occur as two main northeast-striking, 16.5 to 32.8 ft (5 to 10 m) wide dikes in the Lewis zone and possible discontinuous bodies in early eastward-trending compressional faults, e.g., the Lo Fault

    Crowded
    Porphyry
    RDX -

    Volumetrically the most significant intrusive phase. Grey, characterized by a uniformly crowded feldspar porphyry texture. Present as two 164 to 328 ft (50 to 100 m) wide dike zones in the eastern edge of the north to north–northeast mineralized trend of Lewis/South Lewis. RDX is also found as sills throughout ACMA near the basal part of the sill sequence.

    Lath-Rich
    Porphyry
    RDXL -

    Characterized by sparse, elongate plagioclase laths; significant coarser-grained biotite. occurs as two important dikes in the Akivik area that strike south into the centre of the ACMA deposit. In Akivik and ACMA, RDXL occurs as a significant sill immediately below the RDX sill. The RDXL sill continues to the west but pinches out to the east. RDXL dikes are also present within the main Lewis area RDX dike trend, but here they are volumetrically insignificant

    Aphanitic
    Porphyry
    RDA -

    Rhyodacite rock with a salt-and-pepper texture of fine biotite phenocrysts and variable quartz and potassium feldspar phenocrysts. Numerous (up to eight) RDA dikes strike south from the Vortex/Rochelieu (Lewis) area into the East ACMA/ACMA area. The dikes are typically found west of the Vortex Fault but are also present between the Lo and Vortex faults and below the Lo Fault. An extensive sill package of RDA lies immediately above the RDX sills in the ACMA area. In West ACMA, the RDA sills are buttressed against, and locally cross-cut, RDX sills. Another package of RDA sills is found south of the AC Fault, in the Aurora domain.

    Blue
    Porphyry
    RDXB youngest

    Final intrusive event; coarsely porphyritic with large blocky feldspars set in a graphite- and sulphide-rich matrix. locally hosts important high-grade disseminated sulphide material in addition to gold-bearing veins. RDXB occurs as two major dikes, the Lewis Blue Porphyry dike and the Vortex Blue Porphyry dike. Extensive thin RDXB sills are found in the uppermost part of the sill sequence in the South Lewis and ACMA areas, and RDXB sills are present as both distinct sills and co-mingled with RDA in the core of ACMA and in the Aurora domain.

    Note: After Piekenbrock and Petsel (2003)

    7.5.3

    Structure

    The morphology of intrusive rocks in the deposit is largely governed by the rheology of sedimentary rocks and pre-intrusion faults and folds. Faults in the geological model (from earliest to youngest) are the American Creek (AC) Fault, Lo and Rochelieu faults, Vortex Fault, and ACMA Fault. Figure 7-4 shows a plan view of the faulting in the deposit area, and cross-sections through the ACMA and Lewis areas, respectively in Figure 7-5 and Figure 7-6.

         
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    Figure 7-4: 100 m Bench Level Geology

      Note: Oblique view looking north-eastward of the 3D geological model projected on the 328 ft (100 m) pit bench level and the Mineral Reserves (DC9) pit outline. Note: Figure courtesy Donlin Gold.

    Figure 7-5: Lewis Area Section

      Note: Shows intrusive rocks, faults, drill holes, and Mineral Reserves (DC9) pit outline, looking northeast. Note: Figure courtesy Donlin Gold.

         
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    Figure 7-6: ACMA Area Section

      Note: Shows intrusive rocks, faults, drill holes, and Mineral Reserves (DC9) pit, looking southeast. Note: Figure courtesy Donlin Gold.

    7.6

    Deposits

    The Donlin deposits include eleven mineralized areas that exhibit slightly different geological settings but generally fall into two geologically similar deposit areas: ACMA and Lewis. ACMA, or the intrusive sill and shale–siltstone sedimentary setting, includes the Aurora, 400, Akivik, ACMA, and East ACMA mineralized zones. Lewis, or the massive greywacke-hosted intrusive dike setting, includes the South Lewis, Lewis, Vortex, Rochelieu, Queen, and North Akivik mineralized zones.

    Veins in north–northeast-striking, east- or west-dipping faults and fracture zones are the primary control on gold distribution and are ubiquitous in all mineralized areas. Northwest- and northeast-striking veins occur locally but are relatively rare. Veins are narrow (typically <1 cm wide), highly irregular, discontinuous, and generally sparsely distributed, although vein density can locally range up to 2 to 8 per meter in higher-grade zones. Vein zones vary from 6.5 to 100 ft (2 to 35 m) wide and 300 to 1,150 ft (100 to 350 m) long. Individual vein zones generally display limited lateral and vertical continuity; however, swarms of many anastomosing vein zones form larger mineralized corridors characterized by extensive lateral and depth continuity.

    Vein corridors are more apparent in the north–northeast-trending dikes of Lewis than in the west–northwest-trending ACMA sill zone. The greater width of the sill-hosted ACMA mineralized zone makes discreet corridors less obvious (but still present).

         
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    Mineralized zones follow steeply dipping dikes and sills beyond the depth limits of current drilling, or over a vertical range of at least 3,100 ft (945 m).

    Veins are best developed in relatively more brittle intrusive rocks and massive greywacke. Small, irregular, carbonate-altered mafic bodies often host very high grade gold as sulphide dissemination, replacement, and breccia fill. Structural breccias in sedimentary rocks are also favourable sites for high-grade gold. Gold distribution in the deposit closely mimics the intrusive rocks, which contain about 80% of the resource. Structural zones in competent sedimentary units account for the remaining 20%. The more steeply dipping sills in the ACMA sill sequence host the highest-grade and most continuous igneous-hosted mineralized zones, particularly where intersected by northeast-striking “feeder” dikes and faults. Gold grade is directly proportional to vein density and intensity of overlapping disseminated sulphide vein aureoles. The dike-dominant Lewis deposit areas consist of sheeted veins with limited disseminated sulphide in the wall rocks and are characterized by lower-grade and less continuous mineralized zones.

    Gold distribution in the planned pit area is shown in Figure 7-7, as a bench plan.

    7.7

    Mineralization

       

    Gold-bearing zones are coincident with quartz–carbonate–sulphide veins and related disseminated sulphide aureoles in hydrothermally altered rhyodacite bodies and, to a lesser extent, in sedimentary rock near igneous contacts. Continuity and grade of mineralized material within the rhyodacite host rocks varies directly with vein spacing and the amount of vein and disseminated arsenopyrite, the principal gold-bearing mineral. Gold in sedimentary rocks and minor mafic igneous bodies is generally limited to small and discontinuous vein and breccia fill occurrences.

       
    7.7.1

    Vein and Disseminated Mineralization

       

    Veins in the ACMA–Lewis area are subtle in appearance and vary from <1 mm to 20 cm wide, averaging <1 cm. They formed in brittle fractures and are typical of open- spaced fillings with vugs, drusy quartz-lined cavities, vein wall-banded and cockscomb quartz, and bladed carbonate. Veins are composed of gray to clear quartz, white to tan carbonate, and as much as 3% sulphides. Table 7-3 contains a summary of the gold-bearing vein stages.


         
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    Figure 7-7: 100 m Bench Level Gold Distribution (>1 g/t Au grade blocks)


    Note: Figure courtesy Donlin Gold

      Table 7-3: Vein Stages

      Vein Description
      V1

    Thin, irregular, and discontinuous sulphide (>50%) veins with pyrite and trace arsenopyrite, little or no quartz (<30%) or carbonate (<50%). Broad disseminated selvage of pyrite and poorly crystalline illite and Fe–carbonate alteration. Barren or very low grade

      V2

    Thin, discontinuous quartz (>30%) sulphide veins contain variable pyrite and arsenopyrite. May have broad, often pervasive selvages of fine-grained, needle- like arsenopyrite. Broad pyrite aureole may surround the arsenopyrite selvage. Open-space vuggy textures common. Trace amounts stibnite. Have moderate gold grade and strong illite alteration aureoles with variable Fe–carbonate replacement of the host rock.

      V3a

    Higher-grade veins. Thicker, more planar and continuous, open-space quartz veins with Fe-dolomite, pyrite, arsenopyrite, native arsenic, and variable amounts of stibnite. Commonly show broad arsenopyrite-rich selvages with little to no Fe– carbonate as wall rock alteration

      V3b

    Thicker, more continuous, and planar quartz veins with open-space textures and complex mineralogy, including pyrite, arsenopyrite, stibnite, native arsenic, realgar, and trace other sulphides in intensely illite altered material. Gold grades are commonly much higher than the average grade of the deposit

      V4

    Latest vein phase. Barren carbonate-quartz (>50% and <50%, respectively) vein sets that post-date mineralized veins. Primarily barren white and clear quartz veinlets and calcite ± ankerite veinlets with no sulphides.


         
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    Mineralized zones are consistently oriented sub-parallel to the main 01 axis (024) of the compressive structural regime (Piekenbrock and Petsel, 2003). Veins in the ACMA–Lewis resource evolved through a continuum (V1 through V3) of changing mineralogy and increasing gold grade while maintaining a generally consistent NNE strike and SE dip. The final carbonate–quartz vein set (V4) has a broader range of orientation.

    MacNeil (2009) found that the average vein orientation for all veins is 024/71. This orientation is generally consistent across all domains and vein types, which indicates that veins at Donlin formed during the same mineralizing event.

    A comparison by host rock shows that veins in igneous rocks strike more easterly and dip more steeply than veins in sedimentary rocks, probably due to refraction across lithologic contacts.

    Several quartz and carbonate phases have been recognised, including pre-gold-stage Mn–calcite veins and wall rock replacement and cockscomb quartz veins; Fe–dolomite in main gold stage veins; and post-gold-stage clear quartz veins and ankerite stringer veins.

    Euhedral and porous replacement pyrite are the earliest sulphide phases, followed in order by marcasite, arsenopyrite, realgar, and native arsenic. Stibnite is most abundant in later veins. Most accessory sulphides are relatively early, while boulangerite is relatively late. Arsenopyrite occurs as both coarse (up to 1 cm) crystals and very fine (0.1 to 0.2 mm) euhedral grains. Fine-grained arsenopyrite contains five to 10 times more gold than the paragentically earlier coarse-grained phase.

    7.8

    Alteration

    Rhyodacite bodies are ubiquitously altered to an illite–carbonate–kaolinite–chlorite / smectite ± quartz ± graphite assemblage.

    Mafic igneous rocks are strongly altered by carbonate ± fuchsite and contain locally high-grade gold with disseminated, massive replacement or breccia filling sulphide.

    Altered sedimentary rocks consist of relict quartz grains in a matrix of illite, kaolinite, carbonate, hematite, and <1% pyrite and trace sphalerite (Drexler, 2010).

    Pyrite is widespread in all altered rocks (0.5% to 2%) but is more abundant (1% to 4%) in mineralized zones. Alteration is most intense near veins and is typically zoned outward from iIlite ± kaolinite to kaolinite ± illite and then to a distal zone of chlorite ± smectite ± quartz.

         
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    Silica is dominantly restricted to veins in the ACMA–Lewis area and is not generally expressed as pervasive silicification. Vein relationships show an increase in quartz content from early sulphide-dominant veins to late silica-dominant veins. Some increased silicification has been noted in the Queen area (Ebert, 2003b).

    Short-wave infrared reflectance (SWIR) spectroscopy data, collected between 2007 and 2011, are interpreted by Donlin Gold show that higher gold is most strongly correlated with an alteration suite dominated by NH4–illite (ammonia–illite), whereas kaolinite-bearing zones contain lower-grade gold.

    7.9

    Minor Elements

    The most abundant minor elements associated with gold-bearing material are iron (Fe), arsenic (As), antimony (Sb), and sulphur (S). These are contained primarily in the mineral suite associated with hydrothermal deposition of gold, including pyrite (FeS2), arsenopyrite (FeAsS), realgar (AsS), native arsenic (As), and stibnite (Sb2S3). Minor hydrothermal pyrrhotite (Fe 1-x S) and marcasite (FeS2), and syngenetic or sedimentary pyrite, also account for some of the Fe and S.

    Much less abundant elements such as copper (Cu), lead (Pb), and zinc (Zn) are contained in relatively rare or accessory hydrothermal mineral species observed in the deposit, including chalcopyrite (CuFeS2), chalcocite (Cu2S), covellite (CuS), tennantite (Cu12As4S13), tetrahedrite (Cu12Sb4S13), bornite (Cu5FeS4), native copper (Cu), galena (PbS), sphalerite (ZnS), and boulangerite (Pb5Sb4S11). Small amounts of silver (Ag) in the deposit are most likely accommodated within the crystal structures of tetrahedrite and galena, and to a lesser extent in some of the other sulphides. Molybdenum (Mo) occurs in rare molybdenite (MoS2). Very minor nickel (Ni) has been observed in the secondary sulphide mineral millerite (NiS) and minor cobalt (Co) in various secondary minerals in sedimentary rocks. The Ni and Co probably have a sedimentary origin.

    Three elements of particular processing significance are mercury (Hg), chlorine (Cl), and fluorine (F). Graphitic carbon and carbonate minerals also have the potential to negatively affect the metallurgical process.

    7.10

    Comments on Section 7

    In the opinion of the QPs:

    • Knowledge of the deposit settings, lithologies, and structural and alteration controls on mineralization is sufficient to support Mineral Resource and Mineral Reserve estimation.

         
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    8.0

    DEPOSIT TYPES

    According to Donlin Gold, the Donlin gold deposits share characteristics of several gold deposit genetic models. It has been classified as:

    • Granite porphyry-hosted gold polymetallic (Bundtzen and Miller, 1997)

    • Distal or high-level epizonal intrusion-related (Hart et al., 2002)

    • Low-sulphidation epithermal (Ebert et al., 2003a)

    • Orogenic- or intrusion-related (Goldfarb, 2004),

    • Reduced porphyry to sub-epithermal Au–As–Sb–Hg (Ebert et al., 2003c; Hart, 2007).

    Hart (2007) classifies the deposit as a high-level, reduced intrusion-related vein system to account for the reduced ilmenite series intrusions, near contemporaneous age of mineralization, and the apparent genetic relationship to the higher-temperature hydrothermal system at Dome (Drexler, 2010).

    The Lewis–ACMA part of the district is clearly a low sulphidation, reduced intrusion related, epizonal system with both vein and disseminated mineral zones and conforms most closely to the Hart (2007) classification.

    8.1

    Comments on Section 8

    In the opinion of the QPs, deposit models as used in the exploration programs have been appropriate to the style and setting of the mineralization.

         
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    9.0 EXPLORATION
       
    A summary of the exploration programs completed on the Project is summarized in Table 9-1.
       
    9.1 Grids and Surveys
       
    The Project uses Universal Transverse Mercator (UTM) Zone 4 (meters). The map datum is NAD83.
       
    The topographic surface is based on a 2004 survey by Aero-metric. The survey has an accuracy of ±6.6 ft (±2 m) within all key Project areas.
       
    9.2 Geological Mapping
       
    Geological mapping was performed in 1988-1989 (Westgold), and during 1996, 1998, reconnaissance mapping was undertaken by Placer Dome. This was followed in 1999 by Placer Dome completing a 1:10,000 geological mapping program over the entire Project area.
       
    Mapping is generally limited by the poor quality and limited extent of outcrop. Information from the mapping programs was used to support more detailed data obtained from trenches and core drilling.
       
    9.3 Geochemical Sampling
       
    During the period 1988 to 1989, Westgold also collected over 15,000 soil, rock chip and auger samples. Westgold, in 1989, tested biogeochemical sampling, which returned positive results. Teck collected two lines of soil samples in 1993.
       
    Placer Dome, during 2007, collected 600 soil samples in the ACMA and 400 areas, and an additional 646 soil samples and 92 rock samples were collected in 2008. During 2008, Barrick took 1,097 soil, 101 stream sediment, and 66 stream concentrate geochemical samples.
       
      Sampling was used as part of regional prospectivity evaluations.
       
    9.4 Geophysics
       
    Westgold performed an airborne magnetic and VLF survey and ground magnetic surveys during 1998 to 1999. The company also trialled ground-penetrating radar.

         
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    Table 9-1: Work History Summary for Donlin Gold Project

    Year Company Work Performed Results
    1909 to 1956 Various prospectors and placer miners

    Gold discovered on Donlin Creek in 1909. Placer mining by hand, underground, and hydraulic methods.

    Total placer gold production of approximately 30,000 oz.

    1970s to present Robert Lyman and heirs

    Resumed sluice mining in Donlin area and placer mined Snow Gulch.

    First year of mining Snow Gulch was the best ever: 800 oz Au recovered.

    1974, 1975 RAA

    Regional mineral potential evaluation for Calista Corporation. Soil grid and three bulldozer trenches in the Snow Gulch area.

    Anomalous gold values in soil, rock, and vein samples. Trench rock sample results range from 2 ppm Au to 20 ppm Au.

    1984 to 1987 Calista Corporation

    Minor work. Various mining company geologists including Cominco and Kennecott visit property.

    -

    1986 Lyman Resources

    Auger drilling for placer evaluation finds abundant gray, sulphide-rich clay near Quartz Gulch.

    Assays of cuttings average over 7 ppm Au. Initial discovery of Far Side (Carolyn) prospect.

    1987 Calista Corporation

    Rock sampling of ridge tops and auger drill sampling of Far Side prospect.

    Anomalous gold values from auger holes: best result = 9.7 ppm Au.

    1988, 1989 WestGold

    Airborne geophysics, ground geophysics, geological mapping, and soil sampling over most of Project area. Total of 44,362 ft (13,525 m) of D9 bulldozer trenching at all prospects. Over 15,000 soil, rock chip, and auger samples collected. Drilling included 3,106 ft (947 m) of AX core drilling, 1,325 ft (404 m) in 239 auger holes, and 34,187 ft (10,423 m) of RC drilling (125 holes). First metallurgical tests and petrographic work.

    Initial work identified eight prospects with encouraging geology ± Au values (Snow, Dome, Quartz, Carolyn, Queen, Upper Lewis, Lower Lewis, and Rochelieu). Drilling at most of these prospects led to identification of the Lewis areas as having the best bulk-mineable potential. Calculated gold resource of 3 M tons at average grade of 2.50 ppm (218,908 oz) at 1 ppm cut-off. WestGold dissolved by early 1990.

    1993 Teck

    D-9 bulldozer trenching (4,592 ft, 1,400 m) and two 1,640 ft (500 m) soil lines in Lewis area. Petrographic, fluid inclusion, and metallurgical work.

    Identified new mineralized areas and expanded property resource estimate to 3.9 Mt at average grade of 3.15 g Au/t (393,000 oz Au). Metallurgical tests not favourable, Project dropped.

    1995 to 2000 Placer Dome

    286,616 ft (87,383 m) of core, 39,062 ft (11,909 m) of RC drilling, and 27,857 ft (8,493 m) of trenching.

    Drilled the American Creek magnetic anomaly (ACMA), discovered the ACMA deposit. Numerous mineral resource calculations.

    2001, DCJV (Placer

    152,543 ft (46,495 m) of core, 38,022 ft (11,589 m) of RC drilling, 294

    Expanded the ACMA resource.

    2002 Dome/NovaGold)

    ft (89.5 m) of geotechnical drilling, and 879 ft (268 m) of water monitoring holes. Mineral Resource estimate.

    2003 to 2005 DCJV

    83,491 ft (25,448 m) of core and 19,611 ft (5,979 m) of RC drilling. Calcium carbonate exploration drilling; IP lines for facility condemnation studies.

    Infill drilled throughout the resource area demonstrated continuity. Discovered a calcium carbonate resource. Poor quality IP data not useful for facility studies.

    2006 DCJV (Barrick/NovaGold)

    304,475 ft (92,804 m) of core drilling for resource conversion, slope stability, metallurgy, waste rock, carbonate exploration, facilities, and port road studies.

    Geological model and internal resource updates.

    2007 DCJV

    Core drilling totalled 246,906 ft (75,257 m) and included resource delineation, geotechnical and engineering, and carbonate exploration. 13 RC holes for monitor wells and pit pump tests totalled 3,423 ft (1,043 m). Updated Mineral Resource estimate.

    Improved pit slope parameters, positive hydrogeological results, Carbonate exploration was negative.


         
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    Year Company Work Performed Results
    2008 DCLLC (Barrick/NovaGold) 108 core holes totalling 109,663 ft (33,425 m) for exploration and facility related geotechnical and condemnation studies. Metallurgical test work: flotation variability and CN leach. 54 test pits and 37 auger holes completed for overburden characterization. Resource expansion indicated for East ACMA. CN leach resource potential indicated for the main resource area, Snow, and Dome prospects. Facility sites successfully condemned. Updated resource estimates utilizing applicable data through 2007
    2009 DCLLC 19 geotechnical core holes totalling 3,116 ft (950 m) in facility sites and to address hydrology. Mineral Reserve and Mineral Resource estimate update.
    2010 DCLLC Six geotechnical core holes totalling 6,855 ft (2,090 m) to evaluate slope stability of expanded pit. Also drilled 90 auger holes totalling 1,919 ft (585 m) and dug 59 test pits to further evaluate overburden conditions and gravel supplies within TSF area. Mineral Reserve and Mineral Resource estimate update. Pit slope stability of new pit design remained acceptable. Evaluation of construction suitability of surficial materials in TSF is ongoing.

         
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    During 1999, Placer Dome completed 25 line km of max-min (electromagnetic) geophysical survey completed in the ACMA, 400 and southern Lewis areas, and 1,800 line km of aeromagnetic survey was completed at 50 m line spacing and 50 m elevation over the property. In the same year, a total of 17.7 km of IP and resistivity lines were completed, and an additional 41.6 km of IP/resistivity lines were run in 2000. During 2003, IP surveys were undertaken in areas where infrastructure was planned, as part of condemnation evaluations.

       

    Geophysical studies were used to support structural interpretations for geological modelling purposes, exploration targeting, and facilities condemnation drilling

       
    9.5

    Pits and Trenches

       

    During 1988 and 1989, a total of 44,362 ft (13,525 m) of D9 bulldozer trenching was completed by Westgold at all prospects.

       

    Teck completed 4,592 ft, (1,400 m) of D-9 bulldozer trenching during 2003.

       

    The 1996 Placer Dome exploration program included than 8,200 ft (2,500 m) of trenches for sampling and mapping purposes in southeast Lewis area; this was followed in 1997 by 13,852 ft (4,222 m) of trenches in the Lewis area, in 1998 by 6,739 ft (2,054 m) of trenching and geological mapping in the Lewis-Vortex and ACMA areas, and in 1999 by 7,339 ft (2,237 m) of trenching and geologic mapping (Dome, Queen, Far Side, and Vortex).

       

    Pits to provide geotechnical data were excavated in 2005 (22 test pits), 2006 (40 test pits), 2007 (55 test pits), and 2008 (54 test pits).

       
    9.6

    Petrology, Mineralogy, and Research Studies

       

    During 1993, Teck completed petrographic and fluid inclusion studies in support of understanding of the mineralization setting and host rock lithologies.

       
    9.7

    Geotechnical and Hydrological Studies

       

    Geotechnical and hydrological studies have been undertaken in support of mine planning, mine design and environmental considerations.

       
    9.8

    Metallurgical Studies

       

    Metallurgical testwork is primarily based on drill core.


         
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    9.9

    Exploration Potential

       

    Exploration potential in the vicinity of the open pit design in FSU2 includes extensions along strike to the East ACMA, Lewis, and Crooked Creek dike areas. Mineralization remains open at depth uner the current pit limits. Mineralization also remains open to the north of the planned pit and has been tested by shallow trenching and soil sampling, with limited drilling undertaken to date.

       

    The Project also retains exploration potential outside the areas that have been the subject of the mine design in FSU2. Gold mineralization is associated with an overall north–northeasterly-trending high level dike/sill complex that has been outlined in the regional aero-magnetics as a subtle 50 nT magnetic low (Figure 9-1). The zone, approximately 8 km long, and 4 km wide, consists of a northern, dike-dominated area, and a southern, more sill-dominated area (refer to Figure 9-1 and Figure 4-1).

       

    Figure 9-2 shows a gold-in-soils compilation plan of the area indicated by the magnetic low. The ACMA/Lewis area is the southern portion of this plan. No drilling has been performed in the northern portion since initial exploration activities, and some isolated drilling in the 1990s. Exploration targets identified by NovaGold for additional work includes Far Side, Duqum, Snow, Quartz, Queen, Dome, and Ophir (refer to Figure 4- 1). The following summaries of the exploration potential identified by NovaGold are sourced from Buchanan (2009), Chamois (2009) and Francis (2011).

       
    9.9.1

    Far Side

       

    The Far Side prospect has been tested by three NovaGold core drill holes (735 m) along 300 m of strike and 29 RC holes that were drilled by West Gold. Drill results for the core drilling are indicated in Table 9-2. The prospect is situated in an area where dikes of a generally easterly trend intersect the more dominant northeasterly trend.

       
    9.9.2

    Duqum

       

    This prospect is the site of the first recorded lode gold discovery in the Donlin area. It is about 1 km long, and mineralization is associated with a narrow porphyry dike. Three core holes have been drilled (1,043 m) by NovaGold. Drill results are summarized in Table 9-3.


         
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    Figure 9-1: Regional Magnetic Image Showing Magnetic Low Intensity Zone


    Note: Figure courtesy Donlin Gold

    Figure 9-2: Gold-in-Soils Compilation Plan


    Note: Figure courtesy Donlin Gold

         
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    Table 9-2: Far Side

    Hole ID Northing Easting Elevation Azimuth Dip Drill Drilled Gold
                Intercept Thickness Grade
                From (m) (g/t Au)
                (m)    
    DC96-254 6,883,395  542,905 152      320 61 23.2        16.8 4.60
    DC96-255 6,883,398  542,918 68      313 65 122.0        14.0 3.00
    DC96-256 6,883,347  542,843 55      317 61 130.0        15.6 5.86

    Table 9-3: Duqum

    Hole ID Northing Easting Elevation Azimuth Dip Drill Drilled Gold
                Intercept Thickness Grade
                From (m) (g/t Au)
                (m)    
    DC97-387 6,882,477  543,488 -3 281 51 338.0        16.0 2.39
    DC97-388 6,882,690  542,977 47 314 55 210.0        12.0 5.03
    DC97-388 6,882,700  542,967 26 314 55 236.0        10.0 2.29
    DC97-389 6,882,649  543,109 143 39 56 90.0        10.0 3.86
    DC97-389 6,882,696  543,148 49 40 57 202.0        10.0 2.79
    DC97-389 6,882,746  543,188 -53 36 59 320.0        16.0 3.79

    9.9.3

    Snow/Quartz

       

    The Snow and Quartz prospects are hosted in a dike-related, gold-bearing corridor that is about 1.5 km wide and approximately 4 km long. In the area of this dike swarm, the porphyry dikes are 20 m to >100 m wide, discontinuous bodies. Gold mineralization is associated closely with the dikes, and is hosted either within the dikes themselves or in the adjacent sedimentary rocks. Limited drilling has been completed. Better drill results are included in Table 9-4.

       
    9.9.4

    Dome

       

    The Dome prospect is situated under a prominent, rounded hill about 5 km north of the planned ACMA/Lewis pits. Several mineralized felsic porphyries intrude into a greywacke unit and have hornfelsed the sediment over wide intervals. Mineralization consists of stockworks of veinlets containing arsenopyrite, pyrite, pyrrhotite, and minor chalcopyrite. Preliminary metallurgical testwork undertaken by NovaGold indicates that mineralization may be less refractory than that encountered in the ACMA/Lewis area.

       

    Fourteen widely-spaced drill holes have been completed over an area of approximately 500 m x 500 m. Mineralization is open to the north, east, south and to depth, and may be open to the west at depth. Better drill results are included in Table 9-5.


         
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    Table 9-4: Snow/Quartz

    Hole ID Northing Easting Elevation Azimuth Dip Drill Drilled Gold
                Intercept Thickness Grade
                From (m) (g/t Au)
                (m)    
    DC97-383 6,880,381  541,433        210      294 50      16.0        23.0 2.77
    DC97-384 6,880,551  541,531        150      294 52      52.0        10.0 3.34

    Table 9-5: Dome

    Hole ID Northing Easting Elevation Azimuth Dip Drill Drilled Gold
                Intercept Thickness Grade
                From (m) (g/t Au)
                (m)    
    DC08-1785 6,882,528  544,103 207      109 70 139.0        10.0 2.19
    DC08-1785 6,882,525  544,111 185      111 70 163.0        10.0 3.89
    DC08-1785 6,882,518  544,129 134      113 68 211.0        22.6 3.29
    DC08-1785 6,882,512  544,142 98      114 67 248.0        25.0 2.94
    DC97-392 6,882,456  544,082 232      128 65 94.0        52.0 3.21
    DC97-392 6,882,431  544,113 145      128 65 185.0        61.0 3.30
    DC97-392 6,882,418  544,129 100      132 66 258.0        14.0 3.99

    9.9.5

    Ophir

       

    The Ophir Hill is the highest topographic feature in the Donlin district. Surface mapping over an area of about 1.5 km by 750 m indicates Cretaceous sediments have been intruded by felsic to intermediate intrusions, which may be dikes. Surface exposures are completely oxidized, but boxworks after sulphides indicate arsenopyrite, pyrite and other sulfides occur as disseminations and thin veinlets. Soil sampling has identified a strong gold-in-soil anomaly on the southwestern flanks of the hill. No drilling has been undertaken.

       
    9.10

    Comments on Section 9

       

    In the opinion of the QPs:

    • The exploration programs completed to date are appropriate to the style of the deposits and prospects within the Project

    • Additional exploration potential remains in the Project area.

         
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    10.0

    DRILLING

    Approximately 1,678 exploration and development core (89%) and RC (11%) drill holes, totalling 1,206,960 ft (367,886 m), were completed from 1988 through 2007 in at least six separately managed campaigns. Approximately 50% of the core and 40% of the holes were drilled during 2006–2007. Another 108 core holes totalling 109,634 ft (33,425 m) were added in 2008 to explore near-pit expansions and satellite deposits, and for facility-related condemnation and geotechnical studies.

    A total of 1,396 core (89%) and RC (11%) holes totalling 1,114,324 ft (339,733 m), as well as 282 trenches totalling 70,344 ft (21,441 m), were used for the FSU2 resource model.

    Drill summaries for the RC and core drill programs are included in Table 10-1. Drill location plans are provided in Figure 10-1 for the Project, and in Figure 10-2 for the area where Mineral Resources and Mineral Reserves were estimated.

    10.1

    Drill Methods

    West Gold drilling employed a Winkie rig. Drill data from this program are not contained in the Donlin database.

    Boart Longyear was the coring contractor from 1995 through 2010. Core drilling was accomplished exclusively with LF-70 model drills, which were set up in heli-portable configuration, mounted on skids or on self-propelled tracked and low-ground-pressure Nodwell carriers. Standard wire line core retrieval with 5 ft or 10 ft (1.52 or 3.05 m) core barrels was used in all core drilling operations.

    Core sizes used on the Project include: NQ3 (45.1 mm core diameter), NQ (47.6 mm), HQ3 (61.2 mm), HQ (63.5 mm), and PQ (85 mm). Systematic records of core size were not maintained in the database; therefore, an accurate account of HQ and NQ core cannot be easily determined. It can be stated that most of the core drilled since 1995 was HQ size, since all holes were started with HQ tools and reduced to the smaller diameter NQ size as necessary. The relative amount of NQ size core likely increased in recent campaigns as drilling probed deeper in the deposit.

    Depth limits for HQ size holes were 1,560 ft (475 m) for dry conditions and 1,785 ft (545 m) for fluid-filled holes. HQ depth was generally limited to 1,400 ft (426 m) for holes with no planned reduction.

         
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    Table 10-1: RC and Core Drill Summary Table

    Year Company Number Hole Drill Footage Comment
        of Type    
        Drill      
        Holes      
    1988 WestGold 33 Core unknown

    shallow (average 82 ft, or 25 m), AX-diameter

        50 Auger unknown

    shallow (average 26 ft, or 8 m)

    1989 WestGold 125 RC unknown

    31 holes in Far Side, 38 in Snow, 24 in Queen, 8 in Rochelieu, and 24 in Lewis

    1995 Placer Dome 32 core

    30 in Lewis, 1 at Rochelieu Ridge, and 1 near the mouth of Queen Gulch

    1996 Placer Dome 28 RC

    Seventeen of the holes twinned earlier core holes. Four water wells (3 in camp, 1 in Lewis) were drilled with the RC drill, and 5 core holes in the 400 area were pre-collared through deep overburden.

    116 core

    All but 8 of the core holes were drilled in Lewis or Queen. The others were distributed north of the current resource area in the Dome, Far Side, and Snow prospects

    1997 Placer Dome 52 RC

    Lewis, Queen, Rochelieu, ACMA, 400 Area, Vortex, alongside the American Ridge runway, and Snow. Includes two water wells.

        66 core  

    Lewis, Queen, 400 Area, ACMA, and north of the resource area at Quartz, Duqum, and Dome

    1998 Placer Dome 96 core

    The drilling was done in two phases: four holes in the ACMA-400 area in March and April, and 41 closely spaced holes in the Lewis area in June to October to test variography. Resource expansion drilling in the Lewis, Queen, and ACMA areas was also conducted from June to October.

    1999 Placer Dome 33 core

    Twenty-six of these, totalling 21,949 ft (6,690 m), were resource definition holes drilled in ACMA-400

    2000 Placer Dome 7 core

    5 at Dome and 2 at Quartz, for an evaluation of IP anomalies and potential for high-grade deposits

    2001 Placer Dome 42 core

    Evaluation of the potential for significant resource growth in the ACMA area

    2002 NovaGold 146 RC 38,022 ft (11,589 m)

    141 exploration and resource expansion holes in the ACMA, 400, Lewis, Akivik, Rochelieu, Vortex, and Far East prospects. Three water wells were drilled near the mouth of American Creek, and two were drilled in the Low Road on the south face of Lewis

        196 core 128,255 ft (39,092 m)

    Two of the core holes are geotechnical holes in the Anaconda Creek valley.

    2003 Placer Dome 16 RC

    Water monitoring wells

    2004 Placer Dome 17 RC 7,661 ft (2,335 m)

    Condemnation holes in the Anaconda Creek and upper American Creek valleys

        3 core  

    Geotechnical core holes

    2005 Placer Dome 30 RC 11,955 ft (3,644 m)
        90 core 80,696 ft (24,596 m)

    Infill in ACMA and Lewis


         
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    Year Company Number Hole Drill Footage Comment
        of Type    
        Drill      
        Holes      
    2006 DCJV 327 core 304,475 ft (92,804 m) Pit slope stability, metallurgy, waste rock studies, facilities condemnation, and engineering, and calcium carbonate resource bulk sampling, delineation, and exploration
    2007 DCJV 13 RC 3,423 ft (1,043 m) Monitor wells and pit pump tests
        248 core 246,906 ft (75,257 m) Pit resource infill, pit expansion, carbonate exploration, geotechnical, and engineering studies
    2008 DCLLC 108 core 109,663 ft (33,425 m) Exploration, resource infill, condemnation, and geotechnical studies
    2009 DCLLC 19 core 3,116 ft (950 m) Geotechnical and hydrological core holes
    2010 DCLLC 6 core 6,855 ft (2,090 m) Geotechnical core holes

         
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    Figure 10-1: Project Drill Hole Location Plan


    Note: Figure courtesy Donlin Gold.

         
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    Figure 10-2: Resource Area Drill Holes


    Note: Figure courtesy Donlin Gold.

         
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    Drill holes planned to depths greater than 1,560 ft (426 m) were reduced depending on bit changes or logistical scheduling at a depth range of 600 to 800 ft (244 m). Otherwise, depth capacity was dependent on in-hole tools, tool condition (including drill rod condition), ground conditions, drilling techniques, the variable operating capabilities of each individual drill, and crew safety.

    RC down-hole hammer drilling was provided by Tester Drilling in 1989, Dateline Drilling in 1996 and 1997, and TJ Enterprises in 2002, 2004, 2005, and 2007. RC drill rigs mounted on low-ground-pressure, self-propelled tracked carriers equipped with high-volume air compressors, standard 4" (102 mm) dual walled pipe in 20 ft lengths, and down-hole pneumatic hammers with 5¼" (133 mm) carbide button bits were used for RC drilling. Sample discharge and sample splitting equipment consisted of cyclone collectors mounted above Jones splitters for both wet and dry drilling in 1989 and three-tiered Jones splitters for dry samples and pneumatic rotating wet splitters for wet samples in 1996 and subsequent programs.

    RC drilling was used by WestGold in 1988-1989 for its initial exploration, by Placer Dome in 1997 to reduce impact on wetlands areas, and by NovaGold in 2002 to conduct extensive early-stage resource delineation in several areas of the deposit. Since 2002, core drills have been used exclusively for all resource delineation, and RC drilling was relegated to condemnation and hydrology studies.

    10.2

    Geological Logging

    Standard logging conventions were developed by Placer Dome, and refined over the durations of the drilling programs.

    Standard logging and sampling conventions were used to capture information from the drill core and, where applicable, RC chips. The core was logged in detail using paper forms with the resulting data entered into the main database (Access© database) either by the logging geologist or a technician. Five types of data were captured in separate tables: Lithology, Mineralization, Alteration (visual), Structural and Geotechnical. Remarks were also captured. Lithology was recorded in a two to four letter alpha code. The Mineral table captured visual percent veining (by type) and sulphide (pyrite, arsenopyrite, stibnite and realgar). Specific alteration features including FeOx and carbonate alteration were also captured using a qualitative scale. Structural data collected consisted of the type of structure, measurements relative to core axis and oriented core measurements, if applicable. The Geotechnical table recorded percent recovery and RQD for the entire hole, and fracture intensity where warranted.

         
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    RC drilling chips were logged on paper forms and the data entered into an electronic database.

    10.3

    Recovery

       

    A survey of nearly 200,000 core recovery records in the database revealed an overall length-weighted average core recovery of 95%. Average recovery increases from 80 to 95% from 0 to 40 m and then ranges from 95 to 100% below 40 m where overburden and surface weathering effects are generally absent.

       
    10.4

    Collar Surveys

       

    From 1988 through 1993, conventional theodolite survey methods were used to tie drill hole collar and trench locations to a surveyed ground control net. Drill hole collars were surveyed with Brunton compass and hip chain in 1995. A Motorola GPS system was used in early 1996 to establish survey control monuments and to survey some drill collars. Traditional survey methods were subsequently used to locate all 1995 to 1999 and 2001 drill collars and trenches. An Ashtech Promark2 GPS post-processed system consisting of a base unit and up to two roving units was introduced in 2002. The roving Promark2 instruments were operated in the field to collect stationary readings over the drill collars.

       

    Typical reading times for periods of good satellite reception were about 20 minutes and varied from 10 minutes to four hours or more, depending on the number of satellites, roving-camp base unit distance, canopy coverage, and proximity to north- and east-facing slopes. Data collected by the roving unit and base units were downloaded and post-processed through Ashtech Solutions software. The resulting processed drill collar survey data and vector information were checked for accuracy and quality control, and then copied to an Excel survey data file. This in turn was copied to the acQuire database for archival. Based on Ashtech surveys of control points, the approximate maximum horizontal and vertical variances of drill hole collar surveys under optimal conditions were considered by Donlin Gold to be 0.2 and 0.6 m, respectively.

       
    10.5

    Down-hole Surveys

       

    The Sperry Sun single-shot camera method was used through 2000 for directional surveys to determine down-hole deviation. Reflex EZ Shot instrumentation was introduced in 2001. Six parameters—azimuth, inclination, magnetic tool face angle, gravity roll angle, magnetic field strength, and temperature—were measured. Measurements were generally collected at 150 ft (50 m) intervals from 20 ft (6 m) off bottom to within 100 ft (30 m) of the surface. An integrated key pad and LCD display provided for manual operation and data retrieval. Handwritten data were delivered to the geologists with the shift reports for quality control and manual entry into the acQuire database.


         
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    Approximately 60% of the core holes drilled within the resource model area were oriented to collect structural information for geotechnical and geological studies. Core orientation methods included clay impression, EZ Mark, and Reflex ACT instrument. The clay impression method was used through 2005 but proved problematic as average hole depth increased. Clay impression was replaced with EZ Mark and Reflex ACT tools in 2006. The Reflex ACT tool was used almost exclusively for exploration and resource delineation holes, while the EZ Mark method was used as a backup and for some geotechnical holes. Oriented core required the use of HQ3 and NQ3 bits to accommodate thin-walled inner tubes that reduced core rotation and fragmentation. These bits also produced a smaller-diameter core.

    10.6

    Geotechnical and Hydrological Drilling

       

    Geotechnical and hydrological drilling is included in the drill totals in Table 10-1, and are included in the drill location plan in Figure 10-2.

       
    10.7

    Metallurgical Drilling

       

    Specific drill holes were completed for metallurgical testwork. These holes, although not broken out by collar, are included in the drill location plan in Figure 10-2.

       
    10.8

    Condemnation Drilling

       

    Condemnation drilling was performed to sterilize potential infrastructure sites in 2007. During FSU1, some infrastructure was relocated further to the east, as the host mineralized intrusions were identified to be plunging in this direction. Exploration drilling in early 2008 confirmed that the most prospective intrusive sills plunge to the east and therefore allow only modest potential for eastward pit expansion into the 2007 near-pit facility sites. Condemnation drilling in the facility sites that were relocated in 2008 did not identify near-surface mineralized material or a favourable geologic environment within 1,640 ft (500 m) of the surface.

       

    Exploration drill hole DC08-1783, near the relocated lower contact water dam site, intersected moderately south-dipping, mineralized intrusive rocks in two zones between 1,788 ft (545 m) and 2,352 ft (717 m).


         
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    Figure 10-3: Proposed Facility Sites (FSU1 Layout) and Drill Hole Locations


    Note: Figure courtesy Donlin Gold.

    These intercepts indicate that a there may be potential for mineralization that may be able to be extracted using underground mining methods farther south where sedimentary rock bedding dips and the inferred intrusive sills may have a near vertical orientation. This exploration target depth exceeds 2,460 ft (750 m) beneath the surface of American Ridge and the relocated facility sites.

    Locations of the condemnation drill holes are included in Figure 10-2.

         
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    10.9

    Drill Orientations

    Drill hole orientation relative to the contrasting Lewis dike and ACMA sill orientations, combined with the primary north–northeasterly structural control of gold distribution, was investigated by Placer Dome in 1998 and Barrick in 2006.

    Placer Dome conducted variography testing in the North and South Lewis areas. Fourteen core holes oriented approximately normal to the dikes and the north–northeasterly-trending mineralized zones (295° azimuth / -50° dip) and located on a grid spacing of approximately 115 ft (35 m) showed excellent correlation with both the geological and mineralization models. Twenty-six core holes were also drilled in South Lewis in an area of west–northwest-striking, southwest-dipping sills. Nineteen of the 26 variography holes were oriented to optimize drilling across the north–northeasterly-trending mineralized veins (280° azimuth / -50° dip), and seven were oriented specifically to test sill contacts (50° azimuth / -50° dip). All holes were drilled at approximately 115 ft (35 m) spacing. Results of the variography testing in this area showed some variation with the models, although the overall correlation was good (Baker, 1999).

    Barrick (Jutras, 2006) further investigated the possibility of a gold grade bias in the resource model. Five major drill hole orientations totalling 1,298 drill holes were observed:

    • North (340° to 20° azimuth, 220 holes)

    • Northeast (20° to 70° azimuth, 195 holes)

    • Southwest (200° to 260° azimuth, 176 holes)

    • Northwest (265° to 335° azimuth, 656 holes)

    • Vertical (dip of -90°, 51 holes).

    Jutras (2006) found that northeasterly-oriented (sub-parallel to the north–northeasterly-trending mineral zones) drill holes were higher grade than the average grade of the deposit relative to the other orientations, but that this bias was partly caused by some clustering in a high-grade part of the ACMA deposit. The results of the study showed that the northeast-oriented holes did not present a significant bias for the estimation of the mineral resource and that they should be included in future resource estimates. A standard nothwesterly drill hole orientation (300° azimuth, -60° dip to the northwest) “normal” to the north–northeasterly-trending structural control of the gold mineralization was adopted for all resource delineation programs from 2005 onward.

         
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    10.10

    Twin Drilling

       

    Core and RC holes were compared in 1996 when 17 core holes in the Lewis area were twinned with RC holes. This study found that, in most instances, composite assay intervals from the RC holes were thinner, less continuous, and lower in grade than in the twinned core holes (Szumigala, 1997).

       
    10.11

    Drilled Width versus True Thickness

       

    Although the drill holes were designed to intersect the mineralization as perpendicular as possible; reported mineralized intercepts are longer than the true thickness of the mineralization.

       
    10.12

    Summary of Drill Intercepts

       

    A summary of a number of drill hole intercepts from each key area is shown in Table 10-2. Examples of the drill hole geometry, and drill hole intercepts are shown in Figures 10-3 and 10-4 (ACMA) and Figures 10-5 and 10-6 (Lewis), and demonstrate that the drilling was designed to intersect the mineralization as perpendicular as possible.

       
    10.13

    Comments on Section 10

       

    In the opinion of the QPs, the quantity and quality of the lithological, geotechnical, collar and downhole survey data collected in the exploration and infill drill programs completed by Placer Dome, NovaGold, Barrick, the DCJV, and DCLLC are sufficient to support Mineral Resource and Mineral Reserve estimation as follows:

    • Core logging meets industry standards for gold exploration

    • Collar surveys have been performed using industry-standard instrumentation

    • Downhole surveys were performed using industry-standard instrumentation

    • Recovery data from core drill programs are acceptable

    • Geotechnical logging of drill core meets industry standards for planned open pit operations

    • Drill orientations are generally appropriate for the mineralization style, and have been drilled at orientations that are optimal for the orientation of mineralization for the bulk of the deposit area.

         
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    Table 10-2: Drill Hole Intercept Summary Table

    Hole ID Northing Easting Elevation Azimuth Dip Area Drill Drill Drilled Gold
                  Intercept Intercept Thickness Grade
                  From To (m) (g/t Au)
                  (m) (m)    
    DC06-1114 6878385.36 539899.82 127.97 294.85 -65.4 ACMA 178.00 218.19 40.19 4.14
    DC06-1114 6878385.36 539899.82 127.97 294.85 -65.4 ACMA 234.00 304.68 70.68 4.10
    DC06-1114 6878385.36 539899.82 127.97 294.85 -65.4 ACMA 310.28 316.51 6.23 3.79
                    Total 117.10 4.10
    DC06-1115 6878270 540117 133 288.55 -58.2 ACMA 194.00 198.00 4.00 1.37
    DC06-1115 6878270 540117 133 288.55 -58.2 ACMA 232.00 242.00 10.00 4.58
    DC06-1115 6878270 540117 133 288.55 -58.2 ACMA 248.00 252.98 4.98 19.37
    DC06-1115 6878270 540117 133 288.55 -58.2 ACMA 261.00 280.29 19.29 5.64
    DC06-1115 6878270 540117 133 288.55 -58.2 ACMA 286.00 304.00 18.00 2.19
    DC06-1115 6878270 540117 133 288.55 -58.2 ACMA 310.00 332.94 22.94 3.49
    DC06-1115 6878270 540117 133 288.55 -58.2 ACMA 343.05 368.00 24.95 5.87
                    Total 104.16 5.01
    DC06-1120 6878411.6 539846.32 127.22 295.85 -61.2 W. ACMA 145.00 160.00 15.00 2.48
    DC06-1120 6878411.6 539846.32 127.22 295.85 -61.2 W. ACMA 175.00 205.00 30.00 1.11
    DC06-1120 6878411.6 539846.32 127.22 295.85 -61.2 W. ACMA 235.50 271.46 35.96 2.89
                    Total 80.96 2.15
    DC06-1126 6878684.93 539605.65 123.7 299.05 -60.3 W. ACMA 189.04 204.50 15.46 2.56
    DC06-1126 6878684.93 539605.65 123.7 299.05 -60.3 W. ACMA 222.00 229.30 7.30 4.21
    DC06-1126 6878684.93 539605.65 123.7 299.05 -60.3 W. ACMA 259.90 264.00 4.10 1.21
    DC06-1126 6878684.93 539605.65 123.7 299.05 -60.3 W. ACMA 303.00 309.00 6.00 2.32
    DC06-1126 6878684.93 539605.65 123.7 299.05 -60.3 W. ACMA 317.00 335.00 18.00 5.37
    DC06-1126 6878684.93 539605.65 123.7 299.05 -60.3 W. ACMA 345.00 365.00 20.00 3.27
    DC06-1126 6878684.93 539605.65 123.7 299.05 -60.3 W. ACMA 374.00 382.00 8.00 2.40
                    Total 78.86 3.43
    DC06-1134 6879104.23 539709.27 149.47 297.35 -61.3 Akivik 17.00 27.00 10.00 1.64
    DC06-1134 6879104.23 539709.27 149.47 297.35 -61.3 Akivik 35.00 43.00 8.00 4.22
    DC06-1134 6879104.23 539709.27 149.47 297.35 -61.3 Akivik 187.00 201.00 14.00 5.54
                    Total 32.00 3.99
    DC06-1136 6879210 539771.1 150.99 297.55 -59 Akivik 33.00 47.00 14.00 2.90
    DC06-1136 6879210 539771.1 150.99 297.55 -59 Akivik 61.00 69.00 8.00 2.79
                    Total 22.00 2.86
    DC06-1138 6878826.27 539729.18 136.51 298.35 -61.6 Aurora 13.00 40.00 27.00 2.07
    DC06-1138 6878826.27 539729.18 136.51 298.35 -61.6 Aurora 54.63 68.00 13.37 2.70
    DC06-1138 6878826.27 539729.18 136.51 298.35 -61.6 Aurora 104.23 117.00 12.77 1.51
    DC06-1138 6878826.27 539729.18 136.51 298.35 -61.6 Aurora 285.50 288.00 2.50 12.40
    DC06-1138 6878826.27 539729.18 136.51 298.35 -61.6 Aurora 306.00 316.00 10.00 4.05
    DC06-1138 6878826.27 539729.18 136.51 298.35 -61.6 Aurora 405.00 409.00 4.00 4.54
    DC06-1138 6878826.27 539729.18 136.51 298.35 -61.6 Aurora 460.00 474.00 14.00 3.88
    DC06-1138 6878826.27 539729.18 136.51 298.35 -61.6 Aurora 494.00 506.00 12.00 2.19

         
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    Hole ID Northing Easting Elevation Azimuth Dip Area Drill Drill Drilled Gold
                  Intercept Intercept Thickness Grade
                  From To (m) (g/t Au)
                  (m) (m)    
    DC06-1138 6878826.27 539729.18 136.51 298.35 -61.6 Aurora 526.00 530.00 4.00 3.25
                    Total 99.64 2.96
    DC06-1245 6878553.04 540311.42 170.41 301.65 -58.7 E. ACMA 5.33 23.00 17.67 1.62
    DC06-1245 6878553.04 540311.42 170.41 301.65 -58.7 E. ACMA 42.00 62.00 20.00 1.84
    DC06-1245 6878553.04 540311.42 170.41 301.65 -58.7 E. ACMA 94.00 106.00 12.00 5.53
    DC06-1245 6878553.04 540311.42 170.41 301.65 -58.7 E. ACMA 112.00 126.00 14.00 2.33
    DC06-1245 6878553.04 540311.42 170.41 301.65 -58.7 E. ACMA 128.00 148.97 20.97 2.85
    DC06-1245 6878553.04 540311.42 170.41 301.65 -58.7 E. ACMA 166.00 178.00 12.00 1.21
    DC06-1245 6878553.04 540311.42 170.41 301.65 -58.7 E. ACMA 187.00 193.65 6.65 2.33
    DC06-1245 6878553.04 540311.42 170.41 301.65 -58.7 E. ACMA 210.00 245.50 35.50 8.35
    DC06-1245 6878553.04 540311.42 170.41 301.65 -58.7 E. ACMA 254.00 276.00 22.00 1.89
                    Total 160.79 3.68
    DC06-1252 6878663.18 541745.81 302.68 303.35 -59.6 Lewis 45.21 53.21 8.00 3.98
    DC06-1252 6878663.18 541745.81 302.68 303.35 -59.6 Lewis 104.49 108.49 4.00 1.78
    DC06-1252 6878663.18 541745.81 302.68 303.35 -59.6 Lewis 209.10 215.10 6.00 3.07
    DC06-1252 6878663.18 541745.81 302.68 303.35 -59.6 Lewis 240.90 259.00 18.10 2.56
    DC06-1252 6878663.18 541745.81 302.68 303.35 -59.6 Lewis 265.00 281.00 16.00 4.80
    DC06-1252 6878663.18 541745.81 302.68 303.35 -59.6 Lewis 297.00 309.00 12.00 2.73
                    Total 64.10 3.39
    DC06-1253 6879030.63 541736.09 400.88 300.35 -58.9 Lewis 9.00 14.33 5.33 3.00
    DC06-1253 6879030.63 541736.09 400.88 300.35 -58.9 Lewis 26.00 32.50 6.50 1.84
    DC06-1253 6879030.63 541736.09 400.88 300.35 -58.9 Lewis 39.69 42.50 2.81 3.39
    DC06-1253 6879030.63 541736.09 400.88 300.35 -58.9 Lewis 74.50 85.70 11.20 1.17
    DC06-1253 6879030.63 541736.09 400.88 300.35 -58.9 Lewis 90.00 120.00 30.00 1.60
    DC06-1253 6879030.63 541736.09 400.88 300.35 -58.9 Lewis 293.00 305.00 12.00 1.26
    DC06-1253 6879030.63 541736.09 400.88 300.35 -58.9 Lewis 311.00 340.50 29.50 3.29
    DC06-1253 6879030.63 541736.09 400.88 300.35 -58.9 Lewis 349.00 363.00 14.00 1.21
                    Total 111.34 2.05
    DC06-1183 6879518.7 541253.71 240.06 299.65 -61.8 Rochelieu 42.00 51.00 9.00 2.22
    DC06-1183 6879518.7 541253.71 240.06 299.65 -61.8 Rochelieu 60.00 63.00 3.00 2.01
    DC06-1183 6879518.7 541253.71 240.06 299.65 -61.8 Rochelieu 96.00 102.00 6.00 6.72
    DC06-1183 6879518.7 541253.71 240.06 299.65 -61.8 Rochelieu 110.00 116.00 6.00 3.64
                    Total 24.00 3.67
    DC06-1185 6879443.2 541416.5 288.87 295.35 -61.9 Rochelieu 29.80 55.93 26.13 3.61
    DC06-1185 6879443.2 541416.5 288.87 295.35 -61.9 Rochelieu 63.84 78.00 14.16 1.85
    DC06-1185 6879443.2 541416.5 288.87 295.35 -61.9 Rochelieu 196.00 204.77 8.77 6.66
    DC06-1185 6879443.2 541416.5 288.87 295.35 -61.9 Rochelieu 237.00 251.00 14.00 6.25
    DC06-1185 6879443.2 541416.5 288.87 295.35 -61.9 Rochelieu 283.00 289.00 6.00 1.82
                    Total 69.06 4.02
    DC06-1267 6879724.6 542229.44 374.72 299.05 -59.4 Queen 131.67 135.09 3.42 1.54
    DC06-1267 6879724.6 542229.44 374.72 299.05 -59.4 Queen 190.00 206.00 16.00 2.41

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    Hole ID Northing Easting Elevation Azimuth Dip Area Drill Drill Drilled Gold
                  Intercept Intercept Thickness Grade
                  From To (m) (g/t Au)
                  (m) (m)    
    DC06-1267 6879724.6 542229.44 374.72 299.05 -59.4 Queen 212.00 216.00 4.00 4.27
    DC06-1267 6879724.6 542229.44 374.72 299.05 -59.4 Queen 242.00 248.00 6.00 1.46
    DC06-1267 6879724.6 542229.44 374.72 299.05 -59.4 Queen 266.00 272.00 6.00 2.27
    DC06-1267 6879724.6 542229.44 374.72 299.05 -59.4 Queen 316.43 332.80 16.37 2.10
    DC06-1267 6879724.6 542229.44 374.72 299.05 -59.4 Queen 362.00 377.04 15.04 1.47
                    Total 66.83 2.09
    DC06-1268 6879663.53 542219.04 356.56 293.95 -58.2 Queen 159.55 165.00 5.45 2.08
    DC06-1268 6879663.53 542219.04 356.56 293.95 -58.2 Queen 234.00 243.00 9.00 8.06
    DC06-1268 6879663.53 542219.04 356.56 293.95 -58.2 Queen 254.00 257.00 3.00 2.43
    DC06-1268 6879663.53 542219.04 356.56 293.95 -58.2 Queen 288.00 292.34 4.34 2.54
                    Total 21.79 4.69

         
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    Figure 10-4: Example Drill Cross-Section ACMA


    Note: Figure courtesy Donlin Gold

         
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    Figure 10-5: Vertical Cross Section Through ACMA and Lewis Block Model, Looking 315°


         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    Figure 10-6: Example Drill Cross-Section, Lewis


    Note: Figure courtesy Donlin Gold

         
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    Figure 10-7: Vertical Cross Section Through Lewis Block Model, Looking 45°


         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY
    • Drill orientations are shown in the example cross-sections included as Figures 10-3 to 10-6, and can be seen to appropriately test the mineralization

    • Drill hole intercepts as summarized in Table 10-2 appropriately reflect the nature of the gold mineralization. The table demonstrates that sampling is representative of the gold grades in the deposits, reflecting areas of higher and lower grades.

    • In the bottom of the ACMA pit, the preferred orientation of the drill holes and the trend of the mineralization are both northwest. Defining northwest-trending intrusives using northwest-trending drill holes, however, makes establishing the location of the mineralization more subjective than if the mineralization was defined using drill holes perpendicular to the mineralization. Figure 10-5 illustrates that although the contacts of the mineralized intrusions are well defined at higher elevations, the location of the mineralized intrusions are subjective at the bottom of the pit. Figure 10-5 is not demonstrating that the current location of the orebody is incorrect, it is only demonstrating the possibility that the location of the orebody could be different than what is currently in the model. Since non-optimized location of an orebody at the toe of the highwall could have significant economic consequences, AMEC recommends that the location of the orebody at the bottom of the pit be verified by additional drill holes drilled perpendicular to the trend of the mineralization.

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    11.0

    SAMPLE PREPARATION, ANALYSES, AND SECURITY

       
    11.1

    Sampling Methods

       

    Drill hole sampling protocols were developed by Placer Dome, and refined over subsequent drill programs.

       

    Holes are sampled from the top of bedrock to the end of the hole. Overburden, excluding the organic layer, is also sampled if core recovery was good and if the interval was abnormally thick and composed of abundant rock clasts. Core sample intervals are based on rock type, rock type breaks, and presence of visible sulphide/arsenic minerals. The maximum sample length in zones consisting of intrusive rocks or that contain appreciable sulphide/arsenic minerals is 6.6 ft (2 m), whereas sample lengths in sedimentary rock zones that lack appreciable sulphide/arsenic minerals can be 9.8 ft (3 m). A minimum of three additional 6.6 ft (2 m) sample intervals are placed before and after each intrusive rock or mineralized zone. An aluminum tag inscribed with the sample number is stapled to the core box with a same-numbered paper tag at each sample break. A sampling cutting list is generated that also specifies the insertion points for control samples.

       

    The core is then digitally photographed and split in half with an electric rock saw equipped with water-cooled diamond saw blades. Core cutters orient the core in the saw to ensure a representative split. One-half of the core is returned to the core box for storage at site, and the other half is bagged for sample processing.

       

    In December 2006 and January 2007, a total of approximately 39,360 ft (12,000 m) of whole core was shipped to an off-site logging and core-splitting facility in Anchorage. This facility was managed by Alaska Earth Science (AES) and staffed with both AES and Barrick personnel to ensure that logging, sampling, core splitting, and sample shipment procedures were identical to those used at the Donlin site facility

       
    11.2

    Metallurgical Sampling

       

    Typically, metallurgical sampling consisted of taking half-core samples which were used in flotation and pressure oxidation tests. Whole core samples were only taken for drop weight and SMC comminution tests.

       
    11.3

    Density/Specific Gravity Determinations

       

    Historically, only two specific gravity (SG) values were used in tonnage calculations: 2.65 for the intrusive units and 2.71 for the sediment units. Additional SG measurements were collected in 2006 to provide better coverage of deposit rock units and geographic sub-regions. Statistical evaluations of these SG values showed that they were similar to the historical intrusive and sedimentary rock SG values. Therefore, the historic values were used for the Mineral Resource estimate in Section 14.


         
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    The following methodology was used to determine SG:

  •  
  • Samples of whole core approximately 2" to 4" (5 to 10 cm) in length were first weighed dry and then weighed in water. The dry weighing tray assembly was replaced with a wire basket and the sample was submerged in a five-gallon bucket of water. A small tare weight (to compensate for the removed weighing tray) was attached midway up the wire assembly to facilitate alternating wet and dry measurements.

       
  •  
  • The formula for SG calculation was: Weight in Air/(Weight in Air – Weight in Water). The specific gravities were automatically computed in acQuire when the weights were entered into the database.

       
  •  
  • Measurements were collected for all rock types at a minimum frequency of one sample from all logged rock type intervals and one sample every 49.2 to 65.6 ft (15 to 20 m) in the longer rock unit intervals. Mineralized rock takes precedence over unmineralized rock in a given rock type interval, but sufficient measurements of unmineralized material were also collected to document potential variability.

    The weighted average of all SG data points was 2.69. Table 11-1 summarizes the average SG values by rock type. Data points clearly identified as outliers were removed before the average was determined.

    It was noted that SG values were generally similar among the rhyodacite units (2.63 to 2.67); therefore, the SG measurements were re-evaluated for the three main rock groups—rhyodacite, greywacke, and shale. Historically, only two SG values were used in tonnage calculations, 2.65 for the intrusive units and 2.71 for the sedimentary units. Because the grouped average SG values are similar enough to the historical values used in previous estimates, it was determined that the values used in Table 11-2 would be sufficient for tonnage estimation and that a block model of SG estimates was not warranted. In addition, the blocks contained within the overburden model were set with a reasonable SG value of 2.14. No new SG analysis was undertaken for the DC9 block model of 2009.

         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

     

    Table 11-1: Specific Gravity Values by Rock Type

      Rock Code in      
      DH Database Rock Types No. of Samples Specific Gravity
      ARG Argillite 272 2.67
      CGL Conglomerate 9 2.71
      FTZ Fault Zone 25 2.75
      GWK Greywacke 2,368 2.71
      MD Mafic Dyke 473 2.73
      MZD Monzodiorite 2 2.70
      RDA Rhyodacite Aphanitic Porphyry 499 2.64
      RDF Rhyodacite Fine-Grained Porphyry 315 2.67
      RDX Rhyodacite Coarse-Grained Porphyry 1,339 2.66
      RDXB Rhyodacite Coarse-Grained Blue Porphyry 520 2.63
      RDXL Rhyodacite Lath-Rich Porphyry 216 2.64
      SLT Siltstone 838 2.72
      SHL Shale 387 2.70
      All Rock Types   7,370 2.69

     

    Table 11-2: Specific Gravity Values by Grouped Rock Type

      Grouped Rock Type Individual Rock Types No of Samples Specific Gravity
      Intrusive Rocks RDA, RDX, RDXB, RDXL & RDF 2,889 2.65
      Greywacke GWK & CGL 2,377 2.71
      Shale SHL, SLT & ARG 1,497 2.70
      All Rock Types   6,763 2.68

    11.4

    Analytical and Test Laboratories

       

    The primary laboratory for all assaying has been ALS Chemex in Vancouver, BC. During the exploration programs, ALS Chemex held accreditations typical for the time, including, at various times, ISO9001:2000 and ISO 9002, and from 2005, ISO/IEC 17025 accreditation.

       

    Metallurical test facilities have included SGS-Lakefield Research, Hazen Research, and G&T Metallurgical Services (G&T), who are independent, recognized metallurgical testing laboratories. Work has also been performed by test facilities operated by Placer Dome and Barrick. Metallurgical test facilities are not typically accredited.

       
    11.5

    Sample Preparation and Analysis

       

    Most core samples from 2005 onward were crushed at the Donlin camp sample preparation facility and pulverized at the ALS Chemex Vancouver laboratory facility. Samples of 2006 core that were split in Anchorage were shipped to an ALS Chemex preparation laboratory for crushing and pulverizing.


         
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    The Donlin camp preparation laboratory is housed in a heated steel building. The facility was rebuilt before the 2007 drill campaign to improve process flow and to upgrade ventilation and dust control.

    Sample preparation procedures are as follows:

     
  •  
  • The entire bagged sample is dried in an oven heated to 90° to 95°C for 12 hours.

         
     
  •  
  • The sample and sample tag are placed into trays for processing.

       
  •  
  • Blank samples (one of three QA/QC control samples) are inserted into the sample stream.

       
  •  
  • The sample is crushed in a TM Terminator jaw crusher until the end product passes 70% minus 10 mesh (2 mm). Sieve analyses are performed daily to check crush quality, and the crusher jaws are adjusted as necessary. The crushers are cleaned with blank material four times per 12-hour shift and before a new hole is started. Cutting lists also specify special cleaning frequency when unusually sulphide-rich material is processed.

       
  •  
  • Crushed sample is then passed through a riffle splitter four to six times to obtain a nominal 9 oz (250 g) split. This subsample is put into a numbered pulp bag, and the remainder, or coarse reject, is put back into the original sample bag. The splitter and sample pans are cleaned with compressed air.

       
  •  
  • Two additional control samples—standard reference material (SRM) and a duplicate split of crushed sample—are inserted as specified on the cutting list prepared by the geologist. Two of each control sample type, including SRM, duplicates, and blanks, are included in every batch of 78. The blank is prepared by processing a sample from a bin of gravel-size crushed rock through the jaw crusher and riffle-splitting it to ~7 oz (200 g). When a duplicate is required, the crushed core sample is passed once through the riffle splitter, and each half is split repeatedly to obtain a ~7 oz (200 g) sample.

    Final sample preparation and chemical analysis at the ALS Chemex laboratory in Vancouver consisted of the following:

  •  
  • Splits of crushed core were reduced to rock flour or “pulp” (better than 85% passing minus 200 mesh or 75 µm) in a ring-and-puck grinding mill.

       
  •  
  • A 1 oz (30 g) subsample of the pulp was assayed by fire assay-atomic absorption spectroscopy (AAS). Before 2007, the primary gold assay method was Au-AA23, which had an analytical range of 0.005 to 10 g/t Au. The Au-AA25 gold assay method was initiated in 2007 and had an analytical range of 0.01 to 100 g/t Au. This switch was made to reduce the cost and time delay associated with re- assaying samples with values above the 10 g/t Au analytical limit.


         
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    December 2011
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

  •  
  • Samples that exceeded the analytical limit for a given method were re-assayed by fire-assay and gravimetric finish or “ore grade” fire-assay AAS. Significant drill hole assay intercepts for the 2005 through 2007 programs, as tabulated in Appendix B3, were based on a 1 g/t Au cut-off grade with up to 13.12 ft (4 m) internal dilution and a minimum width of 9.8 ft (3 m). Chemex determined the sulphur content of each sample according to the Leco method. The Leco method was also used to analyze samples flagged for acid base accounting (ABA) for carbon content as well as to determine neutralization potential (NP) and acid potential (AP) according to the industry-standard Chemex ABA procedure.

       
  •  
  • Most trace and major element data for drill holes located within the resource model boundary were acquired prior to the 2005 program by various labs using industry-standard acid digestions followed by atomic absorption (AA) or inductively coupled plasma (ICP) instrumental determinations. Subsequent multi-element trace analyses were performed at ALS (Chemex) using aqua regia or four-acid digestions followed by ICP ± mass spectrometry.


    On occasion, if the ALS Chemex Vancouver laboratory was unable to complete the preparation, another ALS Chemex laboratory could be used.

       
    11.6

    Quality Assurance and Quality Control

       
    11.6.1

    1995–2002 QA/QC Protocol

       

    Placer Dome initiated the first QA/QC program during the 1995 drilling campaign. Coarse reject duplicate splits from 10% of the drill hole samples were submitted to an outside lab (Bondar Clegg). Standard reference material (SRM) assay standards and blanks were added in 1996, and an outside lab (Chemex) performed check assays, presumably of coarse reject duplicates. Check assays by a secondary assay lab were apparently discontinued after 1996. A more structured assay QA/QC program, consisting of SRMs, blanks, and duplicates inserted in rotation every 50 ft (15 m) down-hole, was initiated in 1997. This protocol evolved to random and blind insertion of SRMs, blanks, and coarse reject duplicates through the 2002 NovaGold program.

       

    From 1996 to 2002, SRMs and coarse-reject duplicates were inserted at an average rate of one per 24 samples, and blanks were inserted at an average rate of one per 25 samples. Almost all samples associated with SRM and blank control samples that returned values beyond acceptable tolerance limits were re-assayed until the control sample results were either acceptable or validated by duplication.


         
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    DONLIN GOLD PROJECT
    ALASKA, USA
    NI 43-101 TECHNICAL REPORT
    ON SECOND UPDATED FEASIBILITY STUDY

    11.6.2  

    2005–2006 QA/QC Protocol

       

    No resource delineation drilling was conducted in 2003 and 2004. Placer Dome implemented a slightly modified QA/QC protocol in 2005, which Barrick continued in 2006. Three QA/QC samples, consisting of one blank, one coarse reject duplicate, and one SRM, were randomly inserted into every block of 20 sample numbers. Thus, every block of 20 sample numbers contained 17 drill hole samples and 3 QA/QC control samples.

       
    11.6.3

    2007–2010 QA/QC Protocol

       

    The batch size submitted to ALS Chemex was increased from 20 samples to 78 in 2007. To avoid sample mixing with products from other sources in the fusion process, the ALS Chemex protocol was based on a fusion batch size of 84 samples, where the lab added six internal control samples, leaving space for 78 client samples in a given batch.

       

    Each batch of 78 samples shipped to ALS Chemex for sample preparation and analysis contained 9 control samples (12%) consisting of 3 each of standards, blanks, and crushed duplicates. Spacing of the SRMs within the batch was left to the judgement of the geologist. Up to 5% field duplicates (remaining half split of core) were added to the sample batch at the discretion of the geologists for geologic reasons.

       
    11.6.4

    Standard Reference Materials

       

    There is no information available on SRMs used prior to the 2002 drilling campaign. Two standards, (Std-C and Std-D), were used during the 2002 drill campaign.

       

    Standard reference materials remaining from the 2002 campaign were used at the beginning of the 2005 season. Additional reference material was purchased from Analytical Solutions (OREAS 6Pb and OREAS 7Pb) and CDN Laboratories (CDN-GS- 3) when these SRMs were depleted. After the 2005 season, two additional SRMs (Std-G and Std-H) were created from Donlin coarse reject material. These two new standards and CDN-GS-3 were used during the 2006 season.

       

    Nine new “matrix matched” SRMs of varying gold grade were added in early 2007, and the older standards were eventually phased out. The new SRMs were created from coarse reject samples from throughout the deposit. Composites of this material were pulverized and homogenized at CDN Laboratories in Vancouver, BC. A Barrick geochemist certified the 2007 SRMs after industry-accepted round-robin assay and statistical analyses.


         
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    The final SRMs included four each from unoxidized igneous and sedimentary host rocks and one oxidized igneous rock SRM.

       
    11.6.5  

    Blank Materials

       

    Washed river gravel produced by Anchorage Sand and Gravel was used for blanks through early 2006 and then replaced by granite landscape chips purchased from Lowe’s in Anchorage for all subsequent drill campaigns.

       
    11.7

    Databases

       

    The work completed by Placer Dome and predecessors before 2001 was collected and compiled into a main Microsoft Access database. NovaGold compiled the Placer Dome database into an updated Access database and added information from work completed in 2001 and 2002.

       

    Placer Dome contracted ioDigital to convert the Access database to an MS SQL Server database in early 2005 using an acQuire Technology Solutions data model (acQuire). Data obtained after the conversion were imported directly into the acQuire database.

       

    Barrick subsequently used acQuire software to capture 2006 and 2007 drill hole data, which were stored in MS SQL Server. Geologic logs, collar, and down-hole survey data were entered at the Donlin camp using acQuire data entry objects. Assay data were imported directly from electronic files provided by the laboratories. The master Donlin database was moved from the Donlin camp to the Anchorage office mid-year 2006. Assay data were imported directly into the master database in Anchorage for the rest of 2006 and through 2007. Geologic and sample data were entered into the Donlin camp acQuire database and merged into the master database as needed. The acQuire database was converted from the standard acQuire data model to the more robust acQuire “Corp” data model in early 2007.

       

    These database procedures were continued for all subsequent programs.


         
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    11.8

    Sample Security

       

    For all drill programs following the initial involvement of Placer Dome in the Project, core samples are transported from the field and brought to the yard adjacent to the geology office and logging tents at the end of each drill shift.

       

    Core storage is secure because Donlin is a remote camp and access is strictly controlled.

       

    Unauthorized camp personnel have generally been excluded from the core cutting and sample preparation building, but strict access procedures were initiated following a Barrick audit in mid-2006.

       

    Assay splits of prepared core, along with the control samples, are packed in a shipping bag, secured with a numbered security seal, and sealed in boxes for shipment. The coarse rejects and remaining split core are returned to a storage yard south of the airstrip for long-term storage.

       

    The sample shipment procedure is as follows:


  •  
  • Boxed assay splits are flown from the Donlin camp to Aniak airport via Vanderpool Flying Service.

       
  •  
  • Samples are shipped from Aniak via Frontier Flying Service to the ALS Chemex laboratory facility in Fairbanks, Alaska. All sample shipments are accompanied by a Frontier Flying Service waybill. This allows each sample to be tracked from camp to ALS Chemex.

    The samples are logged into the ALS Chemex data system in Fairbanks before shipment to the ALS Chemex Vancouver (or other ALS Chemex facility), where they are pulverized and assayed. The Fairbanks laboratory returns a custody form that reports on the condition of security seals.

    The Anchorage logging and splitting facility was housed in a secure, dedicated, warehouse/office building. Visitor access to the facility was strictly controlled by AES, the facility manager. Advance approval by the Donlin Project Manager was required for any outside visitation for tours or purposes other than daily delivery or pick-up.

    Whole core shipped from camp to the facility was transported by Lynden Air Cargo. Lynden waybills and Barrick custody forms were used to track samples from camp to Lynden’s Anchorage airport facility and from there by Lynden trucks to the Anchorage logging facility. Similar protocols were followed for split core samples shipped from camp to the ALS Chemex Fairbanks laboratory. Bagged split core samples were tied into shipping bags and loaded into palletized supersacks closed with numbered security seals and shipped on Lynden trucks to ALS Chemex in Fairbanks. Waybills aided tracking within the Lynden transport system, and ALS Chemex reported on the condition of security seals in the same manner as shipments from camp.

         
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    11.9

    Comments on Section 11

       

    Sample collection, preparation, analysis and security for all Placer Dome, NovaGold, Barrick, DCJV, and Donlin Gold core drill programs are in line with industry-standard methods for gold deposits:


  •  
  • Drill programs included insertion of blank, duplicate and standard reference material samples

       
  •  
  • QA/QC program results do not indicate any problems with the analytical programs (refer to discussion in Section 12)

       
  •  
  • Data is subject to validation, which includes checks on surveys, collar co-ordinates, lithology data, and assay data. The checks are appropriate, and consistent with industry standards (refer to discussion in Section 12)

       
  •  
  • Independent data audits have been conducted, and indicate that the sample collection and database entry procedures are acceptable

         
     
  •  
  • All core has been catalogued and stored in designated areas

    The QPs are of the opinion that the quality of the gold analytical data from the Placer Dome, NovaGold, Barrick, DCJV, and Donlin Gold drill programs are sufficiently reliable to support Mineral Resource and Mineral reserve estimation without limitations on Mineral Resource confidence categories.

         
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    12.0

    DATA VERIFICATION

       
    12.1.1  

    AMEC (2002)

       

    As a test of data integrity, the data used to estimate the January 2002 Donlin Creek Mineral Resources reported in the February and March 2002 Technical Reports (Juras, 2002, and Juras and Hodgson, 2002) were validated. AMEC concluded that the assay and survey database used for the Donlin Mineral Resource estimation at that time was sufficiently free of error to be adequate for support of Mineral Resource estimation.

       
    12.1.2

    NovaGold (2005)

       

    NovaGold conducted a 100% check of 2005 drill hole gold assays within the 2005 resource area against electronic assay certificates. An error rate of less than 1.5% was noted.

       

    NovaGold also checked the collar and down-hole survey data. Electronic down-hole survey files were read for the drill holes and compared to those stored in the resource database.

       

    As a result of the verification, NovaGold in 2005 considered that the database at the time was adequate to support Mineral Resource estimation.

       
    12.1.3

    NovaGold (2008)

       

    In support of preparation of a technical report on the Project in early 2008, NovaGold undertook a data review of the 2006 and 2007 drilling. Data reviewed included:


  •  
  • Drill collar locations: The Ashtech output files and geologic logs were compared to 5 percent of the electronic collar surveys. There was one unexplained 20-cm discrepancy between the elevation file and the database. Although a number of errors were noted in the geological survey tables, and attributed to likely use of proposed, rather than final, collar co-ordinates, NovaGold concluded that the collar surveys from the Ashtech data files were sufficiently error free to be used for support of Mineral Resource estimation.

       
  •  
  • Down hole surveys: 10% of the drill holes were checked and an error rate of 4.4% was measured. NovaGold recommended that DCLLC review their down-hole survey transcription protocols and complete a 100% check of the down-hole survey database. Despite the high error rate, the magnitude of the errors was small; therefore, in NovaGold’s opinion the impact on the estimation of grade was likely to be minimal


         
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  •  
  • Assay data: For 2006 drilling, 70% of the assays were compared and an acceptable discrepancy rate of 0.4% was measured. For 2007, 99% of the assays were compared to the electronic assay certificates and a discrepancy rate of 1% was measured. NovaGold recommended that the source of the discrepancies be identified. NovaGold believed that the assay database was sufficiently error free to be used for Mineral Resource estimation.


    12.2

    AMEC (2011)

       

    AMEC reviewed the March 2008 through December 2010 QAQC information for the drill hole data used to construct the DC9 model from with the following results:


  •  
  • Certified Reference Materials Results: AMEC reviewed the results for 691 CRMs from 2,078 samples and calculated that the relative bias for all CRMs is within the acceptable limits of G5% relative bias (mean/best value)/1).

       
  •  
  • Blank Material Results: AMEC reviewed the results for 694 blank samples submitted for analysis. There were 4 samples which returned higher than allowed gold assays which may be due in part to mislabelled samples. This is an infrequent occurrence and will not affect project assay results, thus AMEC considers that there is no significant risk to the resource estimate.

       
  •  
  • Coarse Reject Duplicate Results: The absolute value of relative differences (AVRD = |pair difference|/pair mean) of the crush duplicate pairs that have pair means greater than 1 ppm Au show that 90% of these pairs agree within 15 percent; this indicates acceptable precision is achieved thus the current sample preparation procedure (crushing and pulverizing).


    12.3

    Comments on Section 12

       

    AMEC considers that a reasonable level of verification has been completed during the 2011 data review, and that, between this review, and reviews completed by NovaGold in 2007 and 2005, and AMEC in 2002, no material issues would have been left unidentified from the verification programs undertaken.

       

    The QPs, who rely upon this work, have reviewed the appropriate reports, and are of the opinion that the data verification programs undertaken on the data collected from the Project adequately reflect deposit dimensions, true widths of mineralization, and the style of the deposits, and adequately support the geological interpretations, the analytical and database quality.


         
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    No problems with the database, sampling protocols, flowsheets, check analysis program, or data storage were identified that were sufficient to preclude the use of the database for estimation purposes.

    Drill data are typically verified prior to Mineral Resource estimation by comparing data in the Project database to data in original sources. For most of the data, the original sources are electronic data files; therefore, the majority of the comparisons were performed using software tools.

    AMEC recommends that Donlin Gold performs a comparison between the trench samples and core/rotary samples to determine if there is any bias that may affect the resource estimation, as trench sampling is used to support the estimate.

         
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    13.0

    MINERAL PROCESSING AND METALLURGICAL TESTING

       
    13.1

    Metallurgical Testwork

       
    13.1.1  

    Domains

       

    Lithological and Geological Domains

       

    Mineralization at Donlin is temporally and spatially associated with rhyodacite dikes and sills intruded into sediments of the mid-Cretaceous Kuskokwim Group. The sediments are predominantly inter-bedded greywacke and shale. Six intrusive phases and two major sedimentary phases have been recognized. The lithological domains and their respective percentages within the Donlin deposits are summarized in Table 13-1. The listed intrusive lithologies are ranked in order of relative age.

       

    Due to poor continuity and minor occurrence within the deposit, the Mafic Dike (MD) intrusive lithology is not separately defined within the geological model, and therefore its relative content within the ore is not accurately defined. Based upon lengths of ore intercepts within the exploration drilling, the content of MD in the ore is estimated to be 0.2% of the ore tonnes.

       

    Two main pits have been identified within the current Donlin deposits, Lewis and ACMA, each subdivided into spatial zones typically separated by faults or other key geological structures.

       

    The Lewis pit is dike dominated and contains more sedimentary ore than ACMA. The ACMA pit is sill dominated and typically contains less sedimentary ore and more intrusive ore than Lewis. With regard to the intrusives, the Lewis pit is dominated by RDX and RDXB compared to the ACMA pit area, which is dominated by the RDA, RDX, and RDXL intrusive lithologies.

       

    For the purpose of defining discreet geological domains within the deposit, the Donlin geological team is using a multi-type separation of the deposit based on a combination of both lithology and spatial location. This is undertaken by independently recognizing the main pit spatial domains (Lewis, ACMA, Akivik, 400, Aurora, and Vortex) for the intrusives, but separating out the sedimentary lithologies from every spatial area, and differentiating into the two sedimentary global domains, greywacke (GWK) and shale (SHL). However, for metallurgical interpretation all different types of domain categorization are considered, used, and applied as appropriate. Table 13-2 provides an estimate of the ore tonnes and gold ounces distribution on the basis of geological domain break-up.


         
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    Table 13-1: Intrusive and Sedimentary Lithologies of the Donlin Gold Project

          Relative ~% Au Ounces LOM ~% Ore Tonnes LOM
      Lithology Description Abbrev. Intrusive Age (%) (%)
      Intrusive Phases        
         Blue Porphyry RDXB Youngest 13.9 15.2
         Aphanitic Flow-banded Porphyry RDA   22.8 22.5
         Lathe-rich Porphyry RDXL   7.1 7.2
         Crystalline (Crowded) Porphyry RDX   28.7 27.6
         Fine-grained Porphyry RDF   0.4 1.4
         Mafic Dikes MD Oldest 1.5* 0.2*
      Sedimentary Phases        
         Greywacke GWK   22.7 20.4
         Shale SHL   4.3 5.4
      Total     100.0 100.0

    Note: MD component percentage is an estimate only – and is therefore not included within the total composition

     

    Table 13-2: Major Geological Domains of the Donlin Gold Project

     
        ~% Ounces LOM ~% Ore Tonnes LOM  
      Description (%) (%)  
      Lewis Intrusive 23.1 26.7  
      AMCA Intrusive 21.5 16.8  
      Aurora Intrusive 5.2 4.4  
      Akivik Intrusive 4.0 4.8  
      Vortex Intrusive 10.0 12.0  
      400 Intrusive 6.3 5.9  
      Oxide 4.0 4.0  
      Greywacke (GWK) 20.5 20.0  
      Shale (SHL) 5.4 5.4  
      Total 100.0 100.0  

    Figure 13-1 is a graphical representation of the geological domains in the Donlin deposits.

    Gold Mineralization

    There are four primary vein types in the deposit area with variable mineralization as summarized in Table 13-3.

    Sulphide Mineralization

    A number of mineralogical investigations were undertaken in 2004 to 2007, including work carried out by Amtel, Hazen Research, G&T Metallurgy, and Barrick Technology Centre (BTC). Typical sulphide mineralization presence based on these investigations is summarized in Table 13-4.

         
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    Figure 13-1: Donlin Gold Project Geological Domains

     

     

    Table 13-3: Vein Types

      Vein Type Dominant Mineralogy Grade ~(g/t) Average Orientation Relative Age
      V1 Sulphide 2.7 020/67 Oldest
      V2 Qtz-Sulphide 3.9 022/68  
      V3 NA, St, Re 7.4 028/72  
      V4 Carbonate 0.6 028/65 Youngest

     

    Table 13-4: Typical Sulphide and Metals Mineralization in the Donlin Ores

      Typical Occurrence Sulphides Chemical Formula
      Major to Minor Pyrite FeS2
        Marcasite FeS2
        Arsenopyrite FeAsS
      Minor to Trace Native Arsenic As
        Realgar AsS
        Stibnite Sb2S3
      Trace Chalcopyrite CuFeS2
        Sphalerite ZnS
        Tetrahedrite (Cu,Fe)12Sb4S13
        Galena PbS
        Pyrrhotite FeS
        Molybdenite MoS2
        Bornite Cu5FeS4
        Covellite CuS
        Mercury Sulphide HgS in pyrite
        Native Copper Cu

         
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    Pyrite is the dominantly occurring sulphide within the deposit. Marcasite is also present, at an approximate ratio of 1:7 as marcasite to pyrite. Arsenopyrite is the main carrier of arsenic within the deposit. Stibnite is the main carrier of antimony.

       
    13.1.2  

    Gold Deportment

       

    The primary host mineral of gold within the Donlin ores is arsenopyrite, which carries 80% to 90% of the gold as “solid solution” gold atomically distributed within the arsenopyrite crystal. Fine arsenopyrite is typically highest in gold grade at 500 to 1,500 g/t (mineral gold content), compared to coarse arsenopyrite at ~100 to 500 g/t. A proportion of the arsenopyrite (particularly coarser-grained arsenopyrite) has been shown to have a preferential concentration of gold around the rim of the crystal.

       

    Pyrite is the second most important gold carrier, hosting 10% to 20% of the contained gold, also in “solid solution” form. Typical gold grades of pyrite are 1 g/t to 50 g/t. Similarly, marcasite is also a gold carrier.

       

    Free gold is a minor source of gold (rare occurrence) within the deposit at less than 1% of the contained gold being free liberated particles less than 20 µm in diameter. Native arsenic is an insignificant carrier of gold in terms of both grade and quantity.

       
    13.1.3

    Mercury, Chlorine, Carbonates and Organic Carbon Deportment

       

    The Donlin ore hosts mercury at a grade of 1 g/t to 3 g/t. For occupational health and environmental considerations, mercury is an important species considered for the process plant design.

       

    Detailed deportment of mercury was undertaken by Amtel in 2007, which indicated that, based upon the concentrate samples tested, pyrite is the principal carrier (66%) of mercury as HgS in “solid solution” within the sulphide mineral matrix, followed by the sulphide marcasite (18%). The Amtel study also indicated that fine-grained pyrite, marcasite, and stibnite are relatively enriched in mercury content. Arsenopyrite has relatively low mercury content and is an insignificant carrier. No specific mercury minerals have been found or identified in any of the samples examined by Amtel in the deportment work completed.

       

    The Donlin ore contains chloride with an average concentration of approximately 22 ppm. Detailed testing of chloride deportment carried out by Amtel in 2007 indicated that the principal carrier of chloride in flotation concentrate was muscovite (white mica) as KAl2(Si3Al)O10(OH,F,Cl)2, followed by hydroxylapatite, Ca5(PO4)3(OH,F,Cl).


         
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    The form of carbonates within the ores at Donlin is an important consideration - carbonate within the flotation tails stream is used to neutralize the acidic liquor coming from the autoclave. The three forms of carbonate identified by Amtel in 2007 were calcite (CaCO3), siderite (FeCO3), and dolomite (CaMg(CO3)2). The predominant carbonate species remaining in flotation tails was dolomite, followed by ankerite. Calcite was only identified in two test samples and siderite in only one test sample. For the samples tested, the Aurora geological domain consistently contains calcite as the dominate carbonate.

       

    The Donlin ores host organic carbon in both the intrusive and sedimentary lithologies. The sedimentary lithologies are relatively abundant in organic carbon. The RDF and RDXB intrusive lithologies are characteristically high in graphitic carbon content.

       
    13.1.4  

    Samples

       

    The samples used for the metallurgical testwork during 2006 to 2007 were obtained from a variety of sources. Samples for metallurgical testwork were selected to represent the various geological lithologies and geological domains.

       
    13.1.5

    Comminution

       

    Grinding testwork for the Project has been conducted periodically since Project inception. The initial grinding testwork was undertaken in the 1990s and very little specific information is available on this work. Subsequently, in 2002-2003 additional work was completed at Hazen managed by NovaGold.

       

    Placer Dome initiated some further work by SGS Lakefield in 2004, which tested ACMA material. During this program, JK and Bond testwork were completed in conjunction with testing the applicability of high pressure grinding rolls (HPGRs) to the deposit. During 2006 Barrick initiated three major testwork programs at SGS Lakefield:


     
  •  
  • SGS Lakefield 2006 HQ half-core testwork

         
     
  •  
  • SGS Lakefield 2006 PQ whole-core testwork

         
     
  •  
  • SGS Lakefield 2007 HQ half-core testwork.

    The work completed in 2007 for the feasibility study indicated that grinding testwork should preferably be carried out on freshly drilled core that has been protected from the Alaskan weather. It is apparent that the physical hardness properties of the drill core are affected upon exposure to the Alaskan environment while in storage (i.e., reduced in competency, believed to be through a cryogenic weathering effect) and therefore could lead to biased low hardness test results.

         
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    Consequently, the test results from the core recovered from fresh exploration drilling undertaken in 2006 were used preferentially as the basis for the FSU design of the grinding circuit.

    Testwork Period 2002–2003

    A summary of results of the testwork undertaken in 20022003 is shown in Table 13-5. The testwork summary indicates that the material tested at that time was moderately hard. These results align with more recent testwork.

    Testwork Period 2004–2005

    Placer Dome initiated a testing program at SGS Lakefield in 2004. This work was performed on two large samples from the Donlin deposits with the objective of comparing the power efficiency of using high-pressure grinding rolls (HPGR) as opposed to semi-autogenous grinding to prepare the ore ahead of ball milling. The results of the JK and Bond work are summarized in Table 13-6.

    The following key items were identified from the testwork:

  •  
  • The sedimentary sample was described as moderately hard with respect to resistance to impact, as measured by the impact work index (CWI) and drop- weight test (A x b).

       
  •  
  • The intrusive sample measured as hard in terms of low-energy impact and medium in the drop-weight test.

       
  •  
  • The sedimentary sample was moderately hard with respect to resistance to abrasion breakage (ta), while the intrusive sample was hard.

       
  •  
  • Both samples can be categorized as medium in terms of Bond rod mill (RWI) and ball mill work indices (BWI).
       
  •  
  • Both samples were only mildly abrasive (Ai).


         
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    Table 13-5: Grinding Testwork Results from Hazen Research

     
            BWI  
      Pit Locale Ore Type (kWh/t)  
      ACMA - Intrusive 14.3  
        - Sediment 13.0  
      Lewis Rochelieu Intrusive 14.1  
        Rochelieu Sediment 13.3  
      Lewis North Intrusive 13.9  
        North Sediment 13.1  
      Lewis South Intrusive 15.1  
        South Sediment 12.1  

     

    Table 13-6: Summary of Grindability Testing

        CWI     Ore Density RWI BWI Ai
      Sample Composite (kWh/t) A x b ta (g/cm3 ) (kWh/t) (kWh/t) (g)
      ACMA Sedimentary   9.9 38.7 0.39 2.76 14.7 14.0 0.205
      ACMA Intrusive 11.3 52.8 0.31 2.69 13.5 14.7 0.181

  •  
  • The two samples were also submitted to a series of bench-scale HPGR tests. The tests showed that the HPGR successively reduced the ½ material feed to ball mill feed size with an energy input of 2.07 and 1.94 kWh/t for the sedimentary and intrusive samples, respectively. The BWIs of the resultant samples were tested at 75 µm and produced values of 12.5 and 13.3 kWh/t, respectively, indicating that a reduction (in the BWI) was attributable to the HPGR processing.

    SGS-Lakefield recommended further HPGR work at Polysius with the REGRO unit accompanied by ATWAL abrasion testing. The testwork by Polysius judged that a specific energy input in a range of 1.5 to 2.0 kWh/t was the optimum for HPGR comminution of the provided test samples. A specific throughput rate of approximately 250 ts/hm3 was achieved with the grinding force selected.

    The abrasion testwork indicated an ATWAL wear index of 8.3 g/t for sedimentary material and 24.8 g/t for intrusive material. This testwork was done at the standard feed moisture of 1%. This wear is medium abrasive compared with Polysius’ database at the time.

    SGS Lakefield 2006 HQ Core Testwork

    In 2006, SGS Mineral Services (SGS) conducted an extensive test program (Appendix D3-7) to determine grinding parameters for the Donlin ores. The samples (Appendix D2-8) used were HQ drill core taken from the 1999 and 2002 drilling campaigns, which had been stored at the exploration site. Parameters were obtained from the following tests:

         
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  •  
  • Minnovex SAG power index (SPI), crusher index (Ci), and modified Bond ball mill work index (Modified Bond test)

         
     
  •  
  • SMC drop-weight index test (DWI)

       
  •  
  • Bond low-energy impact (CWI), rod mill work index (RWI), ball mill work index (BWI), and abrasion index (Ai)

         
     
  •  
  • High-pressure grinding roll energy test.

    SGS Lakefield 2006 Whole PQ Core Testwork

    HQ drill core was too small in diameter for full drop weight tests for JKSimMet modelling. Larger diameter PQ holes were drilled in a 2006 drilling campaign, targeting bulk mineralized areas of the deposit, and covering the full range of lithologies. The samples from these drill holes were processed to develop JK grinding parameters, in addition to conventional Bond ball mill and rod mill work index numbers. This information was important for use in checking the grinding parameters developed from the HQ testwork. It was seen by comparing work index properties within lithologies from the PQ core results and the 2006 HQ core results that the freshly drilled PQ core was consistently harder than the HQ core samples drilled in 1999 and 2001. At that point in the study, it was recommended that a second set of HQ samples be selected from freshly drilled core that became available during 2006.

    SGS Lakefield 2007 HQ Core Testwork

    With the availability of additional fresh HQ drill core from the 2006 exploration program, a second phase of variability testing was initiated in early 2007. A total of 149 additional samples were tested.

    In parallel with this testwork AMEC undertook grinding circuit design trade-off studies in 2006, which indicated that a semi autogenous–ball mill-crushing circuit (SABC) circuit design was the preferred option. Therefore, the test program was designed to maximize generation of hardness properties relating to the required parameters for the SABC circuit. The parameters tested in the program were then restricted to:

     
  •  
  • Minnovex SAG power index (SPI)

         
     
  •  
  • Minnovex crusher index (Ci)

         
     
  •  
  • Minnovex modified Bond ball mill work index


         
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  •  
  • Rod mill work index (RWI)

         
     
  •  
  • Ball mill work index (BWI).

    Once again, it became apparent that within lithology, the results from the 2007 test program indicated consistently harder comminution properties than the 2006 test program.

    Therefore, it was recommended that comminution data obtained from core predating 2006 be normalized to fresh rock conditions for design calculations when it is known that the core has been exposed for an extended period to Alaskan climate. Such normalization was actually carried out on the 1999 and 2001 HQ core data.

    Comparison of 2006 Drilled and 1999-2001 Drilled Half HQ Core

    Table 13-7 directly compares the average results by lithology between the two crushing and grinding hardness data from the 2006 (core drilled 1999/2001) and 2007 (core drilled 2006) testwork programs.

    It should be noted that for the test JKTech parameters Axb and Crushing Index, ore hardness is inversely proportional to the test result (i.e., lower numerical result is physically harder to process, requiring more power).

    Given that the 2007 program used the freshly drilled 2006 HQ core, these results were regarded as being a more reliable predictor of the hardness of the deposit. However, the 2006 test results were preserved to provide more test samples to augment the variance analysis and subsequent population of the geological model. Minnovex adjusted the 2006 variability results to match the hardness distribution of the later (harder) test results.

    Table 13-8 is a summary of the adjustments for each lithology, derived by splitting the data into separate lithological groups, comparing the frequency distribution of the 2006 and 2007 sample data, and then determining a simple adjustment to move the 2006 data to fit the form of the 2007 distribution.

    After the adjustment to each of the lithologies within the 2006 test data, the two 2006 and 2007 data sets (Ci, SPI, BWI) were combined and forwarded to AMEC for geostatistical review and modelling, to provide a block-by-block mill feed hardness schedule for the parameters of Ci, SPI, and BWI. With these parameters, it is possible to use a Minnovex CEET (comminution economic evaluation tool) grinding model to predict the milling capacity and power requirements for each ore block in the designed Donlin circuit.

         
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    Table 13-7: Comparison of Average Results from 2006 and 2007 Test Programs

        JKMRC, A x b SPI, Minutes Crushing Index  RWI, kWh/t BWI, kWh/t
      Lithology 2006 2007 2006 2007 2006 2007 2006 2007 2006 2007
      RDX 46.0 42.7 47.6 90.5 17.2 14.9 13.3 15.1 14.6 15.6
      RDA 41.0 41.1 53.8 90.1 17.3 14.6 14.0 15.0 12.4 13.8
      RDXL 54.2 48.5 44.6 72.3 22.5 21.3 13.6 14.0 13.6 14.3
      RDXB 50.0 49.4 53.7 91.9 18.8 15.3 13.7 15.5 14.6 16.4
      GWK 41.8 34.4 55.9 94.7 16.6 12.5 14.5 15.2 13.3 14.7
      SHL 50.2 42.2 48.4 58.1 21.7 16.0 13.4 16.7 13.5 14.4
      RDF 37.0 30.1 61.9 100.8 18.3 14.2 13.9 14.8 15.1 15.3
      Average 45.7 41.2 52.3 85.5 18.9 15.5 13.8 15.2 13.9 14.9

     

    Table 13-8: Adjustments Made to the 2006 Test Program Data

        Ci SPI BWI
      Lithology 2006 Correction 2006 Correction 2006 Correction
      RDA x/1.2 1.8*x+6 1.09*x
      RDX x/1.1 1.35*x+18 1.07*x
      RDXB x/1.1 1.1*x+33 1.17*x
      RDXL x 1.2*x-18 1.07*x
      RDF x/1.3 1.2*x+27 1.05*x
      GWK x/1.3 1.8*x+2 1.12*x
      SHL x/1.5 1.3*x 1.07*x
      MD x/1.8 1.8*x 1.09*x

    Note: where * = multiplication

    Effect of Grind Size on BWI

    The potential adoption of an MCF2 flowsheet would result in two different ball mill product size distributions. The products from the primary ball milling circuit would target P80 120 to 150 µm, and the secondary ball mill would target P80 50 µm. From previous work by others, it is known that in some circumstances the measured BWI of a test sample will vary according to the target P80 of the BWI test.

    To determine if this is the case for Donlin, and to quantify that effect, a number of additional BWI tests at different final product sizes were undertaken by SGS Lakefield. For the feasibility grinding circuit modelling and design, an overall adjustment model was applied to the BWI number used, based on these results.

    Using the blended pilot-plant feed sample, an additional set of BWI tests was undertaken at varying product sizes.

    From the results obtained, it was seen that as the target product P80 size increased in fineness, the measured BWI of the test sample increases. On the blended pilot-plant sample a significant increase in BWI occurs between P80 54 µm and 42 µm. Figure 13-2 is a graphical representation of the pilot-plant blend feed sample results.

         
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    Feasibility Comminution Circuit Selection

    In 2006, Barrick contracted Orway Mineral Consultants (OMC), Australia, and SGS Lakefield (Minnovex), Canada, both of which have specialist groups in grinding circuit modelling and design, to perform appraisals of the various comminution options. The consultants were asked to examine the testwork information and provide capital and operating cost alternatives for four options:

     
  •  
  • Option 1 – autogenous milling with ball milling and crushing of pebble reject (ABC)

         
     
  •  
  • Option 2 – semi-autogenous milling with ball milling and pebble crushing (SABC)

         
     
  •  
  • Option 3 – coarse crushing followed by HPGR with ball milling

         
     
  •  
  • Option 4 – fine crushing followed by ball milling.

    Both consultants created grinding models based on their internal databases and calculated capital and operating costs.

    AMEC then performed a trade-off study based on the OMC recommendations to investigate the economics of various options as well as non-economic factors. AMEC also performed a check analysis on the various models to ensure that these were reasonable in their execution. It should also be noted that since the OMC and SGS work was done, it has become apparent that the samples from the 1999 and 2002 drill campaigns may have weathered sufficiently to affect the primary mill grindability testing. Consequently, AMEC also performed a separate analysis using only the results from the samples obtained in the 2006 drilling campaign.

    The SABC circuit, Option 2, was selected as the comminution circuit for Donlin, for several reasons:

     
  •  
  • Lowest capital cost

         
     
  •  
  • Ability to cope with the clay fraction in the ore

         
     
  •  
  • Ability to cope with the climatic conditions

         
     
  •  
  • General ease of operation and maintenance

         
     
  •  
  • Flexibility in throughput rates

         
     
  •  
  • Widely applied technology in the milling industry

         
     
  •  
  • Barrick’s extensive experience in SABC circuit application.


         
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    Figure 13-2: Test P80 vs. Measured BWI Results on Blend Composite Sample

    Although the lower operating costs ascribed to Options 3 and 4 indicate a long-term financial advantage, these options are considered unwarranted in use because of uncertainty with regard to their ability to achieve the operating costs and stated availabilities. In particular, the natural high clay content of the Donlin ores would hinder the performance of these two types of circuits, particularly as ore moisture increases.

    Geostatistical Assessment

    To better define the hardness characteristics of the scheduled mill feed, the compiled comminution data set for Ci, SPI, and BWI was provided to AMEC (U.S.) to build a metallurgical model, which served as the basis for the geostatistical assessment. The purpose of the metallurgical model was to develop the supportable relationships that might exist between the ore sample hardness, rock lithology, spatial location, ore grades (Au, S, As, Mg, Sb), and RQD (rock quality designation), and to use these relationships to populate the geological block model with the ore hardness properties Ci, SPI, and BWI. The key results of the geotechnical assessment are summarized below.

         
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    No correlations could be determined between these three grinding hardness properties, and therefore each parameter had to be assessed individually.

  •  
  • Variability of the crushing index data is quite high within and between lithologies, and is dominated by lithology as indicated in Table 13-9.

       
  •  
  • The SPI results are dominated by lithology, with the differences between lithologies being significant. Based on analysis of the variance of the data, four separate categories were defined and used to populate the block model. These are summarized in Table 13-10.

       
  •  
  • Lithology was determined to be the significant variable influencing ball work index (Table 13-11). Within a lithology, the variance in results was quite low, with the exception of the SHL and MD lithologies, due to a small number of test results.

    MCF2 Grinding Circuit Design Method and Capacity Definition

    During the first quarter of 2007, pilot flotation testing was carried out at SGS Lakefield to assess the potential of an mill chemical float, twice-style (MCF2-style) flowsheet to improve confidence in the overall gold recovery and economic value of the Project (Figure 13-3). The MCF2 flowsheet incorporates two separate stages of grinding and flotation. An economic evaluation was completed and it was decided to use the MCF2 flowsheet as the basis of the circuit design for the 2007 FS and FSU, utilizing the SABC configuration as the primary grinding step.

    In 2007, a throughput confirmation study carried out by the mining team, indicated that a Project optimum ore production rate was in the order of 49,600 stpd (45,000 t/d). In 2008, an updated resource model based on additional drilling was completed and the mill throughput was increased to 59,000 stpd (53,500 t/d).

    Mineralogical assessment, flotation bench-scale, and flotation pilot-scale testwork have indicated that the primary rougher feed particle 80% passing size distribution (P80) should be in the range 120 to150 µm, and that the secondary rougher feed P80 should be in the range 50 to 60 µm, to achieve high flotation gold recoveries. These assumptions were applied to the grinding circuit modelling and design.

    Based on a common hardness data set and the conventional SABC circuit, three different grinding circuit modelling techniques were used (JKSimMet, CEET, and a power-based model referred to as DJB) to “calibrate” a CEET model to represent an agreed design model set-up.

         
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    Table 13-9: Orebody Estimation of Crushing Index (Ci)

      Domain Ci Mean Std. Dev. Sample Count
      RDXL 21.9 8.2   33
      Non-RDXL West (<541,000) 14.2 6.7 158
      Non-RDXL East (>541,000) 14.8 6.9   81
      All 15.3 7.4 272

      Table 13-10: Orebody Estimation of SAG Power Index (SPI)
      Domain SPI Mean Std. Dev. Sample Count
      SHL 58 17     9
      RDXL 71 14   33
      RDX 85 22   79
      Remaining Lithologies 95 21 152
      All 88 23 273

      Table 13-11: Orebody Estimation of Bond Ball Work Index (BWI)  
      Domain BWI Mean Std. Dev. Sample Count  
      RDX + RDF + MD 15.5 1.3 107  
      GWK + SHL 14.6 1.5   48  
      RDA 13.8 1.3   46  
      RDXL 14.4 1.1   33  
      RDXB 16.4 1.2   39  
      All 15.0 1.5 273  

         
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    Figure 13-3: Illustration of MCF2 Generic Flowsheet


    At the 2008 target throughput of 59,000 stpd (53,500 t/d), the design parameters collated in the steps above suggested that large mills and power input would be required. The mill size was approaching the production capacity limitations of the largest existing and operating SAG mill (40 ft diameter) and ball mill (18 MW).

    Based on the adjusted CEET grinding model and using the “calibration” dataset, a single 38 ft SAG mill (20 MW) followed by a single 18 MW ball mill could process ~59,000 stpd (53,500 t/d) to produce a primary flotation feed P80 of 120 to 150 µm. This could then be followed by another 18 MW ball mill to produce the required secondary rougher flotation feed of P80 of 50 µm.

    In the design moving forward, therefore, a smaller 38 ft diameter SAG mill with a corresponding motor size of 20 MW was adopted, while still retaining two 18 MW ball mills.

    Plant Ramp-Up

    The Project ramp-up schedule was defined on the basis of operating data available to Barrick Gold Corporation from other sites. The sites used for the comparison were selected on the basis of being large SABC circuits with some additional downstream process sections, and where reasonable ramp-up data were available.

    This ramp-up schedule provides an additional constraint to the CEET model for capacity scheduling. Figure 13-4 and Figure 13-5 show the assumed feasibility ramp-up schedules for the plant utilization and throughput, as a percentage of design.

    Downstream Autoclave Productivity Limits

    Given that the milling circuit provides feed to flotation, and that the flotation concentrate generated is then processed through a site-based pressure oxidation unit, CEET must consider downstream constraints to production. The pressure oxidation facility has a design capacity of 714 stpd (648 t/d) of sulphur.

    Therefore, for example, during periods of high-grade sulphur content in the mill feed, the pressure oxidation units limit the upstream (grinding circuit) plant throughput. Sixty days of downtime for complete autoclave re-lining have been incorporated into the LOM schedule every six to seven years.

         
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    Figure 13-4: Plant Utilization Ramp-up Schedule for Donlin Feasibility Compared to Other Available Commissioned Sites

    Figure 13-5: Plant Throughput Ramp-up Schedule for Donlin Feasibility Compared to Other Available Sites

         
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    Mill Feed Schedule Definition

    With the CEET model set up with the recommended grinding equipment sizes, the sulphur productivity limits, and the ramp-up schedule, it is then possible to model the mill feed schedule from mining on a block-by-block basis to determine the operating capacity of the milling circuit to process the ore on a period-by-period basis, taking into account all the plant constraints.

         
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    Continuous Improvement Assumptions

       

    Nominal continuous improvement targets have been defined over the life of mine to account for anticipated productivity gains earned through operating experience and promoting a continuous improvement culture at the mine. These increases are not significant in size and are considered to be within the existing plant design limitations. The fixed constraints of the designed and installed grinding circuit power and oxygen plant capacity are not exceeded within any given period regardless of these improvements.

       

    Table 13-12 summarizes the productivity improvement assumptions used for the scheduling of FSU2.

       
    13.1.6

    Flotation

       

    Introduction

       

    Extensive bench and pilot flotation testwork has been carried out on samples of the Donlin ores from 1995 through to mid-2007.

       

    The objective for the flotation circuit within the flowsheet is to provide high gold recovery (+90%) to a high-grade sulphur concentrate (greater than 6.5% total sulphur content) for pressure oxidation feed. Once the feed contains sufficient grade of sulphide sulphur as fuel to generate heat and achieve the required operating temperatures within the autoclave, there is little overall net benefit in increasing the sulphur grade of the concentrate any further. This is advantageous for Donlin because the mineralogy of the ores produces a gold recovery in flotation concentrate increases significantly as concentrate grade is reduced.


         
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      Table 13-12: Productivity Improvement Assumptions for the FSU
          Assumption
      Production Period Component (% of Design)
      Year 3 & Onwards Milling Circuit Capacity 100.0
      Year 6 & Onwards Milling Circuit Utilization 101.5
      Year 6 & Onwards Autoclave Sulphur Treatment Rate 101.5
      Year 6 & Onwards Autoclave Circuit Utilization 100.0

    For Donlin, a general target of 7% sulphur grade has been selected for the final flotation concentrate, noting that the autogenous grade is determined to be approximately 6.5% sulphur, varying slightly with changes in the solids/liquid ratio, the carbon, and arsenic grade of the autoclave feed; the gap between 6.5% autogenous grade and the target 7% grade is provided to allow for variability in feed and to minimize any requirement to add external heating to the autoclaves.

    AMEC Review of Testwork from 1995 to 2006

    The key conclusions and comments from the AMEC review of the 1995 to 2006 flotation testwork are as follows:

    Bench Testing

  •  
  • The testwork results indicated that producing a bulk concentrate was the optimum route to maximize gold recovery

       
  •  
  • A variety of reagent schemes were attempted. The best reagent system was using plant acid, copper sulphate, Xanthate, and dispersant

       
  •  
  • Nitrogen-based flotation technology was tested extensively throughout the Project but provided no overall benefit

       
  •  
  • Adequate retention time was found to be very important. To achieve maximum recovery, a long flotation retention time, of 114 minutes, was found to be necessary

       
  •  
  • The required particle grind size reflects the presence of the two different ore types in the feed. While the intrusive ores can tolerate coarser sizes in the range of 75 to 110 µm, the sedimentary ores perform best in the range of 60 to 80 µm for the conventional flotation flowsheet

       
  •  
  • Mass pull from the rougher and scavenger circuits was dictated by entrainment of clays during the long residence time of the flotation. With the use of dispersants, lower flotation feed pulp densities and cleaning, the overall mass pull to final concentrate could be decreased to approximately 15%

       
  •  
  • The process development testwork indicated that froth recovery was a critical factor. Froth recovery can be enhanced by the use of crowding cones in the flotation machine and through launder design

       
  •  
  • Because of the different flotation response of the ores, testwork was performed to assess the outcome of blending the two main ore types. Given adequate reagent dosages and residence times, it proved possible to produce high flotation recoveries with the life-of-mine test blends provided for the testwork


         
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    Flotation Pilot-Plant Testing

  •  
  • Extensive testwork was performed in 2004 and 2006 leading to the demonstration that a recovery of 91% to 92% on a LOM lithology blend was possible, confirming the performance of the conventional rougher / scavenger flowsheet under continuous operation.

    Mineralogy Summary

    In summary, the various mineralogy studies undertaken on many different flotation test samples by a number of different investigators present a relatively consistent themes concerning the aim to achieve high gold recovery (+90%) to concentrate:

     
  •  
  • Fine arsenopyrite must be recovered to final concentrate

       
  •  
  • Many ores (particularly Lewis ores) must be ground finer than P80 of 75 µm to improve the liberation characteristics

       
  •  
  • Pyrite hosts a significant portion of the gold as solid solution gold within the crystal matrix, and therefore must also be recovered to concentrate

       
  •  
  • Over-grinding of the liberated sulphides to less than 10 µm diameter particles would be detrimental to flotation recovery

       
  •  
  • The liberation properties of the sulphides within the Donlin ores are somewhat variable

       
  •  
  • The expected concentrate sulphur grade from a Donlin flotation circuit producing high gold recovery (+90%) is going to be relatively low, less than 10% sulphur content, because of the presence of low-grade composite particles of sulphide with gangue, floatable gangue (such as carbon and carbon/clay binaries), and non- floatable gangue as entrainment.

    Pilot-Plant Testing (G&T Metallurgy Q3 2006)

    Testwork was performed at G&T in Q3, 2006 to confirm process parameters on a blend (50% Lewis Intrusive, 25% ACMA Intrusive, and 25% Sedimentary). This pilot flotation testing did not produce flotation gold recoveries (at required grade) that were comparable to those achieved from a bench test undertaken on the same sample. Pilot results showed recovery in the range of 83% to 85% in the blended sample, compared to the bench flotation test at 91% to 93%.

         
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    Subsequent testing was undertaken to explain this discrepancy through exhaustive retesting of both the bench and pilot-plant flotation cells under different operating and test conditions. This work led to the conclusion that the froth conditions generated by the G&T Metallurgy pilot plant were hindering recovery of sulphide composite particles to the concentrate. As a consequence the importance of froth recovery was emphasized in design.

    Pilot-Plant Testing (SGS Lakefield December 2006)

    The SGS pilot run in December 2006 was initiated after the G&T Metallurgical work was completed in 2006, when it was reasonably understood which key items were affecting gold recovery under pilot conditions.

    The purpose of this pilot run was to demonstrate the ability of closing the gap between bench and pilot performance and to understand the performance of the new proposed flowsheet. This pilot-plant testwork was performed on available material and later followed up by another pilot run using freshly drilled core to eliminate any issues related to sample weathering.

    It should be noted that the generally poorer flotation results were attributed to partially geologically oxidized ore in some of the upper areas of the deposit, as evident from the 2007 variability testing. This was not understood at that time, and therefore the December 2006 composite sample included material that was geologically oxidized.

    From the pilot-plant run approximately 91% gold recovery to a 7% sulphur concentrate was achieved compared to a bench test result of 93% recovery on a LOM lithology blend. This was a significant improvement from the earlier pilot work in 2006.

    At this point in the study, to further improve the overall performance of the flotation circuit design, an alternative MCF2 flotation configuration (Figure 13-6) was considered, and some comparative bench testing of this option commenced.

         
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    Figure 13-6: Comparison of SGS Lakefield Dec 2006 Key Bench and Pilot-Plant Results

    Bench Testing of MCF2 Alternative Flowsheet

    In early 2007, a series of bench flotation tests were initiated to explore the potential benefit of the MCF2 circuit configuration.

    Initial tests were conducted by SGS Lakefield at a nominal grind of P80 40 µm for the second stage product and with varying primary stage grind sizes. The results showed that the alternative MCF2 grind/flotation configuration was realizing a measurable improvement in gold recovery of approximately 2% at the same final concentrate grade as the conventional grind/float configuration.

    These tests also suggested that the primary grind size selection was not a key parameter and that there could be some flexibility in the final size selection for the stage of grinding.

    Subsequent bench tests were then undertaken to attempt to quantify the effect of the secondary grind size on gold recovery. The results did not clearly define an optimal secondary grind target but a nominal 50 µm (P80) was selected for the subsequent MCF2 pilot run based on trends identified through previous mineralogical work.

    Pilot-Plant Testing, (SGS Lakefield February 2007)

    A second SGS Lakefield pilot-plant campaign was undertaken February to March 2007, with the primary aim to confirm pilot-plant recovery of the conventional flotation circuit on a LOM lithological blend of freshly-drilled core.

    The pilot runs produced on average 92.8% recovery to a 7% sulphur concentrate grade compared to the bench tests, which indicated ~94.0% recovery. The second series of pilot runs was undertaken and confirmed that a mild steel primary mill with high chrome media could be used in lieu of the original stainless steel mill with high chrome media. A third series of pilot tests was undertaken where the scavenger concentrate was reground before cleaning. A slight improvement in the grade/recovery profile was evident. A fourth series of pilot tests was conducted to evaluate the MCF2 configuration. A clear trend of improved recovery is evident. At a concentrate of 7% sulphur grade, the recovery difference between MCF2 and conventional flotation is ~1.8% . This compares reasonably well to the ~2% recovery improvement indicated by the initial MCF2 screening bench test results.

         
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    Based on the Phase 2 piloting at SGS Lakefield, it was recommended that the MCF2 option be selected evaluated as a potential base case for the Donlin feasibility grinding/flotation circuit design.

    Variability Testing (SGS Lakefield, Q2, 2007)

    The flotation variability testwork program consisted of a total of 149 flotation tests using 102 different test samples selected to cover a range of different lithologies and geological domains from core drilled throughout 2006. The number of samples selected for each lithology was based on the proportion of the orebody represented by that lithology

    Of the 102 samples, 22 were characterized as having some form of partial geological oxidation, and the remaining 80 were considered unoxidized fresh rock. Two types of variability bench-scale flotation tests were carried out, a modified Minnovex Flotation Test (MFT), and a conventional bench flotation test (CFT).

    Analysis of test data indicated:

  •  
  • RDX, RDA, RDXL, and MD lithological domains appear to behave similarly in terms of average and standard deviation, and together have an average flotation recovery of ~96%.

       
  •  
  • GWK and SHL recoveries are, on average, lower (91.5% and 89.8%, respectively), with a relatively large variation in performance.

       
  •  
  • RDXB lithology is the worst-performing intrusive, with an average recovery of 94.8%. This is characteristic for RDXB, which has relatively high graphitic carbon content compared to the other intrusives.

       
  •  
  • RDF is the best-performing intrusive, with an average gold recovery of 97.7% and relatively low variance in performance.

       
  •  
  • The GWK domain dataset does exhibit a weak correlation of flotation recovery with both arsenic and gold grade, noting that there is a natural strong relationship between gold and arsenic head grades. Given the relatively poor correlation with the GWK dataset this type of relationship is not recommended to predict recovery for the GWK ores. Instead, a non-weighted average of the test data should be used.


         
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  •  
  • For the SHL test results, no obvious correlations are evident to improve recovery predictions, and a non-weighted recovery average should be used here as well.

    Another methodology for characterizing the test samples is via geological domain, which is mainly based on physical location rather than rock type.

    Examining the statistical representation of recovery data based on geological domain suggests:

  •  
  • Samples from 400, ACMA, Akivik, and Aurora behave fairly similarly, with an overall average recovery of ~96.5%.

       
  •  
  • GWK and SHL recoveries are, on average, lower (91.5%, 89.8%, respectively), with a relatively large variation in performance.

         
     
  •  
  • Lewis is the worst-performing intrusive, with an average recovery of 94.3%.

         
     
  •  
  • Vortex is the medium-performing intrusive, with an average recovery of 95.2%.

    The Vortex geological domain is adjacent the Lewis domain; the Lewis geological domain has a high content of RDXB intrusive, and conversely, RDXB is concentrated mainly in the Lewis area of the deposit.

    Effect of Geological Oxidation on Flotation Performance

    The upper portion of the Donlin orebody contains ore that shows some sign of geological oxidation or weathering. To quantify the potential impact of this oxidation on flotation at Donlin, a series of 22 samples were selected. Results indicated that presence of some form of geological oxidation significantly affects the flotation performance. The average gold recovery is 72%, with a relatively high standard deviation of 22% recovery.

    Where there was insufficient sulphur content in the test sample to generate a 7% concentrate, gold recovery was determined at the 15% mass-to-concentrate point instead.

         
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    Flotation Circuit Scale-up and Modelling for Feasibility Design

       

    The conventional approach to designing flotation circuits focuses on the use of scale- up factors from bench-scale testwork to determine the residence time required for a full-scale plant. For Donlin, a modelling approach was adopted.

       

    Two flotation simulators were tested. One is JKSimFloat developed by the AMIRA P9 Project, Australia, and the second is FLEET (Flotation Economic Evaluation Tool) developed by Minnovex (now SGS).

       

    JKSimFloat was the simulator used for Donlin flotation circuit design because of its robustness based on rigorous validation at various operations, including those with refractory and PGM ore types relevant to Donlin. The JKSimFloat simulations predicted the Donlin pilot-plant results fairly accurately, especially with the addition of a cleaning circuit in the secondary rougher circuit.

       

    Simulated results for a 59,000 stpd (53,500 t/d) flotation circuit using the floatability parameters suggested a gold recovery of about 94% at a concentrate grade of 7.2% S for a two-row circuit configuration treating a throughput with head grade of 1.12% S and 2.37 g/t of gold. Flotation cell sizes of 300 m3 were selected based on lip loading data.

       
    13.1.7

    Pressure Oxidation

       

    Chemistry of Pressure Oxidation and Hot Cure

       

    Pressure oxidation in gold processing generally refers to the oxidation of gold-bearing sulphide minerals to metal sulphates using a combination of heat (typically 200°C to 230°C), acid, and oxygen sparging in a specifically designed pressure vessel. The breakdown of the sulphide particles effectively releases the gold locked within the mineral matrix, rendering it amenable to leaching by cyanidation.

       

    Batch Autoclave Testing (Dynatec 2004)

       

    During 2004, Dynatec carried out bench-scale autoclave testing of four composite samples, ACMA Intrusive, ACMA Sediment, Lewis Intrusive, and Lewis Sediment. The scope of the test program included kinetic and locked-cycle mass balance pressure oxidation tests on the concentrates, followed by neutralization tests on the pressure oxidation discharge liquors and carbon-in-leach (CIL) cyanidation tests.


         
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    The concentrates were relatively fine, with P80 levels of 33 to 41 µm (82% to 89% minus 44 µm), and were tested without further size reduction for comparison purposes. Direct CIL cyanide leaching of the unoxidized feeds yielded gold extractions between 3% (ACMA Sedimentary) and 11% (Lewis Intrusive).

    Higher autoclave oxidation kinetics were observed at 210°C and 220°C than at 200°C. Gold extractions were highest from the solids oxidized at 220°C. All subsequent pressure oxidation testwork on all four concentrates were, therefore, conducted at 220°C.

    As shown in Figure 13-7, the sulphide sulphur oxidation kinetics was rapid, with more than 98% oxidation achieved within 30 minutes.

    Gold extractions from the oxidized concentrates, as shown in Figure 13-8, were correspondingly high after 30 minutes of pressure oxidation and improved marginally to their maximum values of between 95.1% (ACMA Sedimentary) and 98.5% (ACMA Intrusive) after 45 minutes of oxidation. With extended pressure oxidation time, however, the gold extractions declined, most markedly for the sedimentary concentrates, which had relatively high organic carbon content.

    A retention time of 45 minutes was selected for the pressure oxidation in the subsequent material balance and locked-cycle testwork.

    In the locked-cycle testwork, the pressure oxidation tests were conducted at pulp densities approaching those anticipated for the discharge slurries in commercial autoclave operation. The extents of sulphide sulphur oxidation with the 45 minute pressure oxidation retention time at 220°C exceeded 98%, with more than half over 99%. There was no systematic change in the oxidation extent with increasing cycle number, indicating that the recycled solution did not affect the sulphide oxidation.

         
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    Figure 13-7: Sulphide Oxidation Pressure Oxidation Kinetics at 220°C

    Figure 13-8: Gold Recovery Profiles from Pressure Oxidation at 220°C

    Analysis of selected solution samples for gold and silver indicated that these were below their respective detection limits of 0.2 mg/L and 1 mg/L in the pressure oxidation discharge solutions.

    Stirred tank CIL cyanide leach gold extractions varied from 90.3% to 98.8% for the four oxidized concentrates, with median extractions of 93.5% to 97.4% .

    Subsequent testing has indicated that oxidation rates and performance of the batch tests are strongly affected by the decision to pre-acidify the concentrate sample charge, or not, both at pilot scale and bench scale.

         
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    Batch Autoclave Testing (Barrick Technology Centre 2006)

    A summary of the various tests undertaken on composite samples in 2006, indicating the key conclusions, is provided in the following subsections. This discussion is mainly limited to gold recovery performance.

    Baseline Tests

    A series of batch tests was conducted to understand the potential impact of autoclave temperature, oxidation time, pre-acidification, oxygen concentration, and thiocyanate concentration on autoclave performance.

    The bench autoclave test (BTAC) results for the various concentrate samples at a temperature of 220°C. Recoveries in the order of ~95% to 99% were achieved at the 45 minute pressure oxidation time for all ores tested.

    The results indicate that oxidation rate increases with POX temperature and that of the temperatures tested, the optimum in terms of gold recovery profile is 230°C, being marginally better in performance than at 220°C. No definitive explanation can be provided to explain the unusual decrease in performance of the 240°C test result.

    In conclusion, autoclave temperatures of more than 192°C are required to achieve >92% gold recovery within a nominal 1 hour autoclave residence time, and temperatures of 220°C to 230°C provide maximum recovery values.

    Some preliminary BTAC tests were undertaken to evaluate the potential impact of thiocyanate (SCN) dosed into the feed slurry on the CIL gold recovery of the autoclave products. No detrimental impact on recovery was indicated

    From the autoclave test data, it was noted that gold recovery improved with increasing POX residence time. Therefore, experiments were undertaken to test the potential application of higher-temperature POX (230°C to 240°C). It was seen that gold recovery does improve with extended autoclave time and that this improvement is indeed accelerated at 240°C.

    BTAC Testing on 2007 Phase 1 Composite Concentrate Sample

    During January 2007, the Barrick Technology Centre carried out a series of BTAC tests on a sub-sample of the concentrate generated from the SGS Lakefield December 2006 pilot flotation test program. The test program aimed to investigate the pressure oxidation characteristics of the new composite concentrate sample against previous testwork, and to investigate the effect of autoclave on pressure oxidation performance. The best result was at 220°C with 45 minutes of residence time.

         
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    Pilot-Plant Testing (Barrick Technology Centre 2006)

    Four phases of autoclaving pilot tests were carried out during the latter half of 2006. In discussing the gold recovery aspects of the pilot-plant results, all testwork phases undertaken in 2006 are considered as one, for clarity.

    Pilot Campaigns at 220°C and 225°C

    A series of 220°C and one 225°C pilot campaigns were also carried out during the latter part of 2006. The three tests undertaken that generated final CIL gold recoveries of less than 95% were due to oxidation rate performance issues. Based on earlier results, a final run was attempted at 225°C, incorporating pre-acidification of the concentrate feed, with an extended residence time. The purpose was to determine under pilot conditions what recovery could be achieved at more complete oxidation levels. This run successfully demonstrated that the CIL gold recoveries of the autoclave residue in excess of 96% were possible.

    Pilot Campaigns at 240°C

    Previous batch tests carried out at 240°C indicated the potential for operating at this temperature and achieving high gold recovery. As for the 240°C test runs, high gold recoveries were achieved quickly, albeit with faster oxidation kinetics than at 220°C. An improvement in recovery is evident, but the level of improvement is insufficient to achieve high gold recoveries under practical design constraints.

    Pilot-Plant Testing (2007 Phase 1)

    During February 2007, an additional autoclave pilot campaign was undertaken at the Barrick Technology Centre (BTC) to verify the autoclave design parameters and potential gold recovery results for the 2007 FS. This pilot run attempted to explore the three different operating temperatures, 200°C, 210°C, and 220°C. Based on previous successes with pre-acidification in 2006, it was decided to utilize pre-acidification of the concentrate for each of these planned runs.

    The selected residence times for unit operation were purposely set to be higher than design to try to achieve more fully oxidized conditions at the discharge than had been realized in previous campaigns, and to provide a set of data covering a large range of operating residence times. This coverage of residence time is possible by sampling each of the pilot-plant compartments and undertaking CIL tests on each sample.

         
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    A maximum gold recovery of ~96.9% was reached in compartments 3 and 4 for the 225°C test run, representing a residence time of 50 to 55 minutes. For the 220°C run, a maximum recovery of 95% to 96% was achieved in compartment 5 at a residence time of ~55 minutes. It can be seen that CIL gold recovery drops significantly towards the discharge of the autoclave, after peaking at 96.0% to 96.4% near compartments 3 and 4. As the operating temperature of the autoclave is lowered, the required residence time for reaching maximum gold recovery increases.

    Pilot vs. BTAC Testing

    It is well known that the BTAC test has inherent limitations in its ability to replicate a continuous autoclave operation. The continuous introduction of fresh feed into a continuously operating facility, the continuous transfer of new feed to each stage (residence time distribution), and the need to provide the BTAC unit with acid solution in the feed to allow the oxidation reaction to “kick-start” all result in some important differences in the chemistry, and the timing of that chemistry occurring, between a BTAC test and a pilot test.

    Pilot testing is considered the more appropriate test method to replicate the chemistry and kinetics of a production-scale autoclave and is regarded as best practice for definitive autoclave testing.

    Summary

    The main results from 2007 Phase 1 pilot testing indicated that:

  •  
  • CIL gold recovery achieved from the pilot autoclave is sensitive to the autoclave residence time where recovery is lower than optimum when autoclave residence time is either too short or too long.

       
  •  
  • Autoclave operating temperatures of 220°C and 225°C provided the highest gold recoveries with the lowest autoclave residence times, with optimum residence time of around 40 to 55 minutes.

    Considering the target gold recoveries were demonstrated only on autoclave compartmental samples, not actual pilot autoclave discharge samples, it was decided to continue the pilot autoclave testing program into Phase 2.

    Pilot-Plant Testing (2007 Phase 2)

    During June 2007, Phase 2 pilot autoclave testing was carried out at the Barrick Technology Centre. The aim of the Phase 2 testwork was several-fold:

         
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  •  
  • To operate the pilot unit more closely to the design optimum autoclave residence time as determined from Phase 1

       
  •  
  • To demonstrate that target CIL gold recoveries can be achieved on final products from the autoclave, not just on compartmental sub-samples

         
     
  •  
  • To confirm potential downstream CIL gold recoveries

       
  •  
  • To confirm the selected design criteria for the pressure oxidation circuit for the 2007 FS

         
     
  •  
  • To provide more autoclave profile data for oxidation rates and gold recoveries.

    The concentrate feed sample was obtained from material generated from the SGS Lakefield pilot flotation test program undertaken in January 2007.

    Pre-Acidification

    Concentrate pre-acidification reduced the carbonate content (inorganic carbon) of the concentrate reduced from 0.35% to 0.05%, representing approximate 86% dissolution of the contained carbonates.

    Sulphide Oxidation Performance

    The pilot run investigated two different operating temperatures, 220°C and 225°C.

    For test runs W and X, the oxidation rates were reasonably fast and consistent, and that measurable sulphide oxidation was essentially completed by 37 to 42 minutes’ residence time.

    Autoclave Discharge Gold Recovery Performance

    Before the first set of discharge samples was collected, the autoclave was operated for approximately two hours, or approximately 2.5 turnovers, to allow it to come to near steady-state before sampling commenced. The autoclave then continued to operate for another 4 to 5 hours, with regular sampling.

    The pilot autoclave operated at periods of high gold recovery, 96.1% to 97.3%, at times, but then shifted and operated at 93% to 94%, seemingly moving from one “recovery regime” to the other very quickly. It was found that if the autoclave residence time is permitted to exceed 50 minutes, then CIL gold recovery of the AC discharge drops to 93% to 94%. Similarly, if autoclave residence time is too short, then recovery is also lower than optimum at 96% to 97%. Optimum gold recovery, exceeding 97%, is recorded at around 45 to 49 minutes.

         
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    Autoclave Profiles Recovery Performance

    During the operation of pilot campaign runs W and X, four profile samples were collected through the autoclave, approximately every 30 minutes. The same trend of the effect of residence time on CIL gold recovery as seen for the discharge samples is clearly evident. Based on the results from these samples, the optimum operating residence time is confirmed in the range 37 to 47 minutes. Optimum gold recovery from the autoclave appears to correspond with the “just completed” extent of sulphide sulphur oxidation. Excess oxidation after that point is detrimental to gold recovery.

    Carbon Dissolution in Autoclave

    It was seen that the organic carbon content (by assay) has decreased from 0.78% to an average of 0.57%, meaning that ~25% of the organic carbon has oxidized within the autoclave. Further, the graphitic carbon (by assay) shows negligible reduction in assay grade in the autoclave. Inorganic carbon in the feed is reduced to below assay detection limits by the pre-acidification process ahead of the autoclave, and so no apparent trends are discernible.

    Autoclave Vent Gas Sampling

    During the Phase 2 pilot autoclave run, standard gas testing was performed on the main vent gases from the autoclave, ahead of the water scrubber, with the purpose of quantifying the oxygen, carbon dioxide (CO2), carbon monoxide (CO), and total hydrocarbon (THC) generation rates from the autoclave that would be fed to the gas scrubbing system.

    Hot Cure

    An additional testwork series on hot curing optimization was undertaken on the Phase 2 pilot autoclave products to confirm the proposed hot cure section design for the 2007 FS and FSU1, and to provide the latest information for future optimization.

    The analyses of the hot cured solids and liquor components show that sulphur in the solids is indeed dissolving, with a corresponding decrease in solids mass and liquor sulphuric acid concentration and a consequential increase in liquor iron content. It can be seen that a decrease in lime consumption is required for pH adjustment to 11 and that CIL gold recovery improves slightly at the six-hour residence time point. The FSU2 hot cure circuit is designed with a residence time of six hours. Lime consumption of well-washed hot cure solids to pH 9 was relatively constant at around 3 to 4 kg/t of hydrated lime.

         
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    Hot Cured Product Counter-Current Decant Washing

    A large sample of autoclave hot-cured product was collected from the Phase 2 pilot run and subjected to a series of washing tests followed by neutralization to pH 7, 9, and 11. Significant reductions in lime consumption occur up to 98% washing efficiency, the feasibility design target, and the addition of magnesium sulphate substantially increases lime consumption from 8 to 22 kg/t when adjusting pH to 11 but does not greatly affect lime consumption for adjustment to pH ~8.8 to 9.0; in this case, lime consumption increases from ~5.7 to 6.5 kg/t. Note that lime consumption is specified as kilograms of Ca(OH)2.

    The benefit of operating a CIL circuit at pH ~9, in terms of lime consumption, is also demonstrated from this work.

    Summary

    The main results of the 2007 phase 2 pilot autoclave testing program were as follows:

  •  
  • Product CIL gold recoveries of 96.6% can readily be achieved and recoveries of 97% are possible under optimum operating conditions, as indicated by the tests on the discharge samples as well as the autoclave profile samples.

       
  •  
  • CIL gold recovery from the pilot autoclave is sensitive to the autoclave residence time. Gold recovery is slightly lower than optimum when autoclave residence time is too short because of incomplete sulphide sulphur oxidation. Recovery can also be lower than optimum if autoclave residence time is too long and oxidation is excessive.

       
  •  
  • Autoclave operating temperatures of 220°C and 225°C provided good results with optimum residence times of 45 to 49 minutes for CIL gold recovery, based on the autoclave discharge samples.

       
  •  
  • Measurable sulphide sulphur oxidation is essentially completed by 37 to 42 minutes residence time, as indicated by analysis of the autoclave profile samples.

       
  •  
  • The selected hot curing time of six hours as per the feasibility design provides good lime consumption results and CIL gold recovery performance, but dissolution of arsenic is evident, subsequently requiring precipitation in the following neutralization stage.


         
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    Autoclave Mercury Emission Testwork

       

    During a pilot run, two gas streams were sampled for mercury content in the flash pot vent stream and the autoclave vent stream.

       

    The amounts of mercury vented through the flash system and the autoclave vent were 0.21% and 0.026%, respectively, of the calculated mercury head in the concentrate.

       

    The results from a second series of emissions testing indicated the presence of very little mercury emission in the combined gas streams (0.003% of feed mercury content), with the gas and scrubber mercury contents being close to assay detection limit.

       

    Despite the low measurements, a mercury abatement system has been designed to comply with the December 2010 US EPA National Emissions Standard for Hazardous Air Pollutants for gold ore processing and production facilities.

       
    13.1.8

    Neutralization

       

    Introduction

       

    The oxidation of pyrite and other naturally occurring sulphides generates sulphuric acid and other metal sulphates within the autoclave and hot curing circuits. These species need to be neutralized and precipitated into a stable form to ensure that the final tails from the plant have a low soluble metals content, and is also at approximately neutral pH.

       

    Typically limestone (calcium carbonate) and hydrated lime (calcium hydroxide) are utilized for this neutralization duty. Limestone is used for the first part of the pH adjustment while the final pH adjustment to 7 (for acidic liquor neutralization) and 9-11 (for CIL feed) is carried out with hydrated lime as Ca(OH)2, which is a more reactive neutralizing reagent.

       

    Metallurgical studies were undertaken to determine the potential neutralization capacity of the flotation tails stream and also of an identified local natural source of low-grade carbonates known as calcareous sandstone (CSS). This local material was composed of ferroan dolomite, ankerite, calcite and siderite in decreasing levels.

       

    The orebody itself has a relatively high carbonate content of 2.33% reported as CO2 (or 3.18% reported as CO3) over the life of mine compared to a sulphur content of 1.13% S. This represents a stoichiometric ratio of carbonate to sulphur of 1.49. Therefore, there is sufficient, or rather, excess alkali in the ore available to neutralize the sulphates generated from the oxidation of all the contained sulphur. However, this requires effective utilization (reactivity) of the measured neutralization content of the ore for the excess to be valid.


         
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    It has been shown through the testwork carried out, that with the provision of sufficient neutralization residence time, and the elevation of the slurry temperature in neutralization, good utilization of the contained carbonates in the ore is possible, resulting in the minimal amounts of lime needing to be consumed by the plant.

    Further, it has been determined that the use of CSS is not required for neutralization of the acidic autoclave acid, due to effective use of the ore itself, as flotation tails, and at high pH (>5.0), CSS does not compete economically with imported lime. The local CSS resource instead serves as a back-up alkali source for the Donlin Gold Project if, and as, required.

    Dynatec – November 2004

    During 2004 Dynatec tested the neutralization properties of the acidic liquors generated from bench autoclave tests, using flotation tails, limestone, and lime. The neutralization tests conducted with limestone and lime performed well, and at pH 8.0 all metals, with the exception of Mn and Mg, precipitated virtually completely from solution at generally below detection limit grade. The results however were not favourable from a lime consumption perspective.

    Placer Dome Technical Services 2005

    During late 2005 and into 2006, Placer Dome Technical Services (PDTS) investigated the neutralization capacities of CSS and flotation tails. The results show that CSS with high CaO/MgO ratios provided higher carbonate utilization compared to low ratio composites. The grind size did not have a large effect on the neutralization capacity of the CSS composites.

    Testwork on flotation tailings showed higher carbonate usage from the intrusive materials compared to the sedimentary materials. Overall though, the utilization of the carbonate within the tailings was very low.

         
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    Preliminary Batch Neutralization Testing, 2006

    A limited batch neutralization testwork program was initiated in mid-2006, aiming to improve quantification of the neutralization options for Donlin. The results indicated that the flotation tails / lime neutralization option, with extended neutralization residence time, was the most economic with the lowest total cost.

    Pilot Neutralization 2006

    During October 2006 a pilot neutralization testwork program was undertaken. The acid solution used for the pilot test was produced during the pressure oxidation pilot run conducted on 26 Sept 2006, with the autoclave operating at 220°C followed by hot curing of the slurry for at least 12 hours at 95°C.

    Flotation tailings were used as produced from G&T Metallurgical being blended in the ratio of 25% ACMA intrusive, 50% Lewis Intrusive and 25% Sedimentary.

    Profile samples were routinely collected from the pilot-plant run, and then lime added to achieve a final pH of 7.0. At the given ratio of flotation tailings to concentrate (5.135 kg tailings per kg concentrate), neutralization of the dilute acidic pressure oxidation solution using flotation tailings with a carbonate grade of about 1.8% CO2 reached a pH of 4.0 to 4.5 after 12 hours. There was no significant increase in pH at longer retention times. Under the conditions tested, lime consumption after flotation tailings neutralization was approximately 4.5 g quicklime per litre of dilute acid solution, or 24 g of quicklime per kilogram of concentrate. It was also possible to use CSS at the rate of 1.6 kg of CSS per kg of concentrate.

    Neutralization Phase 1 Bench Testing, 2007

    The acidic solution used for this testwork was generated from Donlin concentrate during the continuous pressure oxidation (POX) pilot run on 6 February 2007. The concentrate was pre-acidified to pH of about two with sulphuric acid prior to the oxidation process. The campaign run was at 225°C with a 70 minutes retention time. The discharge slurry was then hot cured for about 24 h immediately following the POX process.

    The hot cure slurry was washed with gypsum saturated water in a pilot counter-current decant (CCD) circuit at a ratio of 2:1. The overflow acidic filtrate was collected and used for batch and continuous tests. Initially, the pH of the diluted POX solution was approximately 1.0.

         
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    The flotation tailings used was a blend from a flotation piloting campaign in December 2006 obtained from SGS Lakefield. The flotation tails used had a carbonate grade of 2.0% CO3 and was sourced from the same pilot float feed sample that provided the concentrate used to generate the acidic liquor, via the pilot autoclave.

    Bench tests at varying temperatures were undertaken to determine the effect of temperature on neutralization rate and utilization. It was seen that a significant improvement in performance occurs as temperature increases. Temperatures of 55°C or greater allowed a pH of 6.0 to 6.5 to be reached, consequently resulting in the reduction of lime consumption to about 1 to 2 g/L of diluted acidic liquor.

    Neutralization Phase 1 Pilot Testing, 2007

    Based on the bench tests, a pilot neutralization campaign was initiated in early 2007, using 70°C and two residence time selections of eight and 12 hours, and using the same acidic liquor and flotation tails as used for the bench testing.

    The average pH profiles of both the eight and 12-hour residence time test campaigns are shown in Figure 13-9. It can be seen that final pH levels of greater than 6.0 were reached with the flotation tails, for both the eight hour and 12 hour runs.

    Figure 13-10 shows that lime consumption tests of profile samples taken from the pilot-plant runs. It can be seen that lime addition required is less than 1 g/L of diluted acidic liquor, at the end of the neutralization circuit.

         
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    Neutralization Phase 2 Pilot Testing, 2007

    With the undertaking of the Phase 2 pilot autoclave test program in June 2007, an opportunity arose to confirm the expected performance from the final feasibility neutralization circuit using a suitably designed pilot-plant set-up. In addition, this provided an opportunity to test the neutralization performance of the MCF2 pilot flotation tails. The design of the neutralization pilot circuit closely followed that of the actual feasibility circuit design, with the following key parameters:

     
  •  
  • Four float tails neutralization tanks
         
     
  •  
  • One lime neutralization tank
         
     
  •  
  • Six hours’ residence time ~ optimization based on testwork
         
     
  •  
  • Operating temperature of 55°C ~ based on heat balance

         
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  •  
  • Addition of CIL tails into tank one

    Figure 13-9: 2007 Phase 1 Neutralization Pilot pH Profiles

    Figure 13-10: Lime Demand Test Results of 2007 Phase 1 Pilot Samples, Plotted against Initial pH

     

     
  •  
  • Mixing ratios of diluted acidic liquor and flotation as per the feasibility mass balance.


         
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    In addition to the six-hour residence time run, a second three-hour residence time campaign was undertaken to investigate the potential to reduce the size of the circuit for the detailed design phase.

    The source of the dilute acidic liquor was the overflow stream from the operation of the pilot CCD plant, washing Run W hot cured product. To be conservative, no calcium carbonate was added to this acidic liquor to correct for the pre-acidification process.

    The source of the flotation tails was SGS Lakefield flotation pilot PP-11, which incorporated the MCF2 grinding/flotation flowsheet. The sample used for the pilot test had a carbonate grade of 1.65% (as CO2), compared to the LOM predicted carbonate grade of 2.33% (as CO2).

    The average pH profiles for the two test campaigns are shown in Figure 13-11. It can be seen that with the six-hour test campaign (i.e., matching the feasibility circuit design), a final pH of 6.75 was reached prior to the lime addition step. With the reduction to three hours’ residence time, the final pH was 6.50.

    Figure 13-12 shows the results of the lime demand tests undertaken on profiles from the pilot plant. Lime consumption at the end of the neutralization circuit is less than 0.2 g/L of diluted autoclave acid solution, noting that wash water flows have been increased in the feasibility design (more dilution of the acidic liquor) to improve washing efficiency of the autoclave discharge hot cure product.

    The results of the three-hour test campaign are encouraging, suggesting the potential to decrease the size of the neutralization circuit further. However, for the purposes of managing the potential variability of carbonate content in the ore, and also to provide time for the operations personnel to respond to unplanned grinding or flotation circuit shutdowns, the longer six-hour residence time circuit continues to be the recommended design.

         
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    Figure 13-11: 2007 Phase 2 Neutralization Pilot pH Profiles

    Figure 13-12: Lime Demand Test Results of 2007 Phase 2 Pilot Samples, Plotted against Initial pH

     

         
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    Neutralization Variability Testing

       

    During July to September 2007, a variability neutralization testwork program was initiated at SGS Lakefield. The program had the dual aim of confirming the potential differences in neutralization performance of the varying lithologies and of developing a confident relationship between lime consumption (for final pH trim to 7) and the feed samples carbonate grade.

       

    SGS Lakefield generated a synthetic dilute autoclave acid based on the prediction of key species content from the near-final MetSim model’s prediction of that stream. Flotation tails samples from the recently completed flotation variability program were used as the source of neutralizing solids.

       

    Figure 13-13 is a summary chart of the results of the tests completed.

       

    Prediction of lime consumption for acidic liquor neutralization for the FSU2 operating cost estimate is based upon the relationship between lime demand and flotation feed carbonate grade developed from this test data.

       
    13.1.9

    Carbon-in-Leach (CIL)

       

    The following subsections summarize the processes of cyanidation and carbon adsorption, and then describe the results of the various testwork programs completed on the Donlin ores.

       

    Testwork Introduction

       

    Extensive cyanidation testing has been undertaken on samples of Donlin, at various points in the flowsheet, since 1995.

       

    Cyanidation of unoxidized Donlin Creek ores, with or without the presence of activated carbon, consistently yields very low gold recoveries of 5% to 30%, either as flotation feed, flotation tails, or concentrate. This is characteristic of an ore where gold is predominantly associated with arsenopyrite or pyrite in solid solution form, like Donlin ore.

       

    The bulk of the cyanidation tests carried out to date have largely been on autoclave compartmental and discharge samples, where large numbers of relatively small samples are leached with high concentrations of carbon and cyanide. This is diagnostic tool that enables the performance of various autoclave tests to be established without the added complication of the constraints that could be imposed by attempting to optimize leaching kinetics.


         
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    Figure 13-13: Plot of Neutralization Variability Testing Lime Demand Results at 6 Hours’ Residence Time


    CIL gold recovery has generally shown to be more sensitive to the autoclave operating conditions (residence time, temperature) than to the operating conditions and methods applied in the CIL circuit. The target of the metallurgical design of the CIL circuit is rather to ensure that good CIL recovery performance is achieved on the material presented to it from the autoclave, with optimum reagent (lime and cyanide) usage.

    Key Metallurgical Aspects of Planned Donlin CIL Circuit

    The key metallurgical aspects of the final selected CIL circuit design are summarized as an introduction to the detailed discussions of the specific testwork results presented subsequently. The aim of the testwork programs was to define the operating characteristics and circuit design of the CIL circuit for treatment of CCD-washed autoclave product at Donlin.

    Leach Circuit pH

    The MetSim modelling and metallurgical testing have shown that the Donlin CIL circuit could operate well at a relatively low pH of 9.

    Increasing CIL pH above 9 in the CIL circuit will consume additional lime through the precipitation of magnesium sulphate in solution to magnesium hydroxide. The lime then re-dissolves back into solution when mixed back into the tailings and returns to the plant through tailings water recycle. To achieve the traditional CIL circuit pH levels of 10 to 11, all of the magnesium in the feed solution would need to precipitate completely. In the FSU2, reclaim water used in CIL is treated to remove magnesium and enable operation at conventional pH in CIL.

         
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    Assuming a CIL pH of ~9, lime addition is estimated to be in the order of 5 to 7 kg/t of concentrate. The key component affecting both is the washing efficiency of the autoclave product CCD wash circuit. To maximize washing efficiency, a four-stage CCD circuit with a high wash ratio of 4:1 is used.

    CIL of Flotation Tails

    During late 2006, BTC undertook CIL tests on the flotation tails from the G&T pilot-plant campaigns. Table 13-13 summarizes the results. Based on the low head grade and low gold recovery, it is not economically viable to leach the flotation tails.

    CIL Optimization Testwork – 2006

    A composite of autoclave discharge material from the 2006 autoclave pilot test program was collected and leached under varying conditions. The gold recoveries achieved from this work were limited due to the nature of the product from the particular pilot run. The following conclusions are drawn from the CIL optimization work undertaken on the pilot autoclave product:

     
  •  
  • Cyanidation in the absence of carbon is detrimental to final leach gold recovery.
       
  •  
  • Using higher carbon loadings (pre-loaded with gold) does not adversely affect gold recovery.
         
     
  •  
  • Leaching at pH 9.2 does not negatively affect CIL gold recovery.
       
  •  
  • Increasing the wash efficiency of the CIL feed slurry can significantly reduce reagents consumption rates.
       
  •  
  • Leaching at too high a slurry density can negatively affect CIL gold recovery through unsuitable rheological properties of the slurry.

    Samples of the detoxified CIL products were sent to the University of British Columbia for rheological testing, where it was seen that the slurry solids content (% solids) had a significant impact on the viscosity of the slurry. Densities above 35% are considered potentially problematic.

         
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      Table 13-13: CIL Results from Pilot Flotation Tails
        Head NaCN Lime Float Tails
        Grade Consumption Consumption Gold Recovery
      Ore Type (Au g/t) (kg/t) (kg/t) (%)
      Sediment 0.75 0.88 0.7 16.0
      Lewis Intrusive 0.51 0.56 0.6 33.3
      ACMA Intrusive 0.47 0.61 0.7 27.7
      Blend 0.59 0.79 0.7 30.5

    CIL Pilot-Plant Testing 2007

    In early 2007, the Barrick Technology Centre carried out a pilot CIL test run using CCD-washed pilot autoclave product from the 2007 Phase 1 autoclave pilot testwork program.

    Gold and Silver Recoveries

    The carbon in the pilot CIL circuit was successfully loaded up to 4,000 g/t gold, and two carbon transfers were undertaken.

    A gold balance for the period of the pilot-plant run returned a calculated gold recovery of 93.6% with a tailings grade of 1.39 g/t. At the tank 5 position, ahead of the adverse influence of the higher carbon loadings in tanks 6 and 7, the solids assay was lower, at 1.2 to 1.3 g/t, representing a gold recovery of ~94.0% to 94.5% .

    A bottle roll test of the CIL feed conducted at pH 11 yielded a comparative gold recovery of 93.9%, indicating that the pilot operation provides equivalent recovery performance to that achieved in a batch bottle roll test, even with the pilot plant operating at a relatively low pH of 9 and using profile-loaded carbon.

    Silver recovery from the pilot plant was low, at 29.7%, which is typical of the leaching characteristics of silver from autoclave solids product due to its dissolution and subsequent precipitation as cyanide insoluble Ag-jarosite within the autoclave.

    Leaching residence time of 20 to 24 hours over six tanks is an appropriate design for the Donlin continuous CIL circuit.

    Cyanide Consumption and Addition

    The Donlin CIL feed is relatively free of cyanide consumers. This is characteristic of concentrate that has been subject to pressure oxidation, where sulphur is oxidized completely to sulphate and base metals are dissolved into solution, and then CCD washing, which removes the dissolved metals from the autoclave product ahead of CIL.

         
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    Cyanide addition to the pilot circuit was 1.5 to 1.6 kg/t, with a consumption of 1.1 to 1.3 kg/t. Most of the consumption is likely to be from losses through HCN from the pilot CIL circuit, rather than consumption by species within the ore. HCN losses of this magnitude will not be experienced at full scale, and the HCN that evolves will be recovered via the ventilation and scrubbing system and returned to the CIL circuit feed. Cyanide addition to the pH 11 bottle roll tests on the pilot-plant feed was 1.2 kg/t, with a consumption of 0.05 kg/t.

    Assuming a CIL pH of ~9, cyanide addition of 0.7 to 0.9 kg/t of concentrate is estimated. Consumption of cyanide will be lower at a higher CIL pH of 11.

    CIL Optimization Testwork – 2007

    A series of bench-scale tests was undertaken on products from the 2007 Phase 2 Pilot autoclave test program to attempt to optimize the CIL process.

    Cyanide Addition Optimization

    A series of 24-hour bottle roll CIL tests were conducted at varying cyanide concentrations to determine the relationship between cyanide addition rate and gold recovery. Due to the limitations associated with undertaking low pH CIL tests at laboratory scale, a higher pH of 11 was used. CIL gold recovery was found to be not significantly affected at low cyanide concentration, and that the process is more economically favourable at low CIL cyanide concentration levels.

    Rheology

    Additional rheology testing was undertaken on a detoxified CIL product from the 2007 Phase 2 pilot-plant work. Beyond a level of 35% solids the viscosity of the material was found to climb rapidly.

    Cyanide Detoxification Testwork

    Introduction

    The SO2/Air (sulphur dioxide cyanide destruction) process will be used for cyanide detoxification of the Donlin CIL tailings before this stream is transferred into the neutralization circuit.

         
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    Testwork Summary – 2006

    Cyanide detoxification testwork has been completed on CIL tails slurry generated from the 2006 pilot autoclave test program. Three types of cyanide detoxification methods were tested and found effective:

  •  
  • Prussian blue (iron sulphate precipitation) – consists of adding autoclave discharge acid to detoxify the cyanide complexes
       
  •  
  • AVR (Acidification, Volatilization, Recycle) – consists of acid addition to drive the cyanide off as HCN and capturing the HCN for reuse in the circuit
         
     
  •  
  • SO2/Air testing – uses a combination of SO2 and air to detoxify the cyanide.

    All three test methods were successful in reducing the cyanide levels to expected permit requirements. The SO2/Air method was selected over the Prussian blue and AVR methods for cyanide detoxification at Donlin.

     

     

    For the SO2/Air testing, the CIL tailings slurry was effectively treated in a single stage operating with approximately 60 minutes of retention and an SO2 dosage of 4 g/g CNWAD. A pH of 8.5 was used for all tests since acid addition was required to lower pH levels. The addition of CuSO4 at 10 mg/L Cu2+ was required for effective removal of cyanide that was present in the feed.

     

     

    It was found that the content of arsenic in the liquor phase increased after SO2/Air cyanide detoxification. This solubilized arsenic will be re-precipitated upon mixing the CIL tails into the neutralization circuit as a result of the presence of high levels of dissolved iron in this circuit.

     

     

    13.1.10 

    Thickening and Counter-Current Decantation (CCD) Washing

     

     

    Introduction

     

     

    The feasibility flowsheet includes the following thickening/solids settling operations:


     
  •  
  • Concentrate thickening after flotation
       
  •  
  • CCD washing of pre-acidified concentrate with fresh water to provide optimal oxidation conditions
       
  •  
  • CCD washing of hot cured autoclave product slurry with process water to reduce lime consumption ahead of CIL cyanide leaching
       
  •  
  • Clarification of the portion of hot cure CCD overflow not reporting to pre- acidification to recover entrained gold values

         
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  •  
  • Thickening of flotation tailing prior to neutralization, to minimize dilution during neutralization and reclaim of process water.

    Earlier flowsheets presented for Donlin also included a final tailings thickener to dewater the combined carbon-in-leach (CIL) tailing and neutralization residue prior to discharge to the tailing storage facility. This thickener has been removed from the FSU2 flowsheet.

       

    The flotation tailings, which are a combination of the secondary rougher and the cleaner scavenger tailings, are de-watered before being directed to the pressure oxidation to provide cooling.

       
    13.1.11 

    Environmental Testwork

       

    To provide samples that are reasonably representative of both the complete metallurgical processes, and also the ore, the testing of combined pilot-plant tailings was selected as the preferred testing method.

       

    The final tailings from Donlin consist of a blend of detoxified CIL tails (cyanide leached autoclave and hot cure product) and neutralized autoclave acidic liquor using the flotation tails stream.

       

    The metallurgical process adopted for Donlin is favourable for the establishment of tailings that are not acid producing as a result of near-complete sulphide sulphur oxidation.

       

    The average mill feed grade is ~1.12% sulphur, with no significant sulphate sulphur present. The mill feed averages 2.51% carbonate as CO2, which is a molar excess of 37% to the contained sulphur in the mill feed after Year 2, meaning that the ore has excess carbonate content to sulphur content.

       

    Mineralogy undertaken by SGS Lakefield indicates that up to 23% of the sulphate sulphur in the 2006 pilot final tails sample is in the form of jarosite, with 7% in the 2007 Phase 1 pilot-plant final tails and 8% in the 2007 Phase 2 pilot-plant final tails. Modifying the calculated ABA parameters, assuming that jarosite is an acid-forming component of the sulphate, indicates that the tailings will still contain an excess of neutralization capacity.

       

    Pressure oxidation of arsenopyrite in the presence of excess iron is generally considered a best-practice process for generation of stable arsenic precipitates, in forms such as scorodite, for disposal into a tailings storage facility. Promoting the formation of stable precipitates is particularly favoured when molecular ratio of iron to arsenic ratio in the applicable process solutions exceeds 4. Within the plant feed for Donlin, there is sufficient iron to provide the recommended molar ratio of 4 of iron to arsenic. It should be noted that the actual assay grade of iron typically is double the iron content that is accounted for by arsenopyrite and pyrite alone and is more typically at grades of 15,000 to 40,000 ppm.


         
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    The cyanide within the CIL circuit dissolves a portion of the mercury in the solids feed to the circuit. A portion of this dissolved mercury in the CIL circuit is adsorbed onto the circuit carbon, and is then recovered from the carbon via stripping and carbon regeneration. However, the capacity of the circuit carbon to completely adsorb the mercury is limited and therefore a component of the soluble mercury remains in the CIL tails solution. This remaining soluble mercury is then blended with the detoxified CIL tails into the neutralization circuit, which then reports to the tailings storage facility.

       

    Reductions in soluble mercury content in recirculating plant waters can be achieved by addition of mercury precipitation reagents, which convert soluble mercury to a stable mercury sulphide product. This is currently practised using the Cherokee Chemical UNR reagent suite at operating mine sites in the U.S.

       

    Based on the testwork completed, it is recommended that the process plant design includes a dosage facility for Cherokee reagent UNR 829 to permit addition to a recirculating water stream for precipitation of mercury in solution into a stable HgS solid. Doing so will eliminate potential build-up of mercury in the process water circuit.

       
    13.2

    Recovery Estimates

       

    There are two components to defining the final recovery of gold to bullion from the proposed Donlin processing facility:


     
  •  
  • Gold recovered from the flotation circuit to the flotation concentrate

       
  •  
  • Gold recovered through leaching/adsorption (CIL) of the pressure oxidized (Autoclaved) flotation concentrate.

    Due to the refractory nature of the Donlin ores and the relatively low grade of the flotation tails stream, it is not economically viable to recover gold from the flotation tails stream. Therefore gold not recovered to the flotation concentrate is directed to plant tails and represents a final gold loss, along with the CIL tail residue post cyanide destruction.

    The following sections will discuss both these aspects of the definition of final gold recovery from the proposed Donlin processing facility.

         
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    Please note for the purposes of this document the terminology MCF2 and MF2 are interchangeable as they pertain to the key equipment within the Donlin flowsheet.

       
    13.2.1

    Flotation

       

    The MCF2 flowsheet is assumed to be the basis of the flotation circuit design and for all recovery figures referenced in this section, both as bench testing and pilot testing.

       

    In late 2006, sections of the Donlin deposit were identified as being geologically weathered (altered). While the extent of the geological oxidation is relatively low, there is still potential for significant effects on flotation recovery. Good flotation performance relies on the existence of un-oxidized sulphide surfaces for flotation collector adherence. Minor surface oxidation of sulphide particles can strongly affect recovery of gold to flotation concentrate performance. This is particularly the case for Donlin, where the fine nature of the arsenopyrite mineralization means that there is a potentially greater exposure of the sulphides to minor particle surface oxidation.

       

    In anticipation of this potential issue, flotation pilot and variability testwork was undertaken separately on fresh ores or oxidized (partially) ores. The pilot flotation testing was undertaken exclusively on non-oxidized ores, as this represents the majority of the Donlin orebody (93%). All partially oxidized ores were excluded from the pilot composite sample.

       

    Variability bench flotation testing was then used to define the relative performance of the (partially) oxidized ores so that the overall deposit flotation recovery could be corrected to account for this smaller oxidized component of the orebody.

       

    Sulphide Ore MCF2 Pilot-Plant Results

       

    The MCF2 pilot-plant testwork campaign is recommended as the basis for selecting the overall recovery for flotation. The MCF2 pilot flotation test program simulates the entire FSU MCF2 flotation flowsheet, incorporating the cleaning circuit, cleaner scavenger, and recirculation of the cleaner scavenger tails back to the secondary mill. This scope exceeds the work done in the batch flotation variability tests, which for practicality consist of roughing stages only.

       

    The design of the flotation circuit is directly scaled from, and based upon, the pilot- plant flowsheet, incorporating an identical cleaner and cleaner-scavenger recycle circuit configuration.

       

    The sample selected for the MCF2 pilot campaign was sourced from newly recovered HQ core drilled in 2006 and was compiled as a blend to represent the known LOM composition at the time of compositing based on rock lithology. Notably the sample compositing was not based on geological domain categorization, and known geologically oxidized affected core samples were purposely excluded from this composite sample.


         
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    The establishment of gold recovery from the MCF2 pilot program is achieved by means of fitting a linear regression line through all the MCF2 pilot survey results.

    No MCF2 test results are excluded from the data set used for the linear regression fit (Figure 13-14), and the gold recovery survey calculations incorporate both the primary rougher concentrate and the secondary rougher cleaner concentrate. Cleaner scavenger concentrate was recirculated to the feed of the secondary rougher.

    At the design concentrate target of 7% (total) sulphur, based upon linear regression fit to the MCF2 pilot-plant results, gold recovery of the blended composite sample tested is 94.64% . This recovery forms the basis of the flotation gold recovery estimate but must be adjusted to account for effects of geological domain and alteration (oxidation extent).

    Variability Testing – Unoxidized Ores

    Two different methods of compiling and assessing the non-oxidized/intrusive ore variability flotation recoveries are discussed in this section, being based upon either lithology or geological domain. It is recommended that the intrusive variability samples be grouped on the basis of geological domain, for the following reasons:

  •  
  • Statistical variance of the variability test flotation recovery results grouped by geological domain is lower than sets of results grouped by lithology. Average statistical variance by lithology is 2.48%R, compared to lithology groupings at 2.64%R.


         
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    Figure 13-14: MCF2 Pilot-Plant Campaign Survey Results

     

  •  
  • Geological domain assignment is spatially based, so that sample grouping by geological domain would better correct the orebody recovery predictions on a spatial basis to overcome any potential bias through variability sample selection being higher in density in better performing areas of the deposit (i.e., the AMCA deposit area).

       
  •  
  • Calculation of overall flotation recovery by geological domain results in a more conservative estimate of the orebody average recovery compared to that based on lithological grouping.

    The recommended recoveries per geological domain are summarized in Table 13-14.

    Adjusting the MCF2 pilot-plant recovery based on the geological domain variability flotation performance results in a minor upward adjustment of 0.16% for the unaltered ore types. The gold and sulphur head grades of the MCF2 pilot-plant composite are very close to the orebody average.

         
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      Table 13-14: Summary of Average Flotation Recovery in Variability Testwork Program, by Geological Domain
          Average Variability Flotation
        % Tonnes in Orebody Recovery
      Geological Domain (%) (%)
      AKIVIK   4.6 97.61
      400   5.7 96.97
      ACMA 16.2 96.45
      AURORA   4.2 96.22
      VORTEX 11.5 95.19
      LEWIS 25.7 94.87
      GWK 19.2 91.45
      SHL   5.2 89.99
      OXIDE   7.7 81.45
      Overall 100.00 93.65

    Variability Testing – Oxidized Ores

    There is a large variation in flotation test results (i.e., gold recoveries to target concentrate grades) of the ores affected by oxidation. Therefore, to improve the accuracy of the application of the test data to the deposit characteristics, it is recommended that the relationship between sulphur grade and flotation recovery be used as the basis of recovery estimation.

    Oxidation Wireframe Model and Mine Block Model

    To better clarify the tonnes and grades of oxidation-affected ores within the deposit, a geological wireframe was developed by the Donlin geological team. They then used this model to categorize each mining block as either oxidized-affected ore or non-oxidized ore. This was undertaken on the 6 m x 6 m x 6 m block size, where a block was flagged as oxidized if 50% of the block or greater was located within the oxidation wireframe model.

    With the mine blocks defined as such, it was then possible to allocate tonnes and grade of oxidized ore into a mill feed schedule. The oxidized-affected tonnage portion of the deposit is estimated by wireframe modelling to be 7.7.

    Mine Model and Stockpile Oxidation Allowance

    Ores will be stockpiled and subsequently reclaimed for mill feed occur during the course of the mine life. A sulphur degradation and flotation recovery factor of -5% was applied to reclaimed material that was stockpiled for longer than one year.

         
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    Final Flotation Recovery Model Definition

    As discussed previously, it is recommended that the adjusted MCF2 pilot-plant performance results be used for the un-altered (un-oxidized) areas of the deposit. The performance of the altered oxides is not changed. Because the pilot-plant sample was originally composited on the basis of lithological domain, rather than geological domain, the results have been adjusted slightly to account for the variation in content between the pilot-plant sample and the latest estimate for the orebody.

    The results from the variability testwork program can be used to assign different geological domains within the mine plan to improve the estimation of time-based cash flow from the mine. However, a slight adjustment is again required to match the MCF2 pilot result.

    The proportionally adjusted flotation recoveries by geological domain are summarized in Table 13-15; these numbers have been adjusted slightly from FSU1 to reflect the ore release sequence in the FSU2 mine plan. It is recommended that these recovery values be used within the mine plan where the geological domains are separately defined on a period-by-period basis.

    No clear relationships between gold, arsenic, or sulphur head grades or in flotation recovery were able to be identified in the variability testwork.

    Considering the flotation recoveries of non-oxidized ores by geological domain, of the separately defined oxidized ores, and the impact of stockpiling, the entire flotation circuit can be determined. The overall LOM flotation recovery, based on the ore feed delivery schedule also taking into account feed grade, is calculated to be 93.0% .

    The production plan has been optimized to maximize the feed grade and cash flow in the early years of processing. This has been accomplished through the use of sequential mining of both the Lewis and ACMA pits along with the use of stockpiles and associated ore rehandling,

    Figure 13-15 shows the flotation recovery trend across the mine life. Recovery drops in 2043 due to an increase in oxide content in addition to the Lewis intrusive component in the mill feed. Recoveries steadily improve from 2019 through to 2027 due to decreasing oxide content in the mill feed. Flotation recoveries trend downwards from 2028 to the end of the mine life in 2045 due to increased content of oxide and Lewis intrusive material and a decrease in ACMA intrusive content. In 2022 a dip in flotation recovery occurs due to a significant increase in processing of oxidized stockpiled material.

         
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      Table 13-15: Summary of Flotation Recovery in Variability Testwork Program by Geological Domain and Adjusted to MCF2 Pilot Result
        % Tonnes in Orebody Adjusted Recovery to MCF2 Pilot Result
      Geological Domain (%) (%)
      AKIVIK   4.6 97.77
      400   5.7 97.13
      ACMA 16.2 96.61
      AURORA   4.2 96.38
      VORTEX 11.5 95.34
      LEWIS 25.7 95.03
      GWK 19.2 91.61
      SHL   5.2 89.99
      OXIDE   7.7 81.45
      Overall 100.0 93.81

    Figure 13-15: Flotation Recovery Trend throughout Mine Life

    13.2.2

    Pressure Oxidation

       

    Introduction

       

    Considering the varying nature of test results from the different test methodologies (bench, semi-continuous, or continuous), it is recommended that the continuous testing method be used as the basis of estimating gold recovery performances from the Donlin ores. Continuous pilot testing is considered to best represent a real continuous autoclave operation and is the basis that Hatch has used for evaluating the expected gold recovery from the pressure oxidation/CIL circuit. The Hatch evaluation predicts that an overall 96.6% recovery can be achieved from the POX/CIL circuit under the testwork conditions.


         
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    To undertake variability testing on individual ore types, a test method that consumes less sample, such as bench-scale testing (BTAC) or semi-continuous testing (SCAC), must be used.

    It is the degree of oxidation, the residence time required to achieve full oxidation, and the ability to control the autoclave oxidation level that influence the chemistry in the test autoclave vessel and therefore the final gold recovery.

    Pilot-Plant Testing and Design Considerations

    Hatch has reviewed the pilot autoclave testwork completed to date on the Donlin Gold Project and has concluded that, based upon the proposed plant design, an overall gold recovery of 96.6% can be achieved through the POX/CIL circuits on a continuous and long-term basis. This evaluation assumes that the concentrate sample used for piloting during the 2007 Phase 2 test program is representative of the overall orebody composition and that the conditions of the testwork are maintained in the final design. The result of this review has been carried forward to FSU2.

    The deleterious components in the feed to pressure oxidation, based on the historical BTAC testwork, are the sedimentary ores themselves (shale and greywacke rock lithologies).

    The proposed autoclave design for Donlin incorporates a level control system on the slurry content of the pressure vessel to permit direct control of operating residence time in the autoclave. Therefore, based on the pilot testwork and the flowsheet design, Hatch recommends that an overall POX and CIL recovery of target of 96.6% be adopted and be applied equally to all ore types fed to the autoclave.

    The nature of the orebody within the deposit is such that sedimentary ores will always be blended into the mill feed. The actual lithologies of the intrusives present within the blend may change on a macroscopic basis, but the sedimentary content from greywacke and shale will remain an ongoing part of the blend. The inventory within concentrate storage tanks ahead of the autoclave feed tank will provide a mechanism to smooth short-term variability of sedimentary content.

         
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    13.2.3  

    Overall Plant Gold Recovery

       

    To determine the overall plant recovery, both pressure oxidation and flotation need to be considered together. The overall plant recovery averages 89.83% over the LOM.

       
    13.3

    Metallurgical Variability

       

    Variability testing undertaken for the Project is discussed in Sections 13.1 and 13.2 under the various testwork programs.

       
    13.4

    Deleterious Elements

       

    The likely deleterious elements identified in the metallurgical testwork programs are discussed in Sections 13.1 and 13.2 under the various testwork program headings.

       
    13.5

    Comments on Section 13

       

    In the opinion of the AMEC QPs, the following conclusions are appropriate:


  •  
  • The metallurgical test results and process design described in this report are essentially the same as those presented in the 2008 Technical Report prepared on the Project.

       
  •  
  • Metallurgical testwork and associated analytical procedures were performed by recognized testing facilities, and the tests performed were appropriate to the mineralization type

       
  •  
  • Samples selected for testing were representative of the various types and styles of mineralization at Donlin. Samples were selected from a range of depths within the deposit. Sufficient samples were taken so that tests were performed on sufficient sample mass

       
  •  
  • Mineralogical studies have shown that the gold is not visible. Testwork analysis indicates a high level of association of gold with arsenopyrite. Other sulphides such as pyrite and marcasite are also present, with reduced tenors of gold.

       
  •  
  • Organic carbon, a potential preg robber, is present in the sedimentary ore. It is also present at lower levels in the intrusive ores, believed to be in the form of well- ordered graphite. This form of organic carbon is possibly less likely to preg-rob

       
  •  
  • Testwork completed by SGS-Lakefield Research, Hazen Research, and G&T Metallurgical Services (G&T) under Barrick’s supervision has shown that the Donlin ore requires pre-treatment prior to cyanidation to recover the gold. Process development work has determined that pressure oxidation is the preferred method of pre-treatment. Extensive testwork on composites has shown that acceptable gold recoveries can be produced through a combination of flotation pre- concentration, POX, and CIL cyanidation.


         
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  •  
  • Alternative flowsheets to flotation-POX-CIL were considered, including whole ore pressure oxidation, roasting a flotation concentrate, and bio-oxidation (BIOX). None of these proved to be a viable economic alternative to the flotation-POX-CIL route

         
     
  •  
  • No new metallurgical testwork programs have been carried out.

       
  •  
  • The average Bond work index for the ore is in the range of 15 kWh/t. Flotation work has shown that kinetics are initially rapid, but to achieve high recoveries, a combined primary and secondary rougher residence time over 100 minutes, together with a high reagent loading in the system, is required. Clay-like minerals will affect slurry viscosity and settling. Slurry density in the underflow will be less than 50% solids for the concentrate thickeners.

       
  •  
  • Partially geologically oxidized (altered) ore in the deposit, up to 7% of the mill feed, is the key non-performing ore type in the flotation circuit. Degradation of the sulphide ore via oxidation in the stockpile will also affect the flotation recovery, applied as 5% recovery loss within flotation on all ores stockpiled for longer than one year.

       
  •  
  • Pressure oxidation (POX) has been shown to be successful in releasing the valuable constituents, under certain conditions. To optimize oxidation conditions, the water systems design has been modified to use the highest-quality water in the oxidation circuit. The autoclave design incorporates variable level control to provide better control over operating residence time.

       
  •  
  • Areas of design modification from FSU1 include detailing the mercury abatement systems in the gold cyanidation, elution, and refining circuits, and also the treatment of off-gases from the pressure oxidation (POX) process to meet more stringent air emissions legislation

       
  •  
  • Metal production and recoveries from the flotation process have been adjusted upward slightly to account for changes in the new mine plan related to ore-type sequencing, and the Prussian blue process has been removed as a backup system to the SO2/Air method for cyanide detoxification

       
  •  
  • Air flotation using the MCF2 flowsheet provides an estimated life-of-mine (LOM) average of 93.0% recovery, with CIL recoveries after POX at approximately 96.6% for an estimated combined plant total gold recovery of 89.8%. The concentrate pull will vary from 15% to 17% and that will result in a concentrate grade of 13.0 to 12.7 g/t Au.


         
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    14.0

    MINERAL RESOURCE ESTIMATES

       
    14.1

    Key Assumptions/Basis of Estimate

       

    The geologic model and resource model (DC9) are based on all drilling through the 2009 drilling campaign. The cut-off date for the DC9 model was 1 November 2009, and no new information was added after that time. Note that the geologic model update includes only the felsic dikes and sills and the overburden model. Shale wireframes were not updated. Also of note is that additional assays for arsenic were provided 18 July 2008, with an updated block model estimate for arsenic only provided on the basis of that data.

       

    The mineral estimate was prepared by Barrick and audited by AMEC. Composites and 3D solid models were constructed utilizing Vulcan® commercial mine modelling software. The models extend a total of 13,200 ft (4,020 m) in the north–south direction, 13,200 ft (4,020 m) in the east–west direction and a total of 3,150 ft (960 m) in the vertical direction. The block model was created with a constant block size of 20 ft x 20 ft x 20 ft (6 m x 6 m x 6 m).

       

    The coordinate system used for resource modelling is NAD83. Resource estimation uses a topographic surface derived from a 2004 survey by Aero-metric. The survey has an accuracy of ±6.6 ft (±2 m).

       

    The Mineral Resource estimates were prepared by Mr. Chris Valorose of Barrick with reference to the Canadian Institute of Mining Metallurgy and Petroleum (CIM) Definition Standards (2010) and CIM Best Practice Guidelines.

       
    14.2

    Geological Models

       

    Three-dimensional solids for the geological model were constructed from polygons resulting from geologic interpretation of cross-section and level plans. Tools available in Vulcan® commercial mine design software were used to create the polygons representing the geometry of the intrusive sills and dikes. These were digitized directly on a computer screen snapping to drill holes in section.

       

    Once digitized, the polygons were used to develop 3D wireframe solids to incorporate geologic control into the grade model for the intrusive rocks. The solids were validated and checked for crossing errors, consistency, and closure prior to use. The solids were used to assign the corresponding geological code to the 3D block model.


         
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    To limit the size of the model, blocks were assigned a default code of greywacke (ROCK = 93) and were then overprinted with rock values according to the established priorities. Rocks assigned a greywacke code had the lowest priority value.

       

    Rock codes are held as three variable types in the block model. The ‘ROCK’ variable is assigned with values that include codes for lithology, overburden, and air. The ‘ROCK_EST’ variable, although similar to ‘ROCK’ does not include overburden and topography in order to allow unrestricted estimation of blocks at or near the topographic surface. The ‘ROCK_MINE’ variable holds a simplified rock code nomenclature.

       

    Nine mineral and geological domains were assigned to the database as indicated in Figure 14-1.

       

    The geotechnical domain zone codes were input into the resource model, as required for the LG pit optimization, using domain solids provided by BGC on 27 June 2008.

       

    A waste rock management category (WRMC) model was coded to identify overburden from the other WRMC codes.

       
    14.3

    Exploratory Data Analysis

       

    To better understand the deposit, AMEC performed several additional EDA studies with the following conclusions:


  •  
  • Gold mineralization appears to reflect a single population that approximates a log normal distribution and includes a high percentage of sub-economic grades.

       
  •  
  • Arsenic histograms suggest two separate near-log normal populations of mineralization, one less than 300 ppm and one greater than 300 ppm.

       
  •  
  • Mercury shows a near-log normal distribution with two distinct kinks that probably reflect different detection limits.

       
  •  
  • Arsenic shows a near-log normal distribution with two distinct kinks that probably reflect different detection limits.