Boddington Operations Western Australia NI 43-101 Technical Report

CERTIFICATE OF QUALIFIED PERSON

6363 South Fiddlers Green Circle Greenwood Village, Colorado, USA Phone: 001 303 708 4599

I, Donald Charles Doe, am employed as Group Executive, Reserves with Newmont Mining Corporation (Newmont). This certificate applies to the technical report titled “Boddington Operations, Western Australia, Australia” with an effective date of 31 December 2018 (the “technical report”). I am a Registered Member of the Society for Mining, Metallurgy and Exploration, 4044636 and a Professional Engineer in Alberta, 44399. I graduated from Mining Engineering at the University of Alberta in 1986 (B.Sc.) and in 1991 (M.Sc.). I have practiced my profession for over 31 years. I have been directly involved in mine engineering, mine operations, mine design, mineral reserve estimation, mineral reserve audits, in consulting and corporate positions within the mining industry in Canada, the United States, Peru, Australia, Ghana, Suriname, New Zealand and Indonesia. In my current senior technical management role at Newmont, I am accountable for Newmont’s governance system for Mineral Resources and Mineral Reserves, including the multi-discipline inputs to those estimates, and I approve the annual estimates for Mineral Resources and Mineral Reserves provided by Newmont sites and projects, along with their compliance to Newmont’s internal policies and the required controls, standards and guidelines for the Securities Regulatory requirements under which Newmont reports. As a result of my experience and qualifications, I am a Qualified Person for the content in the technical report, as the term Qualified Person is defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101). I visited the Boddington site most recently on 18 April 2018. I am responsible for all sections of the technical report. I am not independent of Newmont as independence is described by Section 1.5 of NI 43–101. I have been involved with the Boddington property in a technical review and support role from 2009 to 2013 and in a reserves governance role since 2014. I have read NI 43–101 and all sections of the technical report have been prepared in compliance with that Instrument. As of the effective date of the technical report, to the best of my knowledge, information and belief, all sections of the technical report contain all scientific and technical information that is required to be disclosed to make the technical report not misleading. Dated: 4 March 2019 “Signed and sealed” Donald Charles Doe, Registered Member 4044636 SME

Boddington Operations Western Australia NI 43-101 Technical Report

C ONTENTS

1.0 SUMMARY ...... 1 1.1 Introduction ...... 1 1.2 Project Setting...... 1 1.3 Development History ...... 1 1.4 Mineral Tenure and Surface Rights ...... 2 1.4.1 Tenure History ...... 2 1.4.2 Mineral Tenure ...... 2 1.4.3 Surface Rights ...... 3 1.4.4 Native Title ...... 3 1.5 Royalties ...... 3 1.6 Geology and Mineralization ...... 3 1.7 Exploration ...... 4 1.8 Drilling and Sampling ...... 4 1.9 Data Verification ...... 5 1.10 Metallurgical Testwork ...... 6 1.11 Mineral Resource Estimates ...... 6 1.12 Mineral Resource Statement ...... 7 1.13 Mineral Reserves Estimates ...... 8 1.14 Mineral Reserves Statement ...... 9 1.15 Mine Plan ...... 11 1.16 Recovery Plan...... 11 1.17 Infrastructure ...... 12 1.18 Marketing ...... 12 1.19 Environmental, Permitting and Social Considerations ...... 12 1.20 Sustaining Capital Costs ...... 12 1.21 Operating Costs ...... 13 1.22 Economic Analysis ...... 13 1.23 Interpretation and Conclusions ...... 13 1.24 Recommendations ...... 15 2.0 INTRODUCTION ...... 16 2.1 Terms of Reference ...... 17 2.2 Qualified Person ...... 17 2.3 Site Visits and Scope of Personal Inspection ...... 17 2.4 Effective Dates ...... 18 2.5 Information Sources and References ...... 18 2.6 Previous Technical Reports ...... 19 3.0 RELIANCE ON OTHER EXPERTS ...... 20 4.0 PROPERTY DESCRIPTION AND LOCATION...... 21 4.1 Location ...... 21 4.2 Tenure History ...... 21 4.3 Property Agreements ...... 22 4.3.1 Background ...... 22

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4.3.2 Management Agreements ...... 22 4.4 Current Mineral Tenure ...... 23 4.5 Surface Rights ...... 27 4.6 Royalties and Encumbrances ...... 27 4.7 Native Title ...... 27 4.8 Permits ...... 30 4.9 Environmental Considerations ...... 30 4.10 Social Considerations ...... 30 4.11 Project Risks ...... 30 4.12 Comments on Property Description and Location ...... 31 5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...... 32 5.1 Accessibility ...... 32 5.1.1 Road ...... 32 5.1.2 Port ...... 32 5.2 Climate ...... 32 5.3 Local Resources and Infrastructure ...... 32 5.4 Physiography ...... 33 5.5 Comments on Accessibility, Climate, Local Resources, Infrastructure and Physiography ...... 34 6.0 HISTORY ...... 35 7.0 GEOLOGICAL SETTING AND MINERALIZATION ...... 37 7.1 Regional Geology ...... 37 7.2 Project and Deposit Geology ...... 39 7.2.1 Wandoo South ...... 39 7.2.2 Wandoo North ...... 42 7.3 Mineralization ...... 43 7.4 Comment on Geological Setting and Mineralization ...... 45 8.0 DEPOSIT TYPES ...... 46 8.1 Gold Porphyry Deposits ...... 46 8.2 Shear-zone Hosted Deposits ...... 47 8.3 Comment on Deposit Types ...... 47 9.0 EXPLORATION ...... 48 9.1 Grids and Surveys ...... 48 9.2 Geological Mapping ...... 48 9.3 Geochemical Sampling ...... 48 9.4 Geophysics ...... 50 9.5 Pits and Trenches ...... 50 9.6 Petrology, Mineralogy and Other Research Studies ...... 50 9.7 Exploration Potential ...... 53 9.8 Comment on Exploration ...... 53 10.0 DRILLING ...... 54 10.1 Drilling Methods ...... 54 10.1.1 Vacuum Drilling ...... 54

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10.1.2 RAB Drilling ...... 54 10.1.3 AC Drilling ...... 59 10.1.4 RC Drilling ...... 59 10.1.5 Core Drilling ...... 62 10.2 Geotechnical and Hydrogeological Drilling ...... 64 10.3 Geological Logging ...... 64 10.3.1 RC Drilling ...... 64 10.3.2 Core Drilling ...... 64 10.4 Recovery ...... 65 10.4.1 RC Drilling ...... 65 10.4.2 Core Drilling ...... 65 10.5 Collar Surveys...... 65 10.5.1 RC Drilling ...... 65 10.5.2 Core Drilling ...... 65 10.6 Downhole Surveys ...... 65 10.6.1 RC Drilling ...... 65 10.6.2 Core Drilling ...... 66 10.6.3 Declinations ...... 66 10.7 Drilling Used in Mineral Resource Estimation ...... 66 10.8 Sample Length/True Thickness ...... 66 10.9 Comments on Drilling ...... 70 11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY ...... 71 11.1 Sampling Methods ...... 71 11.1.1 Geochemical Sampling ...... 71 11.1.2 Trench Sampling ...... 71 11.1.3 Vacuum Sampling ...... 71 11.1.4 Rotary Air Blast Sampling ...... 71 11.1.5 AC Sampling ...... 71 11.1.6 Reverse Circulation Sampling ...... 72 11.1.7 Core Sampling ...... 72 11.2 Density Determinations ...... 73 11.2.1 Historical Determinations ...... 73 11.2.2 Recent Determinations ...... 74 11.3 Analytical and Test Laboratories ...... 75 11.4 Sample Preparation and Analysis ...... 75 11.4.1 Sample Preparation ...... 75 11.4.2 Analyses ...... 76 11.4.3 Additional Analytical Determinations ...... 77 11.5 QA/QC ...... 78 11.5.1 Historical QA/QC ...... 78 11.5.2 Current QA/QC (2006 to date) ...... 79 11.6 Databases ...... 79 11.7 Sample Security ...... 79 11.8 Sample Storage ...... 79 11.9 Comments on Sample Analysis, QA/QC and Security ...... 80 12.0 DATA VERIFICATION ...... 81 12.1 Site Visits ...... 81

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12.2 Laboratory Visits ...... 81 12.3 Analytical Reviews ...... 81 12.3.1 Historical Round Robins ...... 81 12.3.2 Round Robins, 1988 to 1990 ...... 82 12.3.3 Round Robins, 1991 to 1995 ...... 82 12.3.4 Round Robins, 1995 to June 2000 ...... 83 12.3.5 Geostats Pty Ltd Round Robins, 1993 to 2000 ...... 83 12.3.6 Analytical Repeatability Reviews ...... 83 12.4 Data Reviews ...... 83 12.4.1 External Reviews ...... 83 12.4.2 Internal Reviews ...... 85 12.4.3 Data Validation ...... 86 12.4.4 Sample Preparation ...... 86 12.4.5 Crushing and Grinding Performance ...... 86 12.4.6 Drill Bias Review ...... 86 12.5 Comments on Data Verification ...... 87 13.0 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 89 13.1 Metallurgical Testwork History ...... 89 13.1.1 Variability Composites ...... 89 13.1.2 Comminution Testwork ...... 91 13.1.3 Tailings Characterization and Assessments ...... 91 13.1.4 Metallurgical Performance ...... 92 13.1.5 Gold Revenue...... 93 13.1.6 Testwork and Feasibility Study Conclusions ...... 93 13.1.7 Metallurgical Testwork – Post Feasibility Study ...... 94 13.2 Mill Throughput Modelling ...... 99 13.3 Throughput Assumptions ...... 99 13.4 Recovery Estimates ...... 100 13.5 Metallurgical Variability ...... 100 13.6 Deleterious Elements ...... 101 13.7 Comment on Mineral Processing and Metallurgical Testwork ...... 101 14.0 MINERAL RESOURCE ESTIMATES ...... 102 14.1 Introduction ...... 102 14.2 Geological and Mineralization Models ...... 102 14.3 Bulk Density Model ...... 102 14.4 Geotechnical Model ...... 104 14.5 Acid Rock Drainage Model ...... 104 14.6 Composites ...... 105 14.7 Statistical and Exploratory Data Analysis ...... 105 14.8 Variography ...... 106 14.9 Estimation/Interpolation Methods ...... 107 14.10 Post-processing ...... 108 14.11 Block Model Validation ...... 109 14.12 Classification of Mineral Resources ...... 109 14.13 Reasonable Prospects of Eventual Economic Extraction ...... 110 14.14 Mineral Resource Statement ...... 111

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14.15 Factors That May Affect the Mineral Resource Estimate ...... 113 14.16 Comments on Mineral Resource Estimates ...... 113 15.0 MINERAL RESERVES ESTIMATES ...... 114 15.1 Mineral Reserves Schedule ...... 114 15.2 Mineral Reserves Statement ...... 114 15.3 Factors That May Affect the Mineral Reserves Estimate ...... 116 15.4 Caution Regarding Forward-Looking Information ...... 116 15.5 Pit Optimization ...... 116 15.5.1 Metallurgical Recovery Assumptions ...... 116 15.5.2 Geotechnical Assumptions and Pit Slope Configuration ...... 118 15.5.3 Dilution and Mining Losses ...... 122 15.5.4 NSR Cut-off ...... 123 15.6 Pit Optimisation ...... 124 15.7 Pit Designs ...... 126 15.8 Comments on Mineral Reserves ...... 127 16.0 MINING METHODS ...... 128 16.1 Description of Mining Method ...... 128 16.2 Hydrology ...... 129 16.3 Production Schedule ...... 130 16.4 Drilling and Blasting ...... 131 16.5 Mining Equipment ...... 131 16.6 Mine Plan Considerations ...... 132 16.7 Comments on Mining Methods ...... 132 17.0 RECOVERY METHODS ...... 133 17.1 Process Flow Sheet ...... 133 17.2 Plant Design ...... 134 17.2.1 Coarse Crushing (in mining area) ...... 135 17.2.2 Fine Crushing and Screening (process plant) ...... 135 17.2.3 Gold and Copper Recovery to Concentrate ...... 136 17.2.4 Gold Smelting and Bullion Production ...... 136 17.3 Plant Performance and Process Optimisation ...... 136 17.4 Product/Materials Handling ...... 140 17.5 Energy, Water and Process Materials Requirements ...... 141 17.6 Comments on Recovery Methods ...... 141 18.0 PROJECT INFRASTRUCTURE ...... 142 18.1 Overview ...... 142 18.2 Roads and Logistics ...... 142 18.3 Waste Dump Facilities ...... 144 18.4 Tailings Storage Facilities ...... 146 18.5 Water Management ...... 146 18.6 Camps and Accommodation ...... 147 18.7 Power and Electrical ...... 147 18.8 Communications ...... 147 18.9 Comments on Infrastructure ...... 147

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19.0 MARKET STUDIES AND CONTRACTS ...... 148 19.1 Market Studies and Contracts ...... 148 19.2 Commodity Price Projections ...... 148 19.3 Comments on Market Studies and Contracts ...... 149 20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT ...... 150 20.1 Baseline Studies ...... 150 20.2 Environmental Considerations ...... 150 20.3 Closure Plan ...... 151 20.4 Permitting ...... 153 20.4.1 Initial Permit ...... 153 20.4.2 Interim Permit ...... 154 20.4.3 Boddington LOM Expansion Project ...... 154 20.5 Considerations of Social and Community Impacts ...... 154 20.6 Comment on Environmental Studies, Permitting and Social or Community Impact ..... 155 21.0 CAPITAL AND OPERATING COSTS ...... 156 21.1 Operating Cost ...... 156 21.2 Capital Expenditure ...... 156 21.2.1 Mining Capital Expenditure ...... 156 21.2.2 Processing Capital Expenditure ...... 157 21.2.3 G&A Capital Expenditure ...... 157 21.3 Closure Costs ...... 158 21.4 Comments on Capital and Operating Costs ...... 158 22.0 ECONOMIC ANALYSIS ...... 159 23.0 ADJACENT PROPERTIES ...... 160 24.0 OTHER RELEVANT DATA AND INFORMATION ...... 161 25.0 INTERPRETATION AND CONCLUSIONS ...... 162 26.0 RECOMMENDATIONS ...... 164 27.0 REFERENCES ...... 165 27.1 Bibliography ...... 165 27.2 Glossary of Abbreviations, Symbols and Units ...... 170

T ABLES Table 1-1: 2018 Mineral Resource Economic Parameters and Cut-offs ...... 7 Table 1-2: Mineral Resource – Gold and Copper at the Effective Date of 31 December 2018 ...... 7 Table 1-3: 2018 Mineral Reserves Economic Parameters and Cut-offs ...... 8 Table 1-4: Mineral Reserves – Gold and Copper at the Effective Date of 31 December 2018 ...... 10 Table 4-1: Transactions Through Which Newmont Acquired Its Interest in the BGMJV ...... 22 Table 4-2: Mineral Tenure Table ...... 24 Table 9-1: Geophysical Surveys ...... 51 Table 10-1: Project Drilling by Ownership ...... 55 Table 10-2: Project Drilling Supporting the Mineral Resource Estimate by Year Drilled ...... 55 Table 10-3: Drill Holes Used in Mineral Resource Estimation ...... 67 Table 11-1: South Pit Density Values Assigned ...... 74

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Table 11-2: North Pit Density Values Assigned ...... 74 Table 11-3: Analytical Laboratory Dates ...... 76 Table 11-4: Sampling Protocol for Drill Core ...... 76 Table 11-5: Lower Detection Limits for Analytical Suite ...... 78 Table 13-1: Variability Composite Results ...... 89 Table 13-2: Variability Composite Results (LSK Composites) ...... 90 Table 13-3: Comminution Results ...... 91 Table 14-1: High-grade Cuts By Domain Code ...... 105 Table 14-2: 2018 Mineral Resources Economic Parameters and Cut-offs ...... 110 Table 14-3: Mineral Resource – Gold and Copper at the Effective Date of 31 December 2018 ...... 112 Table 15-1: Mineral Reserves – Gold and Copper at the Effective Date of 31 December 2018 ...... 115 Table 15-2: Mineral Reserves Economic Assumptions ...... 116 Table 15-3: Gold and Copper Metallurgical Recovery Functions ...... 118 Table 15-4: Feasibility Study Pit Slope Design Parameters ...... 118 Table 15-5: 2018 Mineral Reserves Economic Parameters and Cut-offs ...... 123 Table 16-1: 2018 Mineral Reserves Mine Schedule by Pits (kt) ...... 130 Table 17-1: Processed 2018 Full Year Results ...... 137 Table 17-2: Current and Projected Mill Utilization and Throughput ...... 138 Table 17-3: Projected Gold and Copper Recoveries and Copper Concentrate Grades ...... 140 Table 18-1: Waste Dump Capacity ...... 145 Table 21-1: Operating and Capital Unit Cost ...... 156 Table 21-2: 2019BP LOM Capital Expenditure Summary ...... 156 Table 21-3: 2019BP Mining Sustaining Capital Summary ...... 156 Table 21-4: 2019BP Processing Sustaining Capital Summary ...... 157 Table 21-5: 2019BP G&A Sustaining Capital Summary ...... 157

F IGURES Figure 2-1: Project Location Plan ...... 16 Figure 4-1: Mineral Tenure Map ...... 26 Figure 4-2: Surface Rights Map (Regional) ...... 28 Figure 4-3: Land Ownership Map (Mining Operation) ...... 29 Figure 7-1: Regional Geological Map ...... 38 Figure 7-2: Geology Map at 150 mRL ...... 40 Figure 7-3: Interpretative Scaled Cross-Section across the South Pit Area ...... 41 Figure 9-1: Geochemical Sample Locations ...... 49 Figure 9-2: Location Map of Major Geophysical Surveys ...... 52 Figure 10-1: Vacuum Drill Collar Location Map...... 57 Figure 10-2: RAB Drill Collar Location Map ...... 58 Figure 10-3: AC Drill Collar Location Plan ...... 60 Figure 10-4: RC Drill Collar Location Map ...... 61 Figure 10-5: Core Drill Collar Location Map ...... 63 Figure 10-6: Collar Locations of Drilling Supporting Mineral Resource Estimates ...... 68 Figure 10-7: Geological Cross-Section at 10600N (looking towards 290°) ...... 69 Figure 13-1: Metallurgical Domains ...... 90 Figure 13-2: 2011 Long Section Showing North Pit Metallurgical Sample Locations (looking east) ...... 95 Figure 13-3: 2018 Gold Recovery YTD Reconciliation vs Budget ...... 97 Figure 13-4: 2018 Copper Recovery YTD Reconciliation vs Budget ...... 98 Figure 13-5: 2018 Copper Recovery Model Performance (Actual EOM Reconciled Data vs 2016 Model) 98

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Figure 13-6: 2018 Gold Recovery Model Performance (Actual EOM Reconciled Data vs 2016 Model) .... 99 Figure 14-1: Gold Grade Plan at 0mRL ...... 103 Figure 14-2: Cross-section North Pit (oblique section A-A’ looking west) ...... 103 Figure 14-3: Cross-section South Pit (oblique section B-B’ looking west) ...... 103 Figure 15-1: Actual versus Model Recovery Variance ...... 117 Figure 15-2: Geotechnical Domains – North Pit ...... 119 Figure 15-3: Geotechnical Domains – South Pit ...... 120 Figure 15-4: Recommended Slope Profiles for North Pit Area ...... 121 Figure 15-5: Recommended Slope Profiles for South Pit Area ...... 122 Figure 15-6: Mineral Reserves North Pit Optimisation Curves ...... 125 Figure 15-7: Mineral Reserves South Pit Optimisation Curves ...... 125 Figure 15-8: 2018 Mineral Reserves and Mineral Resource Pits...... 127 Figure 16-1: Layout of Project Mining Area ...... 129 Figure 17-1: Overall Process Flowsheet ...... 134 Figure 17-2: Mill Monthly Throughput Since Start-up ...... 137 Figure 17-3: Mill Monthly Utilisation Since Start-up ...... 138 Figure 17-4: Total Gold Recovery Showing Split Between Concentrate and Leach...... 139 Figure 17-5: Total Copper Recovery ...... 139 Figure 17-6: Monthly Copper Concentrate Cu Grade ...... 140 Figure 18-1: Site Infrastructure Layout ...... 143 Figure 18-2: Processing Plant Major Infrastructure and Layout ...... 144

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1.0 SUMMARY

1.1 Introduction Newmont Mining Corporation (Newmont) has prepared a Technical Report entitled “Technical Report, Boddington Gold–Copper Project, Western Australia” (the Report) for the Boddington gold–copper operations (referred to as Boddington or the Project), located in Western Australia (WA). Mr. Donald Doe, Registered Member of the Society for Mining, Metallurgy & Exploration (RM- SME), Group Executive, Reserves at Newmont, is the Qualified Person (QP) for the Report, as defined in National Instrument (NI) 43-101 Standards of Disclosure for Mineral Projects. This Report supports declared Mineral Resources and Mineral Reserves and summarizes the Project development and current operations. Newmont will use this Report in support of disclosure and filing requirements with the Canadian securities regulators as specified in Section 4.2 (1) (c) of NI 43-101. This Report will be filed under Newmont’s System for Electronic Document Analysis and Retrieval (SEDAR) profile. The Project is owned and operated by a joint venture (JV) between two indirectly wholly- owned Australian subsidiaries, Newmont Boddington Pty Ltd (66 and 2/3% interest in the Project) and Saddleback Investments Pty Ltd (Saddleback) [33 and 1/3% interest in the Project]. The manager of the JV is also an indirectly wholly-owned company, Newmont Boddington Gold Pty Ltd (NBG). In this Report, the name Newmont is used interchangeably for the parent and subsidiary companies.

1.2 Project Setting The Project is located approximately 130 kilometers (km) southeast of the city of Perth and 17 km northwest of the township of Boddington. The Project is readily accessible from public roads via a bituminized access road installed for the earlier oxide operation. Major roads allow access to and from Perth, Mandurah, and Bunbury facilitating relatively easy access for staff, materials, and for transport of concentrate and doré production from the mine. The climate is Mediterranean, with hot, dry summers and cool, wet winters. Mining operations are conducted year-round.

1.3 Development History Gold–copper mineralization was identified at Boddington in 1980. Open pit mining began in March 1987, with production commencing in August 1987 at a design rate of 3 million tonnes per annum (Mtpa), rising to 4.5 Mtpa in October 1987. A plant expansion to 6 Mtpa was completed in 1989. Increases in production continued, and by 1996 oxide (laterite) and bedrock ores were processed at a rate of 8.6 Mtpa. Oxide resources were depleted in November 2001 and processing ceased on 30 November 2001. In December 2001, the mine was placed on care and maintenance. The plant and infrastructure were decommissioned, and redundant equipment was sold and removed from site. Several feasibility studies were undertaken during the late 1990s to devise an economic method of extracting the gold mineralization from the Wandoo and other bedrock sources. A decision to proceed with Project development was made in the first quarter of 2006. Commercial production began in 2009, and in March 2011 the operation produced its first one

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million ounces (Moz) of gold. At the end of December 2018, the Project had produced approximately 6.90 Moz of gold and 347 thousand tonnes (kt) of copper.

1.4 Mineral Tenure and Surface Rights 1.4.1 Tenure History In 1980, Reynolds Australia Alumina Ltd. (Reynolds), The Shell Company of Australia Ltd. (Shell), BHP Minerals Ltd. (BHP), and Kobe Alumina Associates (Australia) Pty Ltd (Kobe). formed a consortium known as the Worsley Alumina Joint Venture (Worsley JV) to mine bauxite and refine alumina near the township of Boddington. Worsley Alumina Pty Ltd (Worsley) was established to manage the Worsley JV. Separated to the west by a tenement boundary, but covering the same laterite body, were bauxite deposits operated by Alcoa of Australia Ltd. (Alcoa). Both are covered by provisions under State Agreement Acts. Following the discovery of gold mineralization near Boddington, the Worsley JV created the Boddington Gold Mine Joint Venture (BGMJV). The BGMJV Agreement was executed on 31 March 1987. On formation of the BGMJV the manager and participants in the Worsley JV and BGMJV were identical. However, ownership of the BGMJV has changed over time and the Worsley JV and BGMJV now have different and unrelated ownership. The current parties to the BGMJV are Newmont Boddington Pty Ltd (66 and 2/3%) and Saddleback (33 and 1/3%). Both companies are indirectly- wholly owned Newmont subsidiaries. The key tenements within the Project area are held by the Worsley JV and sub-leased to the BGMJV. Under the sub-leases and contractual arrangements with the Worsley JV, the BGMJV is granted the rights to all minerals, other than bauxite within the project area. Given the proximity of the Worsley JV operations and the BGMJV operations and the Worsley JVs rights to bauxite, the BGMJV and Worsley JV are party to a cross-operation agreement to regulate certain aspects of the interactions between the two JVs. 1.4.2 Mineral Tenure The Project has an interest in a total of 89 mining tenements in the Boddington area, including one granted Exploration Licence, five Miscellaneous Licences, three General Purpose Leases, 41 Mining Leases and five Mineral Leases. The remaining 26 leases are at an application stage. The total granted area is approximately 21,249 hectares (ha) and under application area is approximately 60,734 ha. The primary Project area is covered by 13 Mining Leases granted pursuant to the Mining Act 1978 (Western Australia): M70/21–26, M70/564, M70/799, ML264SA, M70/1031, G70/215, and G70/218–219, which together cover an area of approximately 7,599 ha. The five main mining tenements (M70/21–25) were renewed in 2007 for an additional 21-year term. M70/21-25, M70/564 and M70/799 are sub-leased from the Worsley Joint Venturers, while rights for ML264SA(2) and M70/1031 were acquired from Hedges Gold Pty Ltd (Hedges) in 1998. The leases are held in the name of Newmont Boddington Pty Ltd and Saddleback. The region surrounding the Project is a well-known bauxite-alumina mining and production area now held predominantly by either the Worsley JV or the Alcoa Group. The major part of the Mining Act tenure (including M70/21-25, M70/564 and M70/799)

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lies within a State Agreement Area known as the Worsley State Agreement, defined by special lease ML258SA. The remainder of the tenure (including ML264SA and M70/1031) lies within the original boundaries of a State Agreement Area known as Alcoa State Agreement defined by special lease ML1SA. 1.4.3 Surface Rights Newmont has freehold ownership to the land title for most of the Project, covering the eastern and central area of planned operations and infrastructure. Within this freehold land are all the existing residue disposal areas, the plant site except the primary crusher and portion of the overland conveyor, almost all the area of the main open pit from the former oxide operation and all but one of the smaller satellite open pits from the former operation. The western portion of the operational area, adjacent to Newmont’s freehold title, is forest land owned by the State. Mining operations can be conducted in this area with known and manageable additional requirements. Most of the land requirements fall within current or historically freehold land; however, a small part of the Project lies within Crown land (State Forest). 1.4.4 Native Title The Project falls within an area of a native title land claim registered under the Native Title Act (NTA) and referred to as the Gnaarla Karla Booja claim. All relevant requirements for the Project were secured under a voluntary Agreement with the claimant group in 2006. The Moorditj Booja Community Partnership Agreement (MBCPA) has an end date of 31 December 2025.

1.5 Royalties Production royalties are payable to the WA government and are included in the NSR cut-off determination. Royalty payments were first incurred in the second half of 2009, and comprise:  Copper royalty of 5% of the realized copper value, calculated in United States Dollars (US$) and payable in Australian Dollars (AU$);  Silver royalty of 2.5% of the realized silver value, calculated in US$ and payable in AU$;  Gold royalty of 2.5% of the gold value, except that no gold royalty is payable in respect of the first 2,500 ounces (oz) of payable gold produced in any financial year. Silver is not estimated in the Mineral Resource estimate and does not make a material contribution to the economics of the Project.

1.6 Geology and Mineralization The Boddington deposit is hosted within the Saddleback Greenstone Belt, which lies in the southeastern corner of the Archaean Yilgarn Craton. The greenstone belt comprises a steeply-dipping and extensively faulted sequence of sedimentary, felsic to mafic volcanic and pyroclastic rocks that have been metamorphosed to greenschist–amphibolite facies. The Boddington deposit lies within a 6 km strike length of the Wells Formation, consisting of felsic to intermediate volcanic and related intrusive rocks. The Boddington deposit is divided into two portions, Wandoo South and Wandoo North. Wandoo South is centered on a composite diorite stock, the Central Diorite, which has a known strike length of approximately 1,200 m and thicknesses varying from 300 m to 600 m. Wandoo North is dominated by diorites, with lesser fragmental volcanic rocks. The structural

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setting for the gold–copper mineralization is a major regional sinistral-strike-slip regime controlled by east–southeast to west–northwest compression. Alteration types associated with gold and copper mineralization are clinozoisite–biotite– actinolite-sulfide ± silica veins/veinlets/fractures/clots and late actinolite–sulfide veins with clinozoisite–biotite and/or albite alteration haloes. These vein types form the basis of the stockwork mineralization of the Wandoo deposits. The bulk of the gold mineralization is associated with the clinozoisite–biotite–sulfide alteration event. However, the second mineralizing alteration style of late actinolite–sulfide veining contains generally higher gold concentrations.

1.7 Exploration Exploration on the Project has included establishing detailed grids and topographic surveys, geological mapping in small oxide pits, geochemical soil and stream sediment sampling, airborne and ground geophysical surveys, structural, petrology, mineralogy, lithogeochemical and research studies, geotechnical, mining, metallurgical and hydrological studies. These have outlined a large, low-grade bedrock Au–Cu–Ag resource at Wandoo, and several geochemical, geophysical and structural targets that remain to be tested.

1.8 Drilling and Sampling Approximately 32,716 drill holes have been completed for approximately 2.44 million meters (Mm) of drilling in Rotary Air Blast (RAB), Aircore (AC), Reverse Circulation (RC) and core drill holes (refer to Table 10-1). Supporting Mineral Resource estimation are 7,236 drill holes for approximately 1.39 Mm, comprising 2,463 RC, 1,218 grade control RC, and 3,555 core drill holes (refer to Table 10-2). Drill holes are geologically logged and lithology, alteration, structure, texture, mineralization and alteration minerals and intensity, and veins are recorded. Drill holes have digitally- surveyed drill collars, and generally, downhole surveys at intervals ranging from 30 m to 50 m. Recoveries were not routinely measured for RC drilling. However, visual checks during 2006 to 2008 of operating rigs indicated RC recoveries were very good. Diamond core recoveries are routinely close to 100%. RC drilling is typically sampled on 1 m intervals via a cyclone before sample reduction utilizing a riffle splitter or inverted cone splitter. The samples were composited to 2 m intervals of 4 to 6 kilograms (kg). Typically, NQ (47.6 mm diameter pre-1999) drill core was sampled on 2 m intervals and HQ (63.5 mm) core sampled on 1 m intervals. Since 1999, the predominant diamond coring size has been NQ2 (50.7 mm). Geology-based intervals are used for specific areas such as sampling edges of dolerites and large quartz or actinolite veins. Density determinations were collected in the periods 1994 to 2000 and 2006 to 2011, using water-immersion methods (Archimedes principle) [refer to Section 11.2]. Historically, sample analysis has been performed by several independent laboratories, including Classic Comlabs, Genalysis Laboratory (Genalysis), Amdel Laboratory (Amdel), Bureau Veritas Kalassay, and Australian Assay Laboratories (AAL) in Perth, and AAL in the township of Boddington. It is not known if the laboratories were certified at the time. Primary assays were completed by a contract mine laboratory based at Boddington during the period 1985 to 2001. In 1995, the mine laboratory became ISO 9002 accredited.

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Approximately 80% of the pre-2001 analytical data was completed by the mine laboratory. There was no drilling in the period 2001 to end-2005. Since 2006, assays have been undertaken by Genalysis (Intertek) as the primary laboratory, and UltraTrace as the secondary laboratory. Both Perth-based laboratories are currently accredited to ISO/IEC 17025. The Kalassay laboratory, which has been used for grade control analysis since 2008, achieved ISO/IEC 17025 accreditation during 2010. Analysis for gold of the RC samples was by fire assay with atomic absorption spectrophotometry (AAS) of dissolved prill. Copper analysis was by either single acid digest or three-acid digest followed by AAS of solution. Multi-element determination was not undertaken routinely but rather on selected drill holes as part of detailed geological investigations. When used, the multi-element analytical suite requested typically consisted of Ag, As, Bi, Ce, Mo, Ni, Pb, Sb, Ti, W, Y, Zn, and Zr. In 2011-12, regional exploration drill samples were assayed for a suite of 40 elements. The typical analytical suite requested for drill core samples comprised fire assay gold with AAS finish on 50-gram (g) charge and a multi-element suite using four-acid digest with inductively-coupled plasma mass spectrometry (ICP-MS) finish for Cu, S, As, Bi, Mo, Sb, Cd, and Co. Newmont has used a Quality Assurance and Quality Control (QA/QC) program comprising blank, standard and pulp reject samples. Newmont’s QA/QC submission rate meets industry- accepted standards of insertion rates. Standard reference materials were used to monitor the performance of gold and copper analysis. Overall the results were good, with most erroneous results appearing to be due to sample swaps either at Boddington or at the primary laboratory. Blank sample submission indicates limited contamination of samples during the analytical process. Crushing and grinding performance has generally been acceptable. Duplicate samples are not normally included in the sample batch. Typically, NBG reanalyzes pulp rejects, rather than submitting field duplicates. Check assay data have been acceptable for the 2006 to 2015 programs. Extensive review of assay data collected during the operation of the Boddington site laboratory indicated a potential minor bias with pre-2000 analytical results; however, the biases were consistently shown to be minor, and no adjustments have been made to the assay database. Formal systems of continuously monitoring assay quality control by NBG have been in place since January 1989. Systems include review of laboratory performance using methods commissioned by NBG as well as review of the mine laboratory’s internal systems.

1.9 Data Verification The QP has conducted personal inspections of the Project as part of his data verification. Mr. Doe has visited the Project numerous times during his career at Newmont, most recently on 18 April 2018. An extensive program of data verification has been part of the development of the Project, including audits and reviews by internal staff, external consultants, and JV partners during the JV period. The data verification programs undertaken on the data collected provide sufficient confidence in the geological interpretations, the analytical and database quality, and therefore support the use of the data in Mineral Resource and Mineral Reserves estimation.

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1.10 Metallurgical Testwork During feasibility-stage studies from 1997 to 2003, several programs of metallurgical testwork were completed on the Boddington deposit. Eight metallurgical ore types form the basis for ore characterization studies. The annual gold recovery fluctuates between 77.6% and 85.2% due to variations in the ore type being mined. Over the life of the mine the gold recovery averages 83%. The average grade over the life of the mine is 0.67 grams per tonne (g/t) Au. The annual average copper head grade remains constant (between 0.10% and 0.12%) until 2022, when the high-grade Southern Diorite Deeps ore type is processed. The mean grade over the life of the mine is 0.10% Cu. The annual average copper recovery fluctuates between 78.0% and 81.0% due to variations in the type of ore being mined. Over the life of the mine the copper recovery averages 78.0%. Approximately 75% of the Project gold revenue is obtained from gold contained in the exported copper concentrate, while the remaining 25% is obtained from gold bullion produced on site.

The main concentrate penalty charge has been from Al2O3 + MgO, while a low but consistent arsenic penalty exists from ores mined in most areas of the pit. Penalties also apply to bismuth which is intermittent throughout the orebody. Most of the contained bismuth was mined during the first five years of operation.

1.11 Mineral Resource Estimates The geological model developed for the Project has used standard industry methods. Models incorporated lithology, structure, alteration, density, mineralization, geotechnical, acid-rock drainage and metallurgical characteristics, using Vulcan® and MineSight® and Leapfrog Geo® modeling software. The Mineral Resource process included section and plan interpretation with validation of geology and mineralization solids; compositing of analytical data; exploratory data analysis (EDA); variography and establishment of spatial continuity parameters; investigation of, and applicability of, grade caps; examination of interpolation boundaries; establishment and review of waste and metallurgical classifications; selection and evaluation of selective mining unit (SMU) sizes; dilution of estimate by non-mineralized dolerite dikes and sills that have either a non-selective mining width or orientation. Mineral Resources were classified primarily based on the average distance to the nearest three drill holes, as follows:  Measured Resource: three drill holes within average distance ≤ 25 m;  Indicated Resource: three drill holes within average distance ≤ 50 m;  Inferred Resource: three drill holes within average distance ≤ 100 m. Reasonable prospects for eventual economic extraction was addressed by applying a resource shell defined using a pit optimization to identify mineralization that could reasonably be economically extracted, using a cut-off value based on a net smelter return (NSR) approach. Mineral Resources are reported within a detailed crest and toe Mineral Resource pit design exclusive of Mineral Reserves. The Mineral Resource reports the Measured, Indicated and

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Inferred Resource within the Mineral Resource pit shell above an NSR cut-off of AU$15.43/t, less reported Mineral Reserves. Table 1-1 presents the economic parameters and cut-offs used for the 2018 Mineral Resource pit optimization. Table 1-1: 2018 Mineral Resource Economic Parameters and Cut-offs Cut-off for 2018 Mineral Resource Units At Cost Basis of 2019BP Gold Price AU$/oz 1,750 Copper Price AU$/lb 4.00 Exchange Rate US$0.80 = AU$1.00 Gold Royalty % 2.5 Copper Royalty % 5.0 Mill Throughput Mtpa 40.5 Mill Recovery Gold (Average recovery % at LOM grade 0.67 g/t Au) % 83 Mill Recovery Copper (Average recovery % at LOM grade 0.1% Cu) % 78 Base Processing Cost-without rehandle AU$/t milled 9.71 Sustaining Capital (Plant and G&A) AU$/t mined 1.14 incl. Capital Recovery Factor (CRF) = 1.10 Site and Regional G&A (exclude CAPEX) AU$/t milled 2.09 Incremental Ore, Resource Conversions and Closure (LOM-FASB) AU$/t milled 0.22 Breakeven Mill Cut-off (BMCO) AU$/t milled 13.17

Stockpile Rehandling AU$/t rhdled 1.48 Recoveries Degradation AU$/t milled 0.79 Breakeven Stockpile Cut-off (BSCO) AU$/t milled 15.43

LOM Operating Mining Cost AU$/t mined 4.22 Mining Capital AU$/t mined 0.62 incl. CRF = 1.11

Note: 2019BP = 2019 Business Plan

1.12 Mineral Resource Statement Mineral Resources are exclusive of those Mineral Resources that were converted to Mineral Reserves and are reported at gold price of US$1,400 or AU$1,750/oz, and a copper price of US$3.25/pound (lb) or AU$4.00/lb on a 100% basis, with an effective date of 31 December 2018. Newmont cautions that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. The Mineral Resource estimates were prepared by Mr. Rohan McCormack, Member of the Australian Institute of Geoscientists (MAIG), Senior Resource Geologist and Newmont employee, under the supervision of the QP. Table 1-2 presents the total gold and copper Mineral Resource for the Project.

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Table 1-2: Mineral Resource – Gold and Copper at the Effective Date of 31 December 2018

Measured Resource Indicated Resource Measured + Indicated Resource Inferred Resource Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal (kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt) Boddington OP 95,200 0.55 1,680 0.11 100 253,800 0.55 4,510 0.12 300 349,000 0.55 6,190 0.12 400 5,100 0.49 80 0.09 0 Total 95,200 0.55 1,680 0.11 100 253,800 0.55 4,510 0.12 300 349,000 0.55 6,190 0.12 400 5,100 0.49 80 0.09 0

Notes to accompany the Mineral Resource tables:  OP = open pit;  Mineral Resources have an effective date of 31 December 2018;  Mineral Resources are reported exclusive of Mineral Reserves, and are reported on a 100% basis;  Mineral Resources are estimated within designed pits generated based on optimised pit shell using Whittle®;  Mineral Resources are reported using a gold price of US$1,400/oz, and a copper price of US$3.25/lb, equivalent to AU$1,750/oz, and AU$4.00/lb at an exchange rate of US$0.80 = AU$1.00;  Mineral Resources contain material that is above a NSR cut-off of AU$15.43/t within the Mineral Resource ultimate pit design, exclusive of Mineral Reserves. Table 1-1 presents the economic parameters and cut-offs used for the 2018 Mineral Resource pit optimization;  Royalties are considered in the NSR cut-off determination;  Metallurgical recovery is considered in the NSR cut-off for in situ and stockpiled ore. Since the metallurgical recovery of stockpiled ore degrades over time, weighted average LOM recovery (refer to Section 13.4) is calculated using a combination of in situ recovery less 3% and direct ROM feed in situ recovery;  Tonnage and grade measurements are in metric units. Gold ounces are reported as troy ounces. Tonnages include allowances for losses resulting from mining methods. Tonnages are rounded to the nearest 100,000 tonnes. Ounces are estimates of metal contained in the Mineral Resource and do not include allowances for processing losses. Gold ounces are rounded to the nearest 10,000 ounces and copper metal tonnage is rounded to the nearest 10,000 tonnes;  Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grade and contained metal content. Copper metal tonnage for Inferred Resource is less than 5 kt and hence is not presented in the table to maintain consistency in rounding.

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1.13 Mineral Reserves Estimates The cut-off for the Project, used for both Mineral Resources and Mineral Reserves is defined by revenue, due to the Project being a poly-metallic deposit with two product streams, gold doré and copper concentrate. The block revenue is calculated on an NSR basis. The NSR block revenue reflects the dollar return expected from the sale of the concentrate produced from a tonne of in situ material. NSR calculations account for concentrate shipping, smelting and refining costs, and royalties. Mineral Reserves are reported within a detailed pit design. Only Measured and Indicated Mineral Resources within the Mineral Reserves pit design are directly converted to Proven and Probable Mineral Reserves respectively. The Mineral Reserves are reported within the optimized pit design based on the following price inputs; US$1,200/oz or AU$1,600/oz for gold and US$2.50/lb or AU$3.35/lb for copper metal prices, providing an NSR cut-off of AU$15.32/t. The break-even NSR accounts for processing costs, incremental ore mining costs, sustaining capital process, metallurgical recovery, smelter-related costs, and residue dam related rehabilitation costs. Table 1-3 presents the economic parameters and cut-offs used for the 2018 Mineral Reserves pit optimization. Table 1-3: 2018 Mineral Reserves Economic Parameters and Cut-offs Cut-off for 2018 Mineral Reserves Units At Cost Basis of 2019BP Gold Price AU$/oz 1,600 Copper Price AU$/lb 3.35 Exchange Rate US$0.75 = AU$1.00 Gold Royalty % 2.5 Copper Royalty % 5.0 Mill Throughput Mtpa 40.5 Mill Recovery Gold (Average recovery % at LOM grade 0.67 g/t Au) % 83 Mill Recovery Copper (Average recovery % at LOM grade 0.1% Cu) % 78 Base Processing Cost-without rehandle AU$/t milled 9.71 Sustaining Capital (Plant and G&A) AU$/t mined 1.14 incl. CRF = 1.10 Site and Regional G&A (exclude CAPEX) AU$/t milled 2.09 Incremental Ore, Resource Conversions and Closure (LOM-FASB) AU$/t milled 0.21 Breakeven Mill Cut-off (BMCO) AU$/t milled 13.16

Stockpile Rehandling AU$/t rhdled 1.48 Recoveries Degradation AU$/t milled 0.68 Breakeven Stockpile Cut-off (BSCO) AU$/t milled 15.32

LOM Operating Mining Cost AU$/t mined 4.22 Mining Capital AU$/t mined 0.62 Incl. CRF = 1.11

Note: 2019BP = 2019 Business Plan

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A mine schedule was developed with the mining sequence according to the current business plan. The economic viability of Mineral Resource and Mineral Reserves addition is tested by comparing undiscounted cash flows. In assessing economic viability, the expected net cash flows from Mineral Reserves had a 5% discount rate applied in accordance with Newmont financial and business planning parameters.

1.14 Mineral Reserves Statement Proven and Probable Mineral Reserves were declared for the open pit mineralisation and stockpiles and are reported using the 2014 CIM Definition Standards for Mineral Resources and Mineral Reserves (the CIM Definition Standards). The surface stockpiles of ore produced from North Pit and South Pit are classified as Proven and Probable Mineral Reserves as their volume is established by survey pickup and their grade estimated by production designs based on Measured and Indicated Mineral Resources respectively. The Mineral Reserves estimate was prepared by Mr. Eka Setiawan Lim, Mine Engineering Superintendent and Newmont employee, under the supervision of the QP. Mineral Reserves have an effective date of 31 December 2018 and are reported to a gold price of $US1,200/oz or AU$1,600/oz, and a copper price of US$2.50/lb or AU$3.35/lb.

At the effective date of the Report, the exchange rate was US$1.00 = approximately AU$1.421. Table 1-4 presents the total gold and copper Mineral Reserves for the Project.

1 Source: Reserve Bank of Australia (RBA) website (https://rba.gov.au/statistics/frequency/exchange-rates.html)

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Table 1-4: Mineral Reserves – Gold and Copper at the Effective Date of 31 December 2018

Proven Reserves Probable Reserves Proven + Probable Reserves Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal

(kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt)

Boddington OP 240,400 0.71 5,520 0.09 230 240,300 0.71 5,470 0.11 260 480,700 0.71 10,990 0.10 490

Stockpiles 6,900 0.67 150 0.08 0 86,100 0.44 1,210 0.08 70 93,000 0.45 1,360 0.08 70

Total 247,300 0.71 5,670 0.09 230 326,400 0.64 6,680 0.10 330 573,700 0.67 12,350 0.10 560

Notes to accompany the Mineral Reserves tables:  OP = open pit;  Mineral Reserves have an effective date of 31 December 2018;  Mineral Reserves are reported on a 100% basis;  Mineral Reserves are estimated within designed pits generated based on optimised pit shell using Whittle®;  Mineral Reserves are reported to a gold price of US$1,200/oz, and a copper price of US$2.50/lb, equivalent to AU$1,600/oz, and AU$3.35/lb at an exchange rate of US$0.75 = AU$1.00;  Mineral Reserves contain material that is above a NSR cut-off of AU$15.32/t within the Mineral Reserves ultimate pit design. Table 1-3 presents the economic parameters and cut-offs used for the 2018 Mineral Reserves pit optimization;  Royalties are considered in the NSR cut-off determination;  Metallurgical recovery is considered in the NSR cut-off for in situ and stockpiled ore. Since the metallurgical recovery of stockpiled ore degrades over time, weighted average LOM recovery (refer to Section 13.4) is calculated using a combination of in situ recovery less 3% and direct ROM feed in situ recovery;  Tonnage and grade measurements are in metric units. Gold ounces are reported as troy ounces. Tonnages include allowances for losses resulting from mining methods. Tonnages are rounded to the nearest 100,000 tonnes. Ounces are estimates of metal contained in the Mineral Reserves and do not include allowances for processing losses. Contained gold ounces are rounded to the nearest 10,000 ounces and copper metal tonnage is rounded to the nearest 10,000 tonnes;  Rounding of tonnes as required by reporting guidelines may result in apparent summation differences between tonnes, grade and contained metal content. Copper metal tonnage for Proven Reserves is less than 5 kt and hence is not presented in the table to maintain consistency in rounding;  The Mineral Reserves are forward-looking information and actual results may vary.

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While the assumptions in the Mineral Reserves estimates are appropriate to the date of estimation, areas of uncertainty that may materially impact the Mineral Reserves estimates include changes to long-term metal price assumptions, and changes to input cost assumptions such as consumables, labor costs, or taxation rates. The Mineral Reserves are forward-looking information and actual results may vary. The risks regarding Mineral Reserves are summarized in the Report (refer to Section 15.3 and Section 25.0). The assumptions used in the Mineral Reserves estimates are summarized in the footnotes of the Mineral Reserves tables, and in Section 15.0 of the Report.

1.15 Mine Plan The Project is mined as a conventional truck-and-shovel operation. Equipment is owner- operated, using a large mining truck (240-tonne class) and shovel (electric rope) fleet, support equipment fleet and drill fleet, mining predominantly 12 m benches. Mine start-up required an 18-month mining ramp-up that included contractor pre-stripping between August 2006 and September 2007 of approximately 12 Mt of predominantly waste ahead of process plant commissioning. In October 2007, owner-operated mining commenced, continuing the pre-stripping operation to access the bedrock mineralization. A total of 912.7 Mt of material has been mined up to the end of December 2018. The Life of Mine Plan (LOMP) currently envisages mining at an average rate of approximately 70 Mtpa for 13 years and peaking at 84 Mtpa in 2026, with a maximum rate of advance by pit stage of seven benches per annum and an average of five benches (60 m) per year. NBG currently expects that mine production life will extend into approximately year 2032 with material mined from the two open pit sources. Milling will cease in the same year after treatment of stockpiled ore. Two major phases (laybacks) have been completed to date; two more are underway. These pits form the basis for the current mine plan. A fifth layback, S05A, which was granted the approval for capital expenditures, commenced mining in 2016.S05A required a waste dump (WD) expansion to accommodate the waste generated by S05A mining over and above the currently constructed WD capacity. Approximately three-quarters of the mining fleet is located in the higher-priority South Pit, mining an ore face and a waste push-back simultaneously. The remaining one-quarter of the fleet is working the North Pit, mining a mixture of ore and waste faces.

1.16 Recovery Plan The process consists of primary crushing, closed circuit secondary and High Pressure Grinding Rolls (HPGR) tertiary crushing, ball milling, and hydrocyclone classification to generate a milled product with a P80 of 150 micrometers (μm) at a slurry density of around 35% solids. The cyclone overflow from the mill circuit is treated in a flotation circuit that produces a copper–gold concentrate for export. Rougher and scavenger flotation concentrates are reground and cleaned to achieve an acceptable final concentrate grade. The concentrate is thickened and filtered before being trucked to port. The cleaner scavenger tailings stream is thickened and leached under elevated cyanide levels. The scavenger tailings are thickened and leached in a conventional leach/adsorption

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circuit. The leached slurry from the cleaner scavenger tailings leach circuit is delivered to the scavenger tailings circuit for combined recovery of gold. Leach residue is pumped to the residue disposal area, and residual cyanide is maintained below a targeted level by a Caros acid cyanide destruction plant. The carbon from the scavenger tailings adsorption circuit is treated by conventional split- Anglo American Research Laboratory method elution and reactivated in horizontal reactivation kilns. Gold recovery from the eluate is by electrowinning, cathode sludge filtration and drying, and smelting. There is a flash flotation and gravity circuit installed in the process plant. These circuits have not been operated and remain decommissioned.

1.17 Infrastructure Infrastructure required to support the LOM plan is complete and constructed on site. An extension to the waste rock facility is required to support a re-designed pit layback. Permits for this facility construction have been received.

1.18 Marketing Newmont has a refining agreement with The Perth Mint in Perth, for refining of gold and silver doré produced from the Project. Newmont’s bullion is sold on the spot market by marketing experts retained in-house by Newmont. Boddington copper concentrate is unique as it contains one of the highest gold contents in the market and a relatively low copper content. Consistently, smelters operating their own precious metal refineries at their copper smelting operations are the subset of smelters that are prepared to contract for Boddington concentrates Copper and gold bearing sulfide concentrates are sold to several overseas smelters, with the majority of concentrates sold under contracts which Newmont have established. Additional concentrate volumes produced are sold on the spot market as required. The terms contained within the sales contracts are typical of and consistent with standard industry practice and contracts for the supply of copper concentrate elsewhere in the world.

1.19 Environmental, Permitting and Social Considerations The Project was subject to a comprehensive environmental impact assessment (EIA) by the WA Environmental Protection Authority (EPA) and other key Government decision-making authorities. The Project operates under various approvals and permits granted by the State Government of WA and in particular, via the Department of Mines, Industry Regulation and Safety (DMIRS) and the Department of Water and Environmental Regulation (DWER). The total LOM rehabilitation costs for the Mineral Reserves schedule are in the order of AU$383 M for the 2019 Business Plan.

1.20 Sustaining Capital Costs Capital cost estimates were developed during 2018, and include funding for infrastructure, pit dewatering, development drilling, and permitting as well as miscellaneous expenditures required to maintain production. Mobile equipment is scheduled for replacement when

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operating hours reach threshold limits. Capital costs are based on previous operating experience, and quotes received from manufacturers during 2018. Sustaining capital costs reflect current price trends. The 2018 Mineral Reserves and Mineral Resources estimates are based on the 2019 Business Plan costs including sustaining capital projects planned in the business plan. Because of the objective of maximizing return on investment, time value of money must be included in the development of sustaining capital cost per unit of mass mined. Equipment capital costs can be recovered by using an amortization rate per unit of mass mined so that the net present value of the future cash flow generated from this charge is equal to the equipment capital cost. The Capital Recovery Factor (CRF) is required to give a 5% pre-tax return on an investment over the life of fleet of facility. The CRF for mining and processing is estimated at 111% and 110% respectively. Including the CRF, the Mining Sustaining Capital is AU$0.62/t mined at CRF 1.11, and Processing and General and Administrative (G&A) sustaining capital is AU$1.14/t milled at CRF 1.10.

1.21 Operating Costs Operating cost estimates were developed by NBG, based on the 2019 Business Plan (budget) developed in Q3 of 2018. The estimated LOM mining cost is AU$4.22/t. Base processing costs are estimated at AU$9.71/t. In addition, site G&A costs are estimated at AU$1.37/t, and the regional G&A back- charge at AU$0.72/t.

1.22 Economic Analysis Newmont is using the allowance for producing issuers, whereby producing issuers may exclude the information required under Item 22 for technical reports on properties currently in production where the Technical Report does not include a material expansion of production. Mineral Reserves declaration is supported by performing an economic test on each pit stage independently on a stand-alone basis. For Mineral Reserves, a pit stage or stockpile must meet a positive discounted cash flow. All pit stages have satisfied the Mineral Reserves qualification and return positive discounted cash flows at a 5% discount rate.

1.23 Interpretation and Conclusions In the opinion of the QP:  Information provided by Newmont’s legal and tenure experts on the mining tenure held by Newmont in the Project area supports that the company has valid title that is sufficient to support declaration of Mineral Resources and Mineral Reserves;  Information provided by Newmont’s legal and tenure experts supports that the Operations hold sufficient surface rights to enable mining operations, and the declaration of Mineral Resources and Mineral Reserves. Appropriate steps, where required, have been taken to lodge either extensions or renewals of tenements as such fall due;  The geological understanding of the deposit settings, lithologies, and structural and alteration controls on mineralization is well understood. The mineralization styles and setting are also well understood;

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 Exploration programs completed to date are appropriate to the different mineralization styles known to occur within the Saddleback Greenstone Belt;  The quantity and quality of the lithological, geotechnical, collar and downhole survey data collected in the exploration, delineation, and grade control drill programs between 1980 and 2016 are adequate to support Mineral Resource and Mineral Reserves estimation and mine planning;  Sampling methods are acceptable and meet industry-standard practices. The quality of the gold and copper analytical data is reliable and sample preparation, analysis, and security are generally performed in accordance with exploration best practices and industry accepted standards;  Data verification programs undertaken on the data collected from the Project adequately support the geological interpretations and the database quality;  Metallurgical testwork completed on the Project has been appropriate to establish optimal processing routes for the different mineralization styles encountered in the deposits. Testwork has been completed on mineralization that is typical of that within the deposits. The mill process and associated recovery factors are considered appropriate to support Mineral Resource and Mineral Reserves estimation, and mine planning;  A detailed review of the metallurgical models and comparisons with process plant actual performance indicated a tendency for the metallurgical models to over-predict recovery at low head grades and under-predict recovery at high head grades. Molybdenum and As grades in copper concentrate were in line with model predictions to date. Ore hardness and abrasiveness of the samples were in line with the orebody average, possibly except for a slight increase in Bond ball mill work index with depth;  Estimates of Mineral Resources and Mineral Reserves for the Project conform with industry standard practices and satisfy the requirements of the CIM Definition Standards. There is some upside for the Project if some or all of the estimated Inferred Mineral Resources can be upgraded to higher confidence Mineral Resource categories;  The mine plan uses conventional mining methods and equipment and is adequately documented in terms of the availability of staff, the existing power, and communications facilities, the methods whereby goods are transported to and from the mine, and any planned modifications or supporting studies are well understood;  Since the effective date of the mine plan, Newmont has regularly undertaken, and will continue to undertake as part of its normal course of business operations, reviews of the mine plan and consideration of alternatives to and variations within the plan. Alternative scenarios and reviews are based on ongoing or future mining considerations, evaluation of different potential input factors and assumptions, and requests made of Project staff by Newmont Corporate;  Based on the current mine plan, some of the existing mine facilities such as waste rock dumps (WRD), tailings dams and water storage are required to be extended to achieve LOMP (2032);  The Environmental Impact Statement (EIS) and the current state of environmental knowledge of the mine area support the Mineral Resources and Mineral Reserves statement can be declared, and the mine plan is achievable;  Closure and remediation requirements have been addressed through the site closure plan and associated environmental bond requirements;

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 Permitting activities have been carried on appropriately to ensure the mining activities can be conducted within the regulatory framework required by the WA and Federal Governments;  At the effective date of this Report, environmental liabilities are limited to those that would be expected to be associated with a gold and copper mine of comparable scale, including roads, open pits, site infrastructure, waste and tailings disposal facilities. Newmont has appropriately addressed the potential and actual environmental impacts of the operation sufficient to support Mineral Resources and Mineral Reserves estimation, and mine planning;  The mill process is conventional, producing doré and copper-gold concentrates. The mill process and associated recovery factors are considered appropriate to support Mineral Resource and Mineral Reserves estimation, and mine planning;  All required infrastructure has been established and is operational;  Capital and operating cost parameters are considered appropriate to support Mineral Resource and Mineral Reserves estimation, and mine planning;  Review of the environmental, permitting, legal, title, taxation, socio-economic, marketing, and political information on the Project supports the assumptions used in the economic analysis, which is positive, and supports the Mineral Reserves;  Other than what is presented in this Report, there are no other significant factors and risks that may affect access, title, or the right or ability to perform work on the property.

1.24 Recommendations The QP makes the following recommendations:  Completion of metallurgical testwork to determine the extent of recovery degradation on long term stockpiles. It is estimated that over a period of five years, a total of 12 samples will be prepared and tested multiple times at a cost of approximately US$30k per test for a total program cost of approximately US$840k;  Infill drilling to increase Mineral Resource confidence in the later pit stages and lower benches of the current Mineral Reserves. Over the next five years, it is anticipated that a total of approximately 44,631 m (average cost/m of approximately $AU177.00/m inclusive of drilling, assaying, and general support costs i.e. diesel, labor and supplies) of core and RC drilling will be completed at an estimated cost of AU$7.9M;  Completion of projects aimed at maximising the capture and use of raw water on site and securing a long-term reliable supply system for the Project. These projects are estimated to cost approximately US$56M over the next three years.

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2.0 INTRODUCTION Newmont has prepared this Report for the Boddington Operations, located in WA (refer to Figure 2-1).

Note: Figure aligned to True North Figure 2-1: Project Location Plan

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2.1 Terms of Reference This Report supports declared Mineral Resources and Mineral Reserves and summarizes the Project development and current operations. Newmont will use this Report in support of disclosure and filing requirements with the Canadian securities regulators as specified in Section 4.2 (1) (c) of NI 43-101. This Report will be filed under Newmont’s SEDAR profile. The Project is owned and operated by a JV between two indirectly wholly-owned Australian subsidiaries of Newmont Mining Corporation; Newmont Boddington Pty Ltd (66 and 2/3% interest in the Project) and Saddleback (33 and 1/3% interest in the Project). The manager of the JV is also an indirectly wholly-owned company, NBG. In this Report, the name Newmont is used interchangeably for the parent and subsidiary companies. All measurement units used in this Report, except for troy ounces, are metric, and currency is expressed in AU$ unless stated otherwise. The Report uses US English. Mineral Reserves are reported to a gold price of US$1,200/oz or AU$1,600/oz. The assumed exchange rate for Mineral Reserves estimation was US$0.75 = AU$1.00. At the effective date of the Report, the exchange rate was US$1.00 = approximately AU$1.422.

2.2 Qualified Person Mr. Doe, RM-SME, Group Executive, Reserves at Newmont, is the QP for the Report, as QP is defined in NI 43-101 Standards of Disclosure for Mineral Projects. Mr. Doe has been Newmont’s Group Executive, Reserves since 2014 and is accountable for Newmont’s governance system for Mineral Resources and Mineral Reserves. Newmont has a system of controls, standards and guidelines designed to support the estimation and reporting of Mineral Resources and Mineral Reserves. On an annual basis, Mr. Doe reviews and approves the estimates for Mineral Resources and Mineral Reserves provided by Newmont sites and projects, along with their compliance to the required controls, standards and guidelines. Mr. Doe also reviews the supporting documentation for Mineral Resources and Mineral Reserves provided by Newmont sites and projects.

2.3 Site Visits and Scope of Personal Inspection Mr. Doe, the QP, has visited the Project numerous times during his career at Newmont, most recently on 18 April 2018. During site visits to the Project, Mr. Doe inspects the operating open pits, and views the process plant and associated general site infrastructure, including the current tailings storage facility (TSF) operations. While on site, he discusses aspects of the operation with site-based staff and assesses the knowledge and abilities of the site staff to carry out their duties as required. These site discussions include the overall approach to the mine plan, anticipated mining conditions, selection of the production target and potential options for improvement. Other areas of discussion include plant operation and recovery forecasts, capital and operating forecasts and results. Mr. Doe receives and reviews monthly reconciliation reports from the mine. These reports include the industry standard reconciliation factors for tonnage, grade and metal; F1 (Mineral Reserve model compared to ore control model), F2 (mine delivered compared to mill received) and F3 (F1 x F2) along with other measures such as compliance of actual production to mine

2 Source: Reserve Bank of Australia (RBA) website (https://rba.gov.au/statistics/frequency/exchange-rates.html)

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plan and polygon mining accuracy. The reconciliation factors are recorded monthly and reported in a quarterly control document. Through the review of these reconciliation factors, the QP is able to ascertain the quality and accuracy of the data and its suitability for use in the assumptions underlying the Mineral Resource and Mineral Reserves estimates. Mr. Doe also reviews Newmont’s processes and internal controls at the mine site with operational staff on the work flow for determining Mineral Resource and Mineral Reserves estimates, mineral process performance, mining costs, and waste management. The following Newmont employees contributed to various aspects of the Report under the supervision of the QP:  Mr. Rohan McCormack, Member of the Australian Institute of Geoscientists (MAIG), Senior Resource Geologist, Newmont – 21 years of experience in mining and resource development and 8.5 years of experience with the Project and Operation;  Mr. Eka Setiawan Lim, RM-SME, Mine Engineering Superintendent at Newmont – 26 years of experience in mining engineering and 10 years of experience with the Project and Operation;  Mr. Steven Hart, Fellow of the Australasian Institute of Mining and Metallurgy (FAusIMM), Principal Advisor Processing, Newmont – 31 years of experience in minerals processing operations, 12 years of experience with Newmont, 6 years with the Project and Operation, and 6 years providing regional support to the operation;  Mr. Unduk Samosir, Member of the Australasian Institute of Mining and Metallurgy (MAusIMM), Senior Long-Term Planning, Newmont – 22 years of experience in mining engineering and 7 years of experience with the Project and Operation;  Mr. Paul Petrucci, MAusIMM, Senior Metallurgist, Newmont – 16 years of experience in mining and resource development and 10 years of experience with the Project and Operation.

2.4 Effective Dates This Report has an effective date of 31 December 2018, which is the date of the Mineral Resource and Mineral Reserves estimates and the economic parameters presented in this Report. There have been no material changes to the information on the Project between the effective date and the signature date of the Report.

2.5 Information Sources and References Unless otherwise noted, all figures and tables were prepared by Newmont for the purposes of this Report. As part of this Report, and in addition to day-to-day operations, supplementary background data were reviewed and new studies were completed as required, under the supervision of the QP. Information that supports this Report has been obtained from Newmont or external consultants in the relevant field or has been prepared by or under the supervision of the QP. Reference documents are cited in the text as appropriate and summarized in Section 27.0.

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2.6 Previous Technical Reports Newmont has not previously filed a Technical Report on the Project.

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3.0 RELIANCE ON OTHER EXPERTS The QP has relied upon Newmont experts for the information included in this Report on mineral tenure, surface rights, permitting, political, environmental and social considerations, taxation and markets.

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4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 Location The Project is located approximately 130 km southeast of the city of Perth (refer to Figure 2-1), and 17 km northwest of the township of Boddington. The Project centroid is situated at approximately 32°44'15.99"S and 116°21'34.76"E.

4.2 Tenure History The majority of the productive areas of the Project are located within the original boundaries of a single large tenement (ML 258SA) granted under a State Agreement known as the Worsley State Agreement. A State Agreement is an agreement between the State Government and a miner for the development of a resource. ML258SA is held by a JV known as the Worsley Joint Venture (WJV) and the Worsley State Agreement permits the mining of bauxite only. The current participants of the WJV are:  South32 Aluminium (RAA) Pty Ltd 56%;  South32 Aluminium (Worsley) Pty Ltd 30%;  Japan Alumina Associates (Australia) Pty Ltd 10%;  Sojitz Alumina Pty Ltd 4%. In the early 1980s, the WJV discovered a gold resource in Boddington. In order to mine this resource, the Worsley State Agreement was amended to enable the temporary excision of areas from ML258SA. Mining Leases for the excised areas were then granted under the Mining Act of Western Australia 1978. Under the Worsley State Agreement when the Mining Act Mining Leases are relinquished the area reverts to ML 258SA. In order to develop the Boddington gold resource, the WJV established a new JV; the BGMJV. The BGMJV Agreement was entered into on 31 March 1987 and was initially comprised of the same participants as the WJV. The relationship between the bauxite/alumina operations and the gold operations was regulated under a cross-operation agreement which, in a restated form, continues today. The paramount principle regulating the relationship between the WJV and the BGMJV was that bauxite and bauxite operations were to have priority over all other minerals within an area (the ‘Common Area’) defined within the boundaries of ML258SA. The priority of the bauxite operations over that of the gold operations continues today. Consequently, where bauxite is found in an area of the Project Mining Leases where Newmont is active, Newmont is required to mine and stockpile bauxite on behalf of the WJV. Ownership of the BGMJV changed over time so that the participants in the WJV were no longer the same as the BGMJV participants. In order to affect the transfer of ownership to incoming BGMJV participants whilst maintaining rights to bauxite, a series of transactions were entered into, resulting in the present structure whereby the BGMJV participants hold subleases to the Mining Leases on which the gold resource is located. Newmont acquired its interest in the BGMJV through the transactions presented in Table 4-1.

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Table 4-1: Transactions Through Which Newmont Acquired Its Interest in the BGMJV Year Interest Seller Purchaser Newmont Boddington Pty Ltd (then called PosGold 4 and Japan Alumina Associates (Australia) Pty Ltd (then called 1995 (Boddington) Pty Ltd and ultimately held by 4/9% Kobe Alumina Associates (Australia) Pty Limited) Normandy Mining Limited) Newmont Boddington Pty Ltd (then called PosGold 1995 40% Reynolds Australia Alumina Ltd (Boddington) Pty Ltd and ultimately held by Normandy Mining Limited) Newmont Mining Corporation acquires 100% of 2002 Normandy Mining Limited, which was the ultimate owner of PosGold (Boddington) Pty Ltd. 22 and 2006 Newcrest Operations Limited Newmont Boddington Pty Ltd 2/9% 33 and 2009 AngloGold Ashanti Australia Limited Saddleback 3/9%

Since 2009, Newmont has had 100% ownership of the BGMJV but continues to operate it as a JV. The key tenements within the Project area are:  Mining Leases M70/21–26, M70/564, M70/799 – subleased from the WJV;  M70/1031, G70/215 and G70/218–219, M264SA- held directly by Newmont.

4.3 Property Agreements 4.3.1 Background The region surrounding the Project is a well-known bauxite-alumina mining and production area now held predominantly by either the Worsley JV or the Alcoa Group. The major part of the Mining Act tenure (including M70/21–26, M70/564 and M70/799) lies within the original boundaries of a tenement granted under the Worsley State Agreement (ML258SA). The remainder of the tenure (including ML264SA and M70/1031) lies within the original boundaries of another State Agreement Area known as the Alcoa State Agreement (ML1SA). Newmont subleases from the Worsley JV the key mining leases upon which Project operations are located, namely M70/21–26, M70/564 and M70/799. Newmont is entitled to all gold and other non-bauxite mining rights conferred by the lease. The Worsley JV retains the rights to bauxite and rights of access in order to mine and recover such bauxite. 4.3.2 Management Agreements The relationship between the Worsley JVs bauxite operations and the BGMJV gold operations is regulated through a cross-operation agreement. This agreement confers priority on the bauxite operations such that the BGMJV are required to take reasonable measures to conserve bauxite including by mining and stockpiling bauxite on behalf of the Worsley JV. The cross-operation agreement also requires the managers of the respective JVs to keep each other regularly informed on the other’s current and proposed activities in order to alleviate or minimize any potential impact of one operation upon another. Additional information on surface rights is provided in Section 4.5.

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The Project falls within an area of a native title land claim registered under the Native Title Act and referred to as the Gnaala Karla Booja claim. Additional information on Native Title is provided in Section 4.7 of this Report. The Project operates under various approvals and permits granted by the State Government of WA and via DMIRS and DWER.

4.4 Current Mineral Tenure The Project has an interest in a total of 89 tenements in the Boddington area (refer to Table 4-2 and Figure 4-1), 26 of the mining tenements are at an application stage. The total granted area is approximately 21,249 ha and the under-application area is approximately 60,734 ha. The actual mining area is covered by the following 10 WA Mining Act leases; M70/21–26, M70/564, M70/799, M70/1031, G70/215 and G70/218–219, as well as State Agreement M264SA. Mining Leases M70/21–25 are the five tenements under which gold mining activity is concentrated. Through direct lease holding and sub-lease arrangements with the Worsley JV, Newmont holds the rights to minerals other than bauxite in proportion to the NBG ownership percentages. Under the Mining Act of Western Australia 1978, Mining Leases are granted for 21 years and are renewable. The five main mining area tenements of the Project were renewed in March 2007 for a 21-year term.

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Table 4-2: Mineral Tenure Table

Lease Lease Type Lease Status Current Area Application Date Grant Date Expiry Date Commitment (AU$) Shire Rates (AU$) Rent (AU$) Holders

E 70/21491 E Application 28 Blocks 7/12/1998 – – $0.00 $0.00 $0.00 Newcrest Operations Ltd (22.22%), Newmont Boddington Pty Ltd (44.44%), AngloGold Ashanti Australia Ltd (33.33%), E 70/23361 E Application 56 Blocks 25/05/2000 – – $0.00 $0.00 $0.00 Newcrest Operations Ltd (22.22%), Newmont Boddington Pty Ltd (44.44%), AngloGold Ashanti Australia Ltd (33.33%), E 70/25501 E Application 7 Blocks 11/10/2002 – – $0.00 $0.00 $0.00 Newcrest Operations Ltd (22.22%), Newmont Boddington Pty Ltd (44.44%), AngloGold Ashanti Australia Ltd (33.33%), E 70/25621 E Application 8 Blocks 2/12/2002 – – $0.00 $0.00 $0.00 Hedges Gold Pty Ltd (100%), E 70/3750 E Application 40 Blocks 30/11/2009 – – $0.00 $0.00 $0.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), E 70/3982 E Application 6 Blocks 1/10/2010 – – $0.00 $0.00 $0.00 Newmont Boddington Pty Ltd (66.7%), Saddleback Investments Pty Ltd (33.3%), E 70/4018 E Application 3 Blocks 9/12/2010 – – $0.00 $0.00 $0.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), E 70/4019 E Application 6 Blocks 9/12/2010 – – $0.00 $0.00 $0.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), E 70/4235 E Application 2 Blocks 30/09/2011 – – $0.00 $0.00 $0.00 Newmont Boddington Pty Ltd (66.6%), Saddleback Investments Pty Ltd (33.4%), E 70/4301 E Application 8 Blocks 22/02/2012 – – $0.00 $0.00 $0.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), E 70/4302 E Application 6 Blocks 22/02/2012 – – $0.00 $0.00 $0.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/182 M Application 884 Ha 8/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/192 M Application 980 Ha 8/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/272 M Application 747 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/282 M Application 720 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/292 M Application 690 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/302 M Application 690 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/312 M Application 907 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/322 M Application 972 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/332 M Application 856 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/342 M Application 873 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/352 M Application 967 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/362 M Application 639 Ha 14/04/1983 – – $0.00 $0.00 $0.00 BHP Billiton Minerals Pty Ltd (20%), South32 Aluminium (RAA) Pty Ltd (40%), Japan Alumina Associates (Australia) Pty Ltd (10%), The Shell Company of Australia Ltd (30%), M 70/5452 M Application 1000 Ha 13/07/1989 – – $0.00 $0.00 $0.00 South32 Aluminium (RAA) Pty Ltd (50%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (2.5%), The Shell Company of Australia Ltd (37.5%), M 70/9751 M Application 990 Ha 14/01/1997 – – $0.00 $0.00 $0.00 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), P 70/1598 P Application 27.56 Ha 24/06/2010 – – $0.00 $0.00 $0.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), E 70/7103 E Granted 9.44 Km2 27/04/1988 16/01/1989 15/01/1997 $100,000.00 $950.00 $1,830.00 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), G 70/215 G Granted 28.61 Ha 13/05/2005 16/06/2009 15/06/2030 $0.00 $950.00 $478.50 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), G 70/218 G Granted 51.69 Ha 9/02/2006 16/08/2006 15/08/2027 $0.00 $950.00 $858.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), G 70/219 G Granted 9.545 Ha 9/02/2006 13/12/2006 12/12/2027 $0.00 $950.00 $165.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), L 70/152 L Granted 171.68 Ha 6/08/2012 14/03/2013 13/03/2034 $0.00 $0.00 $2,838.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), L 70/165 L Granted 2.128 Ha 15/05/2014 22/09/2014 21/09/2035 $0.00 $0.00 $49.50 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), L 70/28 L Granted 1.2 Ha 16/08/1993 4/11/1993 3/11/2023 $0.00 $0.00 $33.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), L 70/95 L Granted 31 Ha 9/02/2006 5/05/2006 4/05/2027 $0.00 $0.00 $511.50 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), L 70/96 L Granted 6 Ha 13/07/2006 10/11/2006 9/11/2027 $0.00 $0.00 $99.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 264SA(1) M Granted 497.35 Ha 28/12/1987 1/08/1988 31/07/2030 $0.00 $1,340.68 $9,312.60 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 264SA(2) M Granted 408.9 Ha 28/12/1987 1/08/1988 31/07/2030 $0.00 $1,340.69 $7,648.30 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/1031 M Granted 398.9 Ha 8/10/1998 11/10/1999 10/10/2020 $39,900.00 $1,226.44 $7,461.30 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/1102 M Granted 5.2955 Ha 25/11/1983 3/02/1989 2/02/2031 $10,000.00 $1,032.00 $112.20 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/1112 M Granted 121.3 Ha 25/11/1983 3/02/1989 2/02/2031 $10,000.00 $1,032.00 $2,281.40 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/1122 M Granted 29.37 Ha 25/11/1983 3/02/1989 2/02/2031 $12,200.00 $1,032.00 $561.00 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/1132 M Granted 64.485 Ha 25/11/1983 3/02/1989 2/02/2031 $10,000.00 $1,032.00 $1,215.50 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/1142 M Granted 817.8 Ha 25/11/1983 3/02/1989 2/02/2031 $10,000.00 $2,428.25 $15,296.60 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/1152 M Granted 702.2 Ha 25/11/1983 3/02/1989 2/02/2031 $81,800.00 $2,098.40 $13,146.10 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/1162 M Granted 749.3 Ha 25/11/1983 3/02/1989 2/02/2031 $70,300.00 $2,233.20 $14,025.00 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/1220 M Granted 43.335 Ha 4/03/2005 14/11/2012 13/11/2033 $10,000.00 $1,032.00 $822.80 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/1221 M Granted 929.3 Ha 4/03/2005 14/11/2012 13/11/2033 $93,000.00 $2,749.49 $17,391.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/1236 M Granted 946 Ha 17/06/2005 25/11/2013 24/11/2034 $94,600.00 $2,795.39 $17,690.20 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/1237 M Granted 987 Ha 17/06/2005 25/11/2013 24/11/2034 $98,700.00 $2,912.98 $18,456.90 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/1238 M Granted 708 Ha 17/06/2005 25/11/2013 24/11/2034 $70,800.00 $2,112.74 $13,239.60 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/1239 M Granted 960 Ha 17/06/2005 25/11/2013 24/11/2034 $96,000.00 $2,835.54 $17,952.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/213 M Granted 978.05 Ha 14/04/1983 9/04/1986 8/04/2028 $97,900.00 $2,741.93 $18,307.30 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/223 M Granted 984.6 Ha 14/04/1983 9/04/1986 8/04/2028 $98,500.00 $2,758.29 $18,419.50 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/233 M Granted 966.9 Ha 14/04/1983 9/04/1986 8/04/2028 $96,700.00 $2,855.62 $18,082.90 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/243 M Granted 986.15 Ha 14/04/1983 9/04/1986 8/04/2028 $98,700.00 $2,912.98 $18,456.90 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/253 M Granted 968.38 Ha 14/04/1983 9/04/1986 8/04/2028 $96,900.00 $2,861.36 $18,120.30 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/263 M Granted 527.25 Ha 14/04/1983 28/11/2014 27/11/2035 $52,800.00 $1,596.45 $9,873.60 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/462 M Granted 476.25 Ha 1/11/1988 12/10/1989 11/10/2031 $47,700.00 $1,450.17 $8,919.90 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%),

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Lease Lease Type Lease Status Current Area Application Date Grant Date Expiry Date Commitment (AU$) Shire Rates (AU$) Rent (AU$) Holders

M 70/463 M Granted 359.65 Ha 1/11/1988 12/10/1989 11/10/2031 $36,000.00 $1,114.58 $6,732.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/464 M Granted 725.6 Ha 1/11/1988 12/10/1989 11/10/2031 $72,600.00 $2,164.37 $13,576.20 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/465 M Granted 359.55 Ha 1/11/1988 12/10/1989 11/10/2031 $36,000.00 $1,114.58 $6,732.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/466 M Granted 109.5 Ha 1/11/1988 12/10/1989 11/10/2031 $11,000.00 $1,032.00 $2,057.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M70/5542 M Granted 38.61 Ha 13/07/1989 6/04/2004 5/04/2025 $10,000.00 $1,032.00 $729.30 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/5643 M Granted 363.8 Ha 29/08/1989 27/04/1990 26/04/2032 $36,400.00 $1,126.05 $6,806.80 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/588 M Granted 360.15 Ha 31/10/1989 7/06/1990 6/06/2032 $36,100.00 $1,117.15 $6,750.70 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/589 M Granted 120.05 Ha 31/10/1989 7/06/1990 6/06/2032 $12,100.00 $1,032.00 $2,262.70 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/590 M Granted 402.55 Ha 31/10/1989 7/06/1990 6/06/2032 $40,300.00 $1,237.91 $7,536.10 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/591 M Granted 359.9 Ha 31/10/1989 7/06/1990 6/06/2032 $36,000.00 $1,114.58 $6,732.00 Newmont Boddington Pty Ltd (66.67%), Saddleback Investments Pty Ltd (33.33%), M 70/731 M Granted 300 Ha 3/12/1991 26/01/1993 25/01/2035 $30,000.00 $1,032.00 $5,610.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/7993 M Granted 925.4 Ha 21/01/1993 21/09/1993 20/09/2035 $92,600.00 $2,738.02 $17,316.20 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/944 M Granted 1.5305 Ha 21/05/1996 5/12/1996 4/12/2038 $5,000.00 $1,032.00 $37.40 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/945 M Granted 11.76 Ha 21/05/1996 5/12/1996 4/12/2038 $10,000.00 $1,032.00 $224.40 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/946 M Granted 0.3385 Ha 21/05/1996 5/12/1996 4/12/2038 $5,000.00 $1,032.00 $18.70 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/947 M Granted 15.63 Ha 21/05/1996 5/12/1996 4/12/2038 $10,000.00 $1,032.00 $299.20 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/948 M Granted 2.496 Ha 21/05/1996 5/12/1996 4/12/2038 $5,000.00 $1,032.00 $56.10 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/949 M Granted 34.925 Ha 21/05/1996 5/12/1996 4/12/2038 $10,000.00 $1,032.00 $654.50 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/950 M Granted 16.465 Ha 21/05/1996 5/12/1996 4/12/2038 $10,000.00 $1,032.00 $317.90 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/951 M Granted 3.88 Ha 21/05/1996 5/12/1996 4/12/2038 $5,000.00 $1,032.00 $74.80 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/952 M Granted 12.13 Ha 21/05/1996 5/12/1996 4/12/2038 $10,000.00 $1,032.00 $243.10 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/953 M Granted 2.6 Ha 21/05/1996 5/12/1996 4/12/2038 $5,000.00 $1,032.00 $56.10 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/954 M Granted 12.865 Ha 21/05/1996 5/12/1996 4/12/2038 $10,000.00 $1,032.00 $243.10 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/955 M Granted 1.349 Ha 21/05/1996 5/12/1996 4/12/2038 $5,000.00 $1,032.00 $37.40 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), M 70/9761 M Granted 861 Ha 14/01/1997 30/08/2013 29/08/2034 $86,100.00 $2,420.50 $16,100.70 South32 Aluminium (RAA) Pty Ltd (56%), South32 Aluminium (Worsley) Pty Ltd (30%), Japan Alumina Associates (Australia) Pty Ltd (10%), Sojitz Alumina Pty Ltd (4%), M 70/981 M Granted 52.19 Ha 25/03/1997 3/09/1997 2/09/2039 $10,000.00 $1,032.00 $991.10 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), ML 70/662 ML Granted 90 Ha 18/12/1981 1/01/2002 31/12/2022 $10,000.00 $974.00 $1,683.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), ML 70/663 ML Granted 90 Ha 18/11/1981 1/01/2002 31/12/2022 $10,000.00 $1,032.00 $1,683.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), ML 70/751 ML Granted 120 Ha 26/11/1981 1/01/2002 31/12/2022 $12,000.00 $1,032.00 $2,244.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), ML 70/752 ML Granted 120 Ha 26/11/1981 1/01/2002 31/12/2022 $12,000.00 $1,032.00 $2,244.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%), ML 70/753 ML Granted 50 Ha 26/11/1981 1/01/2002 31/12/2022 $10,000.00 $974.00 $935.00 Newmont Boddington Pty Ltd (66.66%), Saddleback Investments Pty Ltd (33.33%),

E = Exploration, G = General Purpose Lease, L = Miscellaneous Licence, M = Mining Lease, ML = Mineral Lease, P = Prospecting Permit 1 Newmont 100% beneficial owner - upon grant can be transferred 2 Newmont holds right to sub-lease 3 Newmont holds a sub-lease

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Note: Figure aligned to True North Figure 4-1: Mineral Tenure Map

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4.5 Surface Rights The Project has freehold ownership of all the eastern and central areas of operations. Within this freehold land are all the existing residue disposal areas, the plant site, almost all of the area of the main open pit from the former oxide operation, and all but one of the smaller satellite open pits from the former operation. The western portion of the operational area that is outside the freehold land is Crown land covered by native forest. Mining operations can be conducted in this area but with certain restrictions imposed by the State Government through the 1978 Mining Act that are applicable to forested Crown lands. The land to the north and to the east was acquired by the JV from Sotico Pty Ltd. (Sotico), a tree-farming company that is a wholly-owned subsidiary of Wesfarmers Ltd., a large WA public company, from December 2006 to May 2008, and September 2011 respectively. To the south and east of the Project is freehold farmland. The largest farm immediately south of the mine, Hotham Farm, was acquired by the NBGJV in December 2011. Figure 4-2 presents the land ownership for the region of the Project, whereas Figure 4-3 displays the land ownership surrounding the mining operation.

4.6 Royalties and Encumbrances Royalties levied are discussed in Section 1.5. Royalty payments for the current operations were first incurred in the latter part of 2009.

4.7 Native Title The Project area of interest is subject to a land claim registered under the NTA and referred to as the Gnaala Karla Booja Claim. The South West Aboriginal Land and Sea Council (SWALSC), which is the statutory body that represents the Native Title claimants, has sought to amalgamate the Gnaarla Karla Booja Native Title Claim with other claims in the southwest of the State to form the Single Noongar Native Title Claim. Newmont and the SWALSC have met regularly and the State has presented a formal proposal to the SWALSC for the resolution of all claims in the southwest.

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Note: NBG = Newmont Boddington Gold; WAPL = WA Prospecting Licence. Figure aligned to True North Figure 4-2: Surface Rights Map (Regional)

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Note: NBG = Newmont Boddington Gold. Figure aligned to True North Figure 4-3: Land Ownership Map (Mining Operation)

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The applicants of a claim have procedural rights under the NTA with respect to any act by another party, such as the JV, that may have an impact on Native Title (a future act) over land tenure where Native Title has not been previously extinguished. Such future acts by the JV where Native Title Claimants may have procedural rights include:  Grant of application for the one GPL sought for the F1 Residue Disposal Area (RDA) extension into Crown Land;  Conversion of Crown Land to freehold as part of the DWER land exchange (land swap). Separate negotiations with the Native Title Claimants, and the SWALSC, have resulted in the signing of several agreements in relation to community engagement and the protection of cultural heritage. NBG continues to enjoy a constructive and respectful relationship with the Gnaala Karla Booja group.

4.8 Permits Information on permitting for the Project is included in Section 20.0.

4.9 Environmental Considerations Information on environmental considerations for the Project is included in Section 20.0.

4.10 Social Considerations Information on social considerations for the Project is included in Section 20.0.

4.11 Project Risks The following significant factors or risks may affect access, title, or right or ability to perform work at the Project:  Accurate closure cost provisioning, management of rehabilitation stockpiles (topsoil, gravels etc.), changes in design of facilities;  WRD Management and generation of acid rock drainage (ARD);  Unapproved clearing of native vegetation;  Incorrect disposal of waste resulting in contamination of local area;  Spread of declared weed species or forest dieback disease;  Failure to obtain approvals and amendments to licences within desired timeframes;  Failure to fully understand regional groundwater interactions and local dewatering activities impact on the local river system;  Breach of commitments with the MBCPA;  Operations impacting sacred site;  Loss of revenue from scrap metal recycling program, which funds local community investment program;  Reputational damage with local community if needs and concerns are not addressed.

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4.12 Comments on Property Description and Location In the opinion of the QP:  Information provided by Newmont’s legal and tenure experts on the mining tenure held by Newmont in the Project area supports that the company has valid title that is sufficient to support the declaration of Mineral Resources and Mineral Reserves;  Information provided by Newmont’s legal and tenure experts supports that the Operations hold sufficient surface rights to enable mining operations and for the declaration of Mineral Resources and Mineral Reserves;  Based on the Environmental Impact Statement and the current state of environmental knowledge of the mine area, Mineral Resources and Mineral Reserves can be declared, and the mine plan is achievable;  At the effective date of this Report, environmental liabilities are limited to those that would be expected to be associated with a gold-copper mine of comparable scale, including roads, open pits, site infrastructure, waste and tailings disposal facilities. Newmont has appropriately addressed the potential and actual environmental impacts of the operation sufficient to support declaration of Mineral Resources and Mineral Reserves;  Closure and remediation requirements (refer to Section 20.0) have been addressed through the site closure plan and associated environmental bond requirements;  To the extent known, there are no other significant factors and risks that may affect access, title, or the right or ability to perform work on the property.

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5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Accessibility 5.1.1 Road The Project is located in the Shire of Boddington in the southwest region of the State of WA, approximately 17 km northwest of the township of Boddington. The township of Boddington is located 130 km southeast of Perth, is 14 km due west of the main Perth–Albany Highway, and is accessed by an all-weather sealed road. From the township of Boddington, access to the Project is northwest along Gold Mine Road off the Marradong–Bannister Road just north of the township of Boddington. All roads to the plant site, including Gold Mine Road, are bitumen-surfaced roads in good condition. Within the mine site, road access has been designated either as a regular trucking route, a regularly-used maintenance route, or an occasional-use route. The road classification affects the width of road, and where practical the final surface, with high- use roads bitumen-sealed where it makes commercial sense. The remaining road types are finished with a gravel surface. 5.1.2 Port The port of Bunbury is approximately 170 km southwest of Perth, and approximately 175 km southwest of the Project. The port is the trans-shipment point for copper concentrates produced from the mine.

5.2 Climate The Project climate is Mediterranean, with hot, dry summers and cool, wet winters. The coldest month is July, with a mean minimum temperature of 4.5 degrees Celsius (ºC) and the hottest month is January, with a mean maximum of 32.1ºC. Rainfall averages approximately 780 millimeters (mm) per year, with most precipitation falling between April and October. Average evaporation is approximately 1,490 mm per year. Mean relative humidity ranges from around 69% at 9 am to around 44% by 3 pm. Prevailing winds are from the east–southeast in summer and west–northwest in winter, and typically range from 10 kilometers per hour (km/h) to 22 km/h. Mining operations are conducted year-round.

5.3 Local Resources and Infrastructure The region is lightly populated. Most of the inhabitants are employed in mining, quarrying, farming, agriculture, and tourism, or in the servicing of these industries. The closest house is located approximately 6 km from the operations. The closest town is Boddington (population 2,226), which provides necessities and acts as a dormitory town for the mine. Most of services are sourced from Perth, which has a large, specialized infrastructure for mining support.

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The Project falls within several administrative jurisdictions, including the local government authority Shire of Boddington, and the State of WA. Surface rights and sufficiency of the rights to support current and planned mining operations is discussed in Section 4.5. Project infrastructure is discussed in Section 18.0.

5.4 Physiography The Project is located on the Darling Plateau in an area of deeply weathered, undulating landscape that ranges from 200 to 500 meters Relative Level (mRL). A series of monadnocks occur in the area, which include Mount Wells, immediately to the northwest of the mine. Local relief is approximately 100 meters (m), with shallow valley floors adjacent to broadly convex hills. The mine is located in the catchment area of Thirty Four Mile Brook, a tributary of the Hotham River, which itself flows into the Murray River and then into the Peel Harvey Inlet. The water quality of the Hotham River and Thirty Four Mile Brook is degraded by inputs of salt and nutrients as they flow through agricultural areas. Flows in the Hotham River and Thirty Four Mile Brook are highly seasonal, with considerable variation from year to year. The Hotham River is degraded by salt (total dissolved solids ranging from 2,000 milligrams per Litre (mg/L) to 8,000 mg/L and nutrient inputs from agricultural areas. During summer, the river is almost reduced to a series of pools. Thirty Four Mile Brook is generally a fresh stream, except where it flows through agricultural areas to the south of the operations area. Pools in the brook may have salinities levels that reach as much as 19,000 mg/L in summer. The existing beneficial uses of these streams are livestock and ecosystem support, which are consistent with the water quality observed in the water bodies. Groundwater in the oxide and bedrock aquifers is also a water supply source although total dissolved solids generally vary from 1,000 mg/L to 5,000 mg/L. The previous mining operation resulted in a substantial drawdown of the oxide aquifer, which began recovering when the previous mining operations ceased. This drawdown has now re-commenced with current dewatering activities. The Mining Leases are located largely on private forested land typical of the eastern Jarrah forest, on Pindalup and Dwellingup landform units. The forests have been subject to selective logging for many decades. The land to the north and east of the mine has historically been managed for timber production under the ownership of Sotico, which retains the right to harvest timber from this area. NBG has recently purchased some of this area to the east and north. Land to the west of the Project area is State Forest, whereas much of the land to the south has been cleared for agriculture and is commonly used for sheep grazing and mixed cropping. Approximately 690 taxa have been identified in the Project area. No protected flora species that are classed as Rare Flora have been identified in botanical surveys. A total of 13 plant community or site-vegetation types have been classified, predominantly varieties of eucalyptus forest. Twelve faunal species of concern have been identified in the area, and include mammals, lizards, snakes, and birds. Jarrah dieback disease is associated with the introduced soil-borne fungus Phytophthora cinnamomi and is considered to be the major disease problem of WA’s native forests. The

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NBG areas have been mapped for dieback incidence, and a comprehensive forest hygiene program has been operating successfully since the commencement of the Project.

5.5 Comments on Accessibility, Climate, Local Resources, Infrastructure and Physiography In the opinion of the QP:  The existing local infrastructure, availability of staff, methods whereby goods could be transported to the Project area are well-established and well understood by Newmont, and can support the declaration of Mineral Resources and Mineral Reserves;  The Project covers an operating mine, and all required infrastructure for the life-of-mine plan discussed in this Report are in place;  Within Newmont’s ground holdings, there is sufficient area to allow construction of any Project infrastructure that may be required in the future;  There are no significant factors and risks that may affect access, title, or the right or ability to perform work on the property.

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6.0 HISTORY A geochemical prospecting program undertaken by the Geological Survey of WA during 1979 identified anomalous Au, As, Cu, Pb, Mo, and Zn in a zone about 5 km long and 500 m wide area within the northern extent of the Saddleback Greenstone Belt. In July 1980, geologists of Reynolds Australia Mines explored the anomaly. Rock chip sampling of the laterite lead to the discovery of a significant gold-mineralized zone in an area within the geochemical anomaly. Exploration drilling activities commenced in July 1982 using RC techniques to drill vertical holes through the oxide profile down to refusal at the bedrock interface and collecting samples at 1 m intervals. In 1984, a Notice of Intent (NOI) was filed with the WA government, outlining details of the proposed Project and assessing environmental impacts. An Environmental Review and Management Program (ERMP) was submitted by the JV in January 1985, leading to Project environmental approval in late 1985. Open pit mining began in March 1987, with production commencing in August 1987 at a design rate of 3 Mtpa, rising to 4.5 Mtpa in October 1987. A plant expansion to 6 Mtpa occurred in 1989. Isolated areas of copper–gold mineralization were discovered within the orebody in 1983. A separate 250,000 tonnes per annum (tpa) flotation/leach plant (referred to as the Supergene or Basement Plant) was commissioned in March 1991, producing a copper–gold concentrate and gold bullion. Processing of copper–gold ores was complete in March 1993. High-grade gold-bearing quartz veins were discovered during 1990 in the northern section of the deposit within the oxide and bedrock zones. The Jarrah Decline underground operation to access these veins commenced in October 1992. The bedrock mineralization was processed through the Supergene Plant up until 1997. An additional large bedrock resource, known as the Wandoo deposit, was identified in 1994. Additional deposits, outside the main deposit area, continued to be identified through exploration programs and six groups of satellite pits were mined between 1993 and 2001. All but one of these pits have now been backfilled and rehabilitated. Increases in production continued, and by 1996 oxide and bedrock ores were processed at a rate of 8.6 Mtpa. In November 1998, the JV acquired the Hedges Gold Mine leases located adjacent to the existing gold operations. Ore from the Hedges Mining Lease was treated at the Oxide and Supergene Plants. Between 1989 and 1991, the Project was the largest gold producer in Australia. Oxide resources were depleted in November 2001 and processing ceased on 30 November 2001. In December 2001, the mine was placed on care and maintenance. The plant and infrastructure were decommissioned, and redundant equipment was sold and removed from site. Several feasibility studies were undertaken during the late 1990s to devise an economic method of extracting the gold mineralization from the Wandoo and other bedrock sources. The 1997 feasibility study focused only on the Wandoo South deposit and was completed under the direction of BHP Billiton Worsley Alumina Pty Ltd, acting as manager for the JV.

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This study was based on conventional semi-autogenous grind (SAG) and ball milling followed by sequential flotation and leaching, with a design throughput of 11.24 Mtpa. Estimated Project economics were considered insufficient to warrant development. Following the purchase of Hedges in 1998, an update to the feasibility study, which used both Wandoo North (Hedges) and Wandoo South, was undertaken in 2000. This study was based on a throughput rate of 22 Mtpa to produce an average of 600,000 oz of gold and 22,500 tonnes (t) of copper per annum over a mine life of at least 15 years. In 2002, management of the JV was restated, with Newmont assuming operator status as NBG. During 2003, an additional update was completed, in three separate phases of work:  Phase 1 Facility Definition – Included review of previous study outcomes, value engineering, trade-off studies, benchmarking studies, infrastructure studies, metallurgical reviews, design criteria and flowsheet definition that resulted in agreed flowsheets and design criteria for Phase 2;  Phase 2 Prefeasibility Study – Focused on the development of the engineering to a level that produced capital (nominally +25 %) and operating (+15 %) cost estimates in support of the prefeasibility study report;  Phase 3 Bankable Feasibility Study – Focused on the development of the engineering to a level that supported capital and operating cost estimates as required to complete a feasibility study report, in sufficient time to allow owner review during 2006. The feasibility study indicated a positive life-of-mine cash flow based on open pit mining and conventional crush-mill-float processing for the copper and cyanide leach processing for the gold. The JV announced a decision to proceed with Project development in the first quarter of 2006. Commercial production began in 2009. In March 2011, the operation produced its first million ounces of gold, reaching 5 Moz of gold production in 2016 and 6.9 Moz by the end of 2018.

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7.0 GEOLOGICAL SETTING AND MINERALIZATION

7.1 Regional Geology The Boddington deposit is hosted within the Saddleback Greenstone Belt, which lies in the southeastern corner of the Archaean Yilgarn Craton (refer to Figure 7-1). The Saddleback Greenstone Belt comprises a steeply-dipping and extensively faulted sequence of sedimentary, felsic to mafic volcanic and pyroclastic rocks that have been metamorphosed to greenschist–amphibolite facies. The belt is approximately 50 km long and 8 km wide and is surrounded by granitic and gneissic rocks. Age dates range from 2,715 million years (Ma) to 2,690 Ma. The greenstones are interpreted to have been emplaced in an island arc setting. Ductile deformation followed, then a second period of supracrustal deposition, again probably in an island arc setting. This second phase was accompanied by coeval granodiorite–tonalite intrusion. Greenschist facies metamorphism followed, and all rocks were affected by brittle– ductile faults. A late monzogranite intrudes the greenstone belt just east of the mine area and is attributed to melting of mid-crustal rocks in an intraplate setting. The Saddleback Greenstone Belt has been subdivided into three formations by Wilde (1976):  Hotham Formation: Metasedimentary rocks; restricted to the southwestern part of the Saddleback Greenstone Belt;  Wells Formation: Felsic to intermediate volcanic rocks and associated granitoid intrusives. This formation is the main host to economic mineralization at the Project;  Marradong Formation: Meta-basaltic lavas and related doleritic/gabbroic intrusives. This formation includes a significant number of ultramafic intrusives in the northern half of the Saddleback Greenstone Belt. The Wells Formation can be informally subdivided into three main units:  ‘Lower’ Wells Formation: A mixed package of intermediate andesites and diorites with mafic basalts and dolerites possibly highlighting the gradational nature between the Marradong and Wells Formations;  ‘Main or Central’ Wells Formation: This is considered by NBG to be the favorable package for mineralization of volcanic andesites and intrusive diorites;  ‘Upper’ Wells Formation: A predominantly mafic package of basalts and dolerites with minor lenses of intermediate and metasedimentary rock types.

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Note: Figure aligned to True North Figure 7-1: Regional Geological Map

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The geological pattern of the regional geology suggests a gradational, rather than a sharp, contact between the mafic volcanic succession of the Marradong Formation and the intermediate volcanic rocks of the (Lower) Wells Formation. These latter andesitic to dacitic volcanic rocks are in turn overlain by an extensive basaltic package that constitutes the bulk of the southern half of the Saddleback greenstone belt. The basalts also contain a few lenses of andesitic and dacitic rocks, again suggesting a more or less gradational contact. A narrow belt of metasedimentary ± felsic volcanic-derived schists (Hotham Formation) is preserved along the southwestern margin of the greenstone belt. The overall stratigraphic trend is at a low angle to the western and eastern granitoid contacts with the greenstones. The Saddleback greenstone belt lies along a major north–northwest- trending zone of Early Paleozoic reactivation of the southwestern portion of the Yilgarn Craton (Libby and DeLaeter, 1998). Several structures have been identified that are controlling elements on the localization and form of mineralization, these being:  Northeast-striking fault corridors, which appear to compartmentalize the deposit. These structures appear to have offset favorable host rocks pre-mineralization;  Intersection of late-stage faults with early ductile quartz–sericite shear zones;  Intersection of west–northwest or northwest-trending faults with structurally-favorable lithologies;  Late brittle–ductile west–northwest- or northwest-trending faults that have subvertical dips, which show elevated mineral abundances, and mineralization-related alteration assemblages.

7.2 Project and Deposit Geology The Boddington deposit lies within a 6 km strike length of the Wells Formation, consisting of felsic to intermediate volcanic rocks and related intrusive rocks. For descriptive purposes the deposit is subdivided into Wandoo South and Wandoo North at approximately 12200 N, subdividing the two main centers of bedrock mineralization. A geology plan for the Project is included as Figure 7-2.

7.2.1 Wandoo South Wandoo South is centered on a composite diorite stock, the Central Diorite, which has a known strike length of approximately 1,200 m and thicknesses varying from 300 m to 600 m. Five intrusive phases are known within the stock. A simplified geological cross-section for the Wandoo South deposit is included as Figure 7-3.

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Note: Figure aligned to Mine Grid Figure 7-2: Geology Map at 150 mRL

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Figure 7-3: Interpretative Scaled Cross-Section across the South Pit Area The southern portion of the Central Diorite strikes north, and dips subvertical and steeply to the west, with an apparent southerly plunge. To the north, the strike of the diorite changes from north to northwest, following the orientation of a transecting dolerite dyke. The dip changes from westerly, to subvertical, to steeply to the southwest. The Central Diorite mineralization becomes patchy towards the southern end of Wandoo South, although the high-grade Southern Diorite Deeps zone is hosted within

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the southern portion of the same diorite. Down-dip extensions of the Central Diorite remain open, but the diorite body appears to pinch out towards the northwest. Volcanic units are mapped in contact with the Central Diorite, and include:  Southern Volcanics: A sequence of porphyritic volcanic rocks in the south and west;  Northern Volcanics: A sequence of tuffaceous volcanic rocks to the north;  Eastern Volcanics: Characterized by glomeroporphyritic plagioclase. Separated from the Central Diorite by the Eastern Shear Zone, a north-striking, steeply west- dipping brittle, ductile tectonic feature. Thin units of fragmental volcaniclastic rocks consisting of angular to well-rounded diorite and andesite clasts ranging from fine ash to agglomerate sizes are common within and around the diorite stock. A series of fine-grained microdiorite dykes, ranging from a few centimeters to several meters wide, cross-cuts andesite, diorite and fragmental lithologies. Diorites and andesites are the preferred host lithologies for mineralization, with lesser mineralization recorded in the fragmental rock types and microdiorites. A suite of Proterozoic dolerite dykes with three prominent orientations cross-cuts the entire mine sequence but do not host any significant mineralization.

7.2.2 Wandoo North Wandoo North is dominated by diorites, with lesser fragmental volcanic rocks. The diorites at Wandoo North are mainly porphyritic, compared to the predominantly aphyric diorites of Wandoo South. Wandoo North and Wandoo South are broadly similar, and formed as part of the same mineralized system; similarities include:  Widespread silica-biotite-actinolite alteration (D1 alteration). Considered to be related to the intrusion of the host diorites;  Quartz–sericite–pyrite ± arsenopyrite alteration. Associated with north–south striking sub-vertical to east-dipping broad ductile shear zones (D2 shears) which overprint earlier, regional silica–biotite alteration;  Quartz–albite–pyrite–epidote-altered brittle/ductile structures (D3), which overprint D2 shears, and have two prominent orientations;  Brittle stockworks (D4) associated with strong biotite–clinozoisite–sulfide (pyrrhotite–chalcopyrite) alteration. D4 overprints all the above alteration assemblages, and mineralization at Wandoo North and Wandoo South seems to be associated with the D4 event. The main differences between Wandoo North and Wandoo South can be summarized as follows:  At Wandoo North, a pervasive chlorite alteration is observed to overprint D2 shear zones. This alteration strengthens in intensity towards the Western Sediments;  Chlorite and carbonate are seen in the alteration assemblage (retrograde?) at Wandoo North, but are absent in Wandoo South;  Actinolite veins and albite-rich veins, common in Wandoo South, are less abundant in Wandoo North;  Lithologies are generally more felsic at Wandoo North. A suite of rhyodacitic porphyries are identified at Wandoo North, but rarely observed at Wandoo South;

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 Diorites at Wandoo North have closely packed porphyritic texture (i.e. with relatively large, visible crystals of feldspar, referred to as phenocrysts), compared to the predominant equigranular texture of Wandoo South diorite (which is commonly referred to as aphyric diorite due to a lack of phenocrysts);  Quartz–albite–sulfide veins with coarse molybdenum, a dominant control for molybdenite distribution of the deposit, are found in both areas but are dominant in Wandoo South;  Thin felsic and intensely epidote-altered lithologies that are barren in Au and Cu are seen in the Wandoo North area but not reported from Wandoo South;  Presence of strongly sheared actinolite-bearing barren zone towards the western part of, and beneath the North Wandoo area. In rare cases, Archaean dolerites with anomalous gold grades have intruded the mineralized sequence. However, the vast majority of dolerites in the Project area are identified as post-mineralization Proterozoic dykes.

7.3 Mineralization Alteration types associated with gold and copper mineralization are clinozoisite–biotite– actinolite-sulfide ± silica veins/veinlets/fractures/clots and late actinolite–sulfide veins with clinozoisite–biotite and/or albite alteration haloes. These vein types form the basis of the stockwork mineralization of the Wandoo deposit. The majority of these mineralized late actinolite veins are not large enough to be correlated between drill holes; the exception is the Main Actinolite Vein in Central Diorite. However, broad zones of actinolite and clinozoisite alteration can be traced for a few hundred meters. The interpreted sequence of alteration assemblage is roughly analogous to that which occurs within porphyry-style deposits. However, the alteration is not well zoned as for typical ‘porphyry style’ copper–gold deposits, but rather focused into lithological, structural, or alteration zones. Garnet, commonly found in porphyry deposits, has not been identified in the alteration mineral assemblage in the Boddington mine area. The structural setting for the gold–copper–molybdenum mineralization is a major regional sinistral-strike-slip regime controlled by east–southeast to west–northwest compression. The major mineralized structures in the Wandoo resource area are:  D1 north–south-trending, near vertical, silica–biotite shear zone ± quartz veins; The Jarrah East quartz vein is hosted by a D1 shear zone and mineralized by later cross- cutting D4 deformation;  D2 north–south-trending, near vertical, quartz–sericite ± arsenopyrite shear zone ± quartz veins;  D3 northeast-trending, near vertical and 50º northwest-dipping, quartz–albite ± pyrite shear zone ± quartz veins. The Herring Fault is a D3 type quartz–albite strongly-foliated shear zone (possibly a sheared felsic porphyry) overprinted by a late-D3 quartz ± albite ± epidote ± pyrite shear zone. A D3 shear zone hosts the Jarrah West quartz vein that was later mineralized as a result of cross-cutting D4 structures.  D4 northwest-, north–northwest- and east–west-trending biotite ± actinolite ± clinozoisite ± sulfides (pyrrhotite–chalcopyrite) shear zones. The main northwest- and north– northwest-trending anastomosing D4 corridor is centered on the Wells Formation. The

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main east–west-trending, high-grade gold–copper, A2-type actinolite veins are mainly concentrated in the Wandoo South deposit area, which is interpreted as the regional dilation site. The Wandoo deposit contains D1 and D2 shear zones that are commonly overprinted and reactivated during D4 by biotite-sericite weakly to moderately foliated. All D1, D2, and D4 structures are observed to rotate to some degree into parallelism. Two mineralization stages have been recognized at the Project. The earliest phase consists of widespread silica–biotite alteration and complex quartz + albite + molybdenite ± muscovite ± clinozoisite ± chalcopyrite veins, all of which are variably deformed by ductile shear zones. The second, major, alteration stage cross-cuts the first, and comprises:  Quartz + albite + molybdenite ± muscovite ± biotite ± fluorite ± clinozoisite ± chalcopyrite veining;  Clinozoisite + chalcopyrite + pyrrhotite + quartz + biotite veins that host low-grade Au-Cu mineralization;  Actinolite + chalcopyrite + pyrrhotite ± quartz, carbonate + chlorite veins that host high- grade mineralization;  The bulk of the gold mineralization is associated with the late-stage fracture-centered clinozoisite–actinolite-biotite–sulfide alteration event with gold grades in this alteration being typically less than 3 g/t Au and averaging approximately 0.5 g/t Au to 1 g/t Au. The second mineralizing alteration style of late actinolite sulfide veining (with biotite – clinozoisite salvage) contains generally higher levels of gold, averaging 5 g/t Au to 8 g/t Au, but ranging from 30 g/t Au to 70 g/t Au in the larger veins. Gold distribution typically displays the following characteristics:  Visible gold generally occurs as inclusions and at grain boundaries with grain size generally less than 15 µm;  Visible gold occurs within both silicate and sulfide species. Within the sulfides, visible Au is most common in chalcopyrite, pyrrhotite and pyrite, mainly because these are the most dominant sulfide species in the deposit. Molybdenite, arsenopyrite, and sphalerite also contain visible Au;  There is no preferred association of visible gold with arsenopyrite;  Within the non-sulfide minerals, actinolite and quartz have the most common association with visible Au;  It is very common for visible Au to be associated with bismuth. In decreasing order of abundance, the bismuth species are: (i) native Bi; (ii) maldonite (Au2Bi); (iii) hedleyite (Bi4Te); and (iv) hessite (Ag2Te);  The highest concentration of visible gold is at Jarrah (location of the underground quartz vein). Elsewhere, the visible gold is evenly distributed through the deposit, although still relatively rare.

 The Wandoo resource sulfide distribution is dominated by chalcopyrite (CuFeS2), pyrite (FeS2) and pyrrhotite (FeS), listed in decreasing order of abundance. The distribution of sulfide species has been examined and the following additional observations noted:  The distribution of the three major sulfide species is reasonably consistent across the deposit, with: (i) increased concentrations of chalcopyrite at Southern Diorite Deeps, Blob

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and Pipeline; (ii) pyrrhotite is best-developed at Pipeline and Southern Diorite Deeps along with other higher concentrations at Central Diorite; and (iii) pyrite is best-developed at Pipeline and Southern Diorite Deeps;  Arsenopyrite (FeAsS) is most common within the South Zone, Far North and Blob domains of Wandoo North. This is consistent with the review of the arsenic distribution from the exploration database;

 Molybdenite (MoS2) is mainly within Southern Diorite Deeps, Central Diorite, North Diorite and A Breccia domains, which is consistent with the review of molybdenum distribution from the exploration database;

 Cubanite (CuFe2S3) almost always occurs as exsolution lamellae within chalcopyrite and is distributed throughout the deposit based on petrographic review;  Mackinawite (tetragonal FeS) is common as an accessory mineral throughout the deposit;  Other accessory sulfide minerals in decreasing order of abundance are sphalerite, pentlandite, covellite, bismuthinite, digenite, marcasite and galena.

7.4 Comment on Geological Setting and Mineralization In the opinion of the QP:  The understanding of the lithologies, and geological, structural, and alteration controls on mineralization, and the evaluation of the mineralogy is sufficient to support estimation of Mineral Resources and Mineral Reserves;  The geological knowledge of the area is adequate to reliably inform mine planning.

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8.0 DEPOSIT TYPES The most recent research suggests, that mineralization occurred as a result of interaction or overprinting of two different processes (Stein et al., 2001; McCuaig et al., 2002). The deposit model for the Project has aspects related to the mineralization styles of:  A deformed and metamorphosed porphyry deposit, formed during convergent plate tectonics (Roth, 1992; Barley et al., 1992);  Post-peak metamorphism, shear-zone-hosted deposit post-dating intrusive host rocks by at least 25 Ma (Allibone et al., 1998).

8.1 Gold Porphyry Deposits Porphyry deposits occur throughout the world in a series of extensive, relatively narrow, linear metallogenic provinces. They are predominantly associated with Mesozoic to Cenozoic orogenic belts in western North and South America and around the western margin of the Pacific Basin, particularly within the South East Asian Archipelago. However, major deposits also occur within Paleozoic orogens in Central Asia and eastern North America and, to a lesser extent, within Precambrian terranes (Sinclair, 2006). Porphyry deposits are large and typically contain hundreds of millions of tonnes of mineralization, although they range in size from tens of millions to billions of tonnes. Grades for the different metals vary considerably but generally average less than 1%. In porphyry Cu deposits, Cu grades range from 0.2% to more than 1% Cu; Mo content ranges from approximately 0.005% to approximately 0.03%; Au contents range from 0.004 g/t to 0.35 g/t; and Ag content ranges from 0.2 g/t to 5 g/t. Rhenium is also a significant by-product from some porphyry Cu deposits. Some Au-rich porphyry Cu deposits have relatively high contents of Platinum Group Elements (PGE) (Mutschler and Mooney, 1995; Tarkian and Stribrny, 1999, in Sinclair, 2006). Copper grades in porphyry Cu–Au deposits are comparable to those of the porphyry Cu subtype, but Au contents tend to be consistently higher, averaging between 0.2 to 2.0 g/t Au. Sillitoe (2000) suggested that porphyry Cu deposits should contain >0.4 g Au/t to be called Au-rich. However, Au is an important co-product at grades as low as 0.2 g/t Au. Although the number of deposits in this class is limited, deposits such as Grasberg in Indonesia indicate that porphyry Cu–Au deposits can contain significant Au as well as Cu resources (Sinclair, 2006). Most gold-rich porphyry intrusives consist of a series of both pre- and post-mineralization intrusions. The pre-mineralization intrusions are generally equigranular in texture and genetically related to the porphyry stock, and often intrude along the shoulders of the pre-mineralization intrusion. Post-mineralization dikes and plugs and diatremes are also commonly associated. Various hydrothermal breccias occur as early orthomagmatic (strong K-silicate altered) and/or late phreatic and phreatomagmatic varieties. Copper and gold grades in the early orthomagmatic breccias may be substantially higher than in the surrounding porphyry rocks, while later breccia types are generally of sub-economic grade. Large (>0.5 km wide) low-grade or barren diatreme breccias and minor pebble dikes often conclude the evolution of gold-rich porphyry systems (Sillitoe, 2000).

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Most gold-rich porphyry systems consist of varying quantities of six principal alteration types (Sillitoe, 2000):  Ca–Na silicate alteration;  K–silicate (potassic) alteration;  Propylitic alteration;  Intermediate argillic (sericite–clay–chlorite) alteration;  Sericitic alteration;  Advanced argillic alteration. Most gold in gold-rich porphyry systems is associated with the K–silicate alteration phases. An Os–Re study completed on the Project indicated that an initial mineralization phase dated at 2,700 Ma is associated with an intermediate shallow level intrusion that was likely part of an early arc sequence (Stein et al., 2001). This supports an interpretation of an Archaean analogue to Phanerozoic porphyry-style Au–Cu mineralization (Stein et al., 2001).

8.2 Shear-zone Hosted Deposits Shear-zone deposits can also be described as mesothermal gold, metamorphic gold, gold- only, lode gold, structurally-controlled, greenstone-hosted, and turbidite-hosted deposits. These deposits can form in variably deformed metamorphic terranes that formed during Middle Archaean to younger Precambrian, and continuously throughout the Phanerozoic. Host geological environments are typically volcano–plutonic or clastic sedimentary terranes, but gold deposits can be hosted by any rock type. There is a consistent spatial and temporal association with granitoids of a variety of compositions. Host rocks are metamorphosed to greenschist facies, but locally can achieve amphibolite or granulite facies conditions. Gold deposition occurs adjacent to first-order, deep-crustal fault zones. These first-order faults, which can be hundreds of kilometers long and kilometers wide, show complex structural histories. Economic mineralization typically formed as vein fill of second- and third-order shears and faults, particularly at jogs or changes in strike along the crustal fault zones. Mineralization styles vary from stockworks and breccias in shallow, brittle regimes, through laminated crack-seal veins and sigmoidal vein arrays in brittle-ductile crustal regions, to replacement- and disseminated-type orebodies in deeper, ductile environments. Mineralization can be disseminated, or vein hosted, and displays a timing that is structurally late, and is syn-post-peak metamorphism. At Boddington, the bulk of the mineralization appears to be related to a later-stage mineralizing event, dated at approximately 2,625–2,615 Ma, which is associated with a potassium-rich, post-tectonic magmatic suite (McCuaig et al., 2002). The Os–Re signature is compatible with that of orogenic gold deposits (Stein et al., 2001).

8.3 Comment on Deposit Types In the opinion of the QP:  The understanding of the deposit type was appropriate in guiding initial exploration activities, is suitable for current exploration programs, and is sufficient to support estimation of Mineral Resources and Mineral Reserves.

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9.0 EXPLORATION Exploration commenced on the Project in 1980, and exploration within the Saddleback Greenstone Belt has been ongoing up until 2013. Exploration has been undertaken by either NBG, its precursor companies (e.g. initial gold exploration by Reynolds), or by contractors (e.g. airborne geophysical surveys, hydrological surveys and geotechnical studies).

9.1 Grids and Surveys The difference between NBG mine grid (GN) and magnetic north (MN) as at 31 December 2015 is 39º 54’ 26” and the difference between MN and true north (TN) is 1º 43’ 52”. The difference between TN and MGA94 Zone 50 (MGA GN) is 0º 20’ 57”. The grid reference information uses Australian Geomagnetic Reference Field Computation as at 31 December 2015 based on NBG survey station MP1 (11,704.851N, 9,820.845E NBG mine grid). The topographic surface used to delimit block models is constructed from an as-mined surveyed pickup that is updated on a monthly basis.

9.2 Geological Mapping Very limited amount of bedrock exposure in the Project area restricted surface mapping, while most geological mapping has been derived from logging drill core and drill chip samples.

9.3 Geochemical Sampling Geochemical sampling was completed as part of the initial, first-pass exploration program, and has been superseded by data obtained from drilling and mining operations. Samples collected included stream sediments (bulk-leach extractable gold or BLEG), soil (mobile metal ion or MMI) and rock chip and grab samples. To date, a total of 15,995 geochemical samples have been collected over the Saddleback Greenstone Belt (refer to Figure 9-1). At varying times during the initial oxide mining phase, soil samples were collected on grids that had sample spacing ranging from 50 m x 100 m out to 200 m x 200 m over regional areas of the Saddleback Greenstone Belt.

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Note: Figure aligned to True North. Grid squares are 10 km x 10 km. Although Project tenure outlines are not shown, sampling took place within the confines of the mineral tenure held by the BGMJV at the time that sampling occurred. Current tenure outlines are included in Figure 4-1 Figure 9-1: Geochemical Sample Locations

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A study to characterize the multi-element geochemistry of the Wandoo resource area was conducted in 2004; 9,587 individual sample pulps from nine typical sections representing key mineralized domains were composited into approximately 12 m intervals and analyzed for a suite of multi-elements. The purpose of the study was to enhance the understanding of the geochemical variability within the Wandoo deposit and determine the key drivers in the ore response for flotation, leaching, cyanide consumption, and solution copper levels to assist in sample selection for the metallurgical test-work program.

9.4 Geophysics The Boddington area has been subject to both airborne and ground geophysical surveys, as documented in Table 9-1. The major survey locations are presented in Figure 9-2.

9.5 Pits and Trenches Trenches were dug during the initial development phase of the oxide pit mining to supply additional profile information on the oxide mineralization and distribution of the gold but have since been mined-out.

9.6 Petrology, Mineralogy and Other Research Studies Since 1980, several structural, petrology, mineralogy, lithogeochemical, and research studies have been completed on the Project area. The petrological and mineralogical studies have primarily been completed to quantify mineralization for use in designing appropriate process routes for the Boddington bedrock mineralization. Lithogeochemical studies were completed to address potential contaminant and by-product elements. Three Honors theses, two MSc theses and one PhD thesis have been completed on the deposit:  Monti, R., 1986: The Formation of an Auriferous Laterite Profile at the Boddington Gold Deposit, Western Australia: BSc Honours thesis;  Harrild, A., 1989: The Boddington Gold Deposit – A Review of Current Knowledge, Some Aspects of Greenstone Belts and Porphyry Copper Mineralization: BSc Honours thesis;  Roth, E., 1992: The Nature and Genesis of Archaean Porphyry-style Au–Ag–Cu Mineralization at the Boddington Gold Mine, Western Australia: PhD thesis;  Kalleske, N., 2010: Mineralogical and Petrogenetic Study of Gold Ore from the Boddington Gold Deposit, W.A.: BSc Honours thesis;  Crawford, A.F., 2011: The structural, Chemical and Mineralogical Controls On High Grade Au Mineralisation within the Boddington Deposit, Western Australia; A Petrographic and Microprobe Investigation: MSc thesis;  Jones, H.S., 2011: Controls on High Grade Gold Mineralisation in Southern Diorite Deeps at NBG Mine: MSc thesis.

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Table 9-1: Geophysical Surveys

Survey Date Comment Airborne Kevron 1993 3,220 line km at 50 m and 100 m line spacing, 50 m flying height Aerodata 1996 8,290 line km at 50 m and 100 m line spacing, 60 m flying height Ground 5,243 stations with varying station spacing in five separate data collection Gravity 1995 to 2008 campaigns, comprising 3,030 regional gravity stations and 2,213 detailed stations 97.5 line km of data collection. 50 m and 100 m dipole spacing and 175 to 200 m line separation. Comprised 10 km at Conveyor, 12.5 km at Hume Tank; 6 km at Eastern MIMDAS 2004 to 2006 Southern Diorite Deeps and 6 km at South Southern Diorite Deeps; 12 km plant site and 12.5 km South Southern Diorite Deeps; 38.4 km waste dumps. 2.1 km of dipole-dipole, 100 m electrode spacing. 0.36 km2 coverage of gradient IP 1999 array Fixed loop–Jarrah–13500N–14300N, 9800E–11000E; Moving loop–Jarrah–13550N– TDEM 1999 13750N, 10000E–11000E; Fixed loop–Mallee–12200N–13000N, 11400E, 12125E; Moving loop–Mallee–12450N & 12850N, 11400E–12200E & 11400E–12000E Drill hole 2001 WBD12770, WBD13080, WBD13485, WBD13365 TDEM Three lines surveyed; electrodes at WBD12770–120 m; WBD12770–246 m; Mise-a-la- 2001 WBD13080–312 m. Area of the survey encompassed by 9400E–9800E, 12500N– masse 13300N, approximately 0.24 km2 Natural gamma, magnetic susceptibility, resistivity, EM conductivity. Holes logged: Wireline 2000 WBD12500-008, 300 m; WBD12770-003, 294 m; WBD13985-002, 737 m; Logging WBD13485-001, 880 m; WBD13365-001, 819 m

Note: MIMDAS = Mt. Isa Mining Distributed Acquisition System; IP = induced polarization; TDEM = time-domain electromagnetic

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Note: Figure aligned to True North. Grid squares on map are 10 km x 10 km. Although Project tenure outlines are not shown, sampling has taken place within the confines of the mineral tenure held by the NBG at the time that sampling occurred. Current tenure outlines are included in Figure 4-1 Figure 9-2: Location Map of Major Geophysical Surveys

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9.7 Exploration Potential After discovery of the large Boddington gold deposit, surface exploration programmes involving surface geological mapping, geochemical sampling, assaying of shallow vacuum (bauxite) drill holes, and limited drilling of the oxide profile down to top of bedrock have been conducted. Anomalous gold values have been discovered up to 20 km south, but generally occur in narrow quartz veins, and have limited potential to host mineable quantities of gold. Anomalous gold from surface sampling, albeit smaller in area and weaker in signature than that over the main orebody, warrants follow up exploration, primarily by drilling. The basalts of the Marradong Formation within the Saddleback Greenstone Belt dominate the southern two-thirds of the Saddleback Greenstone Belt and are considered to have low prospectivity.

9.8 Comment on Exploration In the opinion of the QP, the exploration programs completed to date are appropriate to the style of the Project mineralization.

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10.0 DRILLING Approximately 32,716 drill holes have been completed for approximately 2.44 Mm of drilling, in RAB, AC, RC and core drill holes (refer to Table 10-1). Supporting Mineral Resource estimation are a total of 7,236 drill holes for approximately 1.39 Mm, comprising 2,463 RC, 1,218 grade control RC, and 3,555 core drill holes (refer to Table 10-2 and Figure 10-6). The drill holes presented in Table 10-2 are a subset of those presented in Table 10-1. Drill holes were drilled between 1981 and 2018.

10.1 Drilling Methods 10.1.1 Vacuum Drilling Vacuum drilling is a method whereby air is sucked down the drill hole and back up through the center of a single small diameter rod string. Air is sucked up by a vacuum pump through the rods to create a vacuum at the bit face, where the drilled sample is collected. Drill holes are typically small diameter, usually 50 mm, drilled by a tractor- mounted fixed mast drill rig. The drill holes are entered as ‘vertical’ into the database. Vacuum drilling is considered a geochemical sampling tool, and is only undertaken in dry ground conditions, and preferably in dry weather to avoid sample contamination. The method can only drill loose (sand) or highly weathered (laterite) material. A disadvantage of the method is the use of silver-soldered tungsten carbide bits, necessitating careful watching of ground conditions to avoid contamination of Ag, Cu, Zn, and W from the bit in hard drilling conditions. Figure 10-1 presents the locations of the vacuum drill holes. 10.1.2 RAB Drilling RAB drilling, is an open-hole technique in which compressed air is injected down the drill pipe to recover the cuttings up the outside of the drill stem to the surface. This can lead to contamination, particularly in unconsolidated materials. The cuttings are generally piped off and collected rather than being permitted to mound around the drill hole aperture. Rotation (blade), roller, and impact (hammer) drilling methods are employed. Drill hole diameters are typically 9 to 11.5 cm. The drill rig is usually mounted on a four-wheel (dual axle) truck and contains an on-board compressor and all the drill rods. Drill rods can be handled by one person, and the drill mast can be tilted to drill angled drill holes. However, almost all RAB drilling at the Project has been oriented vertically. Figure 10-2 presents the locations of the RAB drill holes.

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Table 10-1: Project Drilling by Ownership

Diamond Hammer AC RAB Total Ownership Holes Meters Holes Meters Holes Meters Holes Meters Holes Meters

Pre-2002 (BGMJV) 3,226 588,672 2,680 254,662 20,992 823,195 2,043 69,164 28,941 1,735,694

Post-2002 (Newmont) 1,073 509,197 488 59,950 1,179 53,943 1,035 85,463 3,775 708,553

Total 4,299 1,097,870 3,168 314,612 22,171 877,138 3,078 154,627 32,716 2,444,247 Note: Details on project history and ownership are presented in Section 4.2 and Table 4-1 Table 10-2: Project Drilling Supporting the Mineral Resource Estimate by Year Drilled Diamond Geotechnical Grade Control Hammer Underground Total Year Holes Meters Holes Meters Holes Meters Holes Meters Holes Meters Holes Meters

1982 – – – – – – 352 36,122 – – 352 36,122

1983 6 1,955 – – – – 1 35 – – 7 1,990

1984 17 2,564 – – – – – – – – 17 2,564

1985 1 95 – – – – – – – – 1 95

1986 10 2,204 – – – – – – – – 10 2,204

1987 95 16,438 – – – – – – – – 95 16,438

1988 200 36,780 – – – – – – – – 200 36,780

1989 547 88,945 – – – – 155 9,736 – – 702 98,681

1990 237 39,674 – – – – 71 4,437 – – 308 44,111

1991 220 37,874 – – – – 76 8,640 – – 296 46,515

1992 196 33,118 3 360 – – 232 20,707 – – 431 54,186

1993 195 34,349 5 1,486 – – 242 22,105 7 600 449 58,540

1994 195 45,247 4 529 – – 235 14,870 47 6,149 481 66,795

1995 127 36,165 – – – – 517 53,704 91 14,316 735 104,184

1996 111 32,660 10 3,448 – – 219 30,021 50 7,488 390 73,617

1997 95 32,500 3 269 – – 59 7,613 – – 157 40,382

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Diamond Geotechnical Grade Control Hammer Underground Total Year Holes Meters Holes Meters Holes Meters Holes Meters Holes Meters Holes Meters

1998 118 45,251 1 285 – – 89 10,313 – – 208 55,849

1999 101 26,117 7 1,531 – – 109 3,629 – – 217 31,277

2000 43 18,345 – – – – 69 6,097 – – 112 24,441

2001 3 253 – – – – 83 5,870 – – 86 6,123

2002 22 4,151 – – – – 14 1,920 – – 36 6,071

2003 11 2,013 – – – – – – – – 11 2,013

2004 8 3,415 – – – – 37 4,523 – – 45 7,938

2005 10 3,577 – – – – – – – – 10 3,577

2006 69 55,274 – – – – – – – – 69 55,274

2007 146 121,800 – – 8 256 – – – – 154 122,056

2008 134 102,071 – – 45 2,756 – – – – 179 104,827

2009 73 45,306 – – 54 3,170 – – – – 127 48,476

2010 115 65,045 – – 403 13,658 – – – – 518 78,703

2011 117 44,299 – – 207 13,715 – – – – 324 58,014

2012 74 22,209 – – 30 3,458 – – – – 104 25,667

2013 – – – – 2 400 1 330 – – 3 730

2014 – – – – 36 3,984 – – – – 36 3,984

2015 – – – – 15 3,129 – – – – 15 3,129

2016 14 5,932 – – 161 27,019 11 4,248 – – 186 37,199

2017 6 3,035 – – 73 13,822 64 4,421 – – 143 21,278

2018 11 4,714 – – 1 96 10 435 – – 22 5,245

Total 3,327 1,013,374 33 7,909 1,035 85,463 2,646 249,776 195 28,553 7,236 1,385,074

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Note: Figure aligned to True North Figure 10-1: Vacuum Drill Collar Location Map

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Note: Figure aligned to True North Figure 10-2: RAB Drill Collar Location Map

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10.1.3 AC Drilling AC drilling is a modification to the basic RAB drilling process, in which the cuttings recovery system is similar to RC, with air blown down the outer tubes of the drill stem leading to “sticks” of semi-consolidated cuttings being delivered up the inner drill tube. There is less chance of contamination that when using the RAB method. Rigs can be small when mounted on a light vehicle (<2 tonne), or large (>30 tonne hammer or core drill rig). The drill rig contains an on-board compressor and all the drill rods. Drill rods can be handled by one person, and the drill mast can be tilted to drill angled drill holes. Rotation (blade), roller, and impact (hammer) drilling methods are employed. Figure 10-3 presents the locations of the AC drilling.

10.1.4 RC Drilling RC drilling was used as a Mineral Resource delineation tool from Project inception to the year 2000 and a Mineral Reserves infill tool from 2009 to 2016. In 2010, a short RC drilling program was completed at the North end of the Saddleback Greenstone Belt for exploration purposes. Figure 10-4 presents the location of the RC drilling. RC drilling employs a downhole percussion hammer on the end of a substantial rod string. Rigs are usually large (20 to 40 tonne), mounted on large 6 to 8-wheel (3 to 4 axle) trucks, and require significant support vehicles such as rod/water truck, booster compressor, and personnel vehicle, and typically a 3-man crew to operate. RC rigs are capable of drilling to at least 800 m depth, but, due to hard ground conditions, the maximum depth at the Project for RC is approximately 400 m. Face-sampling RC drilling was used in exploration after mid-1991. There is a very minor amount of older, pre-1991, RC drilling which was conducted using a conventional RC hammer or “cross-over sub” as opposed to the more reliable face-sampling hammer. Older RC drilling is restricted to the upper part of the Wandoo deposit, and, except for the former Hedges area, is generally spatially intermixed with core drilling. At Wandoo South, RC Mineral Resource definition drilling is limited down to 100 m RL of the deposit. At Wandoo North RC Mineral Resource definition drilling is limited down to 150 m RL.

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Note: Figure aligned to True North Figure 10-3: AC Drill Collar Location Plan

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Note: Figure aligned to True North Figure 10-4: RC Drill Collar Location Map

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Infill RC drilling programs have been completed annually at the Project between 2009 and 2018. Drilling programs have varied from small programs of two drill holes, using spare capacity of hydrological drill rigs, to significant programs exceeding 20,000 m. Data from these programs has been incorporated into the resource modelling data set. Annual infill drilling programs have been scheduled until 2026 to increase confidence ahead of mining in both the North Pit and South Pit. 10.1.5 Core Drilling Core drilling was used to support mining operations, pre-feasibility and feasibility Mineral Resource estimates, and to infill in areas of predominantly RC drilling. Drill hole locations are presented in Figure 10-5. From 1997 to 2000, all core drilling was for geotechnical purposes. A major (456,702 m in 774 drill holes) infill core drilling program to increase Mineral Resource confidence that were then used to support Mineral Reserves estimation for the projected mining operation commenced in May 2006. The infill drilling was completed in April 2012. The resource development and exploration program was designed to drill large areas of the deposit at depth with drill holes planned from the pit edges due to the presence of water in the pits. This meant that the drill holes were generally very deep (ranging from 500 m to 1,200 m with an average of approximately 800 m) with deep pre-collars (as deep as 90 m). Core drilling was considered the only viable drilling technique for the program due to historically poor performance from RC hammer drilling at the Project because of bedrock abrasiveness. Core drill holes were primarily drilled at NQ2 hole diameters (50.6 mm). Historically, HQ (63.5 mm) and NQ (47.6 mm) sized drill core have been completed, with the amount of HQ drilling being variable, depending on ground conditions, requirements for wedge holes and if the hole was required for later installation of piezometers. The drill holes were oriented 30º oblique to the mine grid (i.e. 60/240º mine grid) on an orientation known as “z-grid” with dips generally varying from -50º to -75º. This orientation has been adopted since 1994 for most of the drilling in the main resource area as it minimizes drill bias issues that may be encountered with east–west-striking actinolite veins in the Central Diorite area and is optimal for the orientation of mineralization for the bulk of the main Mineral Resource area.

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Note: Figure aligned to True North Figure 10-5: Core Drill Collar Location Map

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10.2 Geotechnical and Hydrogeological Drilling A total of 191 drill holes (13,041 m) have been drilled in various campaigns using core drilling or RC hammer in the search for groundwater and subsequent construction of production water bores, installation of vibrating wire piezometers (VWP), or simple groundwater monitoring piezometers. Additionally, 74 drill holes (18,880 m) have been drilled for geotechnical investigation purposes. Very few of these drill holes were sampled and assayed and were not used for Mineral Resource estimation.

10.3 Geological Logging 10.3.1 RC Drilling Historically the logging system for RC drilling records lithology, alteration, and mineralization as logged from chip trays in one pass. Current practice is to have intervals logged as geological composites of the 1 m samples. The intensity of the alteration assemblage is defined and then the relative abundances of prescribed alteration minerals are assigned. Mineralization is recorded using the same protocol. A provision for comments is made for each interval. Geological codes used in logging were developed in 1995, and adjusted as necessary for additional lithologies that were intercepted in drilling. Drill chips were logged at the drill site, and a chip tray record of each 1 or 2 m interval retained for reference. Chip trays are catalogued and stored at the Project. Geological logs that were completed prior to digital logging are filed at the Project. 10.3.2 Core Drilling Geological logging at the Project has evolved over the life of the deposit and is specifically tailored to capture the alteration and mineralization types found in the Wandoo resource area. Historically, geological logging of core recorded lithology, alteration, mineralization, and structure in separate ‘passes’ into separate logging templates. Currently alteration and mineralization are recorded in the same template. Lithology is currently recorded on geological intervals and is noted by picking up to six, two-letter codes from a drop-down list. The drop-down list allows only valid codes to be chosen. The two-letter codes provide a description of the major alteration, texture, and/or fabric, and are recorded either before or after the rock type, thus allowing for a level of confidence in naming the rock to be determined during interpretation. Fields for recording the phenocryst size, shape, and percentage are present, and the nature and orientation of the upper contact between each rock unit is also logged. Alteration is logged on geological intervals. The mode of occurrence and the intensity of prescribed alteration types found at the deposit are selected from lists of valid codes. Mineralization occurrences are recorded as intensity associated with the alteration type they accompany. Non-prescribed alteration and mineralization notes are recorded in separate fields allocated for this use. Structure and other features are recorded as a point of occurrence and ‘drill hole’ thickness of the feature. Features are picked from a drop-down list of valid codes with a description field also provided. Structural measurements are recorded as alpha and

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beta measurements from measuring half core. Alpha is the angle between the core axis and ellipse described on the core by the measured feature. Beta is the distance in millimeters between the core orientation line (bottom of drill core marked at NBG) and the bottom of the ellipse and noted as a negative measurement. Where the bottom of the ellipse has been sampled, the top of the ellipse is measured from the orientation line and recorded as 75 minus the distance measured. Distances are measured anticlockwise looking downhole. Logging data is captured through tablet data loggers using an acQuire® software data entry object using the same validation (logging codes etc.) as that used in the main drill database.

10.4 Recovery 10.4.1 RC Drilling Recoveries were not routinely measured for RC drilling. However, spot visual checks during 2006 to 2009 of operating rigs indicated RC recoveries were very good. Analysis of more recent (2007 to 2009) RC drilling indicates that recoveries in the range of 70% to 75% are the norm and that most RC samples were dry. 10.4.2 Core Drilling Core recovery is routinely close to 100%.

10.5 Collar Surveys 10.5.1 RC Drilling All RC drill holes have collar surveys. Collar surveys are taken at drill hole completion. Historically, surveys were based on total station measurements, currently, surveys are picked up using Differential Global Positioning System (DGPS) instruments. 10.5.2 Core Drilling Historically all core drill holes have collar surveys, picked up at hole completion using total station survey instruments. All resource development and exploration drill collars from 2006 onwards have been set out using hand-held Garmin® Global Positioning System (GPS) units by the exploration field supervisor. The nature of the current surface drill program is such that the accuracy of the hand-held GPS units is sufficient for drill hole set out. All collar pick-ups are conducted by the Project mine surveyors using differential GPS surveying. The surveyed collar pick-ups are checked against drill hole set-out and hole name convention.

10.6 Downhole Surveys 10.6.1 RC Drilling Not all RC drill holes were downhole surveyed. Hedges RC drill holes did not have downhole surveys, as was the case with NBG RC drill holes prior to 1995. This early RC drilling was predominantly restricted to shallow drill depths. Drill hole deviation in shallow drilling is minimal, and given the large-scale mining proposed, the QP’s opinion is that the lack of survey data should not have any significant impact on the local Mineral Resource estimate. Early RC drill holes were not down hole surveyed during the initial years and these drill holes generally were drilled to shallow depths (less than 120 m). Dip and azimuth at the collar were assigned from the ‘set-up’ of the drill hole using the same technique

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as the core drilling. RC drill holes that have been drilled between 1995 and 2007 were surveyed every 50 m using a single shot Eastman® camera. Infill drill holes drilled since 2007 have used Reflex® single or multishot electronic survey tools with surveys completed at the collar and every 30 m downhole. 10.6.2 Core Drilling Historically, all core drill holes were typically downhole surveyed at 50 m intervals except where hole deviation requirements meant additional surveys for close monitoring during drilling. Surveying was conducted with a single shot Eastman camera. In a minor number of drill holes, surveys that are adjacent to magnetized dolerites are discounted if they indicated any magnetic effect of these intrusions, and the hole average dip is assigned to that interval downhole. Declination and azimuth at the collar of the drill holes were assigned from the drill hole ‘set-up’. The collar set-up was ‘marked out’ by offsetting from the surveyed mine grid with an ‘optical square’ to provide an azimuth. The dip was determined by the driller, either with an inclinometer or pre-set mast positioning. During 2006, downhole surveys were routinely taken at 50 m intervals using single shot Eastman® camera. In 2007, this was adjusted to 30 m intervals for the first part of the hole until a reasonable hole trace was established and then changed to 42 m intervals by the supervising geologist. A Reflex EZ® digital camera was introduced in 2007. The accuracy of the Eastman® cameras was trialed in a test hole prior to the commencement of each new drill hole. Gyroscopic downhole surveys were also performed on test holes during 2006 to validate the Eastman® recordings and to determine if there was any advantage to drill hole control. This was not adopted as standard practice due to unresolved issues with the gyroscopic methodology. In 2011, selected core drill holes were surveyed for QA/QC purposes using Scientific Drilling’s Keeper GYRO. Results indicate that current downhole surveying method is adequate. Quality control GYRO surveys will continue to maintain higher survey standards. 10.6.3 Declinations Declination corrections have been applied to the downhole survey data as required. The same declination correction factors were used for core as for RC drilling.

10.7 Drilling Used in Mineral Resource Estimation Drill spacing varies by domain within the Wandoo deposits. Typically, drill spacing in the areas that have been classified as Measured Resource is approximately 25 m x 25 m, widening to a 50 m x 50 m spacing for Indicated Resource, and out to approximately 100 m x 100 m in areas classified as Inferred Resource. Table 10-3 presents the drill hole details for the drilling used to support Mineral Resource estimates, and the collar locations for the drill holes are presented in Figure 10-6.

10.8 Sample Length/True Thickness Drilling is normally perpendicular to the strike of the mineralization, but depending on the dip of the drill hole, and the dip of the mineralization, drill intercept widths are typically greater than true widths. Figure 10-7 is a drill section through the Wandoo deposit, showing orientations of the drilling in relation to the mineralization.

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Table 10-3: Drill Holes Used in Mineral Resource Estimation

Drill Type Drill Holes Meters

Diamond Core (Surface) 3,360 1,021,283 Diamond Core (UG) 195 28,553 RC (Surface) 3,681 335,239 Total 7,236 1,385,074

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Note: Figure aligned to True North. Although Project tenure outlines are not shown, sampling has taken place within the confines of the mineral tenure. Current tenure outlines are included in Figure 4-1 Figure 10-6: Collar Locations of Drilling Supporting Mineral Resource Estimates

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Figure 10-7: Geological Cross-Section at 10600N (looking towards 290°)

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10.9 Comments on Drilling In the opinion of the QP, the quantity and quality of the lithological, geotechnical, collar, and downhole survey data collected in the exploration, delineation, and grade control drill programs to support feasibility-level studies and during the previous mining operations at the Project, are sufficient to support Mineral Resource and Mineral Reserves estimation for the following reasons:  Drill hole orientations are appropriate to the orientation of the mineralization;  Drilling is normally perpendicular to the strike of the mineralization, but depending on the dip of the drill hole, and the dip of the mineralization, drill intercept widths are typically greater than true widths;  Drill hole intercepts adequately reflect the wide, low-grade nature of the gold and copper mineralization;  Geological logging meets industry standards for both gold and copper exploration, and for run-of-mine production;  Geotechnical logging meets industry standards for operating open pit mines;  Collar surveys have been performed using industry-standard instrumentation;  Downhole surveys have been performed using industry-standard instrumentation and accurately represent the trajectories of the drill holes;  Drilling has been completed using industry-standard methods yielding a high level of sample recovery.

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11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY Newmont and previous operators’ staff throughout the duration of the Project have been involved with, or responsible for the following:  Sample collection;  Core splitting;  Delivery of samples to the analytical laboratory;  Sample storage;  Sample security. All analytical procedures that support Mineral Resource estimation, including sample preparation and analysis, were performed by independent analytical laboratories.

11.1 Sampling Methods

11.1.1 Geochemical Sampling Soil samples were typically collected from 15 cm to 20 cm depths in the soil profile using a spade. Approximately 2 kg of sample was placed in a pre-numbered calico bag, and the sample location recorded as a grid co-ordinate. BLEG samples were collected from suitable drainages, typically as 2 to 5 kg samples, and placed in pre-numbered calico bags. The sample location was recorded, typically on aerial photographs.

11.1.2 Trench Sampling Trenching machine spoil samples were taken from the side of the trench on 2 m spacings as designated by the supervising geologist.

11.1.3 Vacuum Sampling The drilled sample deposits into a flask, which is part of the sealed vacuum circuit. The sample at intervals from 0.5 m to 1.0 m, depending on area, is then tipped into a vertical riffle splitter and split twice to give a 150 g sample. From approximately the year 2000, the drilled sample has been tipped onto an inverted cone splitter. Reject material spills onto a rubber mat on the ground and is later tipped back into the vacuum hole.

11.1.4 Rotary Air Blast Sampling Samples are blown up through a collar pipe and into a cyclone. The whole sample is collected and poured through a riffle splitter. Samples are collected at intervals ranging from 1 m to 2 m to provide 1 to 2 kg of sample. Reject material is used to backfill the drill hole and collar area.

11.1.5 AC Sampling Drilled samples travel up through the center rod and then through a flexible ‘bull’ hose to the top of a cyclone, beneath which can be mounted a riffle splitter, vertical rotary splitter, or a cone splitter. Samples are usually 2 to 5 kg in size and collected on 1 to

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2 m intervals. Sample reject can be placed in large plastic bags for later logging and resampling.

11.1.6 Reverse Circulation Sampling RC drilling was predominantly conducted using 14 cm face sampling, down the hole hammers with air capacity that averaged 1,100 cubic feet per minute (cfm) at 600 pounds per square inch (psi) for drill holes as deep as 200 m. Prior to 2008, samples were collected in plastic bags on 1 m intervals via a cyclone before sample reduction utilizing a riffle splitter. The samples were composited to 2 m intervals of 4 kg to 6 kg. Sample bags were pre-numbered, and a matching ticket book tag was added to each sample during the sampling process. The drill hole number and depth downhole was written on the ticket book ‘stub’ by the drilling contractor during sampling. Bagged samples were collected in the field by the Project field assistants who recorded the sample numbers used, drill hole depths of each sample and the drill hole number prior to dispatch. For RC drilling programs completed since 2008, samples were collected on 2 m intervals in a drop box and split using cone splitters on the RC rigs. Prenumbered bags, with barcodes attached, were provided to the drillers and the sample number was recorded on a run sheet by the drillers for each interval. The run sheet includes the location of duplicates, standards and blanks to ensure protocols are observed. Bagged samples were collected in the field by the Project Pit Technicians who validate the sample bag sequence, including empty bags for standard and blank insertion, prior to dispatch.

11.1.7 Core Sampling All core drill holes were sampled and analyzed. During drilling, orientation marks were made at 6 m intervals. Drill core was generally recovered in 6 m runs for NQ diameter drill core and 3 m runs for HQ diameter drill core. The drill core was placed into core trays and core blocks marking the depth of each run were inserted. Processing began with the orientation of the drill core and marking of the core into 1 m intervals using a purpose-built rack in the core shed. Drill core was sawn in half for sampling using either a manual core saw or an automatic core saw. The core was cut such that the orientation line was preserved. Typically, NQ core was sampled on 2 m intervals and HQ core sampled on 1 m intervals. Geology-based intervals are used for specific areas such as sampling edges of dolerites and large quartz or actinolite veins. A 2 m NQ sample weighs approximately 4.8 kg and a 1 m HQ sample weighs approximately 4.3 kg. The left-hand side of the core (determined by holding the core such that the orientation line is bottom most while looking down the drill hole) was sampled with the right-hand side, marked with the orientation line, preserved. Proterozoic dolerites of greater than 3 m downhole thickness were not sampled. Where a dolerite of greater than 3 m thickness is encountered in the drill hole, sampling is conducted such that the sampling will stop at the contacts of the dolerites and a one- meter sample, measured from the dolerite contact, is taken from the dolerite margin.

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This ensures that the usually barren Proterozoic dolerite is not mixed with potentially mineralized Archaean rocks. A continuous record of sample intervals matching sample numbers was kept by using ticket books. Each sample interval is written on the ticket stub and its corresponding ticket placed in the sample bag prior to dispatch. For the 2006 to 2013 core drilling program, the core was sampled with the same protocols as the historical core; the difference being the addition of digital photography recordings for all core trays, routine rock quality description (RQD), and specific gravity (SG) measurements i.e. all core was orientated, NQ core was cut and sampled on 2 m intervals and HQ core was cut and sampled on 1 m intervals.

11.2 Density Determinations

11.2.1 Historical Determinations Historically, the Project’s basement in-situ bulk density determinations were collected on a 50 m x 50 m grid across each geological domain. Analabs, an independent analytical laboratory, undertook in-situ bulk density determinations using the immersion method. It is not known if Analabs were certified at the time of analytical determination. The in-situ bulk density was calculated according to the following formula: 푚 휌푑 = 푠 푚푠− 푚푠 푖푛 푤푎푡푒푟 Where:

휌푑 = dry bulk density;

푚푠 = dry mass of sample;

푚푠 푖푛 푤푎푡푒푟 = mass of sample in water. The in-situ bulk densities used were dry bulk densities. Downhole geophysical logging on four drill holes in 1995 and eight drill holes in 2000 confirmed the densities determined using the immersion method. The in-situ bulk density results for the bedrock were generally very consistent. The data were originally analyzed by domain with the mean, median, and mode of the data generally within 0.01 grams per cubic centimeter (g/cm3) plus in each domain between 90% and 95% of the data lie within ± 2.5%. The density values assigned to the South Pit area are presented in Table 11-1. Table 11-2 presents the assigned densities for the North Pit area.

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Table 11-1: South Pit Density Values Assigned In-situ Bulk Gold Estimation Density Gold Estimation Domains with Measured In-situ Bulk Density Domains with Assigned (g/cm3) In-situ Bulk Density Central Diorite, South Volcanics, North Volcanics, NW Volcanics, Hanging-wall, Eastern Blackbutt Upper, Blackbutt Lower, Blackbutt H/G, Pipeline FW Upper, 2.75 Volcanics South, North Pipeline FW Lower, Pipeline H/G1 (excluding Green and White Rock), Volcanics B Central Volcanics Pipeline 2.78 Western Shear Zone 2.79 Southern Diorite Deeps 2.90 Green and White Rock alteration in Pipeline FW Upper 3.00 Proterozoic Dolerites

Table 11-2: North Pit Density Values Assigned In-situ Bulk Gold Estimation Domain Density Gold Estimation Domain with Measured In-situ Bulk Density with Assigned In-situ (g/cm3) Bulk Density Eastern Volcanics Nth, Far Nth, Northern Diorite, South Zone, E2 H/G, Son of Blob, A Breccia, Northern Volcanics A, 2.75 Blob, F-Zone, Eastern Diorite Barren Actinolite, and Jarrah domains

The South Pit in-situ bulk density data show that there is little variation in the in-situ bulk density data between volcanic rocks, diorites and alteration zones (Gleeson et al, 1999). Where variability does occur, it is due to increased and consistent sulfides in Southern Diorite Deeps and intense actinolite alteration found in the Pipeline area. Analysis of the in-situ bulk density for North Pit indicated that the values were very consistent across all domains. As a consequence, an in-situ bulk density of 2.75 g/cm3 was assigned to all domains in the North Pit area. There is greater variability of the in-situ bulk density determinations in the Green and White Rock area of the Pipeline domains, where variable actinolite alteration (in-situ bulk density of 3.00 g/cm3) was responsible for more variable and elevated in-situ bulk density values. The Green and White Rock alteration zone was assigned the mean in- situ bulk density within the zone of 2.90 g/cm3, which is potentially conservative where actinolite alteration is intense. The Green and White Rock alteration constitutes <1% of the Mineral Resource tonnage and grade. Domains that predominantly contained materials that were to be mined as waste were assigned the default in-situ bulk density of 2.75 g/cm3 used for the intrusive and volcanic domains. This was considered reasonable given the overall consistency of the in-situ bulk density in the bedrock where there is no intense actinolite or sulfide alteration. These domains are either predominantly waste material or in the case of the Jarrah domains, already mined out.

11.2.2 Recent Determinations A total of 4,285 bulk density samples were collected and analyzed from resource definition and exploration drilling programs during 2006 to 2011. Samples were collected from 50 m downhole intervals and sent to Genalysis for density determinations. Genalysis is ISO 9000 certified. Starting from 2009 density determinations were completed with an in-house set-up using the immersion method

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(refer to Section 11.2.1) and the data was recorded as point data in the drill hole database. Selected samples were also analyzed at Genalysis for quality control. Analysis of bulk density data by domains indicated a slight variation of in-situ bulk density among domains and consistently higher than historically determined values. Mean, mode and median values are variable between 2.73 g/cm3 to 2.83 g/cm3 except for Pipeline domains which have high mean and median values and very high variability. It also appears that Western Shear, East Volcanic, D-pits and Barren Actinolite domains have slightly elevated in-situ bulk density values. The granitic lithology has the lowest in-situ bulk density. In-situ bulk densities of diorite and andesite lithologies are not significantly different (less than 1% difference), but microdiorite has slightly higher in-situ bulk density values and Proterozoic dolerite has the highest density. Density data for North Pit and South Pit were also analyzed and results indicated that there is no significant difference between the two pits, although the number of samples used for North Pit was 45% less.

11.3 Analytical and Test Laboratories Historically, sample analysis has been performed by several independent laboratories, including Classic Comlabs, Genalysis, Amdel, Bureau Veritas Kalassay, AAL in Perth, and AAL in Boddington. It is not known if the laboratories were certified at the time. The Boddington mine site laboratory, operational between 1985 and 2001, was owned by AAL from 1985 to December 1995. From December 1995 to 2001, the laboratory was owned and operated by Analabs. In 1995, the mine laboratory became ISO 9002 accredited. approximately 80% of the pre-2001 analytical data was completed by the mine laboratory. There was no drilling in the period 2001 to the end of 2005. From 2006 onwards, routine analysis of samples collected during core drilling programs was undertaken at Genalysis. In 2006, Genalysis was accredited to ISO/IEC 17025, version 2005. UltraTrace Geoanalytical Laboratories (UltraTrace), Perth, acts as the umpire laboratory. During periods where sample turnaround time was paramount in 2006 to 2007, UltraTrace also acted as primary laboratory. UltraTrace was part of the Amdel Laboratory group and is accredited to ISO/IEC 17025. UltraTrace and Kalassay are now merged with Bureau Veritas international laboratory group. The Kalassay laboratory used for grade control analysis from 2008 until January 2014 achieved ISO/IEC 17025 accreditation during 2010. Since February 2015 Intertek Genalysis has conducted the blasthole assaying. A summary of laboratory usage timeframes is presented in Table 11-3.

11.4 Sample Preparation and Analysis

11.4.1 Sample Preparation Mine Laboratory Prior to 2000, the sample preparation protocol involved coarse crushing of a 5 kg field sample, and then sub-splitting a 1 kg sample for pulverization to 75 µm. From 2000, the preparation protocol was changed to a larger field sample mass of 12 kg, sub- splitting 1.5 kg for pulverization to 150 µm. Table 11-4 presents the sample preparation for drill core.

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Table 11-3: Analytical Laboratory Dates Year Laboratory 1983 to 1985 Classic Comlabs 1986 to 1987 AAL Boddington 1988 Classic Comlabs 1989 to 1990 Genalysis, AAL Boddington 1991 to February 1992 AAL Boddington, AAL Balcatta February 1992 to January 1993 Analabs Welshpool January 1993 to December 1993 AAL Balcatta, AAL Boddington December 1993 to 2001 AAL/Analabs Boddington 2008 to January 2015 (blasthole) Kalassay 2006 to Current Genalysis 2006 to Current UltraTrace

Table 11-4: Sampling Protocol for Drill Core Mass Size RSD for RSD for RSD for (kg) (cm) 2 g/t Au 3.5 g/t Au 10 g/t Au 5.5 kg primary field sample crushed to P95 3 mm 1.5 10% 13% 18% 1.5 kg (1/4) pulverized to P95 150 µm 0.3 9% 12% 17% 50 g at 150 µm 0.015 6% 8% 11% Overall 15% 19% 27%

Note: RSD = Relative Standard Deviation Genalysis Genalysis uses conventional jaw crushing of samples to nominal minus 10 mm. Subsequently, the sample is fine-pulverized. Genalysis aims to achieve 95% minus 90 µm. UltraTrace UltraTrace also utilizes conventional jaw crushing of samples. Fine pulverization follows to a nominal 95% passing 90 µm. Bureau Veritas Kalassay Bureau Veritas Kalassay crushes samples to a nominal 95% passing 3 mm using a Boyd Crusher and then pulverizes to a nominal 95% passing 90 µm. 11.4.2 Analyses Geochemical Samples Geochemical samples were primarily analyzed using bulk-leach extractable gold methods. Trench and Pit Sampling Trench samples were primarily analyzed for gold by fire assay with AAS of dissolved prill, to quantify mineralization in the oxide zones. Vacuum and Rotary Air Blast Sampling Analysis of gold was by fire assay with atomic absorption spectrophotometry (AAS) of dissolved prill. Copper analysis was by either single acid digestion or three-acid

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digestion followed by AAS of solution. Single acid digestion provides 2 parts per million (ppm) detection limit and a multi-acid digest results in a 1 ppm detection limit result. Reverse Circulation and Core Sampling Analysis of gold was by fire assay with either AAS or inductively-coupled plasma atomic emission spectroscopy (ICP-AES) (also known as inductively-coupled plasma optical emission spectroscopy (ICP-OES)) of dissolved prill. Prior to 2006, copper analysis was by either single acid digestion or three-acid digestion followed by AAS of solution. Single acid digestion provides a 2 ppm detection limit and a multi-acid digest results in a 1 ppm detection limit result. Multi-element determination was not routinely performed but rather performed on selected drill holes as part of detailed geological investigations. When used, the multi- element analytical suite requested typically consisted of Ag, As, Bi, Ce, Mo, Ni, Pb, Sb, Ti, W, Y, Zn, and Zr. Multi-element analysis was performed by Amdel in 2004, using ICP-MS and ICP-AES after triple-acid digestion. A second multi-element program was undertaken in 2007 by UltraTrace. From 2006 onwards, the typical analytical suite requested for core comprised:  Fire assay gold with AAS finish on 50 g charge;  Multi-element suite using four-acid digest with ICPOES-ICPMS finish for Cu, S, As, Bi, Mo, Sb, Cd and Ni. The multi-element suite utilizes an ICP-MS or ICP-OES finish to achieve the acceptable lower detection levels for the elements critical to predicting concentrate quality i.e. As and Bi. Elemental detection limits and methods are presented in Table 11-5. For RC drilling post 2006 the typical analytical suite requested for core comprised:  Fire assay gold with ICP finish on 50 g charge;  Multi-element suite using four-acid digest with ICPOES-ICPMS finish for Cu, S, As, Bi, Mo, and Sb. Since 2013 Mo and Sb analysis has not been conducted.  Between 2009 and 2014 fire assay gold at Bureau Veritas Kalassay was completed with an AAS finish. Assay data are sent electronically to the database manager, and incorporated, following checks, into the master database. To decrease contamination in the sample preparation, both Genalysis, Bureau Veritas Kalassay, and UltraTrace have introduced robotic milling systems. All samples, once crushed down to 3 mm, are processed through Mixermill pulverizers which are completely robotic. To improve sample tracking, both laboratories have also introduced bar coding. Once the samples are received at the laboratory, bar codes are attached, and this number is then used through all steps of the milling and assaying process, minimizing the need for data entry. 11.4.3 Additional Analytical Determinations Minor additional analytical work has been completed and includes whole rock analysis for geochemical rock type identification completed in 1999 to 2000, and as part of the Roth (2002) PhD thesis.

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A full multi-element suite analyzed by neutron activation analysis during a fire assay data check in 1996. The multi-element suite comprised Dy, Eu, In, Lu, Mn, Au, Ho, Ir, Re, Sm, W, Ag, Ar, As, Br, Cl, Co, Cs, Cu, Er, Ga, Hf, I, La, Sb, Sc, Se, Ta, Tb, Th, Tm, U, V, Yb, Al, Ba, Cd, Ce, Cr, Hg, Kr, Gd, Ge, Mo, Na, Nd, Ni, Os, Pd, Rb, Rh, Ru, Sr, Te, Zn, Zr, Bi, Ca, K, Mg, P, Pt, Si, Sn, Ti, Tl, Xe, Y, F, Fe, Nb, Ne, Pb, and S. Table 11-5: Lower Detection Limits for Analytical Suite Element Lower Detection Limit Method Au 0.005 ppm Fire assay on 50 g charge Cu 1 ppm 0.2 g with 4 acid digestion and ICP-OES S 10 ppm 0.2 g with 4 acid digestion and ICP-OES As 1 ppm 0.2 g with 4 acid digestion and ICP-MS Bi 0.05 ppm 0.2 g with 4 acid digestion and ICP-MS Mo 0.1 ppm 0.2 g with 4 acid digestion and ICP-MS Sb 0.1 ppm 0.2 g with 4 acid digestion and ICP-MS

11.5 QA/QC 11.5.1 Historical QA/QC Formal systems of continuously monitoring assay quality control by Worsley have been in place since January 1989 and a continuous record is available for this time. Systems include review of laboratory performance using methods commissioned by Worsley as well as review of the laboratory’s internal systems. The principal monitoring system prior to June 1995 included internal round robins conducted by NBG, which allowed comparison of the Boddington Laboratory against other laboratories and standards. Post June 1995 an updated internal laboratory monitoring system (‘C’ Class) was introduced which allowed electronic capture of quality control data and statistical analysis of results. Formal review of this data on a month by month basis by the Project Exploration and Geological Superintendent and Analabs Management was adopted as the principle method of quality control. Wet screening of sample pulps to ensure 90% passing through an 80 µm mesh screen was recorded as the routine procedure from April 1994 until 2001. Every duplicate sample was screened. Prior to April 1994, no records of routine monitoring of sample preparation are available. Progressive sample preparation requirements, including the separation of mine and exploration sample preparation areas and the use of use of bowl-type grinding equipment (LM5 or similar) led to use of laboratories other than the Boddington Laboratory for most of routine bedrock sample analysis between 1991 and mid-1994, when a separate exploration sample preparation area was established using LM5 grinding equipment. In 1998, a quality control training manual was written; all personnel involved in quality control were required to undertake a test to ensure requirements were understood and met. Prior to 1993, one standard was run routinely with each batch of samples, blanks were run on an intermittent basis, and duplicates chosen on the basis that anomalous results were checked. From 1993 until March 1998, one standard, one blank, and six duplicate samples were present in each fire assay batch of fifty samples. From March 1998, each fire assay batch of fifty included two standards, one blank, two duplicates, and two replicates.

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Standards used by AAL/Analabs Boddington were from two sources, NBG oxide material and commercial standards from Gannet Holdings. Ten different commercial standards were used between April 2006 and August 2007, sourced from either Geostats Pty Ltd or Rocklabs Ltd. Five of these standards are certified for both gold and base metals (As, Bi, Co, Cu, Mo, and Sb). Five of the standards are certified for both gold and sulfur. Certificates for these standards are kept on site at the mine. 11.5.2 Current QA/QC (2006 to date) Analyses were performed at both the Genalysis and UltraTrace laboratories in the current QA/QC program. In the period September 2010 to June 2013 analyses were performed at only the Genalysis laboratory. Check samples were submitted to the UltraTrace laboratory. Standards and blanks are inserted into the batches of samples before dispatch to the laboratory and are targeted, as per Newmont corporate requirements, to be inserted at a rate of at least 5%. For the period August 2011 to July 2012 the standard insertion rate was 11.1%. Sample numbers for the QA/QC programs are pre-calculated using a Microsoft Excel® random number generator before the sampling process begins. Four standards have been created from Boddington material by ORE Research and are certified for both gold and base metals. Six standards are certified for gold, two are certified for gold, copper and sulfur, four are certified for sulfur; and three are certified for base metals. The base metals analyzed are As, Bi, Co, Cu, Mo, and Sb. The standards are 100 g pulps for gold and 30 g pulps for base metals and/or sulfur. Duplicate samples are not normally included in the sample batch. Typically, NBG reanalyzes pulp rejects, rather than submitting field duplicates.

11.6 Databases All drilling-related data are stored in a Microsoft SQL server database which supports multi- user access, using the acQuire® software interface. The database is administered by a dedicated database manager. Survey, geological, topographical and assay data are uploaded in digital format to the database and have been supplied in digital format since the early 1990s. Historical drill data were received in hard-copy format, and manually data-entered. All data are subject to verification checks prior to upload.

11.7 Sample Security Sample security at the Project has not historically been monitored. Sample collection from drill point to laboratory relies upon the fact that samples are either always attended to, or are stored in the locked on-site preparation facility, or are stored in a secure area prior to shipment to the external laboratory. Chain-of-custody procedures consist of filling out sample submittal forms to be sent to the laboratory with sample shipments to ensure that all samples are received by the laboratory.

11.8 Sample Storage RC drill chips are stored in catalogued chip trays on the mine site. A base-of-hole reference sample is taken from all AC drill holes and stored in trays with the RC samples. Drill core is retained and either palletized in a storage lay-down area, or in vertical racks in a shed with a

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covered roof. All core is catalogued. Approximately 30 km of core has been exhausted in metallurgical testwork; these core intervals are marked in the database. Post assaying, all pulps and rejects are retained, and stored in inventoried Kraft bags in designated storage at the Marradong storage facility or within the mine site.

11.9 Comments on Sample Analysis, QA/QC and Security In the opinion of the QP:  The quality of the analytical data is reliable, and that sample preparation, analysis, and security are generally performed in accordance with exploration best practices and industry standards;  Data are collected following mine site-approved sampling protocols;  The sampling methods are acceptable, meet industry-standard practice, and are adequate for Mineral Resource and Mineral Reserves estimation and mine planning purposes;  Sample intervals, broken at lithological and mineralization changes in the core, are typical of sample intervals used for gold and copper mineralization in the industry and is considered adequately representative of the true thicknesses of mineralization;  The density determination procedure is consistent with industry-standard procedures;  Sample preparation for core samples has followed a similar procedure since the early 1990s. The preparation procedure is in line with industry-standard methods;  Newmont has used a QA/QC program comprising blank, standard and duplicate samples. Newmont’s QA/QC submission rate meets industry-accepted standards of insertion rates;  Data that were collected prior to the introduction of digital logging have been subject to validation, using inbuilt program triggers that automatically checked data on upload to the acQuire® software interface;  Verification is performed on all digitally-collected data on upload to the main database, and includes checks on surveys, collar co-ordinates, lithology, and assay data. The checks are appropriate, and consistent with industry standards;  Sample security has relied upon the fact that the samples were always attended or locked in the on-site sample preparation facility. Chain-of-custody procedures consist of filling out sample submittal forms that are sent to the laboratory with sample shipments to make certain that all samples are received by the laboratory;  Current sample storage procedures and storage areas are consistent with industry accepted practices.

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12.0 DATA VERIFICATION

12.1 Site Visits The QP has conducted personal inspections of the Project as part of his data verification (refer to Section 2.3). Mr. Doe has visited the Project numerous times during his career at Newmont, most recently on 18 April 2018. During site visits to the Project, Mr. Doe inspects the operating open pits, and views the process plant and associated general site infrastructure, including the current TSF operations. While on site, he discusses aspects of the operation with site-based staff and assesses the knowledge and abilities of the site staff to carry out their duties as required. These site discussions include the overall approach to the mine plan, anticipated mining conditions, selection of the production target and potential options for improvement. Other areas of discussion include plant operation and recovery forecasts, capital and operating forecasts. Mr. Doe receives and reviews monthly reconciliation reports from the mine. These reports include the industry standard reconciliation factors for tonnage, grade and metal; F1 (Mineral Reserve model compared to ore control model), F2 (mine delivered compared to mill received) and F3 (F1 x F2) along with other measures such as compliance of actual production to mine plan and polygon mining accuracy. The reconciliation factors are recorded monthly and reported in a quarterly control document. Through the review of these reconciliation factors, the QP can ascertain the quality and accuracy of the data and its suitability for use in the assumptions underlying the Mineral Resource and Mineral Reserves estimates. Mr. Doe also reviews Newmont’s processes and internal controls at the mine site with operational staff on the work flow for determining Mineral Resource and Mineral Reserves estimates, mineral process performance, mining costs, and waste management.

12.2 Laboratory Visits NBG staff regularly visit the laboratories to inspect sample preparation and analytical procedures. Observed actions and procedures that are not in conformity with the NBG procedures are recorded in Project files and communicated to laboratory for corrective action to be taken.

12.3 Analytical Reviews 12.3.1 Historical Round Robins A history of quality control monitoring of basement gold assays to support the database for the 2000 feasibility study was compiled during 2000 (Tangney, 2000). Pre-1989 work, representing 7.5% of basement resource data at the time of the 2000 study, had no recorded monitoring of quality; however, it was expected that communications between the laboratory and the BGMJV were taking place as there are internal memorandums stating poor laboratory performance in late 1988. Recorded quality control monitoring between 1989 and 1990 consisted of laboratory interaction and monitoring and external monitoring by round robins. The standards used in these round robins were considered by Tangney to be probably unreliable. Recorded monitoring between 1991 and September 1995 was by laboratory internal

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monitoring and meetings with BGMJV staff. The BGMJV also conducted its own round robins. These round robins showed that the primary laboratory produced poor results in January 1993, January 1994, May 1995 and July 1995, and a period of inconsistent results between June 1991 and May 1992. Monitoring of laboratory performance from June 1995 to 2000 was by monitoring of laboratory results captured by the ‘C’ Class laboratory monitoring system combined with round robins conducted by Geostats Pty Ltd. This monitoring identified periods of positive laboratory bias from June 1995 to October 1995 and from October 1999 to January 2000. Monitoring of blanks indicated that the laboratory showed general improvements with the advent of ‘C’ Class. Recorded monitoring by the Boddington Mine Laboratory of sample preparation by screen size analysis of pulps, conducted since April 1994, shows:  Greater than 70% of screens passing a test of 90% passing 80 µm;  Greater than 90% passing a test of 80% passing 80 µm. Since 1996, most Hedges basement gold assaying has been provided by the Boddington Laboratory. Laboratory protocols have been the same as those used by the BGMJV, but there are no available records of quality control procedures undertaken by Hedges. An exercise of reanalysis using neutron activation analysis was undertaken in September 1996 on a 100 m x 100 m x 100 m grid (x, y, z) over the basement mineralization then held by the BGMJV. This exercise provided a spot check on general levels of gold assay bias between December 1986 and May 1996. The NAA results in totality indicated the likelihood of a slight negative bias in gold assay results; however, periods of negative and positive bias were also shown. Extended periods where positive bias is shown to be likely are mid-1989, early and late in 1993, and mid to late 1995. These data are discussed in the following sub-sections. 12.3.2 Round Robins, 1988 to 1990 Quality control concerns raised in late 1988 regarding the AAL-run Boddington Mine laboratory led to the introduction of external quality control monitoring by Worsley. At the end of the round robin programs, the BGMJV reviewed the results of each program. The review consisted of graphical display of correlation between each laboratory, comment on standard results, and investigation of any anomalous results. No bias adjustments were made to the data. It was concluded that the standards used before December 1989 had been insufficiently typified and were probably unreliable for the estimation of bias. The 0.50 g/t Au and 1.00 g/t Au standards used during 1990 were also considered unreliable, for the same reason. Drill data that were investigated by the 1988 to 1990 round robin programs have since been mined out. 12.3.3 Round Robins, 1991 to 1995 The round robin procedure as used in 1989 and 1990 was modified during the period 1991 to 1995. Instead of submitting pulps to each laboratory (samples with no previous analysis), pulps from samples from recent drill holes were retrieved from the laboratory doing routine analysis. Internal round robins were conducted on a bi-monthly basis. The number of samples for reanalysis was increased to 60 samples with eight standards. Core pulps were selected on the basis that a range of grades, from high to low grade gold, was available for analysis. Generally, pulps were selected from holes

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drilled in the previous month; however, pulps up to a year old were reanalyzed in some cases. Standards used for this procedure were either Gannet Standards or were prepared from basement material. Comparing the round robin results with original assays and assays returned by the routine laboratory produced results within an acceptable range for approximately 97– 98% of the samples. Any errors that occurred were due to accidental sample swaps or mislabeling of samples. Procedures to identify errors, and subsequently limit or rectify the effects of the errors, were put in place. No modifications for analytical bias were made to the data. 12.3.4 Round Robins, 1995 to June 2000 From the use of C-class data capture by the Boddington Mine Laboratory, an assessment of the performance of standards, blanks and assay repeatability was undertaken monthly. Standards and blanks were used to measure the bias performance of the laboratory. 12.3.5 Geostats Pty Ltd Round Robins, 1993 to 2000 The Boddington Mine Laboratory was a subscriber to the Geostats round robin assessments between 1993 and 2000. The round robins indicated that the Boddington Mine laboratory had an acceptable performance. 12.3.6 Analytical Repeatability Reviews A study conducted by BGMJV geologists in 1996 involved selecting a 100 m x 100 m grid of core and hammer drilling over the 1996 Wandoo South area, filtering the samples that had duplicate gold assays and plotting log graphs of original assays versus duplicate assay results. Hammer holes were shown to have excellent repeatability with only a few samples showing any significant deviation from the regression line. Core drill holes showed very good correlation; however, 5 to 7% of the samples showed poor repeatability. In 1999, Snowden Mining Industry Consultants Pty Ltd (Snowden) completed a review of pairs of replicate data for the years 1998 and 1999, as part of an audit of data and procedures used for the preparation of the 1999 Wandoo bedrock Mineral Resource estimates. A total of 3.7% of the relevant data in 1998 were found to have precision outside a ±10% interval. In the 1999 data, 14.9% of the relevant data were found to have a precision outside the ±10% limit. The 1999 number was explained as being related to drilling in areas of relatively coarse gold, with extra repeats done in areas of poor repeatability.

12.4 Data Reviews 12.4.1 External Reviews External database, sampling method and Mineral Resource estimation reviews were undertaken by the following firms and individuals:  Stoker, 2000, 2001, 2002. Independent audit of the Wandoo assay data, subsequent laboratory verification program, review of variation in assay bias between different grade ranges as well as variation between the two verification laboratories. The assay data used to support Mineral Resource estimation appeared reasonable, original gold assays are overall approximately 2% lower than that indicated by the verification laboratories, confirming a slight negative

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bias. Analysis of the copper verification data indicated that the original copper assays were approximately 3% higher than that indicated by the original laboratories. Data were suitable for use in assessing the sensitivity of the geological models to variations in gold and copper grades, but due to the variations noted were not sufficiently robust to support modifications to the original assay data.  Golder Associates Pty Ltd (Golder), 2002. Review of historical assay bias data and resource sensitivity study into the assay bias, incorporating a re-estimation of the Mineral Resource using an assay bias-factored database and pit optimization study.  Quantitative Geoscience, 2003. Audit of the local recoverable resource estimates of gold and copper;  Golder, 2004. Quantitative kriging neighborhood analysis for optimization of the search parameters in the model. This resource model was subsequently used for preliminary mine planning during the 2003 feasibility study prior to the completion of more detailed geological and resource models.  Francois-Bongarcon, 2005. Review of sampling protocols and heterogeneity issues. This included review of base assumptions, gold grain size data and validation using blast hole samples. A new sampling protocol was developed to minimize sampling error reflecting a coarser gold grain distribution than previously, leading to a larger initial field sample mass and change in final grind size (up from 75 µm to 150 µm). The protocol was applied to both drill core and blast hole sampling;  Francois-Bongarcon 2007. The rolling average laboratory mean was monitored to identify sample swap errors (by laboratory or NBG) separately from laboratory bias issues that fluctuated over time in the laboratory. Francois-Bongarcon recommended that the NBG prepare its own matrix-matched standards for future QA/QC programs, as standard deviations within the commercially-purchased standards were larger than ideal for controlling laboratory performance;  CS-2, 2007 and 2008. Review of the methodology to be applied for Mineral Resource estimation of Au at the NBG (Ravenscroft, 2007). The result of the audit was a recommendation to move from the applied method of MIK to the method of uniform conditioning (UC). CS-2 also conducted an audit of the 2007 local recoverable Mineral Resource estimates of gold and copper during 2008 (Masters, 2008). The aim of the audit was to validate the 2007 Mineral Resource and suggest improvements in application;  GeoSystems International (Mario Rossi), 2009. Eternal audit of both the 2009 Mineral Resource and simulation work focusing mainly on Au, to validate the models and identify causes for differences between resource and simulation models. The main recommendation was to reduce the number of samples used in estimation as well as the size of searches.  AMEC, 2010. Review of the 2010 mid-year block model and associated Mineral Resource estimates that formed the basis for business planning, strategic planning and Mineral Resource and Mineral Reserves reporting.  Golder, 2017 and 2018. Review of the 2016 Mineral Resource estimate and 2016 Mineral Reserves estimate and associated mine planning, geotechnical and metallurgical aspects of the Project. Further reviews were conducted in 2017 of

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the 2017 Mineral Resource estimate and 2017 preliminary Mineral Reserves estimate. 12.4.2 Internal Reviews Internal reviews were undertaken by either the JV partners, or BGMJV or NBG staff as follows:  AngloGold Ashanti, 2001 to 2002. A database review as part of a review of the Mineral Resources that supported the 2000 feasibility study. AngloGold Ashanti used a trial mining area study to evaluate the potential for upside (increased tonnes and grade) in the 2000 feasibility study Mineral Resource estimate. The work focused on four trial mining areas located in Wandoo South, Pipeline, Wandoo North, and Wandoo Northwest. The total material in the study areas was more than 50 Mt and represented a significant proportion of the areas that were proposed to be mined during the 2000 feasibility study payback period. AngloGold Ashanti concluded that there were areas of both potential upside as well as downside within the 2000 feasibility study resource model that warranted additional investigation. AngloGold Ashanti also conducted a review of the geological domaining during 2001 to 2002 that aimed to refine the 2000 feasibility study gold/copper domains and incorporate the post 1999 resource drilling.  BGMJV, 2002. Following completion of the AngloGold Ashanti and Golder models in 2002, a combined Golder, Newmont, and AngloGold Ashanti review indicated that additional refinement of the resource model was required, primarily in the areas of restraining high-grade intercepts, and the Golder Chain of Mining methodology applied.  BGMJV, 2003. Audit of the AngloGold Ashanti local Mineral Resource estimate models of gold and copper.  BGMJV, 2004a. Comparison of the local Mineral Resource estimates for gold between the Golder 2003, trial mining conditional simulation, and 2000 feasibility studies. It was concluded that the results were within the expected variance for estimates completed independently by two different estimators using two different methodologies;  Newcrest Mining, 2004. An alternative assay-based interpretation was undertaken that assumed a ‘lode-style’ approach to the mineralization, which used a nominal 0.45 g/t Au cut-off for Wandoo North, and an ordinary-kriged (OK) estimate for selected areas of the Wandoo North and Wandoo South;  BGMJV, 2004b: A sensitivity study using the Newcrest-alternative interpretation for Wandoo North. The study indicated that there were globally no differences between the approaches. The geological database was reviewed and assessed in relation potential sample assay bias, metallurgical sampling, historical BGMJV hard rock production, reconciliation drill spacing, and drill orientation. The review confirmed that there were limited effects on the resource model or mine plan from the items identified;  Newmont, 2012. Internal Mineral Resource and Mineral Reserves Review (3R);  Newmont, 2013. Internal business planning review.;  Newmont, 2014. Internal Mineral Resource and Mineral Reserve Review (3R).

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12.4.3 Data Validation All data imported into the database must go through validation steps before being merged. The validation is set up to automatically run on import, giving error messages when any failures occur. Data cannot be imported until the errors have been corrected. All importing is supervised by the site-appointed Geology Database Administrator. Further validations are run on the data monthly to ensure no data inconsistencies arise. This allows for a consistent approach to data validation. 12.4.4 Sample Preparation The main quality control issues with sample preparation at the core shed are data entry issues associated with the sample numbering and incorrectly labeled standards. Currently, 2 m samples of core are placed into pre-numbered bags. The sample numbers correspond with ticket books, on which the hole number and sample interval are recorded. This process is being adapted to reduce paper data entry and ticket books will be replaced with MS Excel® spreadsheets. To minimize data entry errors in this process, the importation of sample numbers into the database is automatically validated using the acQuire® software interface. 12.4.5 Crushing and Grinding Performance Genalysis completes grind checks on approximately one in 25 samples. This is at both target levels of 3 mm crushed and 90 µm grinding levels. Boddington mineralization is extremely hard. In October 2011 the laboratory begun using a statistical process control to monitor the crusher product in real time and this allowed them to take action immediately as adjustments were required. This improved the overall performance of the crush. The hard Boddington rocks are still an issue for the milling although there has been an improvement from previous years. 12.4.6 Drill Bias Review Examination of bias between the RC and core drill methodologies was undertaken by Peattie (2004) and Douglas (2004). Peattie, 2004 An analysis of potential borehole bias was undertaken for the Boddington deposit, which reviewed borehole diameter, drilling methodology, drilling orientation, sample lengths and assay. The analysis concluded that orientation and drilling methodology does appear to have an influence on grade and therefore will have an influence on the Mineral Resource estimates at the Project. In addition, Peattie (2004) noted:  Around the cut-off grade there is very little difference in the Mineral Resource estimates by excluding the RC samples. At higher grades, the difference resulting from the exclusion of the RC samples become pronounced;  There is no definitive way of selecting the datasets to be used in estimation based on drilling direction. It is difficult to determine which of the drilling orientations is correct, due to the large number of drilling azimuths and dips in Boddington. Any benefit of removing data will often be lost because of the amount of data that is removed in the process.;  The variography and statistics suggests that the 60° and 90° azimuth appears to the best angle to drill for the optimum intersection of mineralization, as confirmed by the geology of the three main directions of mineralization.

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 There is evidence that there is over sampling of the high grades in the RC drilling; RC drilling in 2004 amounted to 24% of the total holes informing the then Mineral Resource estimate. Douglas, 2004 Comparison of paired data for RC versus core and easterly versus westerly-oriented drilling were completed on data supplied by NBGJV. The data were 4 m composites, tagged by wire frames to their respective domains. Douglas (2004) concluded:  Core to RC comparison was made for the whole population and by downhole RC depth. No apparent bias exists between core and RC when the total population is viewed;  Drilling direction showed no apparent bias between easterly and westerly drilled holes when the entire population was viewed;  Drilling direction shows highly variable results when viewed by domain. However, only the Eastern Volcanics north (domain 7) and Northern Volcanics B (domain 14) showed consistent biases of greater than 10% regardless of distance or whether searching from easterly- or westerly-directed drill holes. The study concurred with the findings of earlier reviews that “biases may exist but are inconsistent and not sufficiently developed to be incorporated as factors or adjustments. The use of reasonable geologic domaining and robust geostatistical estimation techniques should reduce the risk of any biases, should they exist.”

12.5 Comments on Data Verification The process of data verification for the Project has been performed by Newmont personnel, staff from Newmont’s predecessors and external consultancies contracted by Newmont. In the opinion of the QP, 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 Reserves estimation, and in mine planning, based on the following:  External database, sampling method, and resource estimation reviews were undertaken annually from 2000 to 2005 and in 2007. NBG has subsequently implemented the recommendations from these reviews, or has placed the areas noted under review;  Internal reviews were completed by the JV partners individually, and by the BGMJV, most notably between 2001 and 2005. Aspects reviewed included the database, geological domaining, geological interpretations, metallurgy, and Mineral Resource estimation processes and results. NBG incorporated results and recommendations from these reviews into the current Mineral Resource estimate as appropriate;  Internal Mineral Resource and Mineral Reserves reviews were completed by Newmont in 2012 and 2014;  External Mineral Resource and Mineral Reserves reviews completed by Golder in 2017 and 2018;  Drill collar data are typically verified prior to data entry into the database, by checking the drilled collar position against the planned collar position;  Standard and blank QA/QC data are checked on a monthly basis. Samples that fail are typically reanalyzed;

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 A check of the density values for lithologies across the different deposits indicates to the QP that there are no major deviations in the density results.

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13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 Metallurgical Testwork History During feasibility-stage studies from 1997 to 2003, several programs of metallurgical testwork were completed on the Boddington deposit. Eight metallurgical domains, based on ore types, formed the basis for ore characterization studies (refer to Figure 13-1). The domains include Blackbutt, Southern Volcanics, Central Diorite, Northern Diorite, Pipeline, Far North, South Zone, and Southern Diorite Deeps (also known as Blob). In 2008, a metallurgical testwork program was carried out on 37 samples selected from seven domains associated with Mineral Reserves/Mineral Resource drilling programs. These samples were submitted for comminution, flotation, and flotation tailings cyanidation testwork, following the flowsheet and procedures developed during the 2000 feasibility study. Further test work was performed on Mineral Resource/Mineral Reserves samples in 2017, targeting an area of the pit with elevated metallurgical risk. The South Pit S09 Pit Area test work confirmed the material performed in line with existing metallurgical models. During 2017, the investigation into the impact of stockpile oxidation on flotation recovery dispatched the first samples for test work following field collection of aged material by geology and the controlled aging of fresh material collected during 2016 by Metallurgy. The results indicated the expected deterioration in copper recovery but no degradation in gold recovery was evident at this stage. 13.1.1 Variability Composites Copper concentrate samples were analyzed to provide a better understanding of the deleterious element content and to generate predictive models, based on feed grades, for use in mine scheduling. Results of the composite testing are presented in Table 13-1. Table 13-1: Variability Composite Results Mass Au Cu Grade Dist Grade Dist % g/t % % % Calculated Head Grade 100 0.87 100 0.11 100 Copper Concentrate 0.62 83.9 59.9 15.3 83.4 Scavenger Tail Leach 96.9 0.138 16.9 0.014 0 Cleaner Scavenger Tail Leach 2.50 0.478 5.3 0.127 0 Final Residue – 0.156 – 0.017 – Total Recovery – – 82.0 – 83.4 Cu Solution Residue – – – 22* –

Note: * solution residue figure in ppm Cu Variability and Large-Scale Kinetic (LSK) composites with poor overall gold recovery and with poor performance in both the flotation and leach areas were subjected to diagnostic analysis. The program focused on the scavenger flotation tailing and the scavenger flotation tailing leach residue, which were the two largest sources of gold loss.

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Note: Figure aligned to Mine Grid Figure 13-1: Metallurgical Domains

Table 13-2: Variability Composite Results (LSK Composites) Mass Au Cu Grade Dist Grade Dist % g/t % % % Calculated Head Grade 100 0.766 100 0.113 100 Copper Concentrate 0.59 77.9 59.9 15.5 77.9 Scavenger Tail Leach 96.1 0.118 15.8 0.016 0 Cleaner Scavenger Tail Leach 3.33 0.45 6.2 0.242 0 Final Residue – 0.138 – 0.024 – Total Recovery – – 82.0 – 77.9 Cu Solution Residue – – 24 –

Note: *solution residue figure in ppm Cu

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13.1.2 Comminution Testwork Composite samples of half NQ core were selected from the Central Diorite and Northern Diorite domains. Parallel pilot-scale comminution tests were then performed, using HPGR and conventional crushing to ball mill and rod mill feed size distributions. Bench-scale assessment of the grinding specific energy requirements was undertaken at various feed size distributions, using bond ball mill and rod mill work index tests and two proprietary tests of expert external entities. Based on these results, a ball mill receiving feed from an HPGR-based crushing circuit will consume approximately 90% of the theoretical specific energy. On the same basis, a ball mill receiving feed from a conventional crushing circuit will consume between 104% and 110% of the theoretical specific energy. The overall HPGR-based comminution circuit specific energy, including crushing, HPGR, and ball mill, but excluding peripherals such as screens and conveyors, can be estimated approximately from the theoretical bond calculations. Final comminution results are as presented in Table 13-3.

Table 13-3: Comminution Results

Parameter 35.2 Mtpa Bond Abrasion Index 0.50 Bond Rod Mill Wi kWh/t 23.4 Bond Ball Mill Wi kWh/t 15.6 UCS Mpa 140 Impact Crushing Wi (+ 19 mm) 8.3 JK Tests Dwi 10.6 A 68.7 b 0.4 A x b (Calculated) 27.3 ta 0.22 t10 (Calculated with Ecs of 1.0kwh/t) 22.0

13.1.3 Tailings Characterization and Assessments Tailings characterization work for the mine tailings consists of the following programs:  Testing was undertaken for Worsley Alumina in late 1999. The samples were tested at 60% solids with some additional tests at 63% solids. The P80 of the samples was approximately 70 µm;  A second set of tailings was tested by Knight Piésold in Nevada in July 2004. The samples were tested at 60% solids. The P80 of the samples was approximately 90 µm;  In 2005, ongoing metallurgical testing resulted in a change in grind size from a P80 of 90 µm to a P80 of 150 µm. As a result, additional tailings testing was undertaken. The tailings were tested at nominally 50%, 60%, and 65% solids;  Cone penetrometer testing in 2014;  Cone penetrometer testing in 2015. Comparison against the 2014 data shows a similar result in 2015. The tailings should behave well under drained conditions,

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which are indicated by an overall friction angle of 30° (Knight Piesold Consulting, 2016);  The recommended undrained shear strength ratio is 0.25 considering lower values obtained from shear vane tests. 13.1.4 Metallurgical Performance Previously the metallurgical performance model was determined by four series of metallurgical recovery functions produced during the 2003 feasibility study. However, the plant has not met these targets, consistently underperforming against these models by approximately 5% for copper and 2% for gold. The reasons for this include:  Copper and gold flotation recovery decreases by 1.1% and 0.9% respectively for a grind size increase of 10 um (Runge, 2012). All the feasibility study tests were conducted at a grind size P80 of 150 µm, therefore the feasibility study models are unable to predict a recovery drop caused by grind size increases experienced in the plant;  The copper concentrate grade achieved in the process plant has been consistently 1 to 2% higher than that predicted by the models, in order to assist with marketing of the low-grade copper concentrate; this has resulted in a reduction in copper recovery;  A significant amount of the mill feed comes from re-handle material that consists of stockpiles built from ore sourced from all mine domains. During the feasibility study functions were developed for each ore domain, and a weighted average of these is assigned to the stockpile. However, owing to the size and variable age of the stockpiles, the performance of this material does not meet the weighted average functions. As a result, a single recovery function was developed to encompass the entire ore body based on previous plant performance. The only exception is the Bond Work Index for each ore domain is used to calculate the grind size as throughput is considered to remain constant with grind size increasing proportionally to ore hardness (Roberts, 2012). The recovery of gold via the Gravity Recovery Gold (GRG) circuit has been excluded from the model as the gravity circuit is currently not operational. The function was expressed in terms of gold, copper and sulfur grade in the feed and were used for mine optimization, mine planning, mine scheduling, and financial modeling. Functions were developed for the following metallurgical parameters:  Mass of copper concentrate;  Grade of copper in copper concentrate;  Grade of gold in copper concentrate;  Recovery of copper;  Total recovery of gold;  Recovery of gold via copper concentrate;  Recovery of gold via leached tailings. Copper recoveries are calculated using the following equation:

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 Recovery of Cu = (Cu Concentrate Grade (%Cu) x Mass)/Cu in Feed. Gold recovery in the copper concentrate is calculated by:  Gold Grade of Cu Concentrate (g/t Au) = Au Rec to Con% * Au Head g/t/Mass of Con%. The total Au extraction can be estimated by subtracting the residue grade from the head grade and dividing by the head grade. Over the remaining life of the mine the gold recovery averages 84%. The average grade over the life of the mine is 0.67 g/t Au. The annual average copper head grade remains reasonably constant between 0.10% and 0.12% until 2022, when the high-grade Southern Diorite Deeps ore type is processed. The mean grade over the life of the mine is 0.10% Cu. The annual average copper recovery fluctuates between 78.0% and 80.0% due to variations in the type of ore being mined. Over the life of the mine the copper recovery averages 78.0%, before tapering off during the last few years as the mine treats low grade stockpiles.

13.1.5 Gold Revenue Approximately 75% of the Project gold revenue is obtained from the gold contained in the exported copper concentrate, while the remaining 25% is obtained from the gold bullion produced on site. Bullion production was originally planned to be derived from two sources, namely leaching and gravity concentration. However, given the gravity recovery circuit has not performed as expected, with throughput reductions experienced in milling due to operation of the flash flotation cells, all gold bullion produced on site currently arises from the two leach circuits that are operating. The flash flotation and gravity circuits were decommissioned in 2016.

13.1.6 Testwork and Feasibility Study Conclusions The following conclusions have been drawn from the collective comminution and metallurgical testwork, pilot tests, value engineering studies, peer and other expert reviews and the 2003 feasibility-level engineering and cost study:  The eight geologically distinct domain rock types identified demonstrate similar metallurgical responses, except possibly for high-grade mineralization in Pipeline, which tends to contain more silicate refractory gold;  The ore types are characterized by similar physical and chemical properties such as slurry viscosity, specific gravity, compressive strength and elemental distributions;  The basement ores are competent with respect to size reduction and are relatively homogeneous between ore type categories, locations and with depth. Laboratory testwork results predicting ore competency have been adequately supported by operational data;  Ball mill feed size distribution, and thus HPGR throughput and product size, does have a significant effect on plant throughput;  Extensive modeling and dynamic simulation of the processing circuit have supported the plant throughput of 39 Mtpa. This corresponds to a circuit utilization of 89% and a ball milling circuit duty of 4,980 tph.

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 There is little difference in the metallurgical response of composite samples from the transitional zone, being the first five meters of bedrock material below the clay zone;  The transitional material does not contain appreciable levels of clay as the oxide zones have been mined to refusal of free digging;  Variability between ore types is low and mostly linked to copper head grade, which ranges from 0.04% Cu to 0.31% Cu. This results in copper concentrate grades ranging from 12% Cu to 22% Cu, but on average a 15.8% Cu concentrate grade is expected for a head grade of 0.11% Cu;  The high gold content and low annual tonnage of concentrate are considered attractive features of the concentrate for smelters;

 The main concentrate penalty charge is for Al2O3 + MgO, with a modest, intermittent, arsenic penalty and a low but more consistent bismuth penalty (particularly when treating ores from the Pipeline domain);  Weak acid-dissociable cyanide (CNwad) destruction requirements are reduced by the adoption of Caro’s acid method for cyanide destruction;  Gold lost in leach residues is locked in silicates, particularly clinozoisite, actinolite and biotite, which contain gold particles below 10 μm and often below 1 μm. 13.1.7 Metallurgical Testwork – Post Feasibility Study In 2008, sample selection for metallurgical testwork was performed on material from the Southern Volcanic, Central Diorite, Northern Diorite, Blackbutt, Pipeline, Far North and Southern Diorite Deep domain. Of the 37 samples selected, 18 samples were submitted for abrasion index, Bond rod mill work index and Bond ball mill work index testing, while all 37 were tested for flotation and leaching performance. In 2011, another sample selection was made from the Far North domain in North Pit, the area converted to Mineral Reserve during 2009 and 2010. This selection covers the sparsely sampled northern area of the Wandoo pit. A total of 29 drill core samples were collected. All selected composite samples have approximate down hole length of 24 m, to obtain a 60 kg composite sample required for flotation test work. Only five samples were selected for comminution testing. All 29 samples were used for Flotation and leach testing. Figure 13-2 presents the 2011 sample locations. Further test work was performed on Mineral Reserve/Mineral Resource samples in 2017. This work targeted an area of the pit where elevated metallurgical risk existed. The S09 area test work confirmed the material performed in line with existing metallurgical models. Comminution Testwork The test work was undertaken at AMMTEC laboratory which had previous experience with similar test wok during feasibility study. The test work at AMMTEC laboratory comprised;  Comminution (Bond ball mill work index, Bond rod mill work index, abrasion index);  Flotation and leaching (locked cycle flotation test, scavenger tail leach, cleaner scavenger tail leach).

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Figure 13-2: 2011 Long Section Showing North Pit Metallurgical Sample Locations (looking east) From the comminution testwork conducted in 2008 results, the abrasion index varied from 0.15 to 0.67 with an arithmetic average abrasion index of 0.48, similar to the 2005 feasibility study update average result of 0.50. Bond abrasion indices for samples collected from Far North domain in 2011 are still within the range of historical data. The Bond ball mill work indices varied from 13.2 kilowatt hours per tonne (kWh/t) to 20.7 kWh/t with an average test result of 16.7 kWh/t, slightly higher than the 2005 feasibility study update average test result of 15.6 kWh/t. This suggests a possible increase in ore hardness with depth. Bond ball indices for recently conducted tested samples from Far North Domain in 2011 are within the range of historical data but higher average values compare to previous data. The Bond rod mill work indices ranged from 19.5 kWh/t to 28.6 kWh/t, with an average test result of 22.8 kWh/t. This was similar to the 2005 feasibility study update average of 23.4 kWh/t. Bond rod mill indices for Far North sample collected in 2011 are within the historic range with an average of 21.91 kWh/t. Flotation and Leach Cyanidation Testwork Result From the flotation locked cycle testwork conducted in 2008, copper calculated head assay for the samples ranged from 0.05% Cu to 0.29% Cu with an average of 0.14% Cu. Copper calculated head assay for Far North samples collected in 2011 varies between 0.04 to 0.26% with an average of 0.12%. The gold calculated head assay varied from 0.4 g/t Au to 2.5 g/t Au with an average of 0.8 g/t Au. For Far North samples tested in 2011, the calculated gold assay varied from 0.17 g/t to 1.15 g/t with an average of 0.54 g/t. Flotation locked cycle testwork resulted in copper concentrate grades ranging from 10.2% Cu to 28.1% Cu with an average of 20.2% Cu, higher than the 2005 feasibility study update average of 15.8% Cu. Copper concentrate grade for Far North domain samples range from 9.73% to 28.3% Cu with an average of 21.45 %Cu. The gold content in copper concentrate ranged from 25 g/t Au to 370 g/t Au with an average of 85 g/t Au. Gold content in copper concentrate for 2011 Far North domain samples range from 23 to 421 g/t Au with an average of 92 g/t Au. Copper recovery ranged from 67% to 91% with an average of 81%. The overall gold recovery varied from 70% to 93% with an average of 83%. Both gold and copper

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recoveries were in line with feasibility study update testwork results. Over all gold recovery for Far North was between 67% to 90% with an average of 78% and copper ranged between 62% and 90% with an average of 78% for the Far North domain. Gold recovery distribution on average was 58% to flotation, 4.5% recovery to the cleaner scavenger tailings leach, and 20.9% to the scavenger tailings leach. The average gold extractions were 81.5% and 59.5% of contained gold from the cleaner scavenger and scavenger tailings leaches respectively. For tested Far North domain samples, the gold recovery distribution averaged 56% in flotation concentrate and averaged 22% from the combined flotation tailings leach. Comparison of Testwork Results with Metallurgical Models Copper concentrate grades were slightly higher compared with the model predicted grade from the feasibility study update. The copper recovery model prediction for the new samples fell within the range predicted from previous models. There was a tendency with both the recent testwork and the feasibility study update test results for the models to over-predict copper recovery at low copper head grades and under- predict recovery at high head grades. The gold to copper upgrade ratios fitted well with the previous model. The model prediction for gold recovery to copper concentrate and overall gold recovery of the new reserve samples were in line with the feasibility study update predictions and current plant performance. Molybdenum and arsenic grades in copper concentrate for the new testwork were in line with model predictions developed during the 2005 feasibility study update and achieved in the operation at that time. Conclusions from Post-feasibility Testwork 1. Ore hardness and abrasiveness of the samples were in line with the orebody average, with the possible exception of a slight increase in Bond ball mill work index with depth. 2. The new test results fall within the model predictions using primary and secondary functions developed during the 2005 feasibility study update for copper concentrate grade, copper recovery, overall gold recovery, gold recovery to copper concentrate, and gold overall recovery to the leach circuits. They also fit within the accuracy of the recently developed plant performance models Metallurgical Models Update The metallurgical models were updated in 2012 based on end of month data from the start of the operation. The results of the 2012 update were: 1. Copper and gold recovery decreases by 1.1% and 0.9% respectively for a grind size increase of 10 um (Runge, 2012). All the FSU tests were conducted at a grind size P80 of 150 µm, therefore the FSU models are unable to predict a recovery drop caused by grind size increases experienced in the plant. 2. The copper concentrate grade achieved in the process plant has been consistently 1-2% higher than that predicted by the models, in order to assist with marketing of the low-grade copper concentrate. 3. A significant amount of the mill feed comes from rehandle material; stockpiles built from ore sourced from all domains of the mine. As the FSU study has developed functions for each ore domain a weighted average is assigned to the stockpile.

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However, owing to the size and variable age of these stockpiles the performance of this material does not meet the weighted average functions. These models were validated against the mid-month and weekly plant data which showed a strong correlation indicating they would be suitable for use in forecast and geological block models. A revision of the plant performance recovery functions was undertaken during June 2014 (Petrucci, 2014) in the lead up to 2014 Mineral Reserves and Mineral Resource development. The methodology outlined in the original report (Metallurgical Performance Models, Roberts, 2012) was again used with more recent plant data. A minor update incorporating process improvements was conducted in 2015 (Petrucci, 2015). A further update was conducted on the copper model in 2016 (Sandi, 2016). The gold model has not been changed since 2015. Changes to the copper recovery model were due to the plant outperforming the existing model based on metallurgical improvements in the Processing Plant. The changes contributing to the plant outperforming the model included: reduced grinding media size which has led to improved mineral liberation through changed grinding conditions, reduced floatation pH which has led to less gold suppression through hydroxide coatings, and improved cleaning and maintenance of flotation launder lips increasing concentrate mass pull and reducing froth transport times.

Figure 13-3: 2018 Gold Recovery YTD Reconciliation vs Budget

The 2018 gold recovery against budget which used the 2015 model showed a small percentage of “Other” showing good conformances against the model.

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Figure 13-4: 2018 Copper Recovery YTD Reconciliation vs Budget

The 2018 copper recovery showed a trend of outperforming the 2016 model used in the 2019 business plan cycle. While some portion of this is attributed to higher than budget feed grades, the operational improvements and Full Potential projects are responsible for most of the difference.

Figure 13-5: 2018 Copper Recovery Model Performance (Actual EOM Reconciled Data vs 2016 Model) The copper metal in tails comparison between model and actuals (refer to Figure 13-5) shows good model performance.

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Figure 13-6: 2018 Gold Recovery Model Performance (Actual EOM Reconciled Data vs 2016 Model) The gold model is based on that generated in 2012 with an update to include later years and improvement assumptions based on improvement projects. The offset between model and actuals is presented in Figure 13-6.

13.2 Mill Throughput Modelling Mill throughput is set by several bottlenecks independent of ore properties as those varied by ore domain are within the capabilities of the ball milling circuit. Bottlenecks include transient causes such as modular maintenance activities in the fine crushing circuit and fixed causes such as conveyor and CIL inter-stage screen capacity. The improvement in average mill throughput over time shown in the 2019BP is based on operating for a larger proportion of time at rates that have already been demonstrated (1,298 tph average milling rate for the 2018 calendar year), as the reliability of items that cause transitory bottlenecks are improved. When formulating the business plan the operating throughput rate is fixed and the grind size is allowed to vary with ore hardness, resulting in differences in recovery by ore domain. Grind size is predicted using the Bond Work Index, with a fixed power input (Metallurgical Performance Model Update 2014; Petrucci, 2014).

13.3 Throughput Assumptions Plant throughput is predicted from a combination of average mill throughput rate and overall mill utilisation for each milling circuit. The mill maintenance schedule has been optimized and changes implemented by the Full Potential team.

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The shutdown footprint comprises of:  2 x 168 hr full plant shut-downs;  1 x 120 hr full plant shut-downs;  Cyclone feed pump rebuilds are being addressed outside the planned shuts above at 12/13-week intervals with ongoing projects aimed at improving life to meet the full plant shut schedule;  Mill re-lines occur outside of the full plant shutdowns. This is to relieve labour constraints within full plant outages and reduce work front interactions. Target Mill throughput for the business plan is taken as the best rolling average 12 months of operation and set for individual mills (i.e. 5,050 for 4 mills ~ 1,263 tph per mill). Maintenance activities for the month are aligned to the throughput rate to produce a monthly milled tonnage. The maintenance activities are based on routine 17/18-week cycle (or multiples thereof) and this is rolled through for subsequent months. Planning is by month for three years and then quarterly for one year then annualised to LOM.

13.4 Recovery Estimates LOM projected average recovery figures are as follows:  Copper: 78%;  Gold: 83%. Recovery factors are appropriate to the mineralization types and the selected process route.

13.5 Metallurgical Variability Samples selected for metallurgical testing during feasibility and development studies were representative of the various types and styles of mineralization within the different deposits. Samples were selected from a range of locations within the deposit zones. Sufficient samples were taken so that tests were performed on sufficient sample mass. A review of the expected variability of the head grade of gold, copper, sulfur, arsenic, bismuth, and molybdenum, using the 35 Mtpa mining schedule, was completed for each major metallurgical domain:  Gold grade variability showed a positively skewed distribution with 80% of the grades below 1.0 g/t Au and the orebody average at ~0.70 g/t Au. Approximately 80% of the grades fall in the range from 0.5 g/t Au to 1.2 g/t Au;  The copper grade variability is normally distributed. Southern Diorite Deeps domain material contains the highest copper grades. There are also minor quantities of high- grade copper ore in the Southern Volcanics and Northern Diorite domains. Approximately 80% of the grades are below 0.12% Cu and the majority fall within the range of 0.02% to 0.20% Cu;  The sulfur grade variability is essentially a positively-skewed distribution about the orebody average of 0.21% S. There is a significant quantity of above-average grade sulfur material in the Southern Diorite Deeps domain. Approximately 80% of the grades are below 0.26% S and 80% fall between 0.08% S and 0.30% S;

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 The arsenic grade distribution is slightly positively skewed. Small quantities of the higher arsenic grades occur in some Wandoo North domains. Approximately 80% of the grades are below 10 ppm As and most fall within the range 0 ppm As to 20 ppm As. The orebody average is 18 ppm As;  The bismuth grade is normally distributed. There is a significant quantity of above- average grade bismuth material in the Central Diorite domain, with lesser quantities in the Southern Volcanics domain, both in Wandoo South. Approximately 80% of the grades are below 4 ppm Bi and approximately 80% fall between 1.0 ppm Bi and 6.5 ppm Bi. The average for the orebody is 3.4 ppm Bi;  The molybdenum grade distribution is a skewed positive distribution around the orebody average of 36 ppm Mo. Central Diorite and Southern Diorite Deeps ore contains the greater proportion of high molybdenum grades. Approximately 80% of the grades are below 40 ppm Mo and 80% fall within the range from 5 ppm Mo to 70 ppm Mo.

13.6 Deleterious Elements Since commissioning in 2009, the operation has actively managed the arsenic level in plant feed and through concentrate blending techniques controlled the level in copper concentrate shipments to below the penalty rate trigger, hence no penalties have been incurred project to date. Bismuth has held a closer association with gold in the Wandoo ores however, so it has resulted in penalty levels being exceeded, particularly in the first 2 years of operation. Most of the high bismuth ores have been processed, resulting in very low to no penalty charges being incurred more recently. Alumina remains the largest penalty element present in the NBG copper concentrate, with shipments regularly exposed to a penalty adjustment. However, at 4-5% Al2O3 the levels are not far off the trigger point of 3% in most contracts and a modification to the process which is due to be completed by Q1 2019 should eliminate these and provide a higher-grade copper concentrate grade for sale.

13.7 Comment on Mineral Processing and Metallurgical Testwork In the opinion of the QP:  The metallurgical testwork completed on the Project has been appropriate to establish optimal processing routes for the different mineralization styles encountered in the deposits;  Testwork was completed on mineralization that is typical of the deposit styles;  The mill process and associated recovery factors are considered appropriate to support Mineral Resource and Mineral Reserves estimation, and mine planning;  A detailed review of the metallurgical models and comparisons with process plant actual performance was conducted in late 2012, with updates completed in 2014, 2015 and 2016 and verification of the model outputs completed in 2018;  The plant will produce variations in recovery due to the day-to-day changes in ore type or combinations of ore type being processed. These variations are expected to trend to the forecast recovery value for monthly or longer reporting periods.

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14.0 MINERAL RESOURCE ESTIMATES

14.1 Introduction The Project database was closed for Mineral Resource estimation purposes on 25 July 2018. The drill hole database contains core drilling information from numerous drilling campaigns beginning in the 1980s through to July 2018. The Mineral Resource estimate is based on the August 2018 Ordinary Kriged (OK) Mineral Resource model and the mining face positions as at 31 December 2018. The latest Mineral Resource incorporated 126 RC and core drill holes completed in 2017-2018 totalling 19,764 m. This resulted in a combined dataset containing 7,236 drill holes for approximately 1.39 Mm (comprising 2,463 RC, 1,218 grade control RC, and 3,555 core drill holes) for Mineral Resource modelling and subsequent estimation (refer to Section 10.0 and Table 10-2). Geological solids for the Project Mineral Resource models were constructed by NBG site geological staff under the direction of Mr. Rohan McCormack, MAIG, Senior Resource Geologist and Newmont employee, under the supervision of the QP.

14.2 Geological and Mineralization Models Wireframe models have been constructed for Au, Cu, S, As, Mo and Bi estimation domains as well as the dolerite lithology and weathering surfaces. The construction of these models considered the logged and mapped lithology, geological structures, alteration, mineralisation and metallurgical characteristics of the deposit. These items were interpreted on section and plan, and reconciled in cross section, long section, and level plan. Figure 14-1 to Figure 14-3 illustrate the estimated gold grades in plan and cross-section views. Copper was modelled within grade shell wireframes constructed by dividing the deposit into four regions based broadly on lithology and, within these regions interpolating high grade (1, 500 ppm) and medium grade (750 ppm) grade shells. The Dolerite lithology wireframes were constructed from coded dolerite intersections from the logged geology and assigning to vein structures by interpretation in section and plan and modelled as 3D wireframes. The wireframe models were used to code the drill hole composites by majority and stored in the domain fields (DOMAU, DOMCU, DOMS, DOMAS, DOMMO and DOMBI) in the composite files. The wireframes were also used to code the block model, with the Au domain recorded in the LITH field.

14.3 Bulk Density Model In-situ bulk density values are interpolated when sufficient data exists. Areas of insufficient data have average values assigned. The 2018 Mineral Resource estimate is based on combined historically and recently collected data. Domains that did not contain any in-situ bulk density data were assigned a mean value of 2.75 g/cm3. An in-situ bulk density of 3.00 g/cm3 was applied to all modelled dolerites. The weathered bulk density values are assigned to regions delineated by the various weathering products.

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Plan view at 0mRL Au grade model and cross-section lines for Figure 14-2 and Figure 14-3. Figure 14-1: Gold Grade Plan at 0mRL

Note: Block model showing Au grades and surface at 31 December 2018 (gray line)

Figure 14-2: Cross-section North Pit (oblique section A-A’ looking west)

Note: Block model showing Au grades and surface at 31 December 2018 (gray line) Figure 14-3: Cross-section South Pit (oblique section B-B’ looking west)

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14.4 Geotechnical Model The geotechnical model for the Project has been defined by geotechnical drilling and logging, laboratory test work, rock mass classification, structural analysis, and stability modelling. The primary geotechnical domains may be summarized as either oxide, transitional or bedrock. An extensive database and history of geotechnical modelling in the oxide domain has been developed over the initial 12 years of operation. The rock mass was characterised and grouped into geotechnical domains for design purposes during a 2014 study, using geotechnical and hydrogeological models as well as operational experience. An updated geotechnical study was completed in October 2016, resulting in an updated pit wall configuration in 2017. A geotechnical domain is an area or region with similar rock mass conditions (structures and fabric, rock strength, fracture frequency, hydrogeology) that is expected to have the same or similar response to excavation, blasting, and slope geometry. The South Pit geotechnical model is split into five domains:

 Domain 1: Andesite and diorite;

 Domain 2: Moderately fractured dolerite;

 Domain 3: Transition zone;

 Domain 4: Broken zone associated with sub-horizontal dolerite sills;

 Domain 5: Oxide. The North Pit geotechnical model has six domains:

 Domain 1: Andesite and diorite;

 Domain 2: Dolerite;

 Domain 3: Transition zone;

 Domain 4: West Shear zone;

 Domain 5: A-Breccia and South-Bowl;

 Domain 6: Oxide. The structural model includes 10 major structures in South Pit, and one major structure in North Pit.

14.5 Acid Rock Drainage Model A total sulphur estimate was used to establish the maximum potential acidity (MPA) on a block by block basis. The net acid-producing potential (NAPP) of the mineralization was also estimated to indicate if the material had the potential to ARD. For classification of the waste into ARD categories the relationship between NAPP and net acid generation (NAG) was examined. Results for NAG versus NAPP were split into North Pit and South Pit. The two areas were examined separately, as the NAPP to NAG calibration was expected to be different due to the presence of carbonate veining in North Pit.

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14.6 Composites Compositing of Au, Cu, As, Bi, Mo, and S was performed in Vulcan® software to a 12 m length. The composite length is equal to the 12 m bench height and is a multiple of the usual 2 m assay interval. Compositing was broken at the start of each domain. Small composite lengths were merged with adjacent composites with a minimum length of 2 m. Length weighting was used during the estimation for all element estimates run in the Vulcan® software. This allowed the use of smaller lengths rather than excluding them as unrepresentative.

14.7 Statistical and Exploratory Data Analysis Statistical analysis of the Au and Cu data sets was performed to calculate cap grades. Contact analysis of Au and Cu was conducted to validate the treatment of data against the contact of adjoining domains. Five domains were treated with soft boundaries with the remainder estimated as hard boundaries. Grade capping (via high-grade cuts) was performed on the composite data, implemented by reduction of the value (not removal) for use in the estimation. High-grade cuts were applied to Au, Cu, As, Bi, Mo, and S, for individual domains. Not all domains had a high-grade cut applied. Depending on the domain, the following ranges of high-grade cuts were used:  Gold: 1 to 22 g/t;  Copper: 3,000 to 20,000 ppm;  Arsenic: 30,000 to 40,000 ppm;  Bismuth: 20 to 200 ppm;  Sulfur: 99,000 ppm (not capped);  Molybdenum: 30,000 ppm. The high-grade cuts applied to Au by domain are presented in Table 14-1.

Table 14-1: High-grade Cuts By Domain Code

Au Domain Cap g/t Au 10 4.0 12 2.0 13 10.0 14 4.0 15 7.0 17 3.0 18 4.0 19 3.0 2 6.0 20 15.0 200 8.0 21 7.0 210 10.0 22 22.0

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Au Domain Cap g/t Au 23 8.0 24 4.0 240 6.0 25 5.0 250 4.0 26 8.0 260 15.0 27 8.0 28 5.0 3 5.0 4 8.0 51 10.0 52 9.0 53 5.0 6 4.0 7 4.0 8 1.0

Reconciliation results between January 2015 and July 2018 indicated the Mineral Resource model was underestimating Au and Cu by up to 5%. The resultant testwork included:  Adjusting the dilution parameters pertaining to the post mineralisation dolerite intrusive;  Adjusting the high-grade cut values in gold domains which had a mean sample/composite grade >= 0.4 g/t Au;  Changing the sample search criteria and removing the quadrant restriction used to estimate a block;  Updating variography for domains with new data;  Changing the percentage of large dolerites on the edge blocks during the post-processing stage. The above testwork confirmed that adjusting the dilution parameters has the largest influence on contained ounces. Recent reconciliation results obtained during 2018 indicated that full year reconciliations yielded +8% for contained gold and +8% for contained copper. No grade adjustments were applied to the gold and copper grade estimates to account for these variances. The Mineral Resource estimates are subject to regular and ongoing reconciliation performance reviews, with corresponding adjustments to Mineral Resource estimation parameters as required.

14.8 Variography Spatial variability of the grades for Au, Cu, As, Bi, Mo and S was modeled through directional variography of 12 m composites with high-grade cuts applied.

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Work had previously been carried out confirming domain stationarity for calculation and modelling of variograms (e.g. SRK, Guibal and Jankowski, 2004a) as well in the audit of the 2007 model (Masters, 2007). More recent audits completed by Golder (Golder 2017, 2018) provided independent support for the variography and assessed the change arising from the adoption of a larger 12 m composite for the 2017 Mineral Resource estimate, An overview of the previously applied modelling approach and methods to select continuity orientations is summarized as:  Examining directional variograms and variogram maps identifying strike, dip and plunge within dip planes;  A check of the three-dimensional (3D) integrity of the variography, to establish whether the models of the principal directions fit experimental variograms in other directions. This gives an indication that the model has identified the appropriate continuity orientations and fits the experimental data in all directions;  Confirming consistency with geological control; indications of continuity trends are checked to be consistent with the geological understanding of mineralisation controls. Nugget variances were modelled from downhole variograms based on a 12 m lag interval. Directional variograms were generally based on a 15 m or 25 m lag interval. Variogram models are progressively updated to reflect additional drill hole data, revisions to composite size (from 6 to 12 m, as changed in 2017) and changes to geological interpretations. The most recent variography supporting the 2018 Mineral Resource estimate had updated Au and Cu variography for domains with new data.

14.9 Estimation/Interpolation Methods Estimation of Au, Cu, As, Bi, Mo, and S was completed using OK in Vulcan® software. The final block size used in the resource block model is a regularised size of 20 x 20 x 12 m. The model is constructed in the Boddington Mine Grid (MG) orientation with no additional rotation or sub blocking applied and matches the size of the SMU. The blocks are coded for dolerite domain and weathering profile percentages to honour the dilution from narrow dykes and oxidised units. The kriged estimate for the metals (Au, Cu, As, Bi, Mo and S) and the assigned specific gravity is conducted in a separate block model with parent block size 10 x 20 x 12 m with sub blocking to 5 x 5 x 6 m. The smaller block model is used for the estimation of the metals and specific gravity as it better reflects the resolution of the geological domains. The estimation results are subsequently combined into the larger final block model via a block regularisation process. No waste masking was applied to remove areas of definable waste during grade estimation. The general estimation methodology involved the following:  Analysis to define the parameters used in searches (including minimum/maximum number of data, distance, octant based);  Domains were treated as hard boundaries, except the contacts between Central Diorite North/South/West and Northern Diorite/Far North Lower/Far North which were treated as soft;

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 Estimation of 10 m x 20 m x 12 m SMU size sub-blocks using OK for Au, Cu, As, Bi, Mo and S and assigned specific gravity;  Search radii and orientations are derived from the variograms second structure ranges and orientations, the first pass search is proportionally scaled up to 75 m, with the second pass scaled up to 400 m;  Re-blocking of grade and density estimates to of 20 m x 20 m x 12 m SMU size blocks;  Grade dilution applied due to unavoidable mining of small dolerite bodies;  The estimation was run in Vulcan® software with weighting by length in the kriging with a minimum composite length of 2 m;  A second pass estimation occurred (with larger search) after a first pass designed to incorporate the bulk of the better-quality estimates;  The minimum number of samples varied according to what helped provide quality estimates in the first pass. In the second pass estimations the minimum number of samples was generally three;  Block discretization was 4 x 4 x 1;  High-grade cuts were applied in most domains;  Post-estimation dilution of grade due to small dolerites.

14.10 Post-processing Maximum potential acidity (MPA) and NAPP were recalculated using the S estimate (SUTOT) and the following relationships:  MPA = 30.6 x (SUTOT/10,000);  NAPP = MPA – ANC. where ANC is the Acid Neutralizing Capacity. The large dolerites (>12 m wide) are accounted for as waste portions within the block, reducing the remaining ore portion accordingly. In addition to the volume of the dolerite within the block, an ore loss factor is added to account for ore loss along the dolerite contact. The ore loss factor is guided by back reconciliation of the adjusted modeled dolerite against dolerite from production. The factor of 10% (equivalent to a 3 m skin) is added to the dolerite percentage and removed from the ore percentage for any block with more than 0% large dolerite by volume up to a maximum 100% dolerite. Dilution for small dolerites (<12 m wide) is applied to the grade variables in the model. The volumetric portion of each parent block within the small dolerite wireframes is calculated and the block grade reduced by this portion. The resulting diluted grade is evaluated the same as any other ore block. The dilution grades of 0.15 ppm Au and 500 ppm Cu were applied, derived from the mean of the data within the small dolerite volume. The 2018 Mineral Resource model has a modifying factor applied to the S block grades. This is applied as a direct factor of 1.1 multiplied by the interpolated S grade for all blocks. This results in most of the domains being factored. The selection of the factor was based on the mill reconciliation data. The application of the factor results in an improved prediction of potential acid forming waste.

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14.11 Block Model Validation A comprehensive model validation process is implemented at several points of the modeling process to ensure that obvious errors were not introduced. The Mineral Resource estimate was validated using the following steps:  Back reconciliation against Ore Control data at the various cut off categories and against mill data at the mill cut off. The ore control back reconciliation is performed on a 12-month period, most recently from July 2017 to July 2018;  Visual validation of the domain coding of both the composite data set and the domain model for all elements modelled in section and plan;  Visual validation of the interpolated elements against the composite data in plan and section;  Statistical comparison of composites versus blocks: Statistics for Au, Cu, As, Bi, Mo, and S generally match those from the block model estimates broken into easting, northing, and RL (elevation). There is some discrepancy for Au at higher RLs, but this is likely to be due to clustering of high grades;  Prior to 2012, sensitivity to block size, data density and variogram model: Various studies were used to determine the block size for estimation, the selective mining unit size, the effect of data density on quality of estimation (and optimizing estimation parameters accordingly), and the sensitivity to the variogram model;  Sensitivity to high-grade-cuts/high values: This was performed with the use of an OK model interpolated using uncut data and including varied high-grade cuts in reported data.

14.12 Classification of Mineral Resources The Mineral Resource classification for the Project is based on drill hole density as defined by the average sample distance to the nearest three drill holes with gold assays. The process involved application of an inverse distance interpolation to store the average distance to the nearest three drill holes. The interpolation used search radii by resource category of 50 x 50 x 12 m (Measured Resource), 100 x 100 x 12 m (Indicated Resource) and 50 x 50 x 12 m (Inferred Resource). Mineral Resource classification was defined as follows:  Measured Resource: three drill holes within average distance ≤ 25 m;  Indicated Resource: three drill holes within average distance ≤ 50 m;  Inferred Resource: three drill holes within average distance ≤ 100 m; The methodology equates approximately to a drill spacing of 40 m for Measured Resource and 80 m for Indicated Resource. No single unsupported drill hole was included in the classifications. Confidence in geological continuity was supported up to 100 m past the last drill hole. Qualitative risk classification of resource confidence was made by NBG using a combination of data density, spatial arrangement of the data and the quality of estimation and geological interpretation. The risk classification approach is used to risk-rank the blocks in the resource model. This incorporates five elements:

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 Drill density;  Data quality;  Geological knowledge;  Quality of the geological modeling;  Estimation confidence. The end result is a high (3), moderate (2), and low (1) risk assignment coded as a separate item in the block model.

14.13 Reasonable Prospects of Eventual Economic Extraction Reasonable prospects of eventual economic extraction was addressed by applying a resource shell defined using a pit optimization to identify mineralization that could reasonably be economically extracted, using a cut-off value based on an NSR approach. The NSR value of AU$15.43/t was applied inside a pit shell determined using a gold price of US$1,400/oz or AU$1,750/oz, and a copper price of US$3.25/lb or AU$4.00/lb. The conceptual parameters for the pit and the NSR derivation are presented in Table 14-2. Mineral Resources for the Project included application of the NSR-based cut-off grade that incorporated mining and recovery parameters, and constraint of the Mineral Resources to a pit shell based on commodity prices. While silver is included in the concentrate calculation at the rate of AU$21/oz, it is not estimated in the Mineral Resource estimate and does not make a material contribution to the economics of the Project. Table 14-2 presents the economic parameters and cut-offs used for the 2018 Mineral Resource pit optimization. Table 14-2: 2018 Mineral Resources Economic Parameters and Cut-offs Cut-off for 2018 Mineral Resource Units At Cost Basis of 2019BP Gold Price AU$/oz 1,750 Copper Price AU$/lb 4.00 Exchange Rate US$0.80 = AU$1.00 Gold Royalty % 2.5 Copper Royalty % 5.0 Mill Throughput Mtpa 40.5 Mill Recovery Gold (Average recovery % at LOM grade 0.67 g/t Au) % 83 Mill Recovery Copper (Average recovery % at LOM grade 0.1% Cu) % 78 Base Processing Cost-without rehandle AU$/t milled 9.71 Sustaining Capital (Plant and G&A) AU$/t mined 1.14 incl. Capital Recovery Factor (CRF) = 1.10 Site and Regional G&A (exclude CAPEX) AU$/t milled 2.09 Incremental Ore, Resource Conversions and Closure (LOM-FASB) AU$/t milled 0.22 Breakeven Mill Cut-off (BMCO) AU$/t milled 13.17

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Cut-off for 2018 Mineral Resource Units At Cost Basis of 2019BP Stockpile Rehandling AU$/t rhdled 1.48 Recoveries Degradation AU$/t milled 0.79 Breakeven Stockpile Cut-off (BSCO) AU$/t milled 15.43

LOM Operating Mining Cost AU$/t mined 4.22 Mining Capital AU$/t mined 0.62 incl. CRF = 1.11

Note: 2019BP = 2019 Business Plan

14.14 Mineral Resource Statement Mineral Resources are exclusive of Mineral Reserves and are reported at gold price of US$1,400 or AU$1,750/oz, and a copper price of US$3.25/lb or AU$4.00/lb on a 100%, basis with an effective date of 31 December 2018. The Mineral Resource estimate was prepared by Mr. Rohan McCormack, MAIG, Senior Resource Geologist and Newmont employee, under the supervision of the QP. Newmont cautions that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. Table 14-3 presents the total gold and copper Mineral Resource for the Project.

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Table 14-3: Mineral Resource – Gold and Copper at the Effective Date of 31 December 2018

Measured Resource Indicated Resource Measured + Indicated Resource Inferred Resource Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal (kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt) Boddington OP 95,200 0.55 1,680 0.11 100 253,800 0.55 4,510 0.12 300 349,000 0.55 6,190 0.12 400 5,100 0.49 80 0.09 0 Total 95,200 0.55 1,680 0.11 100 253,800 0.55 4,510 0.12 300 349,000 0.55 6,190 0.12 400 5,100 0.49 80 0.09 0

Notes to accompany the Mineral Resource tables:  OP = open pit;  Mineral Resources have an effective date of 31 December 2018;  Mineral Resources are reported exclusive of Mineral Reserves, and are reported on a 100% basis;  Mineral Resources are estimated within designed pits generated based on optimised pit shell using Whittle®;  Mineral Resources are reported using a gold price of US$1,400/oz, and a copper price of US$3.25/lb, equivalent to AU$1,750/oz, and AU$4.00/lb at an exchange rate of US$0.80 = AU$1.00;  Mineral Resources contain material that is above a NSR cut-off of AU$15.43/t within the Mineral Resource ultimate pit design, exclusive of Mineral Reserves. Table 14-2 presents the economic parameters and cut-offs used for the 2018 Mineral Resource pit optimization;  Royalties are considered in the NSR cut-off determination;  Metallurgical recovery is considered in the NSR cut-off for in situ and stockpiled ore. Since the metallurgical recovery of stockpiled ore degrades over time, weighted average LOM recovery (refer to Section 13.4) is calculated using a combination of in situ recovery less 3% and direct ROM feed in situ recovery;  Tonnage and grade measurements are in metric units. Gold ounces are reported as troy ounces. Tonnages include allowances for losses resulting from mining methods. Tonnages are rounded to the nearest 100,000 tonnes. Ounces are estimates of metal contained in the Mineral Resource and do not include allowances for processing losses. Gold ounces are rounded to the nearest 10,000 ounces and copper metal tonnage is rounded to the nearest 10,000 tonnes;  Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grade and contained metal content. Copper metal tonnage for Inferred Resource is less than 5 kt and hence is not presented in the table to maintain consistency in rounding.

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14.15 Factors That May Affect the Mineral Resource Estimate Factors which may affect the Mineral Resource estimates include:  Metal price assumptions;  Changes to the assumptions used to generate the NSR cut-off;  Changes to design parameter assumptions that pertain to the conceptual pit shell design that constrain the Mineral Resources, including changes to geotechnical, mining and metallurgical recovery assumptions, and changes to royalties levied and any other relevant parameters that are included in and impact the NSR cut-off determination;  Changes in interpretations of mineralization geometry and continuity of mineralization zones;  Changes to the dilution skin percentages used for large dolerite dykes;  Assumptions as to the continued ability to access the site, retain mineral and surface rights titles, maintain the operation within environmental and other regulatory permits, and retain the social licence to operate.

14.16 Comments on Mineral Resource Estimates In the opinion of the QP, the Mineral Resource estimate for the Project conforms to industry standard practices and satisfies the requirements of the CIM Definition Standards.

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15.0 MINERAL RESERVES ESTIMATES

15.1 Mineral Reserves Schedule The Mineral Reserves LOM schedule was planned using a similar mining sequence to the 2019 Business Plan. The mine schedule was developed using MineSight® reserve engine and spreadsheet-based scheduling tool. The mine plan is based on a 40.5 Mtpa mill throughput. The schedule was developed at an NSR cut-off of AU$15.32/t, incorporating the processing cost, metallurgical recovery, incremental ore mining costs, process sustaining capital and tailings dam related rehabilitation costs. The net revenue calculation assumes a gold price of US$1,200/oz or AU$1,600/oz, and a copper price of US$2.50/lb or AU$3.35/lb. The assumed exchange rate for Mineral Reserves was 0.75 US$:AU$. NBG updates the Strategic Mine Plan each year in preparation for the business plan. All aspects of the plan, including pit stage design and sequencing, cutoff optimization and dump and stockpiling strategy are reviewed. The mill processes higher grade ores delivered from the mine at an elevated cut-off. The ore between the elevated cut-off and the marginal cut-off is stockpiled for later processing at the end of the mine life. Most of the ore is planned to be directly fed to the process plant, however some re-handle is required. Direct feeding to the crusher is constrained by the ore presentation in the pit and crusher availability. Some higher-grade ore is stockpiled and fed back to crusher when required. Approximately 50% of feed is re-handle material from the stockpiles.

15.2 Mineral Reserves Statement Mineral Reserves for the Project are presented in Table 15-1 for gold and copper and are reported using the CIM Definition Standards. Mineral Reserves have an effective date of 31 December 2018 and are reported to a gold price of US$1,200/oz or AU$1,600/oz, and a copper price of US$2.50/lb or AU$3.35/lb. The Mineral Reserves estimate was prepared by Mr. Eka Setiawan Lim, RM-SME, Mine Engineering Superintendent and Newmont employee, under the supervision of the QP. Within the Mineral Reserves pits, there is approximately 720 kt that is currently classified as Inferred Mineral Resources. The Mineral Reserves LOM mine schedule does not include the Inferred Resource material in the production plan and the Inferred Resource material is treated as waste material in the mine plan.

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Table 15-1: Mineral Reserves – Gold and Copper at the Effective Date of 31 December 2018

Proven Reserves Probable Reserves Proven + Probable Reserves Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal Tonnage Au Grade Au Metal Cu Grade Cu Metal

(kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt) (kt) (g/t) (koz) (%) (kt)

Boddington OP 240,400 0.71 5,520 0.09 230 240,300 0.71 5,470 0.11 260 480,700 0.71 10,990 0.10 490

Stockpiles 6,900 0.67 150 0.08 0 86,100 0.44 1,210 0.08 70 93,000 0.45 1,360 0.08 70

Total 247,300 0.71 5,670 0.09 230 326,400 0.64 6,680 0.10 330 573,700 0.67 12,350 0.10 560

Notes to accompany the Mineral Reserves tables:  OP = open pit;  Mineral Reserves have an effective date of 31 December 2018;  Mineral Reserves are reported on a 100% basis;  Mineral Reserves are estimated within designed pits generated based on optimised pit shell using Whittle®;  Mineral Reserves are reported to a gold price of US$1,200/oz, and a copper price of US$2.50/lb, equivalent to AU$1,600/oz, and AU$3.35/lb at an exchange rate of US$0.75 = AU$1.00;  Mineral Reserves contain material that is above a NSR cut-off of AU$15.32/t within the Mineral Reserves ultimate pit design. Table 15-5 presents the economic parameters and cut-offs used for the 2018 Mineral Reserves pit optimization;  Royalties are considered in the NSR cut-off determination;  Metallurgical recovery is considered in the NSR cut-off for in situ and stockpiled ore. Since the metallurgical recovery of stockpiled ore degrades over time, weighted average LOM recovery (refer to Section 13.4) is calculated using a combination of in situ recovery less 3% and direct ROM feed in situ recovery;  Tonnage and grade measurements are in metric units. Gold ounces are reported as troy ounces. Tonnages include allowances for losses resulting from mining methods. Tonnages are rounded to the nearest 100,000 tonnes. Ounces are estimates of metal contained in the Mineral Reserves and do not include allowances for processing losses. Contained gold ounces are rounded to the nearest 10,000 ounces and copper metal tonnage is rounded to the nearest 10,000 tonnes;  Rounding of tonnes as required by reporting guidelines may result in apparent summation differences between tonnes, grade and contained metal content. Copper metal tonnage for Proven Reserves is less than 5 kt and hence is not presented in the table to maintain consistency in rounding;  The Mineral Reserves are forward-looking information and actual results may vary.

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15.3 Factors That May Affect the Mineral Reserves Estimate While the assumptions in the Mineral Reserves estimates are appropriate to the date of estimation, areas of uncertainty that may materially impact the Mineral Reserves estimates include changes to long-term metal price assumptions, and changes to input cost assumptions such as consumables, labor costs, or taxation rates.

15.4 Caution Regarding Forward-Looking Information The Mineral Reserves are forward-looking and actual results may vary. The risks regarding Mineral Reserves are summarized in the Report (refer to Section 15.3 and Section 25.0). The assumptions used in the Mineral Reserves estimates are summarized in the footnotes of the Mineral Reserves table, and in Section 15.0.

15.5 Pit Optimization Lerchs-Grossman (Whittle®) pit optimisations were run using the Mineral Reserves economic inputs and optimum shells were analysed and used to assess the optimality of the 2019 Business Plan (2019BP) pits. The Mineral Reserves economics and costs guidance is presented in Table 15-2. Table 15-2: Mineral Reserves Economic Assumptions At Cost Basis of Units 2019BP Concentrate Related Inputs Units Value

Gold Prices AU$/oz 1,600 Payable Au % 97.50%

Copper Price AU$/lb 3.35 Payable Cu % 92.00%

Silver Price AU$/oz 21.00 Conc. Dust Loss % 0.00%

Base Processing Cost-without rehandle AU$/t milled 9.71 Concentrate Moisture weight % 9.50%

Sustaining Capital Recovery Factor CRF 1.10 Cu TC US$/dmt 85

Sustaining Capital (Plant and G&A) AU$/t milled 1.14 Cu RC US$/lb 0.085

Site G&A (exclude CAPEX) AU$/t milled 1.37 Au RC US$/oz 5.5

Regional G&A Back charge AU$/t milled 0.72 Transport to Port AU$/wmt 27.46

Incremental Closure (LOM - FASB) AU$/t milled 0.23 Transport from Port & Others AU$/wmt 86.59

Increment Ore Mining Cost AU$/t milled -0.02 Penalty AU$/dmt 4.62

Resource Conversion Cost ($4/oz) AU$/t milled – Gold Portion into Concentrate % 74.90%

Stockpile Rehandling AU$/t rhdled 1.48

Recoveries Degradation AU$/t milled 0.68 Royalty Units Value Breakeven Stockpile Cut-off (BSCO) AU$/t ore 15.32 (apply to payable metal)

Au Royalty % 2.5

Cu Royalty % 5.0

While silver is included in the concentrate calculation, it is not estimated in the Mineral Reserve estimate and does not make a material contribution to the economics of the Project. 15.5.1 Metallurgical Recovery Assumptions Metallurgical recovery for gold varies by gold head grade and across the eight ore types in the open pits. The average recovery for gold in the LOMP used to support the

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economic evaluation is 83%. Metallurgical recovery for copper varies by copper head grade in the open pits. The average metallurgical recovery of copper in the LOMP used to support the economic evaluation is 78%. A revision of the plant performance recovery functions was undertaken in June 2014 (Petrucci, 2014) in the lead up to the 2014 Mineral Resource and Mineral Reserves development. The methodology outlined in the original report (Metallurgical Performance Models; Mark Roberts, 2012) was again used with more recent plant data. A minor update incorporating process improvements was conducted in 2015 (Petrucci, 2015). A further update was conducted on the copper model in 2016 (Sandi, 2016). The gold model has not changed since 2015. Changes to the copper recovery model were due to the plant outperforming the existing model based on metallurgical improvements in the processing plant. The changes contributing to the plant outperforming the model included:  Reduced grinding media size, which has led to improved mineral liberation through changed grinding conditions;  Reduced flotation pH, which has led to less gold suppression through hydroxide coatings, and improved cleaning and maintenance of flotation launder lips increasing concentrate mass pull and reducing froth transport times.

Figure 15-1: Actual versus Model Recovery Variance Figure 15-1 compares the actual plant performance with the 2016 model prediction The gold and copper metallurgical recovery functions are presented in Table 15-3.

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Table 15-3: Gold and Copper Metallurgical Recovery Functions

Ore Zone Copper Recovery Overall Gold Recovery Central Diorites [955.668 x Cu HD - (2.607)] / [1180 x Cu HD] [105300 x AU HD - (4306.703)] / [1180 x AU HD] South Zone [955.668 x Cu HD - (2.607)] / [1180 x Cu HD] [105300 x AU HD - (7905.813)] / [1180 x AU HD] Southern Volcanics [955.668 x Cu HD - (2.607)] / [1180 x Cu HD] [105300 x AU HD - (4512.531)] / [1180 x AU HD] Pipeline [955.668 x Cu HD - (2.607)] / [1180 x Cu HD] [105300 x AU HD - (7262.636)] / [1180 x AU HD] Southern Diorite Deeps [955.668 x Cu HD - (2.607)] / [1180 x Cu HD] [105300 x AU HD - (2643.329)] / [1180 x AU HD] Blackbutt [955.668 x Cu HD - (2.607)] / [1180 x Cu HD] [105300 x AU HD - (5283.721)] / [1180 x AU HD] Northern Diorites [955.668 x Cu HD - (2.607)] / [1180 x Cu HD] [105300 x AU HD - (5119.261)] / [1180 x AU HD] Far North [955.668 x Cu HD - (2.607)] / [1180 x Cu HD] [105300 x AU HD - (5336.682)] / [1180 x AU HD]

15.5.2 Geotechnical Assumptions and Pit Slope Configuration Several geotechnical and hydrological studies have been completed to support mining, feasibility, and environmental reports for Boddington. The geotechnical model for the Wandoo deposit has been defined by geotechnical drilling and logging, laboratory test work, rock mass classification, structural analysis, and stability modeling. The hydrological model is based on a three-dimensional flow model, historic pumping rates from the Jarrah Pit, and drill data. Coffey Geosciences were the engineers of record for the feasibility study and performed the final pit design stability review. The design considerations for the pit at the time of the feasibility study are presented in Table 15-4.

Table 15-4: Feasibility Study Pit Slope Design Parameters

Design Criteria Oxide Oxide Bedrock Bench Height (m) 12 12 Batter Height (m) 12 36 Batter Angle (deg) 47 Up to 85 Berm Width (m) 15 11 to 15 Inter-ramp Slope Angle (deg) 25 64 Max. Slope Height at Inter-ramp Slope Angle (m) n/a 180 Catch Berm – Base of Oxide (m) 20 n/a Floating Catch Berm Width at Max Inter-ramp Slope Angle (m) n/a 30

A geotechnical study was completed by SRK Consulting in 2011 to 2012 (SRK, 2012a) to review design parameters of the oxide pit slopes. This study involved drilling several holes, field testing and sampling, laboratory testing, installation of vibrating wire piezometers for groundwater monitoring and slope stability analyses. Results of this study recommended that the Oxide slopes could be designed at a 30o inter-ramp slope angle (IRSA) with the possible exception in areas where relict structures and groundwater may control the design. Historically, a challenge of the hard rock slope design at Boddington has been the inability to consistently achieve the required catch berm width to retain rock falls. A Geotechnical Assessment of fresh rock was conducted by Snowden in 2012 (Snowden, 2012b) and a hard rock study was completed internally in 2014 (Newmont, 2014b). In 2014, a 36 m triple bench (3 x 12 m) design was implemented in all fresh hard rock for an inter-ramp angle of 59º. The successful implementation of this design resulted in crest loss reduced from an average of 4.3 m to 2.3 m in Wandoo South, and in the Wandoo North Andesite/Diorite host rock to 2.4 m.

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The pit wall configuration was updated in 2016. The oxide wall remains unchanged with a 30° IRSA, but the hard rock IRSA has been increased to 60.8°in some areas by steepening the top flitch from 70° to 75° following a study completed in October 2016 (Newmont, 2016) based on trials throughout 2015 and 2016. The rock mass at NBG has been characterized and grouped into geotechnical domains for design purposes using geotechnical and hydrogeological models as well as operational experience. A geotechnical domain is an area or region with similar rock mass conditions (structures and fabric, rock strength, fracture frequency, hydrogeology) that is expected to have the same or similar response to excavation, blasting and slope geometry. Figure 15-2 and Figure 15-3 show the geotechnical domains for both North Pit and South Pit.

Note: Figure aligned to Mine Grid Figure 15-2: Geotechnical Domains – North Pit

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Note: Figure aligned to Mine Grid Figure 15-3: Geotechnical Domains – South Pit

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The recommended slope profiles for the domains in the North Pit and South Pit areas are presented in Figure 15-4 and Figure 15-5.

Figure 15-4: Recommended Slope Profiles for North Pit Area

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Figure 15-5: Recommended Slope Profiles for South Pit Area

15.5.3 Dilution and Mining Losses Dilution is applied to the model using a combination of direct dilution of the block grade values (dilution) and as ore volume reduction (ore loss), which reduces the portion of the block available as ore. The model has small dolerite volumes which are added to the grade variables as dilution, the width of these structures is narrower than the SMU. The portion of each parent block within the small dolerite solids is evaluated and the block grade reduced by this portion. The resulting diluted grade is evaluated the same as any other ore block. This process is unchanged from that employed in the Feasibility Study. The large dolerite volume is applied to the block as a waste portion, this portion is increased by a set amount which represents ore loss against the dolerite contact, the width of these structures is equal to or greater than the SMU; it does not change the interpolated grade but results in the available ore portion being reduced. This portion

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is evaluated as a percentage of the block within the large dolerite solid. This percentage is further adjusted to include a dilution skin/ ore loss by adding a nominal amount to any block. The blocks containing more than 50% oxide material are classified as waste due to the oxide material and have the grade set to zero.

15.5.4 NSR Cut-off The Project’s cut-off is defined by a revenue due to it being a polymetallic deposit with two product streams, gold doré and copper concentrate i.e. the block revenue is calculated on an NSR basis, which is the dollar return expected from the sale of the concentrate produced from a tonne of in-situ material. NSR calculations account for metallurgical recovery, concentrate shipping and smelting and refining costs and royalties. Mineral Reserves are reported within a detailed pit design. Only Measured and Indicated Mineral Resources within the Mineral Reserves pit design are directly converted to Proven and Probable Mineral Reserves respectively. The Mineral Reserves are reported within the optimized pit design based on the following price inputs; US$1,200/oz or AU$1,600/oz for gold and US$2.50/lb or AU$3.35/lb for copper metal prices, providing an NSR cut-off of AU$15.32/t (refer to Table 15-2) NBG follows the Newmont corporate guideline in calculating the NSR cut-off. Table 15-5 presents the economic parameters and cut-offs used for the 2018 Mineral Reserves pit optimization.

Table 15-5: 2018 Mineral Reserves Economic Parameters and Cut-offs

Cut-off for 2018 Mineral Reserves Units At Cost Basis of 2019BP Gold Price AU$/oz 1,600 Copper Price AU$/lb 3.35 Exchange Rate US$0.75 = AU$1.00 Gold Royalty % 2.5 Copper Royalty % 5.0 Mill Throughput Mtpa 40.5 Mill Recovery Gold (Average recovery % at LOM grade 0.67 g/t Au) % 83 Mill Recovery Copper (Average recovery % at LOM grade 0.1% Cu) % 78 Base Processing Cost-without rehandle AU$/t milled 9.71 Sustaining Capital (Plant and G&A) AU$/t mined 1.14 incl. CRF = 1.10 Site and Regional G&A (exclude CAPEX) AU$/t milled 2.09 Incremental Ore, Resource Conversions and Closure (LOM-FASB) AU$/t milled 0.21 Breakeven Mill Cut-off (BMCO) AU$/t milled 13.16

Stockpile Rehandling AU$/t rhdled 1.48 Recoveries Degradation AU$/t milled 0.68 Breakeven Stockpile Cut-off (BSCO) AU$/t milled 15.32

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Cut-off for 2018 Mineral Reserves Units At Cost Basis of 2019BP LOM Operating Mining Cost AU$/t mined 4.22 Mining Capital AU$/t mined 0.62 Incl. CRF = 1.11

Note: 2019BP = 2019 Business Plan

Key parameters in the NSR cut-off calculation are as follows:  Mineral Reserves price assumption is US$1,200/oz or AU$1,600/oz gold and US$2.50/lb or AU$3.35/lb copper. Mineral Resource price assumption is US$1,400 or AU$1,750/oz gold and US$3.20 or AU$4.00/lb copper;  Unit costs used for the cut-off are based on the 2019BP costs with costs escalated to 2019 dollars. CRF was also included both in Economic Tests and cut-offs calculation. The CRF is required to reflect the interest% pre-tax return on an investment over a period;  Metallurgical recovery is considered in the NSR cut-off for in situ and stockpiled ore. Since the metallurgical recovery of stockpiled ore degrades over time, weighted average LOM recovery (refer to Section 13.4) is calculated using a combination of in situ recovery less 3% and direct ROM feed in situ recovery.

15.6 Pit Optimisation Lerch-Grossman (Whittle®) optimisations were run in order to obtain optimum cones to be used as the basis for pit design.

In addition to the inputs above, the Mineral Reserves optimisation used a Cost Adjustment Factor (CAF) to reflect the mining cost difference between pit region and bench. CAF is determined by the distance from pit to destination and the elevation of the bench. Figure 15-6 and Figure 15-7 display the optimality of the selected ultimate pits for North Pit Mineral Reserves and South Pit Mineral Reserves respectively. Figure 15-6 highlights that the pit size of North Pit (79.2 Mt ore) is significantly smaller than the optimum Mineral Reserves Shell (Shell#27) [167 Mt ore]. This is due to the optimum shell expanding to the east and west beyond the pit design walls.

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NP total reserve 95Mt @ $15.19/t)

Figure 15-6: Mineral Reserves North Pit Optimisation Curves In Figure 15-7, the pit size of South Pit (401 Mt ore) is quite close to the size of optimum Mineral Reserves Shell #27 (430 Mt ore). The difference is mainly due to the exclusion of some of the optimum shell in the west wall from Mineral Reserves pit design where the mining width is not enough to design a mineable layback.

Figure 15-7: Mineral Reserves South Pit Optimisation Curves

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Aligning with Newmont 2018 Mineral Reserves Guidelines, all the newly redesigned cutbacks were tested using the Mineral Reserves economic guidance for the economic viability. Economic tests were performed on pit stages using 2019BP unit costs at Mineral Reserves prices. Each pit stage was tested independently on a stand-alone basis. A pit stage needs to satisfy the discounted cash flow and be positive to be eligible for Mineral Reserves reporting and has a positive undiscounted cash flow to be eligible for Mineral Resource reporting. Pit stages that did not meet this hurdle were redesigned under the standard design criteria i.e. minimum operating width, access adequacy, etc.

15.7 Pit Designs The Mineral Reserves pit designs are full crest and toe detailed designs with final ramps based on the selected optimum Whittle® cones. Pit designs have honoured geotechnical guidelines with 15.2 m catch berms. Economic tests were done to all laybacks designed using 2019BP unit costs at the Mineral Reserves prices (US$1,200/oz or AU$1,600/oz gold and US$4.00/lb or AU$3.35/lb copper). The Mineral Reserves North Pit and South Pit positions along with the Mineral Resource expansion regions are presented in Figure 15-8. Mineral Reserves pits located at the north and south of the primary crusher conveyor lane. Figure 15-8 presents the Mineral Reserves pits to the north as blue, green and red colored solids and to the south as purple, orange, cyan, green, red and magenta colored solids.

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Note: Figure aligned to Mine Grid Figure 15-8: 2018 Mineral Reserves and Mineral Resource Pits

15.8 Comments on Mineral Reserves In the opinion of the QP, the Mineral Resources and Mineral Reserves have been performed using industry-accepted practices and conform to the requirements of the CIM Definition Standards. The Mineral Reserves are adequate to support mine planning.

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16.0 MINING METHODS Mine start-up required an 18-month mining ramp-up that included contractor pre-stripping between August 2006 and September 2007 of approximately 12 Mt of predominantly waste ahead of process plant commissioning. In October 2007, Owner-operated mining commenced, continuing the pre-stripping operation to access the bedrock mineralization. Plant start-up was achieved in September 2009. The Project was brought into full operational status in November 2009. A total of 912.7 Mt of material has been mined to the end of December 2018.

16.1 Description of Mining Method Boddington is mined by a conventional truck-and-shovel operation. Equipment is owner- operated and includes a large mining truck fleet (240-tonne class), electric rope shovels, support equipment, and drills. Mining is done predominantly on 12 m benches. The LOM plan currently envisages mining at an average rate of approximately 70 Mtpa for 13 years and peaking at 87 Mtpa in 2026 with a maximum rate of advance by pit stage of seven benches per annum and an average of five benches (60 m) per year. NBG currently expects that mine production life will extend into approximately year 2032 with material mined from the two open pit sources. Milling will cease in the same year after treatment of stockpiled ore. Approximately three-quarters of the mine fleet will be located in the higher-priority South Pit, mining an ore face and a waste push-back simultaneously. The remaining one-quarter of the fleet will be working the North Pit, mining a mixture of ore and waste faces. A layout of mining operations is presented as Figure 16-1, showing the locations of the pit layback/phase stages and the waste rock storage facilities.

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Note: Figure aligned to mine grid Figure 16-1: Layout of Project Mining Area

16.2 Hydrology Detailed engineering, procurement, and construction of the pit dewatering system Phase 1 to 3 have been completed by Transeng Design and Construct. Sinclair Knight Mertz (SKM) completed a feasibility study update design optimization study. The designs of pit dewatering system Phase 1 to 3 were based on the findings of the SKM optimization study. Local and regional monitoring bores and vibrating wire piezometers (VWP) were progressively installed during 2007 under the supervision of Golder. Three deep bores, consisting of eleven VWPs commenced monitoring of groundwater heads in basement rocks in 2007.

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Detailed engineering, procurement, and construction of the pit dewatering system Phase 4 to 9 has been completed by NBG, with progressive upgrades to system performance and capacity. The ongoing hydrogeological study indicates that the pit dewatering system will continuously receive large volumes of groundwater and surface runoff over the life of mine. The water management strategy is to maximize the use of the groundwater within the process plant and the loss of excess water by evaporation from the residue disposal area. It is expected that this will provide satisfactory water management for much of the mine life. Surface water management structures have been designed to:  Collect and store water generated on the mine site for use in the process plant;  Separate clean (non-impacted) and potentially impacted water sources to provide flexibility in the site global water management strategy;  Minimize the movement of sediment loads;  Limit surface erosion and maintain the integrity and effectiveness of the post-closure store and release dump cover. The LOM pit dewatering plan strategically upgrades the In-Pit Sump Pump system (passive dewatering) and dewatering-bores (active dewatering) to cope with various operational needs. Groundwater extracted through pit dewatering will be contained on site and used by the process plant. The site water management plan accounts for the water from Active and Passive Pit Dewatering as forecasted water sources. A provision has been made for excess surface water to be captured in water storage reservoirs.

16.3 Production Schedule A mine schedule based on the 2019 Mineral Reserves pit was developed using the existing mine planning tools with same parameter assumptions used in the 2019BP for the mine equipment and productivity assumptions and mill throughput, also similar mine and dumping sequences. High grade copper ores from S05A layback sourced from the Southern Diorite Deep ore zone will be mined during 2021 to 2023. Processing the higher feed grades will have an inverse impact on metal recovery. In this case, the milling schedule has applied a copper grade limit of 0.15% by blending higher-grade ores with lower-grade stockpiled ores. Table 16-1 presents the Mineral Reserves mine schedule by pit.

Table 16-1: 2018 Mineral Reserves Mine Schedule by Pits (kt)

Note: Tonnages presented in table represent ex-pit mining only

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16.4 Drilling and Blasting All drilling operations are performed by Newmont with Newmont owned rigs that are maintained by Atlas Copco under a MARC agreement. Production rigs include 7 x Atlas Copco PV231, 3 x Atlas Copco DML, all configured for DTH mode. Wall control rigs; 5 x Atlas Copco D65, all configured for DTH mode. Vertical production holes are drilled vertical (229 mm diameter) with a nominal bench height of 12 m, with 1.5 m subdrill for a 13.5 m overall hole length. Wall control holes are 127 mm diameter for batters and buffers and 115 mm diameter for pre-split. A variety of angles and lengths are used to suit various geotechnical based ground control domains. All rigs are GPS controlled for hole positioning and on-board telemetry systems for recording depth, angle, drill time etc. Production blasting utilises augered and pumped Heavy ANFO blends, with a nominal approximately 500 kg per hole. Crushed aggregate, produced on site, is used as stemming for the top 3.7 m of each hole. Electronic detonators are used for ore blasts whereas nonel detonators are used on waste blasts. Nominal Powder Factors on production blasts vary from 1.18 to 1.40 kg/m3. The wall control configuration features a triple bench design; a 70° or 75° angled batter on the top bench is followed by a single pass 24 m vertical pre-split for the second and third benches. All first bench blasts are fully free faced. On average, the site fires blasts approximately twice per week. All blasts are monitored for airblast overpressure and ground vibration to ensure adherence to applicable environmental limits.

16.5 Mining Equipment The equipment assumptions are based on the business plan 2016BP. The equipment list is as followings:  Major Equipment: o 2 x Bucyrus 7495; o 1 x Hydraulic Shovel; o 1 x Terex RH340; o 39 x Cat 793D/F; o 3 x Cat 785C; o 3 x Atlas DML; o 7 x Atlas PV235; o 4 x Atlas D65.  Ancillary Equipment: o 4 x Cat D11R/T; o 2 x Cat D10T; o 4 x Cat 854G WDZ; o 1 x Cat 834H WDZ; o 3 x 24H/M; o 1 x 16H; o 3 x Hitachi EX3600/2600;

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o 3 x Cat 994H/F; o 3 x Cat 785D/C WC.

16.6 Mine Plan Considerations Newmont has regularly undertaken and will continue to undertake as part of its normal course of business operations, reviews of the mine plan and consideration of alternatives to and variations within the plan. Alternative scenarios and reviews are based on ongoing or future mining considerations, evaluation of different potential input factors and assumptions, and requests made by Project staff and Newmont Corporate. Such iterations can include where appropriate, but are not limited to:  Changes to Mineral Resource and/or Mineral Reserves estimation methodologies;  Changes to dilution and reconciliation strategies;  Changes to metal price assumptions;  Changes in allocations of planned drilling, or drilling locations, that can be used to increase Mineral Resource confidence and support conversion of Mineral Resources to Mineral Reserves;  Changes to deposit sequencing;  Changes to production rates;  Changes in mining equipment strategies;  Alternate pit configurations, including laybacks or pit wall slope changes;  Changes to geotechnical or hydrological assumptions;  Changes in short-term production;  Mill throughput reviews and potential mill modifications;  Process flowsheet modifications and potential recovery improvements;  Stockpile throughput, allocations, and planned depletion rates;  Optimization of cash flows and review of different cash flow scenarios;  Changes to allocations of capital expenditures to different years within the mine plan;  Modifications to sustaining capital and operating cost assumptions;  Changes to accounting and taxation assumptions.

16.7 Comments on Mining Methods In the opinion of the QP, the mine plan is achievable, and the equipment fleet is adequate to support the production schedule.

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17.0 RECOVERY METHODS

17.1 Process Flow Sheet Based on feasibility assessments, the selected process consists of primary crushing, closed circuit secondary and HPGR tertiary crushing, ball milling, and hydrocyclone classification to generate a milled product with a P80 of 150 μm at a slurry density of 35 to 38% solids. Flash flotation facilities are included to treat a portion of the mill discharge stream, and trials are planned to assess the economics of bringing the flash flotation circuit into operation. There are no plans to bring the gravity circuit into operation. Cyclone overflow from the mill circuit is treated in a flotation circuit that produces a copper– gold concentrate for export. Rougher and scavenger flotation concentrates are reground and cleaned to achieve an acceptable final concentrate grade. Concentrate is thickened and filtered before being trucked to the port of Bunbury. The cleaner scavenger tailings stream is thickened and leached under elevated cyanide levels. Scavenger tailings are thickened and leached in a conventional leach/adsorption circuit. Leached slurry from the cleaner scavenger tailings leach circuit is delivered to the scavenger tailings circuit for combined recovery of gold.

Leach residue is pumped to the residue disposal area, and residual CNwad is maintained below a targeted level by a Caro’s acid cyanide destruction plant. This facility can treat the following streams:  Decant water returning to the plant so that cyanide levels do not inhibit flotation;

 Decant water recycling to the decant pond to maintain CNwad levels in the pond at an average of 30 ppm and a not-to-exceed level of 50 ppm;  Residue slurry from the plant to protect the decant pond from excursions caused by short- term variability in the copper head grade. The carbon from the scavenger tailings adsorption circuit is treated by conventional split- Anglo American Research Laboratory (AARL) method elution and reactivated in horizontal reactivation kilns. Gold recovery from the eluate is by electrowinning, cathode sludge filtration and drying, and smelting. A flowsheet is presented in Figure 17-1. Production from the plant was initially estimated using utilization estimates of 75% for the primary crushing circuit, 88% for the secondary crushing circuit, and 95% for the milling circuit.

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Figure 17-1: Overall Process Flowsheet

17.2 Plant Design The Project process plant incorporates the following major equipment:  2.3 km overland conveyor;  Two ROM stockpiles, two medium-grade stockpiles, three waste dump areas;  Two 60-110 primary crushers;  Six MP1000 secondary crushers;  Four 3.6 m x 8.5 m coarse screens;  4 x 5.6 MW HPGR capacity (8,000 tph);

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 Four 15 MW ball mills (7.9 m x 13.4 m);  4 x Flash flotation SK1800 Outotec cells (to be decommissioned);  Roughers/scavengers 3 x parallel trains consisting of 2 x 150 m3 and 6 x 200 m3 Outotec tank cells;  19 m diameter Outotec regrind high-rate thickener;  Regrind mills 2 x Vtm 1250 950 kW Vertimills;  First cleaners/cleaner scavengers 9 x 100 m3 Outotec tank cells;  Coarse cleaners 3 x 30 m3 Outotec tank cells;  Second and third cleaners 5 x 8 m3 Outotec U-shape cells;  27 m diameter Outotec high-rate concentrate thickener;  108 m2 area Larox concentrate pressure filter;  19 m diameter Outotec cleaner scavenger tails high- rate thickener;  74 m diameter Outotec scavenger tails high-rate thickener;  2 x 450 m3 and 7 x 175 m3 cleaner scavenger tails (CSt) leach tanks;  Carbon-in-leach, 2 x parallel trains (each 12 leach tanks); total 60,500 m3 capacity. 17.2.1 Coarse Crushing (in mining area) The mine haul trucks dump ore to two primary crushers (60/113 MK-II gyratory crusher). Crushed ore is transferred via an overland conveyor to a 230,000 t capacity (40,000 t live capacity) stockpile adjacent to the processing plant. Dozers operating on the coarse ore stockpile can increase the total storage capacity up to 400,000 t. 17.2.2 Fine Crushing and Screening (process plant) Three apron feeders reclaim ore from beneath the coarse ore stockpile and delivers the ore to the secondary crusher feed conveyor and the six secondary crushers (MP1000 cone crusher, five original with a sixth crusher installed in late 2010), which are operated in closed circuit with four coarse screens (three original with a fourth coarse screen installed in 2010). Oversize material returns to secondary crushing and the fine material reports to the tertiary crushing plant that consists of four HPGRs. The tertiary product is stored in a 20,000 t fine ore bin ahead of the ball milling circuit. Fine ore is reclaimed from the bin via eight reclaim belt feeders (two per four parallel milling trains) and delivered to the fine screens ahead of ball milling. Each line consists of two screening units. Undersize material from the screens reports to one of four cyclone feed hoppers and the oversize returns to the HPGRs for additional crushing. Cyclone clusters classify the finely-crushed particles, with the finer cyclone overflow material (80% passing 150 µm) reporting to the flotation distribution box and the coarse cyclone underflow material to a split between the ball mills or flash flotation cell (for free gold recovery). Concentrate from the flash flotation cell feeds a gravity circuit where gold is concentrated prior to intensive cyanidation, electrowinning and smelting. The flash flotation circuit is currently not operating, due to water balance issues associated with the grinding circuit. Tails from the flash flotation cells are recycled back to the ball mills for further grinding. The target final product grind size from the milling circuit is 80%

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passing 150 µm, although increased throughput rates in 2011 and 2012 resulted in grind size P80 coarsening to 170 to 190 µm at times. Improved process control and increased ball mill operating power draws have reduced the grind size back to around 150 to 160 µm at the elevated throughput rates. 17.2.3 Gold and Copper Recovery to Concentrate The flotation distribution box transfers cyclone overflow product from each of the four parallel grinding lines in to three parallel trains comprising, eight-unit flotation cells. Concentrates produced by cells #1 and #2 report to the coarse cleaner cells for final cleaning, and product from the other cells reports to regrind thickening. The regrind plant consists of two Verti-mills (one duty and one standby) with product reporting, via cyclone clusters, to the cleaner flotation plant. The cleaner flotation facility has three sequential stages with final product being transferred to the concentrate thickener, then storage in two, 1,000 m3 tanks before being sent to the filtration plant. Concentrate is trucked to the port of Bunbury to be exported by sea. The cleaner circuit has a scavenger circuit consisting of six cells. Cleaner scavenger tailings are leached in a dedicated circuit of nine leach tanks with the product combining, for further leaching, with the tails from the eight-unit flotation cells. 17.2.4 Gold Smelting and Bullion Production Scavenger flotation tails report via a flotation tailings thickener to two five-unit leach tank trains. From the leach tanks, it transfers to two trains of seven carbon-in-leach (CIL) tanks and finally transferring to the residue disposal area (RDA). Activated carbon used in the leaching process to adsorb gold leached into solution reports to a two-train elution plant. Each elution plant has two elution columns, two heat exchangers and two elution heaters. The barren carbon product from the elution columns is transferred to the carbon reactivation plant before transfer back to the CIL train, and the gold solution transfers to the electrowinning circuit. The final product from the eight-unit electrowinning circuit reports to the gold furnace for smelting.

17.3 Plant Performance and Process Optimisation Mill throughput reached design in 2014 with 35.4 Mt milled at an overall availability of 85.2%. Major improvements in mill availability can be attributed to the restructuring of the shutdown plan with the elimination of many smaller ones and creating two larger, longer shuts. Significant improvements in equipment reliability that were initiated in prior years were also fully realized in 2014. Conveyor belt drive systems were upgraded in 2013 and early 2014, surge bins repaired, and the ball mill motors were modified. The average throughput rate for the plant increased to 4,750 tph, as a result of de-bottlenecking exercises completed in the fines crushing circuit as well as increasing the utilisation of power in the HPGR’s and ball mills. Total production to the end of December 2018 is presented in Table 17-1. Figure 17-2 presents the ramp up in mill processing rates from commissioning in 2009 to December 2018. There has clearly been a gradual improvement in performance since 2012 following the initial commissioning period, as plant reliability has improved, and equipment has been optimized. In the longer term, an averaged sustained production rate of 4,800 to 5,000 tph is expected. Work associated with optimizing fine screen panel aperture size and wear life has assisted with the throughput performance improvement as well as increasing the HPGR grinding pressure and process control advanced control systems (expert systems)

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developed in the crushing and milling circuits. Ball mill power draw increases in 2012 from 14.5 MW to 15.6 MW per mill and a decrease in grinding media size from 80 mm to 65 mm have also contributed to throughput gains and assisted with maintaining grind size targets within reasonable levels. Table 17-1: Processed 2018 Full Year Results

Parameter Unit Full Year to December 2018

Crushed Tonnes kt 40,257 Milled Tonnes kt 40,238 Mill Utilization % 88.4 Head Grade (gold) g/t 0.673 Head Grade (copper) ppm 1,180 Gold Recovery % 83.4 Copper Recovery % 79.8 Concentrate Tonnes t 244.339 Concentrate Grade Cu % 15.56 Metal Produced Au oz 728,757 Metal Produced Cu t 38,085

4,500,000 Monthly Tonnes Milled 2009 to 2018 4,000,000

3,500,000

3,000,000

2,500,000

2,000,000 Throughput(t) 1,500,000

1,000,000

500,000

0

Jul-12 Jul-17

Jan-10 Jan-15

Jun-10 Jun-15

Oct-13 Oct-18

Apr-11 Apr-16

Feb-17 Sep-11 Feb-12 Sep-16

Dec-12 Dec-17

Aug-09 Aug-14

Nov-10 Nov-15

Mar-14 May-18 May-13

Figure 17-2: Mill Monthly Throughput Since Start-up Figure 17-3 presents the ramp up in plant utilization from commissioning in 2009 to December 2018. Total plant utilization averaged 88.4% for 2018 and further work is ongoing to improve the overall plant utilization to 89%.

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120 Monthly Utilisation 2009 to 2018 100

80

60 Utilisation Utilisation (%) 40

20

0

Jul-12 Jul-17

Jan-10 Jan-15

Jun-10 Jun-15

Oct-13 Oct-18

Apr-11 Apr-16

Feb-12 Sep-11 Sep-16 Feb-17

Dec-12 Dec-17

Aug-09 Aug-14

Nov-10 Nov-15

Mar-14 May-18 May-13 Figure 17-3: Mill Monthly Utilisation Since Start-up The projected average mill utilisation and tonnes per operating hour (defined at the ball milling circuit) for 2018 to 2022 is presented in Table 17-2.

Table 17-2: Current and Projected Mill Utilization and Throughput

Actual Projected

2018 2019 2020 2021 2022 Average Mill Utilisation % 88.4 89.0 89.0 89.0 89.0 Tonnes per Operating Hour tpoh 5,198 5,198 5,198 5,198 5,198 Annual Milled Tonnes Mtpa 40.2 40.5 40.6 40.5 40.5

Figure 17-4 presents the average overall gold recovery and split between concentrate and leach since 2009. There was a slight reduction in gold recovery since 2010, which is mostly due to higher throughput rates achieved, and slightly lower average head grades; however gold recovery has consistently remained near 80% in recent years. An improvement in performance in late 2013 back towards 82% is noted. Gold recovery has been the subject of several continuous improvement projects since 2013. Figure 17-5 presents a similar performance for copper recovery, although in general copper recovery was lower than expected at between 75% and 77% during 2012 to 2014, there has been a notable improvement back to between 78% and 80% since then. Figure 17-6 presents the copper concentrate grade which is consistent at around 17% copper. Coarse particle flotation and the reduction of slimes losses are key areas for further investigation and opportunities for improvement in the copper recovery circuit.

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100% Gold Recovery 2009 to 2018 90%

80%

70%

60%

50%

Recovery(%) 40%

30%

20%

10%

0%

Jul-12 Jul-17

Jan-10 Jan-15

Jun-10 Jun-15

Oct-18 Oct-13

Apr-11 Apr-16

Sep-11 Feb-12 Sep-16 Feb-17

Dec-12 Dec-17

Aug-09 Aug-14

Nov-10 Nov-15

Mar-14

May-13 May-18

Au Recovery - Flotation (%) Au Recovery - CIL (%)

Figure 17-4: Total Gold Recovery Showing Split Between Concentrate and Leach

100.0 Monthly Copper Recovery 2009 to 2018 90.0

80.0

70.0

60.0

50.0

40.0 CopperRecovery (%) 30.0

20.0

10.0

0.0

Jul-12 Jul-17

Jan-10 Jan-15

Jun-10 Jun-15

Oct-13 Oct-18

Apr-11 Apr-16

Sep-11 Feb-12 Sep-16 Feb-17

Dec-12 Dec-17

Aug-09 Aug-14

Nov-10 Nov-15

Mar-14 May-18 May-13 Figure 17-5: Total Copper Recovery

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25 Monthly Copper Con Cu Grade 2009 to 2018

20

15

10 CopperGrade(%)

5

0

Jul-12 Jul-17

Jan-10 Jan-15

Jun-10 Jun-15

Oct-13 Oct-18

Apr-11 Apr-16

Sep-11 Feb-12 Sep-16 Feb-17

Dec-17 Dec-12

Aug-14 Aug-09

Nov-10 Nov-15

Mar-14

May-13 May-18

Figure 17-6: Monthly Copper Concentrate Cu Grade The projected gold and copper recoveries and copper concentrate grade expectations for 2017 to 2020 are presented in Table 17-3.

Table 17-3: Projected Gold and Copper Recoveries and Copper Concentrate Grades

Actual Projected

2018 2019 2020 2021 2022 Gold Recovery % 83.4 82.8 84.1 85.6 86.1 Copper Recovery % 79.8 76.0 78.8 80.0 80.5 Copper Concentrate % 15.56 14.7 13.47 17.3 19.6

17.4 Product/Materials Handling Copper concentrate filtered on site is stored in a 5,000 t capacity storage shed at the site. From there the concentrate is loaded into dual road train side tipper trucks for transportation to the port of Bunbury, approximately 175 km from the site. The road trains carry on average 64 tonnes of concentrate per load, with 24 hour a day operation possible. The concentrate is tipped at the port and rehandled onto concentrate load out stockpiles within a nominal 30,000 tonne capacity storage shed owned by NBG. The concentrate is stored in the shed until required for load out and sales to an overseas smelter. The concentrate storage shed has the capacity for reclaim and conveying of the concentrate onto the Bunbury Port Authority (BPA), shiploading conveyor system, and from there into the hold of a bulk cargo vessel. The system has a capacity to load up to 500 tph of concentrate, with vessels carrying between 5,000 t and 15,000 t of material depending on the sales agreement.

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17.5 Energy, Water and Process Materials Requirements Power supply to the Project is via the local grid system with a coal fired power station built at Collie providing the additional demand for the operation as well as supplementing the existing grid. The power supply is generally very reliable with only minor load shedding required since start up and one production interruption occurring in early 2015 when bush fires resulted in intermittent loss of power to the operation for a few days (production loss <600 kt). Water supply to the operation comes from a variety of sources, the majority being recovered water from the Hotham river (during winter only) which is stored as Raw Water for use in subsequent months. This is supplemented by pit dewatering water, borefield water adjacent to the pits, rainfall run off and recovered water from the thickeners and tailings dam (Residue Disposal Area [RDA]). Water is reticulated around the site as either Process Water, Raw Water or Potable Water for various uses in the recovery process. At present the site relies on seasonal flow in the Hotham River. Decreased rainfall during winter could impact the mill production if sufficient water cannot be drawn from the river. Reagents and chemicals used in the processing include; Primary collector (thionocarbamate), secondary collector (Xanthate) and frother for copper and gold recovery in flotation, lime for pH modification, flocculant for thickening, cyanide, caustic and hydrochloric acid for leaching and gold recovery circuits as well as sulphuric acid and peroxide for the caros acid cyanide destruction process. Most chemicals are delivered to site in bulk containers with large storage tanks able to support the operation for several weeks if the supply chain is compromised. Grinding media is also delivered and stored in bulk on site.

17.6 Comments on Recovery Methods In the opinion of the QP, the mill process and associated recovery factors are considered appropriate to support Mineral Resource and Mineral Reserves estimation, and mine planning.

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18.0 PROJECT INFRASTRUCTURE

18.1 Overview The site plan covering existing facilities for the Project includes:  2.3 km overland conveyor;  Two ROM stockpiles, two medium-grade stockpiles, three waste rock facilities (WRFs);  One active RDA;  Four major water management facilities, including a decant water pond and water recycling facility;  Electrical sub-station (132 kV) to receive power from the south-west interconnected grid;  Concentrate storage shed and load-out facility;  Accommodation village. Site infrastructure layout plan are included as Figure 18-1 and Figure 18-2. The site also has numerous support facilities including truck shops, warehouses, and offices. Other support facilities include on-site core storage and a sample pulp storage at Marradong, approximately 20 km from the mine site.

18.2 Roads and Logistics The general site access to the Project is discussed in Section 5.0. The Project is accessed by an all-weather road network from Perth. Various access routes are used depending on the commodity being transported:  Construction and commodity supplies from Perth, Fremantle, Kwinana or Henderson typically use the Albany Highway;  Supplies and materials sourced from the local Peel Region and Bunbury area use the South West Highway;  Over-size loads traveling from Bunbury arrive via Collie using the Albany Highway. Copper concentrates are trucked using the Pinjarra–Williams Road and the Southwest Highway or Forrest Freeway to the Port of Bunbury.

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Note: Figure aligned to True North

Figure 18-1: Site Infrastructure Layout

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Note: Figure aligned to True North

Figure 18-2: Processing Plant Major Infrastructure and Layout

The filtered copper concentrate is stockpiled in the concentrate shed using a front-end loader before being loaded onto trucks for transport to Bunbury Port. Once at the port, the concentrate is tipped onto the receivable area before being stockpiled via a frontend loader ready for shipment. Loading of a ship is undertaken by reclaiming the stockpiled concentrate into a dump hopper. The concentrate is then fed from the hopper onto the out-load conveyor via a belt feeder and transfer conveyor. The concentrate is sampled by a two-stage cross belt sampling system for moisture and assay determination. The concentrate then discharges onto the port authority transfer conveyor system and eventually the port shiploading conveyor.

18.3 Waste Dump Facilities A plan showing the waste dump (WD) layout is presented in Figure 16-1. Projected capacities for the WDs are presented in Table 18-1.

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Table 18-1: Waste Dump Capacity

Capacity

Dump Stage 3 1,000 x m kt Oxide 7WD 8,008 12,012 9WD 7,187 10,781 10WD-W 8,717 13,076 11WD 10,446 15,670 Total 8,008 12,012 Rock 10WD-E 57,317 115,054 10WD-W 42,786 85,884 11WD 15,001 30,111 D4WD 128,390 257,718 Total 243,494 488,767

The following parameters have been assumed for WD slope design for the Project:  A lift height of 30 m for fresh rock and 15 m for oxide, with the final rehabilitated slope at a maximum of 30 m high;  A berm width of 83 m for every 30 m of vertical lift;  WD rill slope of 37º;  WD will be constructed to allow for a final overall dump slope of 15º;  Haul ramp gradient of 10%;  Minimum haul ramp width of 38 m including drainage features. The waste management strategy comprises:  WDs are planned to minimize the visual impact on the surrounding environment by limiting WD heights to the existing natural topography;  Encapsulation of potentially acid-forming waste with progressive WD rehabilitation and application of a store-and-release dump cover comprising at least 2 m of non-acid forming oxide regolith covered by a nominal 300 mm of root-bearing and waste gravels and 100 mm of topsoil;  Shallow seepage emanating from the base of WRFs will preferentially report to the impacted water blanket that will be constructed along the course of the 34 Mile Brook under the WD10;  A wet well constructed at the bottom (south) end of WD10, along the current 34 Mile Brook. The wet well dike will prevent uncontrolled migration of impacted water seepage downstream and will be combined with a low-permeability cut-off barrier below;  A mine water storage pond constructed at the low point of WD10 at the downstream of the wet well;  Preferential storing of potentially acid-forming waste in zones of WD7, WD8, WD10 and WD11;

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 Potentially-impacted seepage water emanating from the base of the WD9 will be minimized by limiting placement of potentially acid-forming waste in the WD9 and managed to preferentially report to the North Pit and be mixed with pit dewatering sump water. The 2018 Mineral Reserves mine schedule requires 439 Mt of dumping capacity through to the end of the life of mine in 2032, while the existing dumping capacity is approximately 184 Mt. It requires the waste dump expansion (WDX). The final environment permit for the WDX was granted at the end of 2014. The WDX Phase 2 stage-1 construction will start in early 2021 and to be completed by the end of Q3-2021. WDX Phase 2-stage 2 in the area of D4 dam will commence in Q1-2023 to provide WD capacity for LOM.

18.4 Tailings Storage Facilities Tailings characteristics are as follows:  The residue is split of sand and silt with very small clay amounts and classified as low plasticity or non-plastic;  Sedimentation tests indicated supernatant production of 35% to 40% (undrained) and 20% to 30% (drained). Under-drainage recovery ranged from 15% to 25%;  The testing indicated rapid settling achieving densities of approximately 1.30 tonnes per cubic meter (t/m3) to 1.40 t/m3 in less than one day. The air-drying tests indicated a maximum dry density of approximately 1.45 t/m3 to 1.65 t/m3;  Viscosity testing of the residues indicated that the residue could be pumped using standard pumping equipment up approximately 60% solids;  Compaction and strength testing showed that residue could be used as an embankment material in areas not inundated by the pond. In accordance with Cyanide Code, the level of cyanide in the residue disposal facility area is controlled to no more than 50 ppm (WAD) deposited in the residue at the spigot discharge (target 40 ppm). The residue disposal facility is designed for an ultimate storage capacity of 600 Mt of residue at a deposition rate of 39 Mtpa. At full capacity the embankments are designed to be within the property boundary at the northern and eastern perimeter, with a buffer of approximately 100 m between the toe of the embankment and the property boundaries. The buffer zone contains an access track and seepage monitoring bores. An option report investigating various options to increase residue storage was completed by Knight Piesold in 2013. The Knight Piesold estimate is based on building the RDA to a height of 362.9 m, at which it was assumed that it would be storing 600 Mt; however, subsequent measurement of the actual deposited tailings has indicated that it will likely be holding 640 Mt at this height. To store 750 Mt tailings volume will require the final height to be 369.6 m (Osan, 2014).

18.5 Water Management Process water is supplied direct from the mine pits, from onsite storage reservoirs which have been filled in the winter months by pumping from the Hotham River under a licence from the Department of Water or from regional water bores which are available all year round. Process water is also sourced as reclamation of water from the decant pond at the RDA. Recent increases in the licenced pumping rate from 3,500m3/h to 5,000m3/h and a total abstraction increase from 10GL per annum to 15GL per annum, has improved the site’s resilience to

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water shortages during the dry summer months. Coupled with an additional 3ML water storage dam at the old Hedges operation purchased from Alcoa in 2012, and additional pumping upgrades, the site has been able to maintain production consistently since 2012. Potable water for the camp and mining operation is sourced from two existing, refurbished 550 kiloliters (kL) water storage tanks located on Communications Hill. Water is pumped to the tanks from a reverse osmosis plant that treats water from the brackish D4 water storage dam. The site-wide water balance is managed through a GoldSim® model, with regular water use, abstraction and storage capacity data regularly fed into the model to obtain reliable forecasts of process and raw water. Based on these sources, the QP is of the opinion that potable and process water supplies are sufficient for current and planned development needs.

18.6 Camps and Accommodation A 2,300-person accommodation village was constructed approximately 8 km southeast from the operation to house the majority of the permanent workforce and to provide sufficient capacity to meet the elevated accommodation requirements during planned maintenance events for key infrastructure. The current workforce consists of approximately 1,000 employees and 700 contractors with approximately 25% of these residing locally within a 25 km radius of the operations and the balance residing in the accommodation village.

18.7 Power and Electrical A Power Purchase Agreement was signed on 30 June 2006 with Bluewaters 1 Pty Ltd (formerly Griffin Energy Group Pty Ltd). Power is sourced from the Bluewaters Power Station, a coal-fired power station located 4.5 km northeast of Collie, and approximately 80 km from the mine. The power is transmitted through the State power grid from the power station to the mine site. There is sufficient capacity from the Griffin Energy source to support the life-of-mine requirements.

18.8 Communications The mine is serviced by both a wide-area network, and local-area network. Fiber-optic cables provide a link to the corporate network, VOIP telephony and internet access to both the mine site and the accommodation village. 3G mobile phone coverage is available at the mine site and the accommodation village. An ultra-high frequency (UHF) radio system provides two-way, open, radio channel communications to support the operation.

18.9 Comments on Infrastructure In the opinion of the QP, the infrastructure required for the LOM has been constructed and is in operation and is considered appropriate to support Mineral Resource and Mineral Reserves estimation, and mine planning.

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19.0 MARKET STUDIES AND CONTRACTS

19.1 Market Studies and Contracts Newmont has an agreement with The Perth Mint for refining of gold and silver doré produced from the Project. Newmont’s bullion is sold on the spot market, by marketing experts retained in-house by Newmont. Boddington copper concentrate is unique as it contains one of the highest gold contents in the market and a relatively low copper content. Consistently, smelters operating their own precious metal refineries at their copper smelting operations are the subset of smelters that are prepared to contract for Boddington concentrates. Frame contracts have been negotiated with smelters in Korea, Japan, Philippines, and Germany for a large portion of the mine production of copper concentrate. The remaining production is expected to be sold through independent agents on the spot market. The frame contracts are in line with normal industry practice. The terms contained within the sales contracts are typical of and consistent with standard industry practice and contracts for the supply of copper concentrate elsewhere in the world. Depending on the specific contract, the terms for the sale of the Project’s copper concentrate are either annually negotiated, benchmark-based treatment and refining charges, or a combination of annually negotiated terms and price sharing agreements. The differences between the individual contracts are generally in relative quantity of concentrates that are covered under annually-negotiated treatment and refining charges and that are covered under a price sharing formula. Treatment charges assumed for estimation of Mineral Reserves are based on the blended rates of the existing contracts through the duration of the agreements. The formula used is sensitive to the underlying copper price and is consistent with long-term expectations for copper treatment and refining charges.

19.2 Commodity Price Projections Metal price assumptions are provided by Newmont management and are based on three-year trailing average prices applicable at the time the Mineral Reserves are estimated. Metal pricing and exchange rate assumptions used for the 2018 Mineral Reserves estimates are as follows:  Gold: US$1,200/oz or AU$1,600 /oz;  Copper: US$2.50/lb or AU$3.35/lb;  Exchange Rate: US$0.75 = AU$1.00. Higher metal prices are used for the Mineral Resource estimates to ensure the Mineral Reserves are a sub-set of, and not constrained by, the Mineral Resources, in accordance with industry accepted practice.

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19.3 Comments on Market Studies and Contracts In the opinion of the QP:  The terms contained within the sales contracts are typical and consistent with standard industry practice, and are similar to contracts for the supply of doré and concentrates elsewhere in the world;  Metal prices are set by Newmont management and are appropriate to the commodity and mine life projections.

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20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

20.1 Baseline Studies Several baseline studies were performed as part of the original mining operations, and in support of the current Wandoo pit operations.

20.2 Environmental Considerations The Project’s mine plan was developed in consideration of environmental matters during the feasibility study. Environmental planning includes:  Clearing: Total disturbance footprint at the end of 2018 was 4,123 ha, of which approximately 2,400 ha was previously cleared during the 1987–2001 mining episode. As some of the tenements sit over State Forest, commercial timber removal is conducted prior to clearing operations. Other vegetation is either used in rehabilitation works, burnt, mulched or stockpiled along with topsoil and gravel material which is later reapplied to the area during the rehabilitation process;  Air Quality: Dust levels involved in mining activities are mitigated where possible through various systems which include application of water and use of bag-houses at key points of the mining process. NBG conducts both point source and ambient air quality monitoring which is reported annually to the regulator. No limits are currently enforced through the operational licence;  Drainage: Sediment control measures, including sumps and diversion drains, are in place and monitored according to risk to manage runoff from dump areas and to minimize secondary impacts to surface water on and off the tenements.  Groundwater: Seepage management designs have been established to capture or limit potential impacts from shallow dump seepage at the contact of the waste rock formation (WRF) base. Most of the completed sections of all the WRFs will be progressively rehabilitated to minimize rainfall infiltration into the WRF, reduce the opportunity for acid rock drainage to develop, and to stabilize WRF surfaces to minimize erosion. Some areas that are designated for NAF stockpiles will be rehabilitated at the latter year of mine life when the NAF stockpiles are completely re-handled. The operation also employs a groundwater model supported by regional bores that monitors the impact of pit dewatering on the local hydrology systems. This program undergoes regular review to better understand the potential long-term impacts of the cone of depressurization on wetlands and local river system;  WRF: A substantial amount of waste rock geochemistry characterisation has been completed to develop options for operational and post closure waste rock management. An estimated 62.5% of the waste rock to be generated over the life of the Expansion Project has been characterized as potentially acid-generating (PAG). Of this, approximately 0.25% is high capacity (HC) with the remainder (effectively all) of the PAG- characterized waste rock being low capacity (LC). PAG-HC is defined as material that is likely to rapidly produce unacceptable water quality if left exposed to rainfall. The negligible amount of PAG-HC waste rock to be mined during the Expansion Project is restricted to the South Pit and will be fully encapsulated. The site Waste Rock Management Plan is refined regularly to ensure that all materials with acid-producing potential are appropriately managed and encapsulated;

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 Tailings Residues: The current RDA is approved to a maximum 600 Mt capacity. It will be constructed over 18-stages and is externally, geotechnical assessed annually and monitored through numerous bores and piezometers (Osan, 2014). The completion of each stage warrants the submission of a construction report to the relevant authority noting both compliance with approved engineering documents and any minor alterations. The operation of the Tailings facility is managed according to the site’s ICMI Cyanide Code certification which was last audited in 2018. Tailings into the facility are directed through a Caros Acid Plant which destroys the CNWAD prior to discharge. Return water from the central decant pond on the tailings facility is redirected back through the Caros Acid Plant before being recycled into the processing plant with a 60% recycling rate for water noted. Due to an increase in estimated Mineral Reserves following the initial feasibility study, NBG initiated a Life of Mine Extension Approval to obtain additional footprint for the construction of waste rock, residue disposal and water storage facilities and associated infrastructure (see also discussion in Section 20.4). The approval process required referral to both State and Federal regulators due to the presence Threatened Species / Matters of National Environmental Significance (MNES), most notably the three species of Black Cockatoo (Baudin’s, Carnaby’s and Forest Red Tailed), and the presence of Woylie and Chuditch. The granted approval supports extending the mining and processing activities beyond the current LOM completion date of 2030. The approval identifies the requirement for clearing of up to 1,755 ha of native vegetation. A variety of minor infrastructure will also be constructed to support the operation, including drainage lines, rehabilitation stockpiles, minor water storages and roads. Existing monitoring and regulatory compliance systems have been updated following approval of the project in 2015.

20.3 Closure Plan The 1978 WA Mining Act requires that NBG submits a closure plan every three years that is compliant with the Guidelines for Preparing Mine Closure Plans (Department of Mines and Petroleum & Environmental Protection Authority, 2015). The Boddington Operations Closure Plan (BOCP) was last submitted in December 2016 with the next submission due in December 2019. The plan is supported by the calculation of an approximate closure cost known as the Mine Rehabilitation Fund (MRF). This closure liability is used by the State government to calculate an annual 1% liability levy charged to the site which remains in effect until all tenements have been signed off as rehabilitated. For 2017, based on the existing cleared footprint, NBG contributed in the order of AU$1.37M. Due to a unique circumstance regarding the Bauxite state agreements, Newmont holds one tenement for which the MRF does not apply and a bond must be lodged with the DMP, which totals AU$3.6M. The objectives for closure and rehabilitation outlined within the plan are as follows:  Landforms will be structurally safe and stable;  Landforms will be consistent with the landscape, will facilitate a return to the regional drainage function and will not adversely affect the surrounding natural environment;  Mine waste materials with potential for environmental impact are appropriately contained;

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 Final landform surfaces develop resistance to erosive forces;  The soil profile will be suitable for the development of the target ecosystem;  Vegetation will resemble that of the region and enable integration into the surrounding landscape; and  Rehabilitation areas will be able to be managed as required for the post-mining land use. The Closure Plan includes commitments to rehabilitate the waste rock landforms, residue disposal areas, processing plant and other areas of disturbance. The Wandoo North and Wandoo South open pits will be left at the as-mined angle with an abandonment bund constructed to prevent inadvertent access. The open pits will develop into pit lakes through diversion of the Thirty-Four Mile Brook and inflow from groundwater and surface runoff. A store-and-release type cover is planned for the waste rock landforms, to promote vegetation growth and minimize infiltration of precipitation into the underlying waste rock that comprises a mixture of potentially acid forming and non-acid forming material. The store and release cover is currently costed as comprising:  2 m of non-acid forming regolith (oxide);  300 mm of gravel;  100 mm of topsoil. The cover design for the waste rock landform will be refined through the current and future rehabilitation trials to be conducted in the future with an initial steep slope trial finalized during 2015. Progressive rehabilitation will be undertaken of the waste rock landforms as areas are completed. Seepage from the waste rock landforms during mine operations is managed via collection and preferential recycling to the process plant. The impacted water sump located beneath the No. 8 waste rock landform serves as the primary collection and storage for seepage and is cut-off from external off-site surface water flows. In conjunction with the development of the No. 10 and 11 waste dumps, a new seepage interception infrastructure, called Impacted Water (IW) Blanket, will be constructed beneath the No 10 waste dump. The IW Blanket will serve as the interception & pump-away for seepage. Ultimately, the existing IWS will be integrated with the IW Blanket, so that it will enable collection, recycling & treatment of seepage in a sustainable way through post-closure. Post-closure treatment of seepage may be required until infiltration into the waste rock landform is minimized after construction of the store and release cover. The closure strategy for the residue disposal areas is formation of a water shedding facility to minimize infiltration of precipitation. A cover of 300 mm of gravel and 100 mm of topsoil will be applied over the residue material. Rehabilitation trials have been established to develop an appropriate prescription for revegetation of the basement ore residue surface. The closure strategy for disturbance associated with infrastructure and services is to re-shape disturbance areas to blend in with the surrounding topography and revegetate with local species. All infrastructure will be removed unless a signed agreement is in place regarding post-closure legal responsibility by a third party. Newmont estimates LOM closure costs based on the sum of costs associated with each of 20 different categories representing activities for which closure liabilities exist at a typical mine site. LOM liability represents the total costs to close and fully rehabilitate the site based on current regulatory requirements, closure obligations with key stakeholders and the site’s current mine plan used to establish production projections and budgeted expenditures. The

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LOM estimate considers rehabilitation of the full disturbance footprint that would exist at the end of the operational phase of a project plus contingency to address potential changes to unit rates and/or future assumptions and criteria changes. NBG is also required to calculate costs, both for compliance with Financial Accounting Standards (United States Generally Accepted Accounting Principles, or US GAAP) and for LOM, associated with site closure and are submitted by the Sustainability and External Relations (S&ER) Department annually in collaboration with the Mine Planning group. The US GAAP cost estimate looks at costs associated with closure if the mine was to cease operations early. The 2018 submissions have been based on the existing operations plan continuing out to 2032. The 2018 closure costs calculated the following liabilities based on the 2019 Business Plan. The closure liability estimates for 2018 are AU$383 M for the LOM 2019 Business Plan, and AU$311 M for US GAAP. 2018 cost increases for LOM and US GAAP are associated mostly with increase in unit rates and minor change in Waste Rock Dumps footprints.

20.4 Permitting 20.4.1 Initial Permit The Boddington Expansion Project was subject to comprehensive EIA by the WA EPA and other key Government decision-making authorities. Assessment included:  A Consultative Environmental Review (CER) pursuant to Section 38 of the Environmental Protection Act, for the proposed Extended Basement Operation (EBO) dated August 1996;  Changes to the previously approved mining operations (basement and oxide mining) at the former Hedges gold mine for which the NBG has responsibility (on lease ML264SA Section 2 and M70/1031), and the EBO at NBG subject to Section 46 of the Environmental Protection Act. This assessment resulted in Ministerial Statement No. 591 dated 8 May 2002. As part of the expansion Project, existing Ministerial Statements for the NBG (Nos. 453 and 489) and Hedges Gold Mine (No. 450), were rationalized by consolidating all of the previous NBGJV’s responsibilities for gold mining and processing into one Ministerial Statement (No. 591). Those responsibilities that were Alcoa’s under the existing Ministerial Statement for Hedges remain. The 2002 assessment and subsequent approvals were predicated on a mining operation processing rate of up to 29 Mtpa. Amended permits were applied for based on availability within Section 45C of the Environmental Protection Act for a ±10% change in operating throughputs. The Section 45C amendment report was issued to the EPA in October of 2005. As part of the approval process the JV had to submit the Section 45C report to an independent auditor. Final approval for the amended mining operation was granted on 21 March 2006 under the Environmental Protection Act. A Mining Proposal was prepared and submitted to the Department of Mines and Petroleum in August 2006. The Mining Proposal was approved on 18 September 2006. A submission was made to the Commonwealth Department of Environment and Heritage, under the Environmental Protection and Biodiversity Conservation Act in February 2006 due to the site being within the range of Black Cockatoos. The Minister for Environment and Heritage deemed the Boddington Expansion “is not a controlled action provided it is undertaken in a particular manner” on 3 April 2006.

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Following construction of key components of processing plant and residue disposal areas, Prescribed Premises Licence 8306/1 was issued under Part V of the Environmental Protection Act on 1 May 2009. The latest version of this licence is L8306/2008/2 and was issued on 11 March 2015. 20.4.2 Interim Permit Application was made in November 2011 for some minor expansion works for the Wandoo North and South pits and to obtain additional storage capacity in the No. 7 and 8 WRFs. The Interim Permit was required to be assessed by both the State and Federal environmental regulators and the state Department of Mines and Petroleum and was approved in August 2012 (EPBC 2011/6192). 20.4.3 Boddington LOM Expansion Project With the delineation of additional Mineral Reserves for the Project, there was a decision to seek further approvals for an expanded footprint and extended life of mine. The LOM extension project was referred to the State and Federal Governments in April 2012 and was formally assessed as a Public Environmental Review (PER) due to its considered level of environmental impacts to flora, fauna and regional groundwater. The LOM extension project requested approval for:  Expansion (widening and deepening) of the Wandoo North and Wandoo South open pits;  Increased ore production and increased waste quantities requiring the expansion of existing facilities;  Construction of an additional WRF adjacent to the pit with a final combined capacity for all WRFs of approximately 1,500 Mt;  An increase to size of stockpiles and ancillary infrastructure;  An increase in overall metal production requiring the construction of a second RDA and associated infrastructure to provide additional tailings storage of up to 600 Mt;  Construction of additional water storage dams and associated infrastructure to replace the water storage dam (D4) which would be covered by the new WRF;  Extending mine operations and processing operations beyond the current LOM closure projections of 2032. The PER was approved by the applicable Government agencies in 2014 and has allowed the operation to increase its cleared footprint from 3,680 ha to potentially 6,520 ha (State Ministerial Statement 971; EPBC 2012/6192). The associated Mining Proposal was approved in 2015. The conditions of approval to facilitate the expansion have required NBG to provide offsets to cover the expected residual impacts that cannot be managed through existing systems. Those offsets revolve around the placement of 2000ha of native vegetation into a conservation covenant, the restoration of 470 ha of degraded farmland and the divestment of native vegetation into the State Conservation Estate (to compensate for the loss of State Forest that is proposed to be cleared on the mining tenements).

20.5 Considerations of Social and Community Impacts There is sufficient archaeological evidence to conclude that Aboriginal people seasonally occupied the area around the Boddington mine before European settlement. Since the early 1980s, there have been several archaeological and ethnographic heritage surveys conducted

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around the Boddington area. This work concluded that there were some artifact scatters and mythological sites recorded in and around the mine and proposed expansion. On 28 April 2005, the previous NBGJV obtained approval from the Minister for Indigenous Affairs under Section 18 of the Aboriginal Heritage Act 1972 (WA) for the use of the land described as ML264SA(2), M70/1031 M70/21, M70/22, M70/24 for the purpose of current mining activities and the proposed expansion of the Boddington gold mine. Heritage sites located in the vicinity of the Boddington Operations are managed in accordance with the NBG Heritage Management Plan. Several additional surveys and approvals to support the Life of Mine Extension as well as a number of smaller projects have also been completed. Aside from requirements to monitor when disturbing in specific locations no other requirements or conditions have been imposed.

20.6 Comment on Environmental Studies, Permitting and Social or Community Impact In the opinion of the QP:  The Project holds the necessary permits to operate, and that social and community impacts have been addressed for LOM requirements;  Newmont has received the necessary approvals for a proposed LOM extension project, which will support mining operations continuing beyond the current 2032 closure in the LOM.

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21.0 CAPITAL AND OPERATING COSTS

21.1 Operating Cost 2018 Mineral Resource and Mineral Reserves and estimates were based on the 2019BP costs at 2018 escalation only (2019 Dollar). Table 21-1 presents the operating and capital unit cost of 2018 Mineral Reserves.

Table 21-1: Operating and Capital Unit Cost

2018-RSV Cost Component 2019BP

Base Processing Cost (excl. S/pile Rehandle) AU$/t milled 9.71

Stockpile Rehandle (50% of mill feed) AU$/t rehandle 1.48 Sustaining Capital (Plant and G&A) AU$/t milled 1.14 Site G&A (exclude CAPEX) AU$/t milled 1.37 Regional G&A Back charge AU$/t milled 0.72 LOM Operating Mining Cost AU$/t mined 4.22 LOM Mining Capital Cost AU$/t mined 0.62

21.2 Capital Expenditure Table 21-2 presents the capital expenditure schedule identified across the LOM for 2019BP.

Table 21-2: 2019BP LOM Capital Expenditure Summary

2019BP LOM 2018 2019 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031

Mining (AU$M) 553.9 26.1 57.3 80.8 76.1 67.2 42.4 43.5 35 43.8 36.5 21.6 20.6 3.1

Process (AU$M) 585.3 60.2 42.3 29.8 33.3 24.9 34.7 32.6 38.7 34.6 31.6 39 104.1 41.9 37.7

Site G&A (AU$M) 25 5 1.7 2.2 2.6 1.6 3.2 4.4 1.2 0.8 0.3 0.3 0.2 0.9 0.5

Total (AU$M) 1,164.10 91.3 101.3 112.8 112 93.8 80.2 80.4 74.9 79.1 68.4 61 124.9 45.9 38.1

21.2.1 Mining Capital Expenditure The total life of mine mining sustaining capital is AU$554M. Table 21-3 presents the summary of the capital discussing the major capital projects.

Table 21-3: 2019BP Mining Sustaining Capital Summary

Major Mining Sustaining Capital AU$M

Asset Capitalization 221.9

Mobile Equipment Purchases/Rebuild 283.3

Water Security 43.1

Other Mining 5.7

Total Mining Capital 553.9

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The key mining capital spending in 2019BP includes:  Asset Capitalization: mainly to capitalize equipment engines, shovel tracks and truck trays;  Hydraulic Shovel: purchase of one hydraulic shovel to replace one rope shovel (SH03) to meet the 2019BP target. By cancelling both of the SH03 rebuilds and EX5600 purchase in 2027, the 2019BP gains life of mine benefit in CAPEX spending;  Truck Rebuild: rebuilding 38 x CAT793 trucks with rate of 12 trucks per year rebuild programme during 2020-2022;  Drill Replacement: replacing 7 x PV235 production drills and 4 x D65 wall control drills in 2022 to 2023 to meet to current drills termination lives;  Water Security: part of the pit dewatering and pit drainage construction capital. 21.2.2 Processing Capital Expenditure The total LOM processing sustaining capital is AU$585M. Table 21-4 presents the summary of the capital.

Table 21-4: 2019BP Processing Sustaining Capital Summary

Major Processing Sustaining Capital AU$M

CV1 Structural Upgrade 2.8

Full Potential Projects 35.2

Other Process 59.0

RDA 442.7

Water Storage Construction 45.5

Total Processing CAPEX (AU$M) 585.3

The largest capital item in Processing Capital is the RDA construction cost. Compared to 2018BP the cost is higher by AU$123.4M. 21.2.3 G&A Capital Expenditure The total life of site sustaining capital is AU$25M. Table 21-5 presents the capital allocation.

Table 21-5: 2019BP G&A Sustaining Capital Summary

Major G&A Sustaining Capital AU$M

Village Kitchen Equipment & Units Upgrade 7.2

Others 17.8

Total G&A CAPEX (AU$M) 25.0

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21.3 Closure Costs The closure liability is AU$383M for the LOM 2019BP, and AU$311M for FASB. 2018 cost increases for these liabilities are associated mostly with increases in unit rates and minor changes in WRD footprints.

21.4 Comments on Capital and Operating Costs Capital and operating cost estimates were developed as part of the 2017 Business Plan and are based on vendor quotes and operational experience at the Project, as appropriate. In the opinion of the QP, the cost parameters are considered acceptable to support Mineral Resource and Mineral Reserves estimation, and mine planning.

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22.0 ECONOMIC ANALYSIS This section is not required as the Project is currently in production, Newmont is a producing issuer, and this Report does not include a material expansion of current production.

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23.0 ADJACENT PROPERTIES No current or historical gold or copper mining operations occur in the same region as the Project. However, the adjacent and overlying bauxite mining operations and State Agreements mentioned in the Mineral Tenure and Surface Rights (refer to Section 4.4 and 4.5), Property Agreements (refer to Section 4.3) and Current Mineral Tenure (refer to Section 4.4) are considered during mine planning processes.

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24.0 OTHER RELEVANT DATA AND INFORMATION This section is not relevant to this Report.

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25.0 INTERPRETATION AND CONCLUSIONS In the opinion of the QP:  Information provided by Newmont’s legal and tenure experts on the mining tenure held by Newmont in the Project area supports that the company has valid title that is sufficient to support declaration of Mineral Resources and Mineral Reserves;  Information provided by Newmont’s legal and tenure experts supports that the Operations hold sufficient surface rights to enable mining operations, and the declaration of Mineral Resources and Mineral Reserves. Appropriate steps, where required, have been taken to lodge either extensions or renewals of tenements as such fall due;  The geological understanding of the deposit settings, lithologies, and structural and alteration controls on mineralization is well understood. The mineralization styles and setting are also well understood;  Exploration programs completed to date are appropriate to the different mineralization styles known to occur within the Saddleback Greenstone Belt;  The quantity and quality of the lithological, geotechnical, collar and downhole survey data collected in the exploration, delineation, and grade control drill programs between 1980 and 2018 are adequate to support Mineral Resource and Mineral Reserves estimation and mine planning;  Sampling methods are acceptable and meet industry-standard practices. The quality of the gold and copper analytical data is reliable and sample preparation, analysis, and security are generally performed in accordance with exploration best practices and industry standards;  The process of data verification for the Project has been performed by Newmont personnel, staff from Newmont’s predecessors and external consultancies contracted by Newmont. 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 Reserves estimation, and in mine planning;  Metallurgical testwork completed on the Project has been appropriate to establish optimal processing routes for the different mineralization styles encountered in the deposits. Testwork has been completed on mineralization that is typical of that within the deposits. The mill process and associated recovery factors are considered appropriate to support Mineral Resource and Mineral Reserves estimation, and mine planning;  A detailed review of the metallurgical models and comparisons with process plant actual performance indicated a tendency for the metallurgical models to over-predict recovery at low head grades and under-predict recovery at high head grades. Molybdenum and arsenic grades in copper concentrate were in line with model predictions to date. Ore hardness and abrasiveness of the samples were in line with the orebody average, possibly except for a slight increase in Bond ball mill work index with depth;  Estimates of Mineral Resources and Mineral Reserves for the Project conform with industry standard practices and satisfy the CIM Definition Standards. There is some upside for the Project if some or all the estimated Inferred Mineral Resources can be upgraded to higher confidence Mineral Resource categories;  The mine plan uses conventional mining methods and equipment and is adequately documented in terms of the availability of staff, the existing power, and communications

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facilities, the methods whereby goods are transported to and from the mine, and any planned modifications or supporting studies are well understood;  Since the effective date of the mine plan, Newmont has regularly undertaken, and will continue to undertake as part of its normal course of business operations, reviews of the mine plan and consideration of alternatives to and variations within the plan. Alternative scenarios and reviews are based on ongoing or future mining considerations, evaluation of different potential input factors and assumptions, and requests made of Project staff by Newmont Corporate;  Based on the current mine plan, some of the existing mine facilities such as waste rock dumps, tailings dams and water storage are required to be extended to achieve LOMP (2032);  The EIS and the current state of environmental knowledge of the mine area support the Mineral Resources and Mineral Reserves statement can be declared, and the mine plan is achievable;  Closure and remediation requirements have been addressed through the site closure plan and associated environmental bonding requirements;  Permitting activities have been carried on appropriately to ensure the mining activities can be conducted within the regulatory framework required by the West Australian and Federal Governments;  At the effective date of this Report, environmental liabilities are limited to those that would be expected to be associated with a gold-copper mine of comparable scale, including roads, open pits, site infrastructure, waste and tailings disposal facilities. Newmont has appropriately addressed the potential and actual environmental impacts of the operation sufficient to support Mineral Resources and Mineral Reserves estimation, and mine planning;  The mill process is conventional, producing doré and copper-gold concentrates. The mill process and associated recovery factors are considered appropriate to support Mineral Resource and Mineral Reserves estimation, and mine planning;  All required infrastructure has been established, and is operational;  Capital and operating cost parameters are considered appropriate to support Mineral Resource and Mineral Reserves estimation, and mine planning;  Review of the environmental, permitting, legal, title, taxation, socio-economic, marketing, and political information on the Project supports the assumptions used in the economic analysis, which is positive, and supports the Mineral Reserves;  Other than what is presented in this Report, there are no other significant factors and risks that may affect access, title, or the right or ability to perform work on the property.

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26.0 RECOMMENDATIONS The QP makes the following recommendations:  Completion of metallurgical testwork to determine the extent of recovery degradation on long term stockpiles. It is estimated that over a period of five years, a total of 12 samples will be prepared and tested multiple times at a cost of approximately US$30k per test for a total program cost of approximately US$840k;  Infill drilling to increase Mineral Resource confidence in the later pit stages and lower benches of the current Mineral Reserves. Over the next five years, it is anticipated that a total of approximately 44,631 m (average cost/m of approximately $AU177.00/m inclusive of drilling, assaying, and general support costs i.e. diesel, labor and supplies) of core and RC drilling will be completed at an estimated cost of AU$7.9M;  Completion of projects aimed at maximising the capture and use of raw water on site and securing a long-term reliable supply system for the Project. These projects are estimated to cost approximately US$56M over the next three years.

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27.0 REFERENCES

27.1 Bibliography Newmont Mining Corporation, 2016. GT-GM060-20161026-AGJ-FINAL – 75deg Steeping Trials: Unpublished Internal Report, October 2016. SRK Consulting, 2012a. NEM008: Boddington Gold Mine: Geotechnical Study for Final Saprolite Slopes, February 2012. Snowden Consulting, 2012b. 2235: Geotechnical Assessment of Fresh Rock Pit Cut-back, April 2012. Australian Bureau of Statistics, 2011: 2011 Census of Population and Housing, Basic Community Profile for Boddington (S) (LGA50630),cat no. 2001.0, Australian Bureau of Statistics, Canberra. Allibone, A.H., Windh, J., Etheridge, M.A., Burton, D., Anderson, G., Edwards, P., Miller, A., Graves, C., Fanning, C.M., and Wysoczanski, R., 1998: Timing Relationships and Structural Controls on the Location of Au–Cu Mineralization at the Boddington Gold Mine, Western Australia: Economic Geology, 93, pp. 245–270. AMMTEC, 1989: Assessment of Australian Assay Laboratories Boddington Facility: unpublished internal report by AMMTEC to Boddington Gold Mine. AngloGold Ashanti Australia, 2001: AngloGold Inputs for 1999 Wandoo Resource Assessment and the Derivation of a Factor for the Economic Model: unpublished internal report by AngloGold Ashanti Australia to NBGJV, October 2001. AngloGold Ashanti Australia, 2003: Boddington Joint Venture, 2003 AngloGold–BGM Collaborative Study — Boddington Gold Mine Exploration Data Review and Target Generation: unpublished internal report NBGJV, December 2003. AngloGold Ashanti Australia, 2004a: Boddington – Drill Hole Bias Examination: unpublished internal report by AngloGold Ashanti Australia to NBGJV, February 2004. AngloGold Ashanti Australia, 2004b: Boddington Basement Pits Grade Control Information Examination: unpublished internal report by AngloGold Ashanti Australia to NBGJV, March 2004. AngloGold Ashanti Australia, 2004c: Boddington Geology and Domain Review: unpublished internal report by AngloGold Ashanti Australia to NBGJV, June 2004. Augenstein,C et.al. 2012: Boddington Geological Campaign 2012: unpublished internal report by Jigsaw Geoscience to NBG, November 2012. Barley M.E., Groves D.I. and Blake T.S., 1992: Archaean metal deposits related to tectonics: evidence from Western Australia, Perth, Western Australia, Geology Department and University Extension, University of Western Australia Publication 22, p. 307–324. Boddington Gold Mining Company, 2003: Review of AGAA Estimation Domains: unpublished internal NBGJV report, October 2003. Boddington Gold Mining Company, 2004a: FSU Local Resource Estimate Quality, Mining Dilution and Recovery Review: unpublished internal NBGJV report, January 2004.

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Boddington Gold Mining Company, 2004b: Recommended Configuration of the FSU Phase 3 Local Resource Estimate: unpublished internal NBGJV report, December 2004. Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2003: Estimation of Mineral Resources and Mineral Reserves, Best Practice Guidelines: Canadian Institute of Mining, Metallurgy and Petroleum, November 23, 2003, http://web.cim.org/UserFiles/File/Estimation- Mineral-Resources-Mineral-Reserves-11-23-2003.pdf. Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2014: CIM Definition Standards: Canadian Institute of Mining, Metallurgy and Petroleum, May, 2014, https://mrmr.cim.org/media/1088/cim_definition_standards_may10_2014.pdf. Canadian Securities Administrators (CSA), 2011: National Instrument 43-101, Standards of Disclosure for Mineral Projects, Canadian Securities Administrators. Douglas, I., 2004: Boddington Drill Hole Bias Investigations: unpublished internal memorandum from Newmont Gold Corp. to NBGJV, 4 April, 2004. Fluor Australia Pty Ltd, 2000: Boddington Expansion Feasibility Study Update: unpublished internal report by Fluor Australia Pty Ltd to Boddington Gold Mine, Volume 1 (Executive Summary), Volume 2 (Geology and Resource), Volume 3 (Mining) and Volume 4 (Process). Gleeson, K., Tangney, G., Behn, M., and Hutchin, S., 1999: Boddington Gold Mine — Wandoo Project, Wandoo South and North Geology Report, Volume 1: unpublished internal report by Boddington Gold Mine. Golder Associates, 2002: Report on Geological Modeling and Recoverable Resource Estimation of the Wandoo Deposit: unpublished internal report by Golder Associates to NBGJV, May 2002. Golder Associates, 2002: Review of the Effects of Potential Check Assay Biases: unpublished internal report by Golder Associates to NBGJV, May 2002. Golder Associates, 2003: Report on 2003 Multiple Indicator Kriging Resource Estimate of the Wandoo Deposit, Boddington Gold Mine: unpublished internal report by Golder Associates to NBGJV, May 2003. Golder Associates, 2004a: Kriging Neighborhood Analysis and Re-estimation of the Boddington Gold Mine Expansion Resource Model: unpublished internal report by Golder Associates to NBGJV, April 2004. Golder Associates, 2004b: Review of Acid Rock Drainage Studies for the Boddington Expansion Project: unpublished internal report by Golder Associates to NBGJV, April 2004. Golder Associates, 2005a: Local Resource Estimates for the Boddington Gold Mine Expansion: unpublished internal report by Golder Associates to NBGJV, February 2005. Golder Associates, 2005b: Global Resource Simulation Study for the Boddington Gold Mine Expansion: unpublished internal report by Golder Associates to NBGJV, June 2005. Golder Associates, 2005c: BGME Probability-Based Resource Classification Using Simulations: unpublished internal report by Golder Associates to NBGJV, September 2005. Golder Associates, 2017a: Technical Review of Mineral Resources and Mineral Reserves: unpublished internal report by Golder Associates to NBG, May 2017.

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Golder Associates, 2017b: Newmont Boddington 2017 Preliminary Ore Reserve: unpublished internal report by Golder Associates to NBG, November 2017. Golder Associates, 2018: Technical Review of 2017 Mineral Resource Update: unpublished internal report by Golder Associates to NBG, January 2018. Kenny, K., Tangney, G., and Rowell, A., 2002: Boddington Expansion Project Wandoo Mineral Resource: unpublished internal report by AngloGold Ashanti Australia to NBGJV. Kirkham, R.V., 1972: Porphyry Deposits: in Blackadar, R.G., ed., Report of Activities Part B, November 1971 to March 1972: Geological Survey of Canada, Paper 72-1b, pp. 62–64. Knight Piesold Consulting, 2016: F1/F3 Residue Disposal Area 2015 RDA Cone Penetration Testing: unpublished internal report by Knight Piesold Consulting to NBG, February 2016 Libby, W.G. and DeLaeter, J.R., 1998: Biotite Rb-Sr Age Evidence for Early Palaeozoic Tectonism and the Cratonic Margin in Southwestern Australia, Australian Journal of Earth Sciences, vol 45, pp. 623–632. Masters, S., 2008: Audit of the BGM 2007 Resource Model, Boddington Mine, WA , unpublished internal report byCS-2 Pty Ltd to NBGJV, September 2008. McCuaig, T., and Behn, M., 2001: Boddington Gold Mine: Nature of the Mineralisation in the Wandoo North Basement Resource and Refinements to the Model for Genesis of Mineralisation at Wandoo: unpublished internal report by SRK Consulting to NBGJV. McCuaig, T.C, Behn, M.T., Stein, H., Hagemann, S.G., McNaughton, N., Cassidy, K,F., Champion, D., and Wyborn, L., 2002: The Boddington Gold Mine: a new style of Archaean Au–Cu deposit, in WA Gold Giants, MSc Short Course Notes, Centre for Global Metallogeny, University of WA, pp. 61–64. Miller, A., Behn, M., and Gleeson K., 1996: Wandoo Prospect Geological Report, Volume 1: unpublished internal report by Boddington Gold Mine. Newmont Mining Corporation, 2008a: Form 10K: unpublished statutory report by Newmont Mining Corporation to United States Securities and Exchange Commission, 21 February, 2008: report posted to Newmont website, accessed 4 April 2008, http://ir.newmont.com/phoenix.zhtml?c=66018&p=irol- sec&secCat01.1_rs=21&secCat01.1_rc=10. Newmont Mining Corporation, 2008b: Mineral Resource and Ore Reserves Report as of December 31, 2007, Boddington: unpublished internal report, Newmont Mining Corp, 28 February 2008. Newmont Mining Corporation, 2009a: Form 10K: unpublished statutory report by Newmont Mining Corporation to United States Securities and Exchange Commission, 19 February 2009: report posted to Newmont website, accessed 20 February 2009, http://files.shareholder.com/downloads/NEM/556618887x0xS950134-09- 3236/1164727/filing.pdf. Newmont Mining Corporation, 2009b: Mineral Resource and Ore Reserves Report as of December 31, 2008, Boddington: unpublished internal report, Newmont Mining Corp, February 2009. Newmont Mining Corporation, 2014a: 2015 Budget Pack – 2015 Business Plan Financial, internal report, Newmont Boddington Gold, October 2014.

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Newmont Mining Corporation, 2014b, Geotechnical Study for Hard Rock Slope Design at Newmont Boddington Gold, 20130328-GT-PROJ-AT- Hard Rock Study Report Final, March 2014. Newmont Mining Corporation, 2016, GT-Geotechnical Study for Hard Rock Slope Design at Newmont Boddington Gold, 20130328-GT-PROJ-AT- Hard Rock Study Report Final, March 2014. Newmont Mining Corporation, 2016, 75Deg Steepening Trails. Unpublished internal report, GT-GM060-20161026-AGJ-FINAL, October 2016. Newmont Mining Ltd. 2004a: NML Review of Boddington Project Resource: unpublished internal report by Newmont Mining Ltd to NBGJV, January 2004. Newmont Mining Ltd, 2004b: Boddington Drill Hole Bias Investigations, unpublished internal report by Newmont to NBGJV, April 2004. Osan, M., 2014: 2015 RDA LOM Cost Basis and History, internal report by Newmont Boddington Gold, 30 July 2014. Peattie R., 2004: Boddington Drillhole Bias Examination: unpublished internal memorandum from AngloGold Ashanti to NBGJV, 19 February 2004. Petrucci, P., 2014: Metallurgical Performance Model Update 2014, Newmont Boddington Gold, unpublished internal report by Boddington Gold Mine. Petrucci, P., 2015: 2015 Gold Recovery Function Update, Newmont Boddington Gold, unpublished internal report by Boddington Gold Mine. Quantitative Geoscience, 2003: Audit of the 2003 Boddington Resource Estimate, unpublished internal report by Quantitative Geoscience to NBGJV, October 2003. Ravenscroft, P., 2007: Review of Gold Resource Modeling Methodology, Boddington Mine, WA , unpublished internal report by CS-2 Pty Ltd to NBGJV, August 2007. Roberts, M., 2012: Metallurgical Performance Models, Newmont Boddington Gold, unpublished internal report by Boddington Gold Mine. Rossi, M., 2009: 2009 UC Resource Model Independent Audit Report, Boddington Gold Mine, unpublished internal report by GeoSystems International Inc to NBG, September 2009. Roth, E., 1992: The Nature and Genesis of Archaean Porphyry-Style Cu–Au–Mo Mineralisation at the Boddington Gold Mine, Western Australia: unpublished Ph.D. thesis, University of Western Australia. Runge, K.,2012: Evaluation of Recovery Function Predictions – February 2012, unpublished internal report by Metso Process Technology & Innovation to NBG Sillitoe, R. H., 2000. Gold-Rich Porphyry Deposits: Descriptive and Genetic Models and their Role in Exploration and Discover, in Gold in 2000, Reviews in Economic Geology, Vol. 13, Society of Economic Geologists. Sinclair, W.D., 2006. Consolidation and Synthesis of Mineral Deposits Knowledge - Porphyry Deposits: report posted to Natural Resources Canada website 30 January 2006, 14 http://gsc.nrcan.gc.ca/mindep/synth_dep/porph/index_e.php>, accessed 4 April 2008.

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SRK Consulting, 2005: Deleterious Elements Modeling: unpublished internal report by SRK Consulting to NBGJV, May 2005. Stein, H.J., Markey, R.J., Morgan, J.W., Selby, D., Creaser, R.A., McCuaig, T.C., and Behn M., 2001: Re-Os Dating of Boddington Molybdenite, SW Yilgarn: Two Au Mineralization Events: report posted to University of Alberta website, accessed 4 April 2008. http://graduate.eas.ualberta.ca/dselby/4IAS.Boddington.pdf. Stoker, P., 2000: Audit Boddington Expansion QA/QC: unpublished internal report to NBGJV, September 2000. Stoker, P., 2001: Review of Additional QA/QC Data, Boddington Expansion: unpublished internal report to NBGJV, February 2001. Stoker, P., 2002: Review of BGM QA/QC Assay Data, Boddington Expansion: unpublished internal report to NBGJV, January 2002. Surman, J., 1999: Boddington Gold Mine Audit of data used in preparation for the Wandoo Bedrock Resource Estimation: unpublished internal report by Snowden Mining Industry Consultants to Boddington Gold Mine. Symons, P.M., Anderson, G., Beard, T.J., Hamilton, L.M., Reynolds, G.D., Robinson, J.M., Staley, R.W., and Thompson, C.M., 1990: Boddington Gold Deposit in: Geology of the Mineral Deposits of Australia and Papua New Guinea, ed. F. E. Hughes, Australasian Institute of Mining and Metallurgy Monograph 14, Volume 1, pp. 165–169. Tangney, G., 2000: Boddington Gold Mine Basement Gold Assay Quality Control: unpublished internal memorandum by Boddington Gold Mine. Wilde, S. A., 1976: The Saddleback Group – A Newly-Discovered Archaean Greenstone Belt in the Southwestern Yilgarn Block: Western Australian Geological Survey Annual Report 1975, pp. 92–95.

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27.2 Glossary of Abbreviations, Symbols and Units Symbol Definition Symbol Definition ' seconds (geographic) Mlbs million pounds ' foot/feet Mm million meters minutes (geographic) mm millimeter/millimeters " inches Moz million ounces # number mRL meters relative level % percent Mt million tonnes / per Mtpa million tonnes per annum < less than MW megawatts > greater than NQ/NQ2 47.6 mm size core µm micrometer (micron) º degrees a annum/ year ºC degrees Celsius Å angstroms oz ounce/ounces (troy ounce) asl above sea level p passing BQ 36.5 mm diameter core measure of the acidity or alkalinity of a pH c. circa solution d day pop population d/wk days per week ppb parts per billion dmt dry metric tonne ppm parts per million fineness parts per thousand of gold in an alloy PQ 85 mm diameter core ft feet t metric tonne g gram tpa tonnes per annum (tonnes per year) g/cm3 grams per cubic centimeter tpd tonnes per day g/dmt grams per dry metric tonne tph tonnes per hour 3 g/m3 grams per cubic meter t/m tonnes per cubic meter Ga billion years ago TDS total dissolved solids ha hectares TSS total suspended solids HP horsepower µm micrometers HQ 63.5 mm diameter core wt% weight percent kg/m3 kilograms per cubic meter kL kiloliters Abbreviation Definition km kilometer ® registered name km2 square kilometers AAS atomic absorption spectroscopy koz thousand ounces AAL Australian Assay Laboratories kt thousand tonnes AC Aircore kV kilovolt Alcoa Alcoa of Australia Ltd kVA kilovolt–ampere Amdel Amdel Laboratory kW kilowatt ANC acid-neutralizing capacity kWh kilowatt hour ANP acid-neutralizing potential kWh/t kilowatt hours per tonne ARD acid-rock drainage lb pound ASX Australian Stock Exchange M million AU$ Australian Dollar m meter AuAA cyanide-soluble gold m3 cubic meter AuEq gold equivalent m3/hr cubic meters per hour AuFA fire assay Ma million years ago AuPR preg-rob gold size based on the number of openings in mesh one inch of screen AuSF screen fire assay Australasian Institute of Mining and mg/L milligrams per liter AusIMM Metallurgy mi mile/miles

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Abbreviation Definition Abbreviation Definition BFA bench face angle Hedges Hedges Gold Pty Ltd BGMJV Boddington Gold Mine Joint Venture HPGR high pressure grinding rolls BHP BHP Minerals Ltd ICP inductively-coupled plasma BLEG bulk leach extractable gold inductively-coupled plasma atomic ICP-AES BLM US Bureau of Land Management emission spectroscopy inductively-coupled plasma mass BMCO breakeven mill cut-off ICP-MS spectrometry BOCP Boddington Operations Closure Plan inductively-coupled plasma optical BPA Bunbury Port Authority ICP-OES emission spectrometry BSCO breakeven stockpile cut-off IRSA Inter-ramp slope angle C.P.G. Certified Professional Geologist IW Impacted Water Capex capital expenditure JCR joint condition rating CAF cost adjustment factor The Joint Ore Reserves Committee of CER Consultative Environmental Review The Australasian Institute of Mining CIL carbon-in-leach JORC and Metallurgy, Australian Institute of Canadian Institute of Mining, Geoscientists and Minerals Council of CIM Metallurgy and Petroleum Australia JV joint venture CNwad Weak acid-dissociable cyanide CRF capital recovery factor KV kriging variance Kobe Alumina Associates (Australia) CRM certified reference material Kobe Pty Ltd CST cleaner scavenger tailings L–G Lerchs–Grossman CTOT carbon total LC low capacity Cu Eq copper equivalent LOA length overall CuCN cyanide-soluble copper LOM life-of-mine Department of Mines, Industry DMIRS Regulation and Safety LSK large-scale kinetic Department of Minerals and Member of Australian Institute of DMP MAIG Petroleum Geoscientists Department of Water and Member of the Australasian Institute DWER MAusIMM Environmental Regulation of Mining and Metallurgy E east MIK multiple-indicator kriging EBO Extended Basement Operation MN magnetic north Moorditj Booja Community EDA exploratory data analysis MBCPA Partnership Agreement EIA Environmental Impact Assessment Matters of National Environmental EIS Environmental Impact Statement MNES Significance EOM end of month MPA maximum potential acidity EOY end of year MRF Mine Rehabilitation Fund EPA Environmental Protection Authority MWMS mine water management system Environmental Review and ERMP MWMT meteoric water mobility testing Management Program N north Fellow of the Australasian Institute of FAusIMM net acid generation/net acid Mining and Metallurgy NAG generating G&A General and administrative NAPP net acid-producing potential Generally Accepted Accounting GAAP Principles NBG Newmont Boddington Gold Genalysis Genalysis Laboratory Newmont Newmont Mining Corporation Golder Golder Associates Pty Ltd Canadian National Instrument 43-101 NI 43-101 “Standards of Disclosure for Mineral GN mine grid north Projects” GPS global positioning system Newmont Boddington Gold Joint NBGJV GRG gravity recovery gold Venture GSM Groupe Spécial Mobile NOI Notice of Intent H horizontal NN nearest-neighbor HC high capacity NNP net neutralizing potential

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Abbreviation Definition Abbreviation Definition NSR net smelter return South West Aboriginal Land and Sea SWALSC NTA Native Title Act Council NW northwest SX-EW solvent extraction–electrowin OK ordinary kriging TF tonnage factor Opex operating expenditure TN true north P.Eng. Professional Engineer Topo topography P.Geol Professional Geologist UC uniform conditioning PAG potentially acid-generating UG underground PER Public Environmental Review UHF ultra-high frequency PGE Platinum Group Elements UltraTrace UltraTrace Geoanalytical Laboratories PLI point load index USGS United States Geologic Survey PoO Plan of Operations US United States PSI pounds per square inch V vertical QA/QC quality assurance and quality control US$ United States Dollar QLT quick leach test VHF very high frequency QP Qualified Person VWP vibrating wire piezometer RAB rotary air blast W west RC reverse circulation WD waste dump RDA Residue Disposal Area XRD X-ray diffraction Reynolds Reynolds Australia Alumina Ltd WA Western Australia Registered Member, Society for WDX waste dump expansion RM SME Mining, Metallurgy and Exploration Worsley Worsley Alumina Pty Ltd RMR rock mass rating Worsley JV Worsley Alumina Joint Venture ROM run-of-mine WRF waste rock formation RPL Environmental Monitoring Plan XRF X-ray fluorescence RQD rock quality designation S south Saddleback Saddleback Investments Pty Ltd SAG semi-autogenous grind S&ER Sustainability and External Relations SE southeast System for Electronic Document SEDAR Analysis and Retrieval Supplemental Environmental Impact SEIS Statement SG specific gravity The Shell Company of Australia Ltd. Shell (Shell) SKM Sinclair Knight Mertz The Society for Mining, Metallurgy & SME Exploration Registered Member of The Society for RM-SME Mining, Metallurgy & Exploration SMU selective mining unit Snowden Mining Industry Consultants Snowden Pty Ltd Sotico Sotico Pty Ltd South32 South32 Worsley Alumina Pty Ltd SRM standard reference material SS sulfide sulfur ST scavenger tailings STOT sulfur total

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Symbol Element Symbol Element Ag silver S sulfur Al aluminum Sb antimony As arsenic Sc scandium Au gold Se selenium B boron Sn tin

Ba barium SO2 sulfur dioxide Be beryllium Sr strontium Bi bismuth Ta tantalum C carbon Te tellurium Ca calcium Th thorium

CaCO3 calcium carbonate Ti titanium CaO calcium oxide Tl thallium

CaSO4•2H2O calcium sulfide dehydrate U uranium Cd cadmium V vanadium Ce cerium W tungsten Cl chlorine Y yttrium CN cyanide Zn zinc CO carbon monoxide Zr zirconium Co cobalt Cr chromium Cs cesium Cu copper Fe iron FeOx iron oxides Ga gallium Ge germanium H hydrogen Hf hafnium Hg mercury In indium K potassium La lanthium Li lithium Mg magnesium Mn manganese

Mn(OH)2 manganous hydroxide

MnO2 manganese dioxide Mo molybdenum N nitrogen Na sodium Nb niobium

NH3 ammonia Ni nickel NOx nitrogen oxide compounds

O2 oxygen P phosphorus Pb lead Pd palladium Pt platinum Rb rubidium Re rhenium

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