NI 43-101 TECHNICAL REPORT Ambatovy Nickel Project, Madagascar
NI 43-101 TECHNICAL REPORT
Ambatovy Nickel Project, Madagascar
CSA Global Report Nº R501.2018 Effective Date: 30 June 2018
www.csaglobal.com
Qualified Persons Michael Elias, FAusIMM (CSA Global) Stewart Lewis, MAusIMM(CP) (IMC Mining) Paul O’Callaghan, FAusIMM (CSA Global) Adrian Martinez, P.Geo (CSA Global) Glen Smith, P. Eng. (Sherritt International Corporation)
SHERRITT INTERNATIONAL CORPORATION AMBATOVY PROJECT – NI 43-101 TECHNICAL REPORT
Report prepared for
Client Name Sherritt International Corporation Project Name/Job Code STO.MSA.01v2 Contact Name Glen Smith Contact Title Consulting Engineer - Technologies Office Address 8301-113 Street, Fort Saskatchewan, Alberta T8L 4K7
Report issued by
CSA Global Office Perth Division Mining Street Address Level 2, 3 Ord Street, West Perth, WA 6005 Postal Address PO Box 141, West Perth, WA 6872 Phone +61 8 9355 1677 Email [email protected]
Report information
File name CSA Global Report R501.2018 NI 43-101 Technical Report - Ambatovy Nickel Project Effective Date 30 June 2018 Report Signature Date 10 January 2019 Report Status Final
Author and QP Signatures
Michael Elias [“Signed”] Coordinating BSc (Hons), FAusIMM(CP) Signature: {Michael Elias} Author CSA Global Pty Ltd at Perth, Australia
Stewart Lewis [“Signed”] Coordinating BEng (Mining), MBA, Signature: {Stewart Lewis} Author MAusIMM(CP) at Brisbane, Australia IMC Mining Pty Ld
Paul O’Callaghan [“Signed”] Coordinating BEng (Mining), FAusIMM Signature: {Paul O’Callaghan} Author CSA Global Pty Ltd at Perth, Australia
Dr. Adrian Martinez Vargas [“Signed & Sealed”] Coordinating PhD., P.Geo. (BC, ON) Signature: {Adrian Martinez Vargas} Author CSA Global Pty Ltd at Ottawa, Ontario, Canada
Glen Smith [“Signed & Sealed”] Coordinating P. Eng., (AB, ON) Signature: {Glen Smith} Author Sherritt International at Fort Saskatchewan, Alberta, Canada Corporation
© Copyright 2019
CSA Global Report Nº R501.2018 I
SHERRITT INTERNATIONAL CORPORATION AMBATOVY PROJECT – NI 43-101 TECHNICAL REPORT
Certificates
Certificate of Qualified Person – Michael Elias As a Qualified Person of this Technical Reporttitled “NI 43-101 Technical Report for the Ambatovy Project, Madagascar”, prepared for Sherritt International Corporation with an effective date of 30 June 2018 (the “Technical Report”), I, Michael Elias do hereby certify that: • I am a Principal Consultant with CSA Global Pty Ltd at its head office at Level 2, 3 Ord Street, West Perth, WA 6005, Australia. • I am a professional Geologist having graduated with a BSc (Hons) Geology from the University of Melbourne (1973). • I am a Fellow of the Australasian Institute of Mining and Metallurgy and a Chartered Professional in the field of Geology. • I have practised my profession as a Geologist for the past 45 years in the mineral resources sector and engaged in the exploration for, assessment, development and operation of numerous mineral projects both within Australia and overseas. • I have read the definition of“ Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43- 101) and past relevant work experience, I fulfil the requirements to be a “Qualified Person” for the purposes of NI 43-101. • I am responsible for Sections 2 to 6 of the Technical Report. • I personally visited the property that is the subject of the Technical Report for five days from 27 to 31 January 2014. • I am independent of the issuer as described in Section 1.5 of NI 43-101. • I have read NI 43-101, and the Technical Reports has been prepared in compliance with NI 43-101. • As of the effective date, the Technical Report, to the best of my knowledge, information, and belief, contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
Effective Date 30 June 2018 Dated at Perth, Australia this 10 January 2019
[“Signed”] {Michael Elias}
Michael Elias, BSc (Hons), FAusIMM(CP) Principal Consultant – Nickel CSA Global Pty Ltd
CSA Global Report Nº R501.2018 II
SHERRITT INTERNATIONAL CORPORATION AMBATOVY PROJECT – NI 43-101 TECHNICAL REPORT
Certificate of Qualified Person – Stewart Lewis I, Stewart Lewis, state that: • I am the CEO and Principal Mining engineer at IMC Mining Pty Ltd, Level 14, 300 Queen Street, Brisbane, Queensland 4000, Australia. • This certificate applies to the technical report titled “National Instrument 43-101 Technical Report for the Ambatovy Nickel Project” with an effective date of30 June 2018 (the “Technical Report”). • I am a “qualified person” for the purposes of National Instrument 43-101 (the Instrument”). My qualifications as a qualified person are as follows: o I am a graduate of the University of NSW with a B.Eng. Mining; o I am a graduate from Curtin University with a B.Eng. Civil; o I am a graduate from the University of Queensland with a Masters of Business Administration (MBA); o I am a member, certified professional (MAusIMM(CP)), of the Australian Institute of Mining and Metallurgy (AusIMM); o My relevant experience includes over 30 years of experience in mining evaluations, mine operations management, feasibility studies and financial analysis of mining operations and mineral projects nationally and internationally in a variety of commodities including more than a decade of industry work experience in laterite style deposits (bauxite, nickel, scandium). • I carried out a site inspection from 18 April 2016 until the 24 April 2016. • I am responsible for Sections 15, 16 and 21 and jointly responsible for Sections 1 and 24 of the Technical Report. • I am independent of the issuer as described in section 1.5 of the instrument. • I have been involved with various studies on Ambatovy since 2016. • I have read National Instrument 43-101. The parts of the Technical Report for which I am responsible have been prepared in compliance with this Instrument. • At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the part of Technical Report for which I am responsible, contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
Effective Date: 30 June 2018 Dated at Brisbane, Australia this 10th Day of January 2019
“original document signed”
Stewart Lewis CEO – IMC Mining Pty Ltd
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Certificate of Qualified Person – Paul O’Callaghan As a Qualified Person of this Technical Report titled“ NI 43-101 Technical Report for the Ambatovy Project, Madagascar”, prepared for Sherritt International Corporation with an effective date of30 June 2018 (the “Technical Report”), I, Paul O’Callaghan do hereby certify that: • I am a Principal Mining Engineer with CSA Global Pty Ltd at its head office at Level 2, 3 Ord Street, West Perth, WA 6005, Australia. • I am a professional mining engineer having graduated with a Bachelor of Engineering (Mining) from the W.A. School of Mines, Kalgoorlie, W.A. (1991). • I am a Fellow of the Australasian Institute of Mining and Metallurgy (AusIMM). • I have practised my profession as a Mining Engineer for the past 25 years in the mineral resources sector and engaged in the assessment, development and operation of numerous mineral projects both within Australia and overseas. • I have read the definition of“ Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43- 101) and past relevant work experience, I fulfil the requirements to be a “Qualified Person” for the purposes of NI 43-101. • I am responsible for Sections 22, 23, 25, 26 and 27 and jointly responsible for Sections 1 and 18 of the Technical Report. • I have not visited the property that is the subject of the Technical Report. • I am independent of the issuer as described in Section 1.5 of NI 43-101. • I have had no prior involvement with the property that is the subject of the Technical Report. • I have read NI 43-101, and the Technical Report has been prepared in compliance with NI 43-101. • As of the effective date, the Technical Report, to the best of my knowledge, information, and belief, contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
Effective Date 30 June 2018 Dated at Perth, Australia this 10 January 2019
[“Signed”] {Paul O’Callaghan}
Paul O’Callaghan, BEng (Mining), FAusIMM Principal Mining Engineer CSA Global Pty Ltd
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Certificate of Qualified Person – Adrian Martinez I, Adrian Martinez Vargas, PhD., P.Geo. (BC), do hereby certify that: • I am employed as a Senior Resource Geologist with the firm of CSA Global Canada Geosciences Ltd located at 365 Bay Street, Suite 501, Toronto, Ontario, Canada M5H 2V1. • I graduated with a degree in Bachelor of Science, Geology, from the Instituto Superior Minero Metalurgico de Moa (ISMM), 2000. I have a Postgraduate Specialization in Geostatistics (CFSG) MINES ParisTech, 2005, and a PhD on Geological Sciences, Geology, from the ISMM in 2006. • I am a Professional Geoscientist (P.Geo.) registered with the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC, Licence # 43008) and the Association of Professional Geoscientists of Ontario(APGO, Membership # 2934). • I have worked as a geologist since my graduation 18 years ago, and I have experience with mineral projects of nickel and cobalt laterites, including Mineral Resource estimation. • I have read the definition of“ qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that because of my education, affiliation with a professional association (as defined in NI43- 101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. • I visited the property that is the subject of the Technical Report for six days from 25 to 30 June 2018. • I am an author of the technical report titled: “NI 43-101 Technical Report for the Ambatovy Project, Madagascar”, prepared for Sherritt International Corporation with an effective date of30 June 2018 (the “Technical Report”). • I am responsible for Sections 7 to 12, and 14 of the Technical Report • I have had no prior involvement with the properties that are the subject of the Report. • As of the effective date of the Report, to the best of my knowledge, information and belief, the Report contains all scientific and technical information that is required to be disclosed to make the Report not misleading. • I am independent of the issuer applying all the tests in section 1.5 of NI 43-101. • I have read NI 43-101 and Form 43-101F1, and the Report has been prepared in compliance with that instrument and form.
Effective Date30 June 2018 Dated at Ottawa, Ontario, Canada this 10 January 2019
[“Signed & Sealed”] {Adrian Martinez Vargas}
Adrian Martinez Vargas, PhD., P. Geo. Senior Resource Geologist CSA Global Canada Geosciences Ltd
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Certificate of Qualified Person – Glen Smith
• I, A. Glen Smith, residing at 4 Charlton Crescent, Sherwood Park, Alberta, am a chemical engineer. • I am a co-author of the technical report “NI 43-101 Technical Report Ambatovy Nickel Project, Madagascar” prepared for Sherritt International Corporation with an effective date of30 June 2018 (the “Technical Report”). • I hold a Bachelor of Chemical Engineering from Lakehead University. I am licensed as a Professional Engineer by the Association of Professional Engineers and Geoscientists of Alberta (#47824) and Professional Engineers Ontario (#43064401). I hold over 25 years of experience in the design and operation of pressure hydrometallurgical plants. I have read National Instrument 43-101‘s definition of “qualified person” and certify that, by reason of my education, registration with a professional association and relevant past work experience, I fulfill the requirements of a qualified person for the purposes of National Instrument 43-101. • I have visited the property that is the subject of the Technical Report. I worked on a daily basis at the Ambatovy Plant Site from August 2009 to May 2016 in the positions of Operations Manager, Refinery and Manager, Technical Services. I continue to support the operation as a Consulting Engineer for Sherritt Technologies. and have visited the mine and plant site in June 2017, December 2017, June-July 2018 and plant only from September to November 2018 for a combined period of approximately four months. • I am responsible for Sections 13, 17, 19 and 20 and jointly responsible for Sections 1, 18 and 24 of the Technical Report. • As an employee of Sherritt International Corporation, I am not independent of the issuer, as defined in Section 1.4 of National Instrument 43-101. • I have worked for Sherritt International Corporation or Dynatec Corporation since 1995; I wasa member of the Ambatovy Owners’ Project Group from 2004 to 2009 covering pre-feasibility to detailed design and have worked at the property that is the subject of this technical report from August 2009 through May 2016. • I have read National Instrument 43-101 and the final Technical Report and confirm the parts of the Technical Report for which I am responsible have been prepared in compliance with National Instrument 43-101. • As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the technical report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
Effective Date 30 June 2018 Dated this 10th day of January 2019, in Fort Saskatchewan, Alberta, Canada.
[“Signed & Sealed”] {Glen Smith}
A. Glen Smith, B. Eng., P. Eng., Sherritt International Corporation
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Contents
Report prepared for ...... I Report issued by ...... I Report information ...... I Author and QP Signatures ...... I CERTIFICATES ...... II Certificate of Qualified Person – Michael Elias ...... II Certificate of Qualified Person – Stewart Lewis ...... III Certificate of Qualified Person – Paul O’Callaghan ...... IV Certificate of Qualified Person – Adrian Martinez ...... V Certificate of Qualified Person – Glen Smith ...... VI 1 SUMMARY ...... 1 1.1 Property Description and Location ...... 1 1.2 Accessibility, Climate, Local Resources, Infrastructure and Physiography ...... 2 1.3 History ...... 3 1.4 Geology and Mineralisation ...... 3 1.5 Exploration and Drilling ...... 4 1.6 Sampling and Analysis and Security of Samples ...... 5 1.7 Resource and Reserve Estimate ...... 6 1.7.1 Mineral Resource Estimate ...... 6 1.7.2 Mineral Reserve Estimate ...... 6 1.8 Mining Operations ...... 8 1.9 Processing Facilities ...... 9 1.10 Capital and Operating Costs...... 10 1.11 Conclusions and Recommendations...... 10 2 INTRODUCTION ...... 12 2.1 Issuer ...... 12 2.2 Terms of Reference ...... 12 2.2.1 CSA Global Terms of Reference ...... 12 2.2.2 IMC Terms of Reference ...... 13 2.3 Principal Sources of Information ...... 13 2.4 Qualified Person Section Responsibility ...... 13 2.5 Qualified Person Site Inspections ...... 14 2.5.1 IMC Mining Pty Ltd ...... 14 2.5.2 CSA Global Pty Ltd ...... 14 2.5.3 Sherritt International...... 15 2.5.4 Current Site Visit ...... 15 3 RELIANCE ON OTHER EXPERTS ...... 16 4 PROPERTY DESCRIPTION AND LOCATION ...... 17 4.1 Location ...... 17 4.2 Property Description ...... 17 4.3 Royalties, Back-in Rights and Other Payments ...... 19 4.4 Permits ...... 20 4.5 Environmental Liabilities ...... 20
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4.6 Other Risks ...... 20 5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...... 22 5.1 Physiography...... 22 5.2 Access ...... 23 5.3 Population Centre ...... 23 5.4 Climate ...... 23 5.5 Local Resources and Infrastructure ...... 24 6 HISTORY ...... 25 6.1 GENiM ...... 25 6.2 PD Madagascar SARL (Phelps Dodge) ...... 25 6.4 2005 Mineral Resource and Mineral Reserve Estimates ...... 26 6.6 2013 Mineral Resource and Mineral Reserve Estimates ...... 29 6.6.1 2013 Mineral Resource Estimate ...... 29 6.6.2 2013 Mineral Reserve Estimate ...... 30 7 GEOLOGICAL SETTING AND MINERALISATION...... 31 7.1 Regional Geology ...... 31 7.2 Local and Property Geology ...... 33 7.2.1 Ambatovy Geology ...... 34 7.2.2 Analamay Geology...... 35 7.2.3 Mineralisation ...... 37 8 DEPOSIT TYPE...... 39 9 EXPLORATION ...... 40 10 DRILLING ...... 41 10.1 Drilling Procedures ...... 41 10.2 Logging and Sampling Methodology ...... 43 10.2.1 Logging and Sampling ...... 43 10.2.2 Density Measurement ...... 44 11 SAMPLE PREPARATION, ANALYSES AND SECURITY ...... 47 11.1 2003 to 2004 Sample Preparation and Analysis ...... 47 11.2 2004 to 2008 Sample Preparation and Analysis ...... 48 11.3 2009 to 2017 Sample Preparation and Analysis ...... 48 11.4 DMSA Validation and QAQC Programs (2003 to 2008) ...... 50 11.4.1 Initial Check of Phelps Dodge Assays ...... 50 11.4.2 Analytical Method and Laboratory Validation ...... 50 11.4.3 Quality Control ...... 53 11.4.4 Interlaboratory Cross-Checks ...... 54 11.5 2009 Round-Robin Laboratory Tests ...... 55 11.6 Validation Test of AMSA ICP Lab ...... 55 11.7 Ambatovy QAQC (2009 to 2017) ...... 55 11.7.1 Standards ...... 56 11.7.2 Duplicates ...... 58 11.7.3 Blanks ...... 59 11.7.4 Summary of 2009–2017 QAQC Results ...... 60 11.8 Inter-Laboratory Pulp Check Samples ...... 60 11.9 Qualified Person’s Opinion and Conclusions ...... 62
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12 DATA VERIFICATION ...... 63 13 MINERAL PROCESSING AND METALLURGICAL TESTING...... 66 13.1 Summary...... 66 13.2 2003 and 2004 Metallurgical Testwork ...... 66 13.2.1 Sample Selection ...... 66 13.2.2 Metallurgical Testwork ...... 66 13.2.3 Ore Variability Testwork ...... 67 13.3 Ore Preparation Pilot Plant Mineral Processing Campaigns ...... 67 13.4 Pressure Leach Testwork ...... 67 13.4.1 Batch Testwork ...... 67 13.4.2 Continuous Testwork ...... 68 13.5 Low-Grade Ore Testwork ...... 69 14 MINERAL RESOURCE ESTIMATES ...... 70 14.1 Introduction ...... 70 14.2 Informing Data ...... 70 14.2.1 Drillhole Data...... 70 14.2.2 Conversion from Laborde to UTM Coordinates ...... 70 14.2.3 Adjustments to Compensate for Bias in some Aluminum and Cobalt Assays ...... 71 14.3 Geological Modelling ...... 73 14.4 Statistical Analysis and Grade Capping ...... 74 14.5 Variograms and Spatial Variability ...... 74 14.6 Estimation of Lithology Proportions ...... 75 14.7 Grade Estimation ...... 76 14.8 Density Estimation ...... 77 14.9 Block Model Validation ...... 79 14.10 Resource Classification ...... 81 14.11 Mineral Resource Estimates ...... 82 14.11.1 Reasonable Prospects of Economic Extraction...... 82 14.11.2 Mineral Resource Reporting...... 83 14.11.3 Factors that may Affect the Mineral Resource...... 83 14.12 Comparison with the previous Mineral Resource Estimates ...... 83 15 MINERAL RESERVE ESTIMATE...... 86 15.1 Introduction ...... 86 15.2 Modifying Factors ...... 86 15.2.1 Run of Mine (ROM) – Run of Preparation (ROP) Modifying Factors ...... 87 15.2.2 Sheeting...... 91 15.2.3 Selection Based on Resource Classification and Domains ...... 92 15.2.4 Stockpile, Block Binning and Cut-Off Grade ...... 92 15.2.5 Block Binning and Block Allocation Introduction ...... 93 15.2.6 Cut-Off Grade ...... 96 15.2.7 Moisture ...... 97 15.3 Optimisation ...... 97 15.4 Mineral Reserve Estimate – Insitu Resources ...... 102 15.5 Mineral Reserve Estimate – Existing Stockpile ...... 104 15.6 Mineral Reserve Statement ...... 106 15.6.1 QP Comments...... 106 16 MINING METHODS ...... 107
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16.1 Introduction ...... 107 16.2 Pit Limits ...... 107 16.3 Pit Design/Stages Definition ...... 107 16.3.1 Five-Year Detailed Pit Designs ...... 107 16.3.2 Tactical (Minesched) Five-Year Plan and Schedule ...... 108 16.3.3 Phase 2 Life of Mine Design and Schedule ...... 109 16.3.4 Stockpile Strategy ...... 110 16.4 Development Plan ...... 111 16.5 Waste Disposal ...... 111 16.6 Grade Control ...... 115 16.7 Mining Fleet ...... 118 16.7.1 Load, Haul and Excavate...... 118 16.7.2 Ancillary and Support Fleet ...... 118 16.7.3 Mining Fleet Allocation (P1/P2 Fleet)...... 118 16.8 Manning...... 119 16.9 Mine Schedule ...... 121 16.9.1 Scheduled Quantities ...... 123 17 RECOVERY METHODS ...... 128 17.1 Ore Preparation Plant ...... 129 17.2 Leach and Sulphide Precipitation Plant ...... 129 17.2.1 Ore Slurry Thickening ...... 129 17.2.2 Pressure Acid Leaching ...... 129 17.2.3 Slurry Neutralisation and Solids Wash Circuit ...... 129 17.2.4 Raw Liquor Neutralisation ...... 129 17.2.5 Sulphide Precipitation ...... 130 17.2.6 Tailings Handling ...... 130 17.3 Nickel and Cobalt Refinery ...... 130 17.3.1 Sulphide Leach and Impurity Removal ...... 130 17.3.2 Solvent Extraction and Zinc Precipitation...... 131 17.3.3 Nickel and Cobalt Reduction ...... 131 17.3.4 End Solution Treatment ...... 131 17.3.5 Utilities and Major Materials...... 131 18 PROJECT INFRASTRUCTURE ...... 132 18.1 Mine Site...... 132 18.1.1 Roads ...... 132 18.1.2 Camps ...... 132 18.1.3 Offices and Workshop ...... 132 18.1.4 Ore Preparation Plant ...... 133 18.1.5 Stockpiles ...... 133 18.1.6 Dumps ...... 133 18.1.7 Mine Runoff Control Dams ...... 133 18.1.8 Pumping Station ...... 133 18.1.9 Power Supply...... 134 18.1.10 Pipeline ...... 134 18.2 Plant Site and Toamasina Area ...... 134 18.2.1 Steam and Power ...... 134 18.2.2 Emergency Power Supply ...... 135 18.2.3 Water Supply ...... 135 18.2.4 Petroleum Products, Supply and Storage ...... 135
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18.2.5 Bulk Materials Supply and Stockpiling...... 135 18.2.6 Ammonia ...... 136 18.2.7 Tailings Disposal ...... 136 18.2.8 Air Strip...... 136 18.2.9 Port ...... 136 18.2.10 Road and Rail ...... 137 18.2.11 Camps and Housing ...... 137 19 MARKET STUDIES AND CONTRACTS ...... 138 19.1 Market Studies ...... 138 19.1.1 Nickel ...... 138 19.1.2 Cobalt ...... 139 19.2 Contracts ...... 140 19.3 QP Comments ...... 140 20 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT ...... 141 20.1 Summary of Environmental Studies and Challenges ...... 141 20.2 Waste and Tailings Disposal, Site Monitoring and Water Management ...... 142 20.2.1 Waste Disposal ...... 142 20.3 Site Monitoring ...... 142 20.4 Water Management ...... 142 20.5 Tailings Disposal ...... 142 20.6 Environmental Permits ...... 143 20.7 Social and Community Requirements ...... 143 20.8 Mine Closure ...... 144 21 CAPITAL AND OPERATING COSTS ...... 145 21.1 Capital Costs ...... 145 21.2 Operating Costs ...... 145 21.2.1 Variable Costs ...... 145 21.2.2 Fixed Costs...... 145 21.3 Mine Operating Costs ...... 146 21.4 Equipment Operating Costs ...... 146 21.5 Administration and Overhead Costs...... 148 21.6 Operating Cost Summary ...... 148 22 ECONOMIC ANALYSIS ...... 151 23 ADJACENT PROPERTIES ...... 152 24 OTHER RELEVANT DATA AND INFORMATION...... 153 24.1 Production Targets ...... 153 24.2 Political Environment ...... 153 25 INTERPRETATION AND CONCLUSIONS ...... 154 25.1 Upside Potential ...... 154 25.2 Downside Risk ...... 154 26 RECOMMENDATIONS ...... 156 26.1 In-Pit Screening of Coarse Saprolite ...... 156 26.2 Ground Penetrating Radar ...... 156 26.3 Mining ...... 156 26.4 Hydrogeology ...... 156
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26.5 Geology...... 156 27 REFERENCES ...... 157
Figures Figure 1: Regional location map...... 18 Figure 2: Property map ...... 19 Figure 3: Aerial image of the mine activity area showing surrounding conservation area ...... 22 Figure 4: Temperature average at Property ...... 23 Figure 5: Rainfall at Property ...... 24 Figure 6: Current site plan, Ambatovy Project ...... 28 Figure 7: Ambatovy Project plant site, tailings management facility and port at Toamasina...... 29 Figure 8: Ambatovy regional geological map ...... 32 Figure 9: Gabbroic magmatic injection ...... 33 Figure 10: Details of gabbroic magmatic injection ...... 33 Figure 11: Locations of the Deposits ...... 34 Figure 12: Ambatovy bedrock geology...... 36 Figure 13: Analamay bedrock geology ...... 37 Figure 14: Ambatovy lateritic profile and vertical grade variation ...... 39 Figure 15: Ambatovy drillhole locations until 2017 ...... 42 Figure 16: Analamay drillhole locations until 2017 ...... 43 Figure 17: Sample preparation flowsheet at Ambatovy laboratory...... 49 Figure 18: Internal standard 1 for nickel (%) from 2009 to 2017 ...... 56 Figure 19: Internal standard 2 for nickel (%) from 2009 to 2017 ...... 57 Figure 20: Standard for cobalt (%) from 2009 to 2017 ...... 57 Figure 21: Coarse duplicate for nickel (%) from 2009 to 2017 ...... 58 Figure 22: Fine duplicate for nickel (%) from 2009 to 2017 ...... 58 Figure 23: Coarse and fine duplicate for cobalt (%) from 2009 to 2017 ...... 59 Figure 24: Fine blank for nickel from 2009 to 2017 ...... 59 Figure 25: Coarse blank for nickel from 2009 to 2017 ...... 60 Figure 26: Internal assays vs external assays for nickel since 2014 ...... 61 Figure 27: Internal assays vs external assays for nickel since 2015 ...... 61 Figure 28: Internal assays vs external assays for cobalt since 2014 ...... 62 Figure 29 Mining operations at the Ambatovy pit and exposure of the nickel laterite profile...... 63 Figure 30 Drilling and sampling logging (upper row and middle left); sampling and logging facility (middle right), and sample storage facility (lower row) ...... 64 Figure 31 Density sample (left) and wet weight determination (right) ...... 65 Figure 32: Drillholes used in the current resource estimation of the Ambatovy Deposit ...... 71 Figure 33: Drillholes used in 2017 Mineral Resource estimation of the Analamay Deposit ...... 72 Figure 34: Schematic vertical cross-section in Ambatovy with the different geological units ...... 73 Figure 35: Schematic representation of Ambatovy block model with the geological codification ...... 73 Figure 36: Ni variogram inside ferralite wireframe for ferralite (above) and gabbro (below) ...... 75 Figure 37: Average nickel grade, calculated over a 10 m moving window, vs height above the base of the ferralite ...... 77 Figure 38: Drillhole and block model section (east-west) with estimated nickel in Ambatovy ...... 79 Figure 39: Drillhole and block model section (east-west) with estimated nickel in Analamay ...... 79 Figure 40: Swath plots (above and lower left) and global change of support (lower right) of kriged nickel grades in the geological domain of limonites+ferralites (and lithology domain/proportion of limonites+ferralites) of the Amabatovy Deposit ...... 80
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Figure 41: Swath plots (above and lower left) and global change of support (lower right) of kriged nickel grades in the geological domain of limonites+ferralites (and lithology domain/proportion of limonites+ferralites) of the Analamay Deposit ...... 81 Figure 42: Left – three passes of kriging; Right – passes manually smoothed to use for resource classification ...... 82 Figure 43: Left – three passes of kriging; Right – passes manually smoothed to use for resource classification ...... 82 Figure 44: Resource block to mining block (gabbro >10%) ...... 86 Figure 45: Resource block to mining block (gabbro <= 10%) ...... 87 Figure 46: Ferralite + Saprolite <= 10% ...... 87 Figure 47: Ambatovy process flow and sampling points ...... 87 Figure 48: Material flow for tonnage reconciliation ...... 89 Figure 49: Sheeting material flowchart ...... 90 Figure 50: Block bins NiEq based ...... 94 Figure 51: Consolidated bin allocation based on revenue/cost ratio vs ROP NiEq revenue ...... 95 Figure 52: Revenue/Cost ratio vs nickel grade ...... 97 Figure 53: Ambatovy optimised staged pit shells – Measured and Indicated only ...... 99 Figure 54: Analamay optimised staged pit shells – Measured and Indicated only ...... 100 Figure 55: Pit floor not saprolite floor – Ambatovy ...... 101 Figure 56: Pit Floor Not Saprolite Floor - Analamay ...... 102 Figure 57: Existing stockpiles ...... 105 Figure 58: Five-year plan total movement by material type ...... 108 Figure 59: Five-year plan ore mined by pit ...... 109 Figure 60: Stockpile balance ...... 111 Figure 61: Ambatovy waste dump locations ...... 113 Figure 62: Analamay dumps and sediment containment ...... 114 Figure 63: Block allocation run ...... 117 Figure 64: Labour roster defintion ...... 120 Figure 65: Ambatovy LOM PAL feed tonnes ...... 122 Figure 66: Ambatovy LOM PAL acid consumption ...... 123 Figure 67: Ambatovy Project mine and process flowsheet ...... 128 Figure 68: Equipment Operating Cost – primary excavator ...... 147 Figure 69: Cost by cost centre ...... 148 Figure 70: Mining Operating Costs per year by category ...... 149
Tables Table 1: Total estimated Mineral Resources for the Ambatovy Project (inclusive of Mineral Reserves) above a cut-off grade of 0.45% Ni (with an effective date of 30 June 2018)...... 6 Table 2: Mineral Reserves (nominal ROM) for the Ambatovy Project as at 30 June 2018 ...... 8 Table 3: Mineral Reserves (ROP) for the Ambatovy Project as at 30 June 2018 ...... 8 Table 4: Waste material for the Ambatovy Project as at 30 June 2018 ...... 8 Table 5: Qualified Person section responsibility ...... 14 Table 6: Measured Mineral Resources at 0.8% Ni cut-off – classified by WGM (October 2004) ...... 26 Table 7: Indicated Mineral Resources at 0.8% Ni cut-off – classified by WGM (October 2004) ...... 26 Table 8: Measured and Indicated Mineral Resources at 0.8% Ni cut-off – classified by WGM (October 2004) ...... 26 Table 9: Inferred Mineral Resources at 0.8% Ni cut-off – classified by WGM (October 2004) ...... 27 Table 10: Total estimated Mineral Resources for the Ambatovy Project (inclusive of Mineral Reserves) above a cut-off grade of 0.6% nickel (with an effective date of 31 December 2013) ...... 30 Table 11: Estimated Mineral Reserves for the Ambatovy Project as at 31 December 2013 ...... 30 Table 12: Ambatovy drillhole statistics until 2017 ...... 41 Table 13: Codes for drill core logging ...... 45 Table 14: Assay results of six laboratories ...... 51 Table 15: Inter-laboratory comparison of assays results ...... 53
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Table 16: Certified reference standard analyses ...... 53 Table 17: Typical analytical report ...... 54 Table 18: Analytical procedures summary ...... 54 Table 19: Internal QAQC ...... 55 Table 20: Internal QAQC – UltraTrace ...... 55 Table 21: Failure from 2009 to 2017...... 60 Table 22: Summary statistics of sample data used for resource estimation in the Ambatovy Deposit ...... 74 Table 23: Summary statistics of sample data used for resource estimation in the Anamalay Deposit ...... 74 Table 24: Variogram models used to interpolate proportions of ferralite and ferricrete, saprolite, bedrock and gabbro ...... 76 Table 25: Search parameters ...... 77 Table 26: Average density values assigned to geological domains and its corresponding lithologies...... 78 Table 27: Average densities by geological domain deduced from geochemical formula ...... 78 Table 28: Average densities by geological domain deduced from density samples collected in Ambatovy ...... 78 Table 29: Reconciliation of tonnage using density assigned as average vs density interpolated in the block ...... 78 Table 30: Mineral Resource estimate1 for Ambatovy Project (inclusive of Mineral Reserves) above a reporting cut-off2 grade of 0.45 %Ni (with an effective date of 30 June 2018) ...... 83 Table 31 Comparison of 31 December 2013 and 30 June 2018 Mineral Resources ...... 85 Table 32: Tonnage Resource – ROP modifying factors (data shown represents million tonnes in reconciliation period) ...... 90 Table 33: 2018 Resource model to ROP (PAL feed) modifying factors ...... 91 Table 34: NiEq and acid consolidated bins ...... 95 Table 35: Optimisation modelling inputs ...... 98 Table 36: Mineral Reserve (nominal ROM – excludes sheeting) ...... 104 Table 37: Mineral Reserve (ROP) ...... 104 Table 38: Waste material within final pit ...... 104 Table 39: Existing stockpile tonnes ...... 105 Table 40: Pit design parameters ...... 107 Table 41: LP LOM production constraints ...... 110 Table 42: Manpower summary – staff numbers...... 121 Table 43: Manpower summary – maximum and minimum Operations and Maintenance numbers...... 121 Table 44: Waste mined ...... 124 Table 45: ROM data (excludes sheeting) ...... 125 Table 46: ROP tonnes (direct) ...... 126 Table 47: PAL feed ...... 127 Table 48: Uninflated annual operating costs (2018–2022) ...... 145 Table 49: Commodity price assumptions used in the cash flow model (average price 2018–2022) ...... 145 Table 50: Total operating cost (2018–2047, uninflated) ...... 146 Table 51: Mining office administration cost centres ...... 148 Table 52: Operating cost by year ...... 150
Appendices Appendix 1: Variogram Models
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1 Summary
The Ambatovy Project is a producing nickel and cobalt project designed for an annual production capacity of 60,000 tonnes (t) of nickel and 5,600 t of cobalt and approximately 210,000 t of ammonium sulphate, with an estimated life of approximately 26 years. Annual production rates are projected to vary over the life of the mine, largely dependent on ore grades. Ambatovy Minerals Société Anonyme (AMSA) and Dynatec Madagascar Société Anonyme (DMSA) (known together as the Ambatovy Joint Venture) are operators of the mine and the plant respectively. Sherritt International Corp. (Sherritt) indirectly holds a 12% interest in each of AMSA and DMSA. Sumitomo Corporation (Sumitomo) and KoreaResources Corporation (KORES) indirectly hold the remaining interest in the Ambatovy Joint Venture. Sherritt acts as the operator of the Ambatovy Joint Venture. This Technical Report provides the scientific and technical information concerning the update of the development and production activities, updated Mineral Resources and Mineral Reserves estimates and updated life of mine (LOM) and production schedule for the Ambatovy Project.
1.1 Property Description and Location The Ambatovy Project’s mine site (“the Property”) is located in Madagascar, 80 km east of the capital city of Antananarivo and 11 km north of the town of Moramanga. The Property covers 143.7 km2 and consists of 368 contiguous 625 m x 625 m blocks. All necessary titles and permits for the Property are held by AMSA. AMSA was granted Exploitation Permit No. 459 (Exploitation Permit 459) effective 7 September 2006 by the Ministère de l’Energie et des Mines. Exploitation Permit 459 grants AMSA the right to extract nickel, cobalt, copper, chrome and platinum from the Property for a period of 40 years dating from 7 September 2006. In order to maintain Exploitation Permit 459 in good standing, AMSA must file an annual report, pay certain annual administration fees of approximately US$46,300 (107,345,600 MGA) payable under the Mining Code, and remain in compliance with its obligations under the Environmental and Social Management Plan (ESMP), which is attached as an appendix to its environmental permit. Under Malagasy law, the Ambatovy Joint Venture is considered a foreign company and therefore is not entitled to purchase land in Madagascar. The surface rights of the Property consist of a 50-year lease with the Malagasy State. The contract for the lease is registered with the Malagasy fiscal authority and the lease is registered at the appropriate local land titles registry. The lease is conditional on the payment of annual fees of US$250,000 and compliance with land usage, which is mining and the establishment of a forest conservation area according to the Ambatovy Joint Venture’s environmental commitment. The Property is subject to the following payments, in addition to those mentioned immediately above: • Annual property tax on land (IFT), which is equivalent to 1% of the market value of the land (land taxes are capped at approximately US$100,000 pursuant to the Loi sur les Grands Investissements Miniers – LGIM). • Annual property tax on buildings, which is equivalent to 1% of the annual rental value of the buildings. • Mining royalties payable on the production which are equivalent to 1% of finished metal and metal sulphide by-product sales pursuant to the Mining Code and the LGIM. The Ambatovy Project’s environmental permit was issued on 1 December 2006 and is conditional upon the implementation of an ESMP and is subject to an annual review process. The ESMP requires the Ambatovy Joint Venture to, among other things, operate a “no net loss” biodiversity strategy based on measures to avoid or minimise impacts. The ESMP also requires the utilisation of offsets to compensate for unavoidable biodiversity loss.
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In addition to the environmental permit, the Ambatovy Joint Venture is also obliged to obtain forest cutting permits for each parcel cleared. The Ambatovy Joint Venture has agreed to work with the Malagasy Government to establish a US$50 million environmental surety which could be accessed to compensate the Government of Madagascar should the Ambatovy Joint Venture fail to correct any material breach of applicable environmental laws, regulations and permits governing the Ambatovy Joint Venture’s environmental obligations, including closure obligations. The Ambatovy Project currently has the necessary exploitation and exploration permits and the infrastructure in place, including bridge access, roads, maintenance shops, power supplies, offices and housing to support its mining operations. Environmental management actions are accompanied by social support and compensation measures. Currently, comprehensive social baseline surveys are being undertaken to provide the basis for monitoring local livelihoods. A reclamation and closure plan has been prepared by consultants and will be put in place for all aspects of the Ambatovy Project. The cost of project remediationand reclamation in connection with the Property is estimated at US$100 million.
1.2 Accessibility, Climate, Local Resources, Infrastructure and Physiography The Ambatovy and Analamay nickel deposits (together, “the Deposits” and each a “Deposit”) are located on a plateau at an elevation of approximately 1,100 m above sea level (ASL). The topography of the Deposits varies from gently undulating hills to a steeply dissected remnant plateau. Locally, the relief is 100 m. The plateau surface is fairly uneven with numerous depressions that form ephemeral pools. Small headwater streams originate in the mine area and flow away in all directions as part of a six-basin configuration. The Property is covered with natural forests. The surrounding area includes intact and degraded forests and scrublands, areas dominated by grasses, eucalyptus plantations, woodlots and rice paddies. The soils in the mine region are generally known as laterites, which are highly weathered iron rich tropical soils. Local resources include a small nearby aggregate quarry, the Mangoro River, which is the water supply for the Property, and a 138 kV powerline which passes approximately 8 km south of the Property. Local infrastructure includes the town of Moramanga, which has a population of approximately 30,000, and a 1 m gauge railway line that runs from Antananarivo to Moramanga and to the port of Toamasina. The port is located approximately 12 km from the Plant Site (as defined below) and is connected to the Plant Site by rail. The port is used to export the refined metals and ammonium sulphate produced by the Ambatovy Project. The Ambatovy Joint Venture has access to the requisite mining personnel through the use of the local population for unskilled and semi skilled labour, as well as having on site residential facilities for necessary expatriate and national senior staff employees. The capital city, Antananarivo, is serviced by commercial air flights from Paris, France, three times per week and daily from Johannesburg, South Africa. Regular flights also run between Antananarivo and Saint Denis, Réunion, Mahébourgh, Mauritius, Nairobi, Kenya and and Addis Ababa, Ethiopia. The Property, located near the town of Moramanga, is approximately 120 km by road from Antananarivo. Major asphalt paved national roads link Antananarivo to Moramanga with a high quality 11 km gravel road leading from Moramanga to the Property. A charter plane, operated on behalf of the Ambatovy Joint Venture, flies regularly between Antananarivo, Toamasina and Moramanga. Buses run daily between Moramanga and the Property to transport employees to the mine. The climate is equatorial to tropical, with average monthly rainfall which historically has exceeded 135 mm in the period from November to March and has been less than 66 mm in the period from April to October,
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evaporation from 600 mm to 800 mm and temperature varying from 5°C to 35°C with an average temperature of 17°C. The cyclonic season occurs from January to March.
1.3 History The presence of the Deposits was first noted by the Malagasy Service Géologique in 1960. Limited pitting and auger drilling were undertaken to assess the surface iron potential. In the early 1970s, a consortium composed of Société Le Nickel, Ugine Kuhlman, Anglo American and the Bureau des Recherches Géologiques et Minières (the BRGM, together “GENiM”) carried out a major exploration program that included 368 vertical diamond drillholes over the Deposits, an aerial photographic survey, geological mapping and geochemical surveys. The next major exploration program was conducted by PD Madagascar SARL (Phelps Dodge) between 1995 and 1998. The work included 369 vertical diamond drillholes, test pits, various surveys and metallurgical testwork for the region known as Ambatovy West. This work culminated in a feasibility study. In 2003, Phelps Dodge and Dynatec Corporation (Dynatec) signed a joint venture agreement to continue the development of the Property, which later resulted in Dynatec receiving a 53% interest in the Ambatovy Joint Venture. Dynatec and Phelps Dodge conducted exploratory drilling and started to develop a thorough feasibility study with a detailed environmental and social impact assessment. In 2005, Dynatec acquired the remaining 47% interest in the Ambatovy Joint Venture from Phelps Dodge and Sumitomo took up a 25% stake in the Ambatovy Joint Venture. In February 2005, a technical report prepared in accordance with NI 43-101 was completed (the 2005 Technical Report). The 2005 Technical Report was revised in May 2005. Definition drilling continued in 2006 and 2007, mostly in the Ambatovy Central sub block, with 100 HQ holes drilled for a total of 5,675 m. A subsequent NI 43-101 Technical Report was filed in October 2011( “the 2011 Technical Report”). In 2007, the Ambatovy Project was certified by the Government of Madagascar under the LGIM. Sherritt acquired Dynatec in June 2007 for approximately US$1.7 billion. Sherritt assumed Dynatec’s ownership position in the Ambatovy Joint Venture and was named project operator. In 2008, the Ambatovy Joint Venture received construction permits for work at the port and Plant Site. Definition drilling started again, mainly on the Ambatovy West sub-block. A total of 168 drillholes for a total of 8,605 m were drilled between June 2008 and November 2009. In 2010, commissioning of the mining equipment commenced. Mining of material for stockpiling started in July. In 2011, commissioning of the ore preparation plant (OPP) at the Property commenced, as well as commissioning of the utilities facilities at the Plant Site. This was followed in 2012 with the start of ore processing in the pressure acid leach (PAL) autoclaves and with mixed sulphides being delivered to the refinery in May 2012. The first finished nickel and cobalt briquettes were produced from the refinery in September 2012, and the Ambatovy Joint Venture continues its ramp up towards full production. In January 2014, Sherritt announced that the requirements for commercial production, defined as 70% of ore throughput nameplate capacity in the PAL circuit, averaged over 30 days, had been achieved by the Ambatovy Project. In September 2014, Sherritt filed another NI 43-101 report which provided the scientific and technical information concerning the development and production activities at that time, including an update of the Mineral Resources and Mineral Reserves estimate. At this time,Sherritt indirectly held a 40%interest in the Ambatovy Joint Venture. Sherritt subsequently reduced its interest in the Joint Venture to 12%, with the remaining interest held by Sumitomo and KORES.
1.4 Geology and Mineralisation The Ambatovy orebody consists of two large, thick, weathered ultramafic lateritic nickel deposits located approximately 3 km apart.
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The regional geological setting is a north-south belt of basic gneisses and migmatites, which are part of the high-grade metamorphic rocks underlying the eastern two-thirds of Madagascar. The dominant feature of the Ambatovy resource is the Antampombato Complex, a large intrusive that cuts the metamorphic rocks. The intrusive is composed of gabbroic to syenitic rocks with two small outer ultramafic bodies rimming the intrusive. Exploration suggests that the complex represents multiple, magmatic intrusions that commenced with ultramafic intrusive, then was followed by gabbroic intrusives and terminated with the more felsic intrusive. The Ambatovy Deposit occurs towards the southern margin of the complex and is approximately 3 km x 2.4 km and oriented in a west-northwest to east-southeast direction. A northwest trending gabbroic intrusive cuts the Ambatovy Deposit resulting in three sub-blocks: Ambatovy West, Central and Southeast. The Analamay Deposit is located at the eastern margin of the complex and is approximately 4 km x 2.8 km, oriented north-south and it also is divided into sub-blocks known as Analamay North, Central and South. Ambatovy West is cut by numerous block faults that strike northwest/southeast with a conjugate set, striking northeast/southwest. Evidence indicates that faulting continued during the laterisation. The Deposits cover an area of about 1,300 ha, and range in thickness from 20 m to 100 m, with the average thickness being approximately 40 m. Within the lateritic profile, there are three distinct zones: • Ferricrete is the uppermost layer, and forms an extremely hard, coherent crust of iron oxides up to 3 m thick and acts as a deterrent to mechanical erosion. • Limonite, referred to locally as ferralite, constitutes more than 90% of the economic grade nickel mineralisation and is predominately a spongy mass with iron concentrations of 40–50%, predominately in goethite. Enriched nickel and cobalt grades are largely achieved by depletion of other elements through the weathering process, rather than additions to the system. The nickel grade of the laterite is influenced by the nickel content of the underlying bedrock. • Saprolite lies at the base of the lateritic zone, on top of the bedrock.
1.5 Exploration and Drilling Since 2009, definition drilling has been continuously carried out by the Ambatovy Joint Venture on the Property, focusing first in completion of 50 m spacing diamond drillholes in Ambatovy West and South East for better Mineral Resource definition and increased information on the surrounding contacts with gabbroic material. Fifty-metre spaced campaigns still need to be implemented for Ambatovy Central to transform Indicated Resources to Measured Resources, but this part of the Ambatovy Deposit is expected to enter the mining sequence in 2021. Since mid-2012, definition drilling has been conducted on the Analamay Deposit, starting in the southern, north-western and north-eastern areas where sedimentation dams need to be established prior to any mining activity. Infill drilling was conducted in the Ambatovy West and Ambatovy South East areas in 2013. This program was designed to better define the orebody in the area where mining will initially take place. From 2013 until 2018, the drilling definition concentrated on infilling the 70 m spaced drillholes in the Ambatovy and Analamay Deposits to enable the conversion of Inferred Resources to Indicated and Measured Resources. Pre-production drilling is also conducted prior to mining. Twenty-metre spaced reverse circulation (RC) drillholes are drilled down to the bedrock and supplementary 10 m spaced RC drillholes are implemented on the active benches for a depth of 16 m to 24 m. This combination of pre-production drilling data is used to design the mining panels, separating the different qualities of ore.
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1.6 Sampling and Analysis and Security of Samples At the Property, the core is measured and the depths marked in metres, photographed and logged according to the principal lithologies and degrees of weathering. Sampling is done at 1 m intervals but broken at sub-metre intervals at significant lithological contacts. The highly weathered core is split in half with a knife. For the boulder saprolite, the sections with– 10 cm boulders and fines are divided in half and the boulders sawn. The +10 cm (fresh) boulders are measured and their percentage of interval recorded, but the boulders are not sampled. Up to 2009, the half-core samples were placed in plastic bags, tagged and shipped in sealed drums to UltraTrace Analytical Laboratories (UltraTrace) in Perth, Australia. The remaining half core was sheathed in polythene tubes, placed in core boxes and sent to an on-site storage area. Since 2009, all analytical measures have been done by the AMSA laboratory, while duplicate samples are sent to UltraTrace for validation. UltraTrace was acquired by the Bureau Veritas Group in 2007 and now operates as Bureau Veritas Minerals. In this report, the laboratory will continue to be referred to as UltraTrace. Since 2009, when the on-site laboratory at the mine site became operational, there has been acontinuous quality assurance and quality control (QAQC) program that uses certified standards and duplicates to monitor the accuracy and precision of the analytical data. In the years when the on-site laboratory at the mine site has been operating, there has been no systematic problem with the reliability of the analytical data, and the data have been deemed reliable for resource and reserve evaluation by the Qualified Persons responsible for Sections 14 and 15 of this Technical Report. Prior to 2009, the QAQC program in place during the drilling coordinated by DMSA was reviewed by the Qualified Person responsible for the 2005 Technical Report, who found that these data were reliable for Mineral Resource and Mineral Reserve estimation. During the 2005 Technical Report, it was established through analysis of duplicates at umpire laboratories that there were small systematic biases in the aluminium and cobalt assays from the Phelps Dodge drilling campaign in the early 2000s. These biases have been corrected by making an across the board adjustment to the aluminium and cobalt assays from the Phelps Dodge drillholes. The reliability of the earliest analytical data, from GENiM in the early 1970s, was checked using closely spaced pairs of old and new drillholes. These closely spaced pairs confirm that there is no systematic bias in the earliest analytical data, and that they can be used for Mineral Resource and Mineral Reserve estimation. The wet weight of the sample is measured, and then the sample is dried for 24 hours at 105°C. Following the measurement of the dry weight, the moisture content is determined from the difference between the weights before and after drying. Density and moisture content are taken every fifth metre down the drillhole. Prior to 2005, when the cores were sent to UltraTrace, the half HQ core samples were crushed, pulverised and split into subsamples with 55% of the subsamples assayed by UltraTrace. The remaining 45%, plus 20% of those subsamples assayed by UltraTrace, were sent to Dynatec Fort Saskatchewan (DYFS) for analysis. In 2005, DMSA started the complete sample preparation on-site at the mine site. The procedure for sample preparation remains equivalent to the UltraTrace procedure described above. The pulps produced during the sample preparation at the mine site were shipped to DYFS in Canada and UltraTrace in Australia. The facility and the procedures were kept the same when Sherritt acquired DMSA. Today, the on-site laboratory at the mine site is the primary lab for the Ambatovy Project and uses the same procedures as described above. The UltraTrace laboratory, now known as Bureau Veritas Minerals is used as the secondary lab. Samples at the Ambatovy Project are also collected for metallurgical pilot plant and batch tests. More than 26,000 total samples were processed at DYFS and UltraTrace with 14 analytical determinations carried out on each sample. A QAQC program has been implemented for the assays from the beginning by DMSA. This specific QAQC program was reviewed by AMEC in 2006. This review led to the introduction of a bench scale which
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directly measures the dry weight of core samples, as well as the weight when immersed in water, and led to improved consistency when calculating the moisture content and dry density of the samples. Early in 2009, an Inductively Coupled Plasma Atomic Emission Spectroscopy (ICP-AES) analytical laboratory was installed at the mine site. A validation test has been completed to ensure that quality from the new analytical labs was satisfactory. The samples are collected and handled at the drill site by Ambatovy Joint Venture personnel. The samples are under the direct control of Ambatovy Joint Venture personnel from the drilling site to the on-site laboratory at the mine site or until they are shipped to UltraTrace. This ensures control of custody by the Ambatovy Joint Venture from the drill sites to the analytical laboratory.
1.7 Resource and Reserve Estimate
1.7.1 Mineral Resource Estimate The Ambatovy Project’s total estimated Mineral Resources are summarised belowin Table 1. The Mineral Resources do not include any of the material that has been mined up to 30 June 2018 with the exception of material in stockpiles that has been set aside for future treatment. A fixed cut-off grade of 0.45% Ni has been applied to both the Ambatovy and Analamay deposits in situ material. Table 1: Total estimated Mineral Resources for the Ambatovy Project (inclusive of Mineral Reserves) above a cut-off grade of 0.45% Ni (with an effective date of 30 June 2018)
Deposit Classification Tonnage (Mt) Ni (%) Co (%) Measured 43.1 1.02 0.08 Indicated 66.3 0.90 0.07 Ambatovy Measured+Indicated 109.4 0.95 0.08 Inferred 27.8 0.80 0.07 Measured 9.5 0.81 0.08 Indicated 63.4 0.93 0.09 Analamay Measured+Indicated 72.9 0.91 0.08 Inferred 41.2 0.88 0.09 Measured 52.6 0.98 0.08 Indicated 129.7 0.91 0.08 ALL DEPOSITS Measured+Indicated 182.3 0.93 0.08 Inferred 69.0 0.85 0.08 Stockpiles Measured 10.7 0.81 0.06 Notes: • Figures have been rounded and hence may not add up exactly to the given totals. • Cut-off grade assumptions are given in Section 14.11. • Cobalt grade does not enter into the definition of the reporting cut-off grade since the vast majority (over 80%) of the Ambatovy Project’s revenue comes from nickel. • Resource classification as defined by the Canadian Institute of Mining, Metallurgy and Petroleum in their document “CIM Definition Standards for Mineral Resources and Mineral Reserves” of 10 May 2014.
1.7.2 Mineral Reserve Estimate The LOM ultimate pit that defined the material to be included in the schedule presented in this Technical Report has been based on mining the portion of the Measured and Indicated Mineral Resource that has been shown to be economically mineable. Within the final pit shell, there are Inferred Mineral Resources which have been included in the scheduled quantities. These Inferred Resources are not included in the Mineral Reserve. Modifying factors have been applied to in-situ resources as outlined in Item 15.
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It is important to note that the run of mine (ROM) Mineral Reserve data has been based on applying some of the modifying factors, but not all, to the Mineral Resource. The tonnage losses during mining have been modelled as reconciliation of past production work has provided a sound basis for the estimate of the diluted tonnage losses during mining. The ROM grades in the Mineral Reserve are the grades of the resource model and do not include the effect of loss and dilution– it is not possible to present the diluted ROM grades as all grade reconciliation is carried out after the material has passed through the OPP. The economic viability of mining the ore has been based on the run of ore preparation plant (ROP) (i.e. OPP product) adjusted material to define the material that ends up as feed to the PAL plant. The OPP, in simple terms, involves a relatively cheap process that enables the rejection of all ROM material that is greater than 0.8 mm in size – with the oversize material being taken to waste dumps or used for road sheeting. Some ore grade material is lost as part of this sizing process. The ROM Mineral Reserve is the portion of the Mineral Resource that has been shown to be the economically mineable part of a Measured or Indicated Mineral Resource as demonstrated by the detailed life of mine study – however the economics are determined on the basis of the ROP characteristics of each block of ore – i.e. the LOM study on which the Mineral Reserve estimate has been based incorporated information on mining, processing, metallurgical, economic and other relevant factors to demonstrate, at the time of reporting, that economic extraction can be justified. An important consideration when reporting the Ambatovy Mineral Reserve in accordance with the NI 43- 101 rules and guidelines is that the ROM Mineral Reserve, as outlined above, includes only some of the diluting materials and has no allowances for material quality characteristics that result from mining loss and dilution. The ROP Mineral Reserve includes the impact of all diluting material and allowances for losses when the material is mined, however it also includes losses and upgrades/downgrades of material through the OPP (i.e. the ROM–ROP factors). This is common to many laterite nickel projects. In order to ensure transparency, it is IMC Mining Pty Ltd’s practice to report Mineral Reserves for laterite nickel projects on the basis of the ROM material and also on the basis of the ROP material. Accordingly, the Mineral Reserve (ROM) is summarised in Table 2 and the Mineral Reserve (ROP) is summarised in Table 3. The ROM data is presented as “nominal” as it does not include the effects of road sheeting. Stockpiled material included in the ROM and ROP are assumed as recoverable material from the total stockpiled resource. The waste material that is mined in order to extract the Mineral Reserve material is summarised in Table 4. The waste has been reported on the basis of directly mined waste (which is called “waste insitu”) and waste that also includes the rehandle of sheeting material that is a necessary aspect of mining at Ambatovy (the total waste including the rehandled sheeting material is called “Waste (insitu+sheeting)”. The material that is the lost material (the difference between the Mineral Reserve (ROM) and the Mineral Reserve (ROP)) has not been reported in the Mineral Reserve statement. This material is primarily the reject material from the OPP. Allowances for the cost of hauling and stockpiling the ROP reject material have been allocated in the mine operating cost model.
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Table 2: Mineral Reserves (nominal ROM) for the Ambatovy Project as at 30 June 2018
Tonnage Ni Co Al Mg Ni metal Co metal Deposit Classification (Mt) (%) (%) (%) (%) (kt) (kt) Proven 35.3 1.01 0.08 4.69 0.60 355.3 29.2 Ambatovy Probable 47.5 0.92 0.07 4.72 1.16 435.2 34.5 Proven+Probable 82.8 0.96 0.08 4.71 0.92 790.4 63.7 Proven 8.3 0.81 0.08 4.08 0.31 67.3 6.7 Analamay Probable 55.1 0.94 0.09 3.90 0.92 516.3 47.4 Proven+Probable 63.4 0.92 0.09 3.92 0.84 583.6 54.1 Proven 43.6 0.97 0.08 4.57 0.55 422.6 35.9 All Deposits Probable 102.6 0.93 0.08 4.28 1.03 951.5 81.9 Mineral Reserve Proven+Probable 146.2 0.94 0.08 4.36 0.89 1,374.1 117.8 (ROM) in situ Stockpiles Proven 8.1 0.81 0.06 6.64 1.82 65.5 4.0 Total Mineral Proven+Probable 154.3 0.93 0.08 4.48 0.94 1,439.6 121.8 Reserve (ROM)
Table 3: Mineral Reserves (ROP) for the Ambatovy Project as at 30 June 2018
Tonnage Ni Co Al Mg Ni metal Co metal Deposit Classification (Mt) (%) (%) (%) (%) (kt) (kt) Proven 31.5 0.96 0.08 4.70 1.21 303.0 26.2 Ambatovy Probable 39.8 0.85 0.07 4.76 1.55 338.9 29.5 Proven+Probable 71.3 0.90 0.08 4.74 1.40 641.8 55.8 Proven 7.5 0.77 0.08 4.07 1.00 58.0 6.1 Analamay Probable 47.8 0.88 0.09 3.92 1.39 421.2 41.4 Proven+Probable 55.3 0.87 0.09 3.94 1.33 479.2 47.5 Proven 39.0 0.93 0.08 4.58 1.17 361.0 32.3 All Deposits Probable 87.6 0.87 0.08 4.30 1.46 760.1 71.0 Mineral Reserve Proven+Probable 126.6 0.89 0.08 4.39 1.37 1,121.1 103.3 (ROP) Stockpiles Proven 4.0 0.77 0.06 6.63 2.52 31.2 2.4 Total Mineral Proven+Probable 130.6 0.89 0.08 4.46 1.41 1,152.3 105.7 Reserve (ROP)
Table 4: Waste material for the Ambatovy Project as at 30 June 2018
Deposit Classification Tonnage (Mt) Waste (insitu) 91.7 Ambatovy Waste (insitu+sheeting) 117.9 Waste (insitu) 60.5 Analamay Waste (insitu+sheeting) 79.0 Waste (insitu) 152.2 All Deposits Waste (insitu+sheeting) 196.9
1.8 Mining Operations The Property’s facilities consist of three camps, mine infrastructure (including offices, workshops, change house, clinic, etc), the Mangoro pumping station, and the OPP and slurry transfer pumping plant. These facilities are located between the Deposits. The average electrical power requirement at the mine site is estimated to be 15 MW, which is supplied by locally installed diesel generators.
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The mining method used by the Ambatovy Joint Venture is open pit mining using an excavator/truck operation. Bench mining is executed in opened deposits using hydraulic backhoe excavators equipped with 5.4 m3 buckets and a combination of rigid and articulated haulage trucks. The mine operates 365 days per year. Ore is either directly fed to the OPP or stockpiled for future processing. Saprolite or ore containing abundant rocks are processed through a dry roller screening plant where the +140 mm rocks are rejected for use as sheeting in the mine and the –140 mm size is taken to the OPP for processing. Ore is processed through the OPP where it is slurried with water in a rotary drum scrubber and the resultant slurry is screened at <0.8 mm to reject partially or un weathered material with a high magnesium content. The screened oversize material is processed through a second scrubber and screening circuit that achieves high recovery of the limonite contained in the ore. The final reject material is used for road construction or is deposited into mined out areas. The product ore slurry is thickened and transported down a 600 mm diameter pipeline that is approximately 220 km to the Processing Plant (as defined below). The route selected for the pipeline is as direct as practical, but some significant deviations were required to avoid environmentally and culturally sensitive areas. The design for the pipeline was prepared by Pipeline Systems Incorporated, a Canadian company. The mine life at the Ambatovy Project is estimated at 26 years.
1.9 Processing Facilities The nickel and cobalt recovery process from lateritic ores that has been selected by the Ambatovy Joint Venture uses Sherritt developed and commercially proven technology which is in operation at other facilities. The processing plant (Processing Plant), located near the port of Toamasina, includes a PAL plant, a metals refinery and associated utility and ancillary plants including: water treatment, steam and power generation, hydrogen, hydrogen sulphide, sulphuric acid, air separation, limestone comminution and lime calcining and slaking. Plant Site facilities include a medical clinic, fire house, training centres, change house, canteen, stores, workshops, fuel storage, laboratory, gatehouse and main offices. The Ambatovy Joint Venture holds several long-term leases for the land on which the Plant Site and nearby tailings management facility (TMF) are located. Such leases are registered at the appropriate local land titles registries. At the Processing Plant, the slurry is thickened and pumped to a PAL circuit consisting of horizontal, mechanically agitated autoclaves. Sulphuric acid is added to the autoclaves to dissolve nickel and cobalt from the slurried ore. The discharged slurry is partially neutralised with limestone and processed through a counter current decantation wash circuit to separate the nickel and cobalt containing solution from the leach residue. The leach residue is impounded in the tailings pond following further neutralisation with limestone and lime. In the nickel and cobalt rich solution, any excess sulphuric acid is neutralised with limestone and the resulting gypsum residue is processed through the wash circuit. Nickel and cobalt are recovered from the solution by precipitation, at elevated temperature and pressure, with hydrogen sulphide gas to produce mixed sulphides. In the refining process, nickel and cobalt present in the mixed sulphide feed are leached with oxygen in autoclaves at elevated temperature and pressure. From the mixed sulphides, cobalt, nickel and other metals are dissolved and the sulphur is oxidised. Following solution purification, nickel and cobalt are separated by solvent extraction. Nickel is then recovered in powder form, and, after washing and drying, is compacted into briquettes. Cobalt is also recovered in powder form and compacted into briquettes or packaged as powder. The remaining essentially metals free solution is forwarded to the ammonium
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sulphate plant where ammonium sulphate salt is produced for sale into the fertiliser market. Other metals present in the feed, such as copper and zinc, are collected in sulphide residues and sold. The TMF, located approximately nine kilometres inland from the Plant Site, has been designed to specific international standards as set out by the Canadian Dam Association, International Commission on Large Dams and the Mining Association of Canada. The design provides for neutralisation and precipitation of the tailings slurry with limestone and lime prior to discharge to the tailings basin. The initial tailings management area comprises two adjoining basins and covers an area of approximately 760 ha. Thirteen dams have been constructed to an elevation of 46 m which have sufficient capacity to sustain operations through to 2021. Stage 3 of the TMF construction plan has commenced to perform additional raises to provide sufficient storage capacity for the remaining mine life. Containment in the tailings basin is achieved by progressive elevation of embankments encompassing the TMF. Groundwater modelling indicates that due to the low permeability of the regional soils and subsequent tailings layer that will be present, seepage losses will be low. A network of groundwater interception wells is in place to identify and prevent any contaminant migration. However, recent hydrological plume studies have indicated that continual pumping is not necessary for the foreseeable future. The entire tailings surface area will continue to be utilised through to the end of the life of the Ambatovy Project, at which point the surface will be contoured, drained and revegetated. Various tests were completed to determine the characteristics of the tailings, and to estimate the in-situ density that will be achieved in the tailings basin after deposition. A dry density of 1.0 t/m3 was subsequently selected for the tailings. Consideration of water management for the tailings basin is essential in this rainfall region. The water management plan involves containing supernatant and surface water run off to allow for solids settlement to permit ultimate discharge to the ocean. A portion of this discharge is recycled back to the Plant Site for reuse. Extensive study produced a water management system design and discharge criteria that are established to result in no adverse effects on the local environment.
1.10 Capital and Operating Costs The capital cost of the Ambatovy Project was US$7.5 billion (“Project Capital Cost”) as at December 2014. The Project Capital Cost includes operational expenditure (OPEX), financing charges, working capital costs and foreign exchange. These costs relate to the pre-production date which were capitalised as part of the Project Capital Cost. Total capital cost from 2018 to end of operations is US$1,246 million based on uninflated price. The estimated uninflated operating costs of the Ambatovy Project from 2018 to 2022 are US$588 million, of which 55% are fixed and 45%are variable costs.
1.11 Conclusions and Recommendations An updated Mineral Resource model has been completed for the Ambatovy project in 2018. The previous Mineral Resource model had been developed in large part prior to the commencement of mining and in the absence of reconciliation data. The earlier Mineral Resource model was focused primarily on the ferralite part of the resource and did not separately model the gabbro dykes. The 2018 iteration of the Mineral Resource model has aimed at simplifying the resource model, while also separately modelling the gabbro dykes (which was important in the context of the Mineral Reserve as it is important to have a good estimate of the aluminum oxide (Al2O3) content for PAL calculations). In addition to the modelling of the gabbro, more attention has been paid to the modelling of saprolite.
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The 2018 Mineral Resource model is much better aligned to the mining and PAL constraints than the earlier model. It is anticipated that the Mineral Resource model will continue to evolve as more resource information is available, through ongoing resource drilling, grade control drilling and reconciliation. The estimation of modifying factors (insitu – PAL) is done through reconciliation against the Mineral Resource (i.e. the modifying factors are relative to the Mineral Resource modelling methodology) accordingly, whilst prima facie the 2018 Mineral Resource is different than the 2013 Mineral Resource. These differences are in large part cancelled out in the process of updating the ROM–ROP factors. Ambatovy is committed to a continued focus on grade control methodology, stockpiling strategies and mining selectivity optimisation.
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2 Introduction
2.1 Issuer This Technical Report has been prepared for Sherritt International Corporation (Sherritt), a producing issuer in Canada, as defined in NI 43-101. The Ambatovy Project is a producing nickel and cobalt project located in Madagascar with a designed annual production capacity of 60,000 t of finished nickel, 5,600 t of finished cobalt, and approximately 210,000 t of ammonium sulphate, and an estimated life of approximately 26 years (out until 2044). Annual production rates are projected to vary over the life of the mine depending largely on ore grades. Ambatovy Minerals Société Anonyme (AMSA) and Dynatec Madagascar Société Anonyme (DMSA) (known together as the Ambatovy Joint Venture) are operators of the Ambatovy Project mine and the plant respectively. Sherritt indirectly holds a 12% interest in the Ambatovy Joint Venture with the remaining interest held by Sumitomo Corporation (Sumitomo) and Korea Resources Corporation (KORES). Sherritt also acts as operator of the project.
2.2 Terms of Reference
2.2.1 CSA Global Terms of Reference Phase 1: • Review of current Mineral Resource estimation models, where these have been prepared by Sherritt (or where these have been reported by previous owners and have not been the subject of recent focus by Sherritt), reasonableness testing, and provision to update and improve on those material projects based on areas of improvement identified and/or as a result of new drilling data, informed as required by grade control and production data reconciliation where appropriate and to report under CIM guidelines and NI 43-101 Technical Reporting requirements. • To review the above and write up the stated sections of the NI 43-101 report. Phase 2: • Critically review the Mineral Reserves prepared by Mining Consultants on behalf of Sherritt, reasonableness testing, review of all available modifying factor data and information in support of formal reporting of Mineral Reserves under CIM guidelines and NI 43-101 Technical Reporting. • Review the mining optimisations, pit shell selection, pit designing, production schedules and economic models and completion of reasonableness testing. Provision to update. • Based on discussions, it is anticipated that CSA Global will need to complete the following: o Review the updated Mineral Resource block models which have been converted into “Mining Models” o Review production data in terms of costs and operations for mining and processing and how it has been employed for usage in the optimisation work completed o Review pit optimisation work, pit shells selected and geotechnical considerations o Check conformance of pit designs to geotechnical guidance and to pit shells in terms of variance o Review of mining schedule, taking into account modifying factors, site constraints, cut-off grades, equipment availability, marketing factors and plant requirements o Review of Financial Model, taking into account the mine schedule, the operating costs and capital costs and cash flow over the mining periods based on an agreed discount rate. • To review the above and write up the stated sections of the NI 43-101 report. Phase 3:
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• Compilation of the technical report, in accordance with NI 43-101 and CIM guidelines, including all of the relevant agreed sections assigned to CSA Global as well as peer review of the parts of the report prepared by Sherritt to be included into the report. The above report to be signed off by the appropriate Qualified Persons.
2.2.2 IMC Terms of Reference IMC Mining Pty Ltd (IMC) has been working with AMSA since 2016 to assist with the development of planning systems, cost modelling, reconciliation, short-term mine planning and long-term mine planning. In 2017, IMC developed a LOM plan for Ambatovy based on the 2013 Resource Model. In the course of that work program, a number of improvement programs were identified for resource modelling that would provide a resource model that is better aligned to the mining strategies at Ambatovy. The decision was made that a new resource model would be developed (the 2018 Mineral Resource model) and consequent to the development of the updated Mineral Resource model, IMC, in conjunction with the AMSA site team, would update the five-year detailed operating plan, LOM plan and Mineral Reserve for Ambatovy. The five-year tactical plan was completed in August 2018 with the updated LOM plan (of which the five- year tactical plan represents the first five years) completed in November 2018 and this subset of the Technical Report (related to Mineral Reserve and mining operations) subsequently compiled in November 2018. Scientific and technical information relating to the Mineral Reserve estimate contained in this Technical Report was prepared under the supervision of, or approved by, Mr Stewart Lewis, CEO, IMC, and a “Qualified Person” within the meaning of NI 43-101.
2.3 Principal Sources of Information The preparation of the Technical Report has been coordinated and completed by CSA Global largely based on information provided by the Owner (AMSA) in conjunction with various specialist, independent consultants. These consultants included: • CSA Global Pty Ltd (CSA Global) – Estimation of Mineral Resources and review of Mineral Reserves • IMC Mining Pty Ltd (IMC) – Estimation of Mineral Reserves • Documents and electronic data files provided by the Ambatovy Joint Venture • Information gathered during visits to the Ambatovy Project by Adrian Martinez Vargas, the Qualified Person for the Mineral Resource estimate • Information gathered during a visit to the Ambatovy Project by Dan Grosso (CSA Global), under the supervision of Paul O’Callaghan • Information gathered from the mining geology literature • Information gathered from SEDAR (System for Electronic Document Analysis and Retrieval). Citations to the relevant reports, articles, documents and websites are provided in Section 27 of this report.
2.4 Qualified Person Section Responsibility This report was prepared by or under the supervision of the Qualified Persons identified in Table 5 for each of the sections of this report.
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Table 5: Qualified Person section responsibility
Section Section title Qualified Person(s) Paul O’Callaghan, Glen Smith, 1 Summary Stewart Lewis 2 Introduction Mick Elias 3 Reliance on Other Experts Mick Elias 4 Property Description and Location Mick Elias 5 Accessibility, Climate, Local Resources, Infrastructure and Physiography Mick Elias 6 History Mick Elias 7 Geological Setting and Mineralisation Adrian Martinez 8 Deposit Types Adrian Martinez 9 Exploration Adrian Martinez 10 Drilling Adrian Martinez 11 Sample Preparation Analyses and Security Adrian Martinez 12 Data Verification Adrian Martinez 13 Mineral Processing and Metallurgical Testing Glen Smith 14 Mineral Resource Estimates Adrian Martinez 15 Mineral Reserve Estimates Stewart Lewis 16 Mining Methods Stewart Lewis 17 Recovery Methods Glen Smith 18 Project Infrastructure Paul O’Callaghan, Glen Smith 19 Market Studies and Contracts Glen Smith 20 Environmental Studies, Permitting, and Social or Community Impact Glen Smith 21 Capital and Operating Costs Stewart Lewis 22 Economic Analysis Paul O’Callaghan 23 Adjacent Properties Paul O’Callaghan 24 Other Relevant Data and Information Stewart Lewis, Glen Smith 25 Interpretation and Conclusions Paul O’Callaghan 26 Recommendations Paul O’Callaghan 27 References Paul O’Callaghan
2.5 Qualified Person Site Inspections
2.5.1 IMC Mining Pty Ltd Stewart Lewis, CEO, IMC undertook a site visit to Ambatovy from 18 April until 26 April 2016 to observe all aspects of the mining operations and the ore preparation plant (OPP).
2.5.2 CSA Global Pty Ltd Michael Elias, Principal Consultant-Nickel, undertook a site visit to Ambatovy from 27 to 31 January 2014 (five days) to observe all aspects of the geology, exploration and mining operations on behalf of Sumitomo Corporation. Adrian Martinez Vargas, Senior Resource Geologist, visited the site from 25 to 30 June 2018 (six days) to review resource estimation procedures from data collection, database compilation, deposit modelling and grade interpolation.
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Paul O’Callaghan (QP and Principal Mining Engineer for CSA Global) has not undertaken a site visit. In his stead, Daniel Grosso, Senior Mining Engineer, visited the site from 25 to 30 June 2018 (six days) to review all aspects of the mining operation, its practices and infrastructure under the supervision and instruction of Paul O’Callaghan.
2.5.3 Sherritt International Glen Smith worked on a daily basis at the Ambatovy Plant Site from August 2009 to May 2016 in the positions of Operations Manager, Refinery and Manager, Technical Services. He continues to support the operation as a Consulting Engineer for Sherritt Technologies. and has visited the mine and plant site in June 2017, December 2017, June-July 2018 and plant only from September to November 2018 for a combined period of approximately four months.
2.5.4 Current Site Visit The Authors consider the site visit by Adrian Martinez Vargas to be a “current” independent site visit under NI 43-101 Section 6.2.
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3 Reliance on Other Experts
The authors of this Technical Report have not undertaken an independent review and assessment of legal, environmental and political considerations. On these issues, the Technical Report relies entirely on information provided by the issuer experts or on documents publicly disclosed by the Issuer. • Section20 of this Technical Report reflects entirely the opinion of Andrew Mackenzie, Environmental Manager of the Ambatovy Project. This information has been provided via email correspondence as at 14 September 2018 with minor amendments thereafter. • Section 19.1 of this Technical Report reflect entirely the opinion of Jack Lowe, Director Marketing of the Ambatovy Project. This information has been provided as at 30 November 2018. • Sections 4.4, 4.5 and 19.2 of this Technical Report reflect entirely the opinion of Hansina Valaydon, Senior Legal Counsel for Ambatovy or Ambatovy Finance Department. The information has been provided via email correspondence as at 17 December 2018. • Sections 0 and 24.2 of this Technical Report are based on the 2017 Sherritt International Corporation Annual Information Form dated 20 March 2018.
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4 Property Description and Location
4.1 Location The Ambatovy Project is located on the island nation of Madagascar. Madagascar is the world’s fourth largest island, covering approximately 587,000 km2. It is located 500 km off the southeast coast of Africa. The project comprises an inland mining operation and a coastal processing plant, connected by an ore slurry pipeline. The mining operation is located approximately 80 km east of the capital city of Antananarivo, and 11 km north of the town of Moramanga in the region of Alaotra Mangoro (Figure 1). The processing plant is located at the coastal port of Toamasina.
4.2 Property Description Sherritt indirectly holds a 12% interest in the AMSA and DMSA Ambatovy Joint Venture. Sumitomo and KORES indirectly hold the remaining interest in the Ambatovy Joint Venture. Sherritt acts as the operator of the Ambatovy Joint Venture. AMSA holds Exploitation Permit 459 (“the Property”) which entitles it to extract nickel, cobalt, copper, platinum and chrome for a period of 40 years dating from 7 September 2006, as well as allowing for the operation of infrastructure at the Property. Exploitation Permit 459 covers the mine area, as illustrated in Figure 2. In order to maintain Exploitation Permit 459 in good standing, AMSA must file an annual report, pay an annual administration fee of approximately US$46,300 (107,345,600 MGA) payable under the Mining Code, and remain in compliance with its obligations under the Environmental and Social Management Plan (ESMP), which is attached as an appendix to the environmental permit. The Property covers 143.75 km2 and consists of 368 contiguous 625 m x 625 m blocks (carrés). Official coordinates of the centre of each block are in the Laborde Geodesic System. The coordinates of the Property have also been converted to the Universal Transverse Mercator (UTM) system. The approximate centre of the Property lies at 212,000E and 7,915,000 N, UTM Zone 39S, WGS 84. The surface rights of the Property consist of a 50-year lease with the Malagasy State which is conditional on the payment of annual fees of US$250,000 and compliance with the stated land usage which is mining and establishment of a forest conservation area according to the ESMP, as discussed in further detail in Section20 of this Technical Report. The processing plant (Processing Plant), located near the port of Toamasina, includes a pressure acid leaching (PAL) plant, a metals refinery and associated utility and ancillary plants including: water treatment, steam and power generation, hydrogen, hydrogen sulphide, sulphuric acid, air separation, limestone comminution and lime calcining and slaking. A tailings management area comprises two adjoining basins and covers an area of approximately 760 ha.
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Figure 1: Regional location map Source: Sherritt International, 2018
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Figure 2: Property map Source: Ambatovy Joint Venture, 2011
4.3 Royalties, Back-in Rights and Other Payments The Property is subject to the following payments, in addition to those referred to in Section 4.1 above: • A property tax on land (IFT) for all communes in Madagascar under the Law on Large Scale Mining Investiments (Loi sur les Grands Investissements Miniers – LGIM), which is equivalent to 1% of the market value of the Property and which is payable annually, but which is capped at 200,000,000 MGA (around US$100,000). • A property tax on the buildings which is equivalent to 1% of the annual rental value of the buildings and which is payable annually. The Ambatovy Joint Venture is entitled to a five-year exemption upon completion of the construction pursuant to the LGIM.
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• Mining royalties payable on production which are equivalent to 1% of the finished metal and by- product metal sulphide sales pursuant to the Mining Code and the LGIM.
4.4 Permits The Ambatovy Project required a number of permits for both its constructionphase, which was completed in 2014, and ongoing operational phase: • The Ambatovy Joint Venture was required to obtain a number of construction permits for facilities at the Property and at the Plant Site, as well as for the resettlement village (as discussed in further detail in Section20 ). All such permits have been obtained. • DMSA received the authorisation of the Ministry of Mines and Ministry of Industry to place the Plant Site into operation in September 2012. • AMSA is currently in the process of updating and renewing the required permits for the disposal of effluents and treated waste water and the pumping of water from the Mangoro and Ivondro rivers with the Autorité nationale de l’eau et de l’assainissement of Madagascar1 (ANDEA). A definitive permit to withdraw water from Mangoro river was delivered by ANDEA on October 2017. • DMSA is working with the Ministry of Energy towards the finalisation of the permits for its diesel generators and power plant, all required technical authorisations have been received. • The Ambatovy Joint Venture has also been granted the necessary environmental permits and authorisations which are discussed in Section 20.
4.5 Environmental Liabilities Environmental permit No. 47/06/MINENVEF/ONE/DG/PE was issued on 1 December 2006 to AMSA/DMSA conditional upon implementation of the ESMP (Ambatovy Project, 2006) and subject to an annual review process. The ESMP made provision for the development of specific management plans relating to project components (mine, pipeline, Plant Site, Tailings Management Facility (TMF) and port extension), environmental thematic plans (air, water, soil, biodiversity) and other specific plans (village relocation, species specific plans, mine closure etc) and for the renewal of plans upon transition from the construction to operation phase of the Ambatovy Project. Studies conducted for the Environmental Impact Assessment (EIA) confirmed the high biodiversity of the proposed mining site, most of which was covered in mid-altitude primary forest growing on a ferralitic substrate and supporting lemurs and other terrestrial animals, birds, amphibians, reptiles and numerous plant species, of which some are endangered. To address the considerable environmental challenges and meet regulator and lender requirements including International Finance Corporation (IFC) Standard 6 (IFC 2006), the Ambatovy Joint Venture operates a “no net loss” biodiversity strategy based on measures to avoid or minimise impacts and uses offsets to compensate for unavoidable biodiversity losses. The environmental and associated risks are discussed in more detail Section 20.
4.6 Other Risks The Ambatovy Project is subjected to certain risks which could affect access, title, or the right or ability to perform work on the Property, pipeline, TMF, Plant Site and/or port. These include, but are not limited to, the following: • Operations at the Ambatovy Project may be affected by the fact that Madagascar’s location potentially exposes it to cyclones and tropical storms of varying intensities. The risk of damage is dependent upon such factors as intensity, footprint, wind direction and the amount of precipitation associated with the storm and tidal surges. While the Ambatovy Joint Venture maintains comprehensive disaster plans
1 National Authority for Water and Sewage
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and its facilities have been constructed to the extent reasonably possible to minimise damage, there can be no guarantee against severe property damage and disruptions to operations. • Madagascar is one of the poorest countries in the world, with low levels of economic activity and high levels of unemployment. These conditions are conducive to social unrest and instability that could, under certain circumstances, have an impact on the Ambatovy Joint Venture’s ability to access, perform work on or result in challenges to the Ambatovy Joint Venture’s title to the Property, pipeline, TMF, Plant Site and/or port. The Ambatovy Joint Venture continues to foster active working relations with relevant Malagasy authorities and civil society to mitigate social risk, maintain its social license, and facilitate operational activities. The political environment and associated risks are discussed in more detail Section24.2 .
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5 Accessibility, Climate, Local Resources, Infrastructure and Physiography
5.1 Physiography The Ambatovy and Analamay Deposits, which are part of the Property, are located on a plateau at an elevation of approximately 1,100 m above sea level (ASL). The topography of the Deposits varies from gently undulating hills to a steeply dissected remnant plateau. Locally, the relief is 100 m. The plateau is flanked to the west by the broad alluvial plain of the Mangoro River and to the east by the Torotorofotsy Wetlands and forested hills. The plateau surface is fairly uneven with numerous depressions that form ephemeral pools. Small headwater streams originate in the mine area and flow away in all directions as part of six basins. The Property is covered with natural forests. The surrounding area includes intact and degraded forests and scrublands, areas dominated by grasses, eucalyptus plantations, woodlots and rice paddies. The soils in the mine region are generically known as laterites, which are highly weathered iron-rich tropical soils. The orebodies are characterised by ferricrete soils with a hard, rock-like surface. This has resulted in the forests on the orebodies being different (azonal) from the surrounding primary forest (zonal). Figure 3 shows an aerial image of the Property area highlighting the mining activity area. The image shows clearly the extent of the forest and the conservation areas surrounding the Property.
Figure 3: Aerial image of the mine activity area showing surrounding conservation area Source: Ambatovy Joint Venture, 2018
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5.2 Access The capital city, Antananarivo, is serviced by commercial air flights from Paris, France, three times per week and daily from Johannesburg, South Africa. Regular flights also run between Antananarivo and Réunion, Mauritius and Kenya. The Property, located near the town of Moramanga, is approximately 120 km by road from Antananarivo. Major asphalt paved national roads link Antananarivo with Moramanga, with a high quality 11 km gravel road leading from Moramanga to the Property. A charter plane, operated on behalf of the Ambatovy Joint Venture, flies regularly between Antananarivo, Toamasina and Moramanga.
5.3 Population Centre The Property is located 11 km north of the town of Moramanga. The town has a population of approximately 30,000. Moramanga is situated on a plateau between the central highlands and the east coast. Project buses run daily between Moramanga and the Property to transport employees to the Property.
5.4 Climate The climate is equatorial to tropical with annual rainfall ranging from 1,000 mm to 2,000 mm; evaporation from 600 mm to 800 mm and temperatures from 5°C to 35°C (Figure 4), with an average temperature of 17°C. The cyclonic season prevails from January to March (Figure 5). The mine operates on a year-round basis.
Figure 4: Temperature average at Property Source: Ambatovy Joint Venture, 2013
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Figure 5: Rainfall at Property Source: Ambatovy Joint Venture, 2013
5.5 Local Resources and Infrastructure For the needs of the Property, the Ambatovy Project has developed bedrock quarries at the Property which provide all the rocks required for road construction and maintenance (sheeting), as well as the mining operation. The quarries are in the Ambatovy main pit. The Mangoro River, approximately 20 km west of the Property, has significant flow throughout the year. The Ambatovy Joint Venture constructed a 600 mm diameter, 24-km long pipeline to bring water from the Mangoro River water to the Property as required for its operations. Local infrastructure includes the town of Moramanga with a population of approximately 30,000. The Ambatovy Project has access to the requisite mining personnel through the use of the local population for unskilled and semi-skilled labour, as well as having on-site residential facilities for necessary expatriate and national senior staff employees (see Section 18.2 of this Technical Report for further detail). The Ambatovy Joint Venture also operates a housing assistance program for national employees from the Plant Site and head office in Antananarivo who are relocated to Moramanga to work at the Property. A one-metre gauge railway line that runs from Antananarivo to Moramanga and to the port of Toamasina was recently privatised and has been upgraded. The ore from the Property is transported to the Plant Site by way of an approximately 220-km pipeline, which is discussed in further detail in Section 17.1 of this Technical Report. A 138-kV powerline, part of the national grid, passes approximately 8 km south of the Property. Please refer to Section18.1.9 of this report for more information regarding the power supply at the Property. The TMF, waste disposal and Processing Plant are located close to the City of Toamasina and are described in greater detail in Section 18 of this Technical Report. Ambatovy Project surface rights are sufficient for mining operations (see Section4.2 ).
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6 History
The presence of the Ambatovy and Analamay Deposits was first noted by the Malagasy Service Géologique in 1960. Limited pitting and auger drilling were undertaken to assess the surface iron potential. In 1962, additional auger drilling (2,500 m) and pitting were undertaken, essentially around the fringes of the deposits, by the French Bureau de Recherches Géologiques et Minières (BRGM) and Ugine Kuhlmann.
6.1 GENiM During 1970–1972, GENiM carried out a major exploration program that included 368 vertical diamond drillholes (totalling 11,600 m) on a 195 m grid, over both the Deposits, an aerial photographic survey of the general area, geological mapping and geochemical surveys of the Deposits and surrounding area.
6.2 PD Madagascar SARL (Phelps Dodge) Phelps Dodge was granted exploration permit no’s 459, 460, 461 and 4416 for the Deposits in 1995. These exploration permits were later transformed into a single exploration permit, exploration permit no. 459, in 2003 pursuant to Decree (Arrêté) No. 10564/2003: Transformation/merger of mining licence. Between 1995 and 1998, Phelps Dodge conducted exploration work which included 369 vertical diamond drillholes (totalling 18,008 m), seven test pits (12–31 m deep), a ground magnetic survey, a soil geochemistry survey and metallurgical testwork on a 5-tonne bulk sample. This work culminated in a feasibility study completed by H.A. Simons Ltd (now AMEC) entitled “Simons Feasibility Study Phelps Dodge – Ambatovy Project” dated April 1998. The Phelps Dodge drilling campaign focused on Ambatovy West and resulted in an unpublished Mineral Resource estimate. This last resource estimate in a Phelps Dodge internal memo included both the ferralite (limonite) and the low magnesium saprolite (LMS) and totalled 56 Mt grading 1.14% Ni and 0.091% Co, at a 0.8% Ni cut-off grade. The historic resource estimate noted in this section is “historical” in nature and not in compliance with NI 43-101. A Qualified Person has not done the work necessary to verify the historical estimate as a current estimate under NI 43-101 and as such they should not be relied upon. The Authors, CSA Global and Sherritt are not treating the historical estimate as a current Mineral Resource; it is instead presented for informational purposes only. The historical resource estimate is superseded by the 2018Mineral Resource estimate presented in Section 14 of this Report. In 2002, the LGIM came into force, providing legal stability and investment incentives in Madagascar for eligible mining companies.
6.3 Dynatec Corporation In 2003, Phelps Dodge and DMSA signed a joint venture agreement to continue the development of the property. DMSA was incorporated in 2003 as well. DMSA conducted exploratory drilling and started to develop a thorough feasibility study with a detailed environmental and social impact assessment. To complete the feasibility study, exploration work included: • The production of topographic maps of the Deposits • The drilling of 545 vertical HQ diamond drillholes (totalling 25,180 m) • Core sampling (essentially at 1 m intervals) and assaying for nickel, cobalt, magnesium, iron, manganese, zinc, copper, aluminum, calcium, cadmium, chromium, sulphur, silicon, and titanium • Density and moisture content determinations on every fifth metre of the drill core for each hole • The collection of samples for pilot plant tests from 6-inch diameter drill core and backhoe trench material totalling 24 tonnes
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• Batch metallurgical testwork on core from 23 HQ drillholes and from a vertical channel sample (the French Cut) totalling approximately one tonne. Furthermore, all available GENiM exploration data was purchased from Société Le Nickel by DMSA. In 2004, DMSA received a 53% interest in the Ambatovy Joint Venture from Phelps Dodge. In 2005, DMSA acquired the remaining 47% interest in the Ambatovy Joint Venture from Phelps Dodge. Sumitomo took up a 25% stake in the Ambatovy Joint Venture which left DMSA as a 75% stakeholder. In 2006, Exploration Permit 459 was converted into Exploitation Permit 459 pursuant to Decree 5623/2006. DMSA was named the project operator and retained 45% ownership – later reduced to 40% when SNC-Lavalin Inc. (SLI) took up 5%. KORES joined the Ambatovy Joint Venture becoming a 27.5% shareholder along with Sumitomo with a 27.5% interest. SLI was awarded the engineering, procurement, construction and management contract for the development of the project. The government agency responsible for the oversight and implementation of Madagascar’s principal environmental legislation including environmental permits, Office National pour l’Environnement (ONE) issued the environmental permit. SLI completed the feasibility study that was published in April 2006. DMSA continued exploration drilling in 2006 and 2007, mostly in Ambatovy Central, with 100 vertical HQ holes drilled for a total of 5,675 m.
6.4 2005 Mineral Resource and Mineral Reserve Estimates The Mineral Resource and Mineral Reserve estimates noted in this section are considered “historical” in nature. A Qualified Person has not done the work necessary to verify the historical estimates as current estimates under NI 43-101 and as such they should not be relied upon. The Authors, CSA Global and Sherritt are not treating the historical estimates as current Mineral Resource and Reserves; theyare instead presented for informational purposes only. The historical resource and reserve estimates are superseded by the 2018 Mineral Resource estimate and Mineral Reserve estimate presented in Sections 14 and 15 of this report respectively. In 2005, Watts Griffis and McOuat (WGM) reported Mineral Resource and Mineral Reserve estimates for the mineralised areas of the Deposits at a 0.8% nickel cut-off (Tables 6, 7, 8 and 9). Table 6: Measured Mineral Resources at 0.8% Ni cut-off– classified by WGM (October 2004)
Area Tonnage (Mt) Ni (%) Co (%) Ambatovy 44.4 1.12 0.103 Analamay - - - Total 44.4 1.12 0.103
Table 7: Indicated Mineral Resources at 0.8% Ni cut-off – classified by WGM (October 2004)
Area Tonnage (Mt) Ni (%) Co (%) Ambatovy 33.4 1.02 0.086 Analamay 56.3 0.99 0.104 Total 89.7 1.00 0.097
Table 8: Measured and Indicated Mineral Resources at 0.8% Ni cut-off– classified by WGM (October 2004)
Area Tonnage (Mt) Ni (%) Co (%) Ambatovy 77.8 1.07 0.096 Analamay 56.3 0.99 0.104 Total 134.1 1.04 0.099
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Table 9: Inferred Mineral Resources at 0.8% Ni cut-off – classified by WGM (October 2004)
Area Tonnage (Mt) Ni (%) Co (%) Ambatovy 14.9 1.01 0.079 Analamay 7.9 0.97 0.099 Total 22.8 0.99 0.086
The Mineral Reserves for the Ambatovy Project as at 1 January 2005 at 0.8% Ni cut-off were: • Proven Reserves: 44.1 Mt at 1.12% Ni and 0.102% Co • Probable Reserves: 80.9 Mt at 0.99% Ni and 0.097% Co • Combined Proven and Probable Mineral Reserves: 125 Mt at 1.04% Ni and 0.099% Co.
6.5 Sherritt International Corporation In 2007, the Ambatovy Project was certified by the Government of Madagascar under the LGIM. Sherritt acquired DMSA for approximately US$1.7 billion. Sherritt assumed DMSA’s ownership position in the Ambatovy Joint Venture and was named project operator. Exploitation Permit 459 was also modified by way of Decree No. 15471/2007 to extend the validity period so that it would be from 40 years starting 7 September 2006, instead of 40 years starting 10 May 1995. For further information regarding Exploitation Permit 459 and surface rights for the Property, please refer to Section 4.2 of this Technical Report. In 2008, DMSA received construction permits for work at the port and Plant Site. During May and June, a new campaign of ore preparation pilot plant tests was undertaken with samples coming from test pits on both orebodies. Definition drilling started again, mainly on the Ambatovy West sub-block. A total of 168 holes for a total of 8,605 m were drilled between June 2008 and November 2009. By the end of 2009, construction of the Ambatovy Project was more than 60% complete. In 2010, commissioning of the mining equipment began. Mining of material for stockpiling started in July. On 31 December 2010, a total of 241,000 t of ore was mined and stockpiled. In 2011, commissioning of the OPP at the Property commenced, as well as commissioning of the utilities facilities at the Plant Site. In 2012, ore processing in the PAL autoclaves began, and mixed sulphides were delivered to the refinery in May 2012, with the first finished nickel and cobalt briquettes being produced from the refinery in September 2012. Since that time, the Ambatovy Project has continued its ramp-up towards full production and as at the end of 2017 it had processed approximately 20,000,000 t producing 187,000 t Ni and 14,800 t Co. Figure 6 is a site plan showing the location of the project’s mine operations and facilities. Figure 7 is a site plan showing the location of the project’s plant site, tailings management facility and port at Toamasina.
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Figure 6: Current site plan, Ambatovy Project
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Figure 7: Ambatovy Project plant site, tailings management facility and port at Toamasina Source: Dynatec Corporation, 2005
6.6 2013 Mineral Resource and Mineral Reserve Estimates Mineral Resources and Reserves, with an effective date of 31 December 2013, were previously reported in accordance with CIM Definitions and Standards in the 2014 Technical Report (Daigle, B. et al., 2014). The 2013 Mineral Resource is discussed in Section 14.12 of this Technical Report. The Mineral Resource and Mineral Reserve estimates noted in this section are now considered“ historical” in nature. A Qualified Person has not done the work necessary to verify the historical estimates as current estimates under NI 43-101 and as such they should not be relied upon. The Authors, CSA Global and Sherritt are not treating the historical estimates as current Mineral Resource and Reserves; theyare instead presented for informational purposes only. The 2013 Mineral Resource and Mineral Reserve estimates are superseded by the 2018 Minearl Resource estimate and Mineral Reserve estimate presented in Sections14 and 15 of this Report respectively.
6.6.1 2013 Mineral Resource Estimate The Ambatovy Project’s total 2013 estimated Mineral Resources are summarisd below in Table 10. These Mineral Resources do not include any all of the material that lies above the bottom of the pit as at 31 December 2013; they therefore reflect only the material that has not yet been mined as of the end of 2013.
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Table 10: Total estimated Mineral Resources for the Ambatovy Project (inclusive of Mineral Reserves) above a cut-off grade of 0.6% nickel (with an ffectivee date of 31 December 2013)
Classification Tonnage (Mt) Ni (%) Co (%) Measured 93.6 0.89 0.08 Indicated 168.2 0.82 0.08 Measured+Indicated 261.9 0.85 0.08 Inferred 62.9 0.76 0.07 Note: Cobalt grade does not enter into the definition of the reporting cut-off grade since the vast majority (over 80%) of the Ambatovy Project’s revenue comes from nickel.
6.6.2 2013 Mineral Reserve Estimate Table 11 provides a summary of the Proven and Probable Reserves for the Ambatovy Project as at 31 December 2013. Table 11: Estimated Mineral Reserves for the Ambatovy Project as at 31 December 2013
Reserve classification (1)(3)(4) Tonnage (Mt) Ni (%) Co (%) Proven 78.4 0.90 0.08 Probable 112.0 0.82 0.07 Total Proven and Probable Reserves 190.4(2) 0.85 0.07 Notes: • Mineral Reserve estimates are based on a cut off grade of 0.6% Ni. All assumptions, parameters and methods used to estimate the Mineral Reserves are disclosed in Section 15 of this Technical Report. The cobalt grade does not enter into the definition of the reporting cut-off grade since the vast majority (over 80%) of the Ambatovy Project’s revenue comes from nickel. • Totals may not sum exactly due to each component number being rounded to its nearest decimal. • Mineral Reserves include materials that have been mined and stockpiled at the mine site, and that, as at 31 December 2013, had not yet been slurried and pumped down the pipeline to the Processing Plant. • Mineral Reserves are based upon an average mill processing rate of 6.5 Mt per year. • Mineral Reserves were calculated on a nickel price of $7.37/lb and a cobalt price of $12.20/lb. The Mineral Reserve estimate for the Project’s open pits was constrained by estimates of nickel and cobalt price, mining dilution, process recovery, refining/transport costs, and royalties. Mining, processing, and general administration costs were also estimated based on expected mill throughput and, along with geotechnical parameters, formed the basis for the open pit optimisation. The mineral inventory block models for each of the Deposits were then used with the Gemcom Whittle - Strategic Mine Planning™ (Whittle) software to determine optimal mining shells. Both Indicated and Measured Mineral Resources developed by R. Mohan Srivastava of FSS Canada Consultants Inc. were included in the pit optimisation process (no Inferred Resources were included). A stockpiled 2.4 Mt of mill feed with a grade of 0.92% Ni and 0.07% Co that was included in the Mineral Reserve estimate.
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7 Geological Setting and Mineralisation
7.1 Regional Geology The dominant regional geological feature of the Ambatovy Project area is a north-south striking belt of basic gneisses and migmatites (Figure 8). These rocks form part of the high-grade metamorphic rocks that underlie the eastern two-thirds of Madagascar. The metamorphism is of late Precambrian age, the same age of metamorphism of the rocks of East Africa, to which Madagascar was joined at that time. As Madagascar broke away from the African continent about 135 Ma ago, the break-up was accompanied by volcanism and internal rifting, the latter forming the horst and graben structural features that are pertinent to the Ambatovy Project. A large intrusive, known as the Antampombato Complex, believed to be of Cretaceous age, cuts the gneissic terrain, and dominates the geological setting of the Ambatovy Project. The intrusive is elliptical in shape and oriented northwest-southeast with the main axis some 12 km in length and the shorter axis about 7 km in length. The Antampombato Complex is considered an Alaskan-type mafic-ultramafic complex, recognised as a distinct class of intrusion. Characteristic petrologic features are the concentric zonation with a central dunite body grading outward into wehrlite, clinopyroxenite and gabbro. All these rocks are broadly cogenetic and are the result of a low-pressure crystal accumulation or fractionation starting from a transitional basalt parental magma (Melluso et al., 2005). During the cooling of ultramafic magmatic intrusion, as the dunite (olivine) crystallized, residual liquid was enriched in pyroxene and then in alumino-silicates, leading to a gabbroic liquid that was trapped in the interstices between blocks of dunite (Figure 9 and Figure 10). The Antampombato Complex occurs on the remnants of a plateau known as the Antampombato Massif. This plateau, a horst structure, is flanked on the east and west by two graben structures. The plateau occurs at an approximate elevation of 1,100 m ASL while the graben structures, represented by basins, occur at elevations of about 900 m ASL. The grabens are filled with40 –70 m of recent alluvial sediments. The broad alluvial plain of the Mangoro River occupies the basin to the west, while the basin to the east is represented by the Torotorofotsy Marshes. In 2004–2007, the World Bank funded a program to improve the governance of mineral resources in Madagascar, through the PGRM (Programme pour la gouvernance des ressources naturelles). This program, implemented via the Ministry of Mines and Energy, funded an airborne geophysical survey (magnetic and radiometric) over the crystalline basement of Madagascar, including the Deposits area. Following this geophysical survey, the PGRM also contracted a major revision of a number of 1:100,000 scale geology maps including the Deposits area. Those maps over the Antampombato intrusive complex did not significantly alter the boundaries and domains of the different formations, but introduced a new tectonic environment, particularly a major previously unrecognised northwest-southeast regional fault system delineated by the aeromagnetic survey which differs from the major north-south trend in Madagascar. This northwest-southeast fault complex is related to the intrusive complex and has a major influence on the regional geology of the Deposits.
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Figure 8: Ambatovy regional geological map Source: Dynatec Corporation, 2005
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Figure 9: Gabbroic magmatic injection Source: Ambatovy Project, 2011
Figure 10: Details of gabbroic magmatic injection Source: Ambatovy Joint Venture, 2011
7.2 Local and Property Geology Within the Antampombato complex two mafic-ultramafic intrusions can be identified, Ambatovy and Analamay (Figure 11). Both consist mainly of ultramafic rocks, dunite and wehrlite, with pyroxenite injections; the ultramafics are surrounded by gabbroic material. Since ultramafic rocks are highly geochemically unstable in a tropical weathering environment, Ambatovy and Analamay present a deep pædologic alteration, with a complete lateritic profile capped by a ferruginous duricrust. The Ambatovy Deposit contains three main zones: Ambatovy West, Central and South East (Figure 12). The Analamay Deposit also contains three main zones: Analamay North, Central and South (Figure 13).
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The ultramafic rocks are continuous between the different zones of the Deposits, resulting in continuous nickel mineralisation.
Figure 11: Locations of the Deposits Source: AMSA, 2011
7.2.1 Ambatovy Geology Figure 12 shows the geological bedrock map of the Ambatovy Deposit based on an interpretation of drillhole bedrock samples, and stream bed boulder and outcrop mapping by GENiM. Ambatovy Northwest is, for the purposes of this Technical Report, treated as an extension to, and part of, Ambatovy West. The principal components are as follows: • The central core area is believed to be an initial intrusion of the Antampombato Complex. It is composed of medium-grained dunite/peridotite with irregular lenses, or possibly segregation of pyroxenite, the majority of which appear to be gently dipping to flat-lying. The olivine in the dunites is the high magnesium variety, forsterite. The peridotite (wehrlite) and pyroxenite both show a decreasing olivine content and an increasing pyroxene content (principally diallage). • Second stage intrusions are peripheral to the core area and consist of a magnetic dunite and pyroxenites, occasionally feldspathic, in the northeast and north, respectively. Gabbroic rocks (some with olivine) occur to the east, west and south in both fault contacts and as peripheral sills intruding into the ultramafic core.
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• Dykes and veins of syenite, dolerite and gabbro, representing end stages of intrusion, have been occasionally intersected by drilling but are generally too narrow to trace or determine their orientation. In keeping with the regional graben setting, Ambatovy West is cut by numerous block faults striking northwest/southeast with a conjugate set, northeast/southwest. Evidence, including sharp changes in laterite thickness and changes in bedrock elevations, together with currently eroding steep valleys, indicates continuing fault movement during laterisation. There appear to be numerous small fault displacements as indicated in the geological cross-sections, but since the interpretation is based essentially on 50 m spaced grid drilling, it is not possible to define their precise orientation. The more significant faults, as interpreted, are indicated on the geological plan (Figure 12).
7.2.2 Analamay Geology The Analamay Deposit occurs some 4 km to the northeast of the Ambatovy Deposit. A geological bedrock map of the Analamay Deposit based on an interpretation of drillhole bedrock samples is shown as Figure 13. Two ultramafic/mafic areas (North and Central-South) occur on plateaus and are separated and enclosed by a later stage gabbroic intrusion. Ultramafic/mafic rocks are mainlymedium -grained dunite associated with peridotite (wehrlite) and minor clinopyroxenite. A more massive pyroxenite zone occurs on the northeastern limit of the North deposit. Dykes and veins of syenite, dolerite and gabbro can also be identified within the ultramafic/mafic rocks. Theses later stage intrusions are generally too narrow to precisely define their attitude but the areas with increased presence are shown on Figure 13. The North and Central areas are gently undulating and generally covered by ferricrete. The South Analamay area is more dissected with less ferricrete capping. In general, the laterite profile of the Analamay Deposit is somewhat thinner than at the Ambatovy Deposit, and the saprolitic horizon within the lateritic profile is much less developed than at the Ambatovy Deposit, suggesting a less fractured and jointed bedrock. The bulk chemistry of the Analamay Deposit laterite is very similar to the Ambatovy Deposit; nickel grade, however, is slightly lower.
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Figure 12: Ambatovy bedrock geology Source: Dynatec Corporation, 2005
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Figure 13: Analamay bedrock geology Source: Dynatec Corporation, 2005
7.2.3 Mineralisation The Ambatovy and Analamay lateritic nickeldeposits overlie the ultrabasics that occur within the intrusive complex. Towards the southern margin of the complex, the Ambatovy Deposit occurs over an ultrabasic body, approximately 3 km x 2.4 km, with the long axis oriented west-northwest-east-southeast. The Ambatovy Deposit is cut by a northwest-trending gabbroic intrusive that results in three sub-blocks known as Ambatovy West, Ambatovy Central and Ambatovy Southeast. However, data gathered since 2005 show continuity between Ambatovy Central and Ambatovy Southeast.
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The Analamay Deposit, located at the eastern margin of the complex, occurs over an ultrabasic body approximately 4 km x 2.8 km oriented north-south. It is also divided into sub-blocks known as Analamay North, Analamay Central and Analamay South. Mineralisation is continuous between Analamay Central and South. Both Deposits occur as residual lateritic soils formed by tropical weathering of ultrabasic rocks. Laterite, as used in this Technical Report, is a general term that includes the entire weathered profile from surface down to the fresh bedrock. The nickel grade of the laterite is generally influenced by the nickel content of the underlying bedrock, with dunite in the range of 0.35% Ni, peridotite 0.25% Ni, pyroxenite 0.15% Ni and olivine gabbro 0.05% Ni. Prolonged weathering has produced a thick mature laterite profile that tends to smooth the nickel distribution in the laterite across the various bedrock types.
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8 Deposit Type
The Ambatovy and Analamay Deposits are typical nickel laterites in which nickel enrichment has occurred in the residual soils formed by tropical weathering of ultramafic bedrock. Prolonged weathering has produced a thick mature laterite profile in which the nickel grades have been enriched from the levels seen in the underlying bedrock. Figure 14 shows the typical laterite profile at the Deposits. From the ground surface down to fresh bedrock, there are three main horizons: (1) an iron-rich crust (“GR”); (2) a ferralite layer (often referred to as “limonite” in other nickel laterites) that is low in magnesium and high in iron, and that shows none of the original texture of the weathered bedrock (“FER”); and (3) a saprolite layer that is high in magnesium and low in iron, and that contains boulder fragments in which the original bedrock textures remain visible (“SAP”).
Figure 14: Ambatovy lateritic profile and vertical grade variation Source: AMSA, 2011 Although weathering smooths the nickel grades in the horizontal direction across the different bedrock types, the horizontal variations in nickel grade are generally influenced by the nickel content of the underlying bedrock: higher over dunites, moderate over peridotites and pyroxenites, and low over gabbros. In the vertical direction, the nickel grades peak near the ferralite-saprolite contact, where the iron:magnesium ratio changes quickly. In the exploration and development drilling programs, holes are drilled to penetrate the entire vertical section of the laterite, i.e. through the ferralite-saprolite contact and as far into the rocky saprolite as possible. The horizontal bedrock geology related variations in nickel grade make dunite the preferred drilling target, with peridotites and pyroxenites being secondary targets. Drilling completed over the gabbroic bedrock that surrounds the ultrabasic intrusions occasionally shows modest mineralisation.
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9 Exploration
Other than drilling (discussed in Section10 of this Technical Report), the most significant exploration work conducted in recent years has focused on surveying and topography issues that needed to be resolved in order to integrate the data from the two early exploration programs of GENiM and Phelps Dodge with DMSA and AMSA data. The official coordinate system used in Madagascar is the Laborde system, a French design based on a Transverse Mercator Projection rotated to align with the long axis of the island of Madagascar. The BRGM, and subsequently GENiM, used the Laborde system, but systematic errors have been found in the drillhole coordinates with some confirmation of surveying problems noted in GENiM correspondence. Phelps Dodge used a separate local system that was converted to the Laborde system, but the transformation formula used to subsequently convert the coordinates to UTM was incorrect for the area. Similar errors occurred with Phelps Dodge’s subsequent differential global positioning system (GPS) survey. DMSA also had difficulty with its own differential GPS survey. After an AMEC audit confirmed large inconsistencies in the survey data, particularly with the Z coordinate (elevation), DMSA hired a Malagasy contractor, SIMTEPHA, to re-survey all identifiable Phelps Dodgeand GENiM hole collars, as well as their own drill collars using Wild theodolites, laser distance measuring instruments and stadia rods. Coordinates were reported in the Laborde coordinate system. Field audits on DMSA and Phelps Dodge hole collars subsequently yielded acceptable accuracy. Some 90 GENiM hole collars and survey stations were re-surveyed, which was considered to be of sufficient number to allow relocation and re-orientation of the GENiM drill grid to fit with the DMSA and Phelps Dodge drill grids. Survey data was also collected at various points along drill roads, and together with the drillhole collar coordinates, was used with the Surfer software program to prepare topographic maps at 5 m contour intervals showing drillhole locations. When the Ambatovy Project 2005 feasibility study returned positive results, it was decided that the development of the Ambatovy Project should be carried out using the standard UTM coordinate system, to allow, among other advantages, the easy use of GPS equipment in the field. Only the operations involving the property boundaries, land occupation or acquisition would continue to use the Laborde official coordinate system to remain compliant with the law in Madagascar. To perform a full conversion of topographic data into the UTM coordinate system, DMSA contracted a LiDAR survey during the third quarter of 2005. The LiDAR survey was performed with a helicopter-borne laser system by LiDAR Services International of Canada, over the Deposits area; along the ore slurry pipeline corridor and the Processing Plant/port and tailings dam area in Toamasina. This LiDAR survey resolved definitively all pending issues related to coordinate inconsistencies. The LiDAR survey, coupled with a digital aerial imagery database and internally developed a geographic information system facility has resulted in the implementation of the UTM coordinate system throughout the Project, from the mine to the Processing Plant. The mine survey team is using the UTM Zone 39S coordinate system projected from the WGS84 ellipsoid, and elevations are referenced to the EGM96 geoid.
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10 Drilling
10.1 Drilling Procedures Drilling was undertaken by several companies since the discovery of Ambatovy deposit. Since the start of the project, 128,049 m was drilled, 89,679 m were drilled in Ambatovy deposit (Figure 15, Table 12) and 38,370 m in Analamay Deposit (Figure 16, Table 12). Between 2004 and 2005, the drilling was undertaken by DMSA’s subsidiary drilling company (Dynatec Drilling Division) based in the USA. The drill rigs provided included: four LongYear LF 70s, skid mounted; one UDR 650, track mounted; and one Explorer, track mounted. Each rig was capable of HQ air core drilling. In summary, 612 vertical holes, for a total of 28,531 m, were drilled by DMSA on the Deposits up to the end of 2005. Since 2007, the Ambatovy Joint Venture used one rig (UDR 650) for the exploration campaign. Table 12: Ambatovy drillhole statistics until 2017
Project Company Campaign No. of holes Depth Ambatovy 2008–2017 907 46,005 DMSA 2003–2007 291 15,772 GENiM 1970–1972 136 6,816 Ambatovy Phelps Dodge 1995–1998 421 20,080 SNC-Lavalin ?? 4 107 Subtotal – Ambatovy 1,759 89,679 Ambatovy 2008–2017 618 19,859 CRS ?? 2 35 Analamay DMSA 2003–2007 321 12,759 GENiM 1970–1972 166 5,718 Subtotal – Analamay 1,107 38,370 TOTAL 2,866 128,049
HQ core was collected, and recoveries were set at a minimum of 90%, with the overall average being 92%. If recovery was less than 50 cm for each metre drilled for three continuous metres, a redrill of the hole was requested. Core losses occurred when using a tricone bit to collar the hole and occasionally when cutting mud zones or perched boulders in the ferralite/LMS units and highly fractured zones in the saprolite. The drillholes were drilled 3 m into fresh rock or 5 m into brecciated bedrock to minimise the chance of stopping the hole prematurely in a large boulder. The core was extruded from the core barrel by hydraulic pressure into polythene tubes to prevent moisture loss. The polythene tubes filled with drill core were then placed in PVC pipe (that had been cut in half) and then placed into core boxes. The drilling interval advances were recorded in meters on wooden blocks at the drill site and placed appropriately in the core box. The drilling activities at Ambatovy were monitored and recorded by the drill geologist present at the drill rig during the drilling (day and night). A daily field report was submitted by the field rig geologist to the geology office (senior geologist) indicating drillhole, depth and recoveries. When the feasibility study drill program was completed in April 2004 by the Dynatec Drilling Division, the Dynatec Drilling Division demobilised all its rigs to the USA except the UDR 650, which was transferred to DMSA. DMSA continued an infill drilling program from mid-2004 up to the end of 2007 as part of the development of the Ambatovy Project, and in order to convert Indicated Resources into Measured Resources. When Sherritt took overDMSA in mid-2007, a new drilling program was designed. This Mineral Resource definition drilling started mid-2008 and continues to the present.
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Figure 15: Ambatovy drillhole locations until 2017 Source: Ambatovy Joint Venture, 2018
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Figure 16: Analamay drillhole locations until 2017 Source: Ambatovy Joint Venture, 2018
10.2 Logging and Sampling Methodology
10.2.1 Logging and Sampling Since 2003, when DMSA started to work on the Ambatovy Project, all logging activities have been conducted on-site. The sampling methodology evolved with time.
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First, during the early feasibility study drilling program (July 2003 to April 2004), the procedure was as follows: at the camp site, the polythene tubes were removed, the core was measured and the depth intervals marked in metres. The core was photographed by digital camera, and when necessary, re- photographed for improved quality. The core was logged according to the principal lithologies and their degree of weathering (Table 13). Sampling was routinely done at 1 m intervals but broken at sub-metre intervals at significant lithological contacts. The highly weathered drill core (ferralite and LMS) was then split in half with a knife. For the bouldery saprolite, the sections with minus 10 cm boulders and fines were split in half and the boulders cut using a diamond core saw when necessary. The plus 10 cm (fresh) boulders were measured but not sampled. The half core samples were placed in plastic bags, tagged and shipped in sealed drums to UltraTrace, the analytical laboratory in Perth, Australia for sample preparation. The remaining half cores were sheathed in polythene tubes and the core boxes sent to the on-site core storage area. More than 26,000 samples were sent to UltraTrace for preparation into subsamples. From April 2004, when DMSA started the infill drilling program, the sampling methodology described above remained in place, but the half core samples were no longer shipped directly to UltraTrace for preparation, instead the sample preparation was performed at the mine site, and the remaining half cores were stored in core boxes in the on-site core storage area. In total 122,901 samples were analysed. UltraTrace is still used as external laboratory for the QAQC program.
10.2.2 Density Measurement Samples of the drill core during drilling were collected for density and moisture content determinations every fifth metre down the drillhole. After weighting the sample, it was wrapped in a thin plastic film and the volume determined by the water displacement in a bucket. The sample size of 50 cm of half core, used at the beginning of the drill program, was changed to 25 cm of full core on 16 March 2004. This was done because of a problem with volume measurement due to air pockets within the plastic film. Subsequently, as determined, a 18% correction was made to all the volume determinations after16 March 2004. Dry weight was determined by weighing the sample after 24 hours of oven drying at a temperature of 105°C. Wet weight, moisture content and dry weight were determined and from this, the density (g/cm3) and moisture content (%) were calculated as shown below:
퐷푟푦 푤푒𝑖𝑔ℎ푡 (𝑔) 퐷푟푦 푏푢푙푘 푑푒푛푠𝑖푡푦 = 푉표푙푢푚푒 (푐푚3)
푊푒𝑖𝑔ℎ푡 푤푎푡푒푟 (𝑔)푥 100 푀표𝑖푠푡푢푟푒 푐표푛푡푒푛푡 (%) = 푊푒푡 푤푒𝑖𝑔ℎ푡 (𝑔)
Where weight of water equals the difference between wet and dry weights in grams. An appropriate portion of the dried sample was combined with the sample interval being sent to Perth for sample preparation and assaying.
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Table 13: Codes for drill core logging
Description Codes Lithological codes Ferricrete (cuirasse) CF Pisolite (grenaille) GR Limonite (undifferentiated) LIM Red ferralite (laterite rouge) LR Red ferralite with Pisolite (laterite rouge a grenaille) LRG Yellow limonite (laterite jaune) LJ Yellow limonite with Pisolite (laterite jaune a grenaille) LJG Low magnesium saprolite (transition zone with relic structure) LMS Peridotite PER Pyroxenite PYR Dunite DN Dolerite DOL Syenite SYE Gabbro GB Serpentinite SER Accessory mineral codes Manganese (WAD) Mn Cobalt Co Chromite Cr Garnierite Ga Secondary silica SiO2 Quartz Qz Weathering codes A scale of 0 to 6 to describe degree of weathering of core: 0 – Fresh rock no alteration 1 – Fresh rock with weathered cracks/joints (typical bedrock) 2 – Rock with advanced weathering in joints and fractures 3 – Hard weathered rock (typical rocky saprolite) core cannot be broken by hand 4 – Soft weathered rock (typical earthy saprolite) core can be broken by hand 5 – Earthy material with visible structures (typical of LMS and transition saprolite) 6 – Complete alteration (typical limonite)
In 2006, a QAQC program was introduced by AMEC, which recommended modifying the procedure of determination for the moisture and density. The new procedure reduced significantly the errors experienced by the previous method. The major improvement was the use of bench scale that can measure directly the weight of a core sample, but also its weight when the core is immersed into water. As a result, there were more consistent results when calculating moisture content and dry density. Ambatovy is still applying the AMEC procedure for density and moisture. Every 5 m in average, 15–20 cm long of core is trimmed directly on the field. The density sample is tightly wrapped with a thin plastic foil and bagged in plastic bags. The sample is weighed in air (wet weight in air: A) and after in water (Wet weight in water: B). The sample will then be unwrapped and dried in the oven for 12 hours at 105°C. The sample is weighed in air (Dry weight in Air: C).
Density= C/(A-B)
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Moisture= (A-C)/C
퐷푟푦 푤푒𝑖𝑔ℎ푡 𝑖푛 푎𝑖푟 퐷푟푦 푑푒푛푠𝑖푡푦 = 푊푒푡 푤푒𝑖𝑔ℎ푡 𝑖푛 푎𝑖푟 − 푊푒푡 푤푒𝑖𝑔ℎ푡 𝑖푛 푤푎푡푒푟
푊푒푡 푤푒𝑖𝑔ℎ푡 𝑖푛 푎𝑖푟 − 퐷푟푦 푤푒𝑖𝑔ℎ푡 𝑖푛 푎𝑖푟 푀표𝑖푠푡푢푟푒 푐표푛푡푒푛푡 (%) = × 100 퐷푟푦 푤푒𝑖𝑔ℎ푡 𝑖푛 푎𝑖푟
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11 Sample Preparation, Analyses and Security
From 2003 to 2017, three generations of assays using inductively coupled plasma (ICP) have been performed for the Ambatovy Project drill core samples: • 2003–2004: o Primary Laboratory: UltraTrace o Secondary Laboratory: DYFS. • 2004–2008: o Primary Laboratory: DYFS o Secondary Laboratory: UltraTrace. • 2009–2017: o Primary Laboratory: AMSA mine site lab o Secondary Laboratory: UltraTrace. The UltraTrace laboratory in Perth, Western Australia is registered with the National Association of Testing Authorities (NATA) of Australia with a laboratory registration number of 14492. This registration imposes critical requirements, not only on the various testing performed, but also on document controls, administration systems and technical competence. UltraTrace meets NATA’s requirements for registration to ISO/IEC 17025 (1999) standards. UltraTrace was acquired by the Bureau Veritas Group in 2007 and now operates as Bureau Veritas Minerals. Bureau Veritas Minerals maintains an ISO9001.2000 quality system for all its laboratories and the Ultratrace lab remains ISO 17025 accredited and NATA registered. In this Technical Report, the laboratory will continue to be referred to as UltraTrace. Ultratrace and its employees are independent from the issuer and the Ambatovy Project. The DYFS laboratory is located in Fort Saskatchewan, Alberta, Canada. Also known as the Sherritt Technical Services Laboratory, DYFS and its employees are not independent from the issuer and the Ambatovy Project. The AMSA Ambatovy laboratory is located at the Ambatovy mine site. The AMSA lab and its employees are not independent from the issuer and the Ambatovy Project.
11.1 2003 to 2004 Sample Preparation and Analysis During the feasibility study, half-core samples from all the drilling conducted for resource estimation purposes were shipped from the Project site in sealed drums to UltraTrace for sample preparation (drying, crushing, pulverising to 75 µm and splitting into subsamples). Fifty-five percent of the subsamples were assayed by UltraTrace (the primary lab) and the remaining 45% were sent to the DYFS lab for analysis (the secondary lab). UltraTrace also sent 20% of the subsamples analysed by them to the secondary lab for analysis, resulting in approximately 10% of all the subsamples being independently check assayed by both UltraTrace and the DYFS secondary lab, which managed the QAQC of the analytical program. A detailed description of the UltraTrace sample preparation procedure is provided in the 2014 NI 43-101 report (Daigle et al., 2014). The analytical method used by UltraTrace was mixed acid digestion fusion/ICP analysis for Ni, Co, Fe, Mg Mn, Zn, Cu, Al, S and Ca and sodium peroxide fusion/ICP analysis for Si, Cr, Ti and Mg.
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11.2 2004 to 2008 Sample Preparation and Analysis After the feasibility study drilling program, DMSA completed sample preparation on site using the following steps: • Drying (with an electric oven) • Crushing (with a jaw crusher) • Pulverising (with a ring pulveriser). This procedure for sample preparation is equivalent to the UltraTrace procedure. The pulps produced during the sample preparation on site were shipped to the DYFS lab (theprimary lab) in Canada and to UltraTrace (the secondary lab) in Australia. The analytical method used by DFS was lithium borate fusion /ICP analysis for Ni, Co, Fe, Mg, Mn, Zn, Cu, Al, S, Ca, Si, Cr and Ti. The analytical method used by secondary lab UltraTrace continued to be mixed acid digestion fusion and sodium peroxide fusion with ICP analysis. In 2008, a temporary laboratory using ICP was set up at the Property, and all samples were subsequently re-directed to this temporary lab.
11.3 2009 to 2017 Sample Preparation and Analysis The AMSA mine-site laboratory became the primary lab of the Ambatovy Project, and the UltraTrace laboratory remained the secondary lab. The AMSA laboratory sample preparation procedure is shown in Figure 17. Steps in the analytical procedure are as follows: • Sample drying: o Weigh the sample: Wet weight o Dry the sample in the oven for 12 hours at 105°C o Weigh the sample: Dry weight o Tag the sample with a barcode to integrate into the Laboratory Information Management System (LIMS) and send the sample to the internal lab for preparation. • Preparation steps: o Crush the sample with a jaw crusher at 2 mm (d95) o Homogenise the sample (three passes) with the Jones Riffle splitter: 100 g is used for pulverisation o Pulverisation at 106 µm: 10 g is used for ICP-OES analysis. • AMSA analytical method: o The sample is melted at 1,050°C with an addition of lithium borate flux and dissolved with dilute HCl o The sample is analysed using a Varian 725-ES ICP-OES instrument o The assays are recorded in the LIMS system and exported in csv format for the end user. o The analytical method used by secondary lab, UltraTrace, continues to be mixed acid digestion fusion and sodium peroxide fusion with ICP analysis.
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Figure 17: Sample preparation flowsheet at Ambatovy laboratory
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11.4 DMSA Validation and QAQC Programs (2003 to 2008) The following sets out the QAQC program designed by DMSA, as taken from the DMSA memo, titled “Quality Assurance and Quality Control for the Ambatovy Drill Core Assaying Project” by Frank Pau, dated 3 May 2004.
11.4.1 Initial Check of Phelps Dodge Assays In October 2003, preparation work began for the assaying of the Ambatovy Project drill core materials. Samples were received from the Property that were previously assayed by Phelps Dodge using both x-ray fluorescence (XRF) and ICP methods. A total of 150 samples were analysed at the DYFS laboratory. The comparison of the DMSA and Phelps Dodge assays indicated good agreement in the two sets of data with the exception of chromium. With the completion of this comparative work, the validation of the analytical method began.
11.4.2 Analytical Method and Laboratory Validation The validation of method included two rounds of inter-laboratory comparison using the Ambatovy Project laterite samples from Phelps Dodge. The first round was conducted in late October 2003. The comparison data include assay results of six laterite samples from six laboratories as shown in Table 14. From these data, it was concluded that all the laboratories were in good agreement with the exception of the ACME Lab. In November 2003, a second round of inter-laboratory comparison was conducted. Six samples and a certified reference standard were analysed by four labs. The comparison data is shown in Table 15. The data indicate that for the most important elements of nickel and cobalt, DMSA, Sherritt and UltraTrace were in good agreement, with SGS being the exception. Based on these comparison studies, it was concluded that DMSA, UltraTrace and Chemex would be good laboratory candidates for any future analytical work involving the Ambatovy Project ore materials. Between the first and second inter-laboratory comparisons, a certified reference standard was acquired. This standard was analysed in replicates as an assurance that the method used at the DMSA laboratory has the accuracy required. Table 14 shows the analytical data.
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Table 14: Assay results of six laboratories
2268-06 2270-08 2272-01 2272-04 2278-07 2268-06 2270-08 2272-01 2272-04 2278-07 Ni Co Dy 0.74 1.48 1.73 1.63 1.63 Dy 0.023 0.111 0.112 0.024 0.136 Ultra ICP 0.785 1.58 1.87 1.72 1.78 Ultra ICP 0.024 0.112 0.119 0.025 0.14 Chemex 0.723 1.52 1.78 1.63 1.67 Chemex 0.024 0.106 0.113 0.024 0.132 Shr 0.744 1.52 1.83 1.68 1.69 Shr 0.0252 0.113 0.123 0.0231 0.142 Acme 0.83 1.72 1.88 1.77 1.90 Acme 0.03 0.11 0.11 0.02 0.14 Ultra XRF 0.772 1.56 1.88 1.75 1.75 Ultra XRF 0.023 0.115 0.120 0.024 0.144 Phelps-Dodge 0.87 1.51 1.73 1.61 1.56 Phelps-Dodge 0.023 0.110 0.109 0.025 0.13 Ultra/Dy 1.06 1.07 1.08 1.06 1.09 1.04 1.01 1.06 1.04 1.03 Cu Zn Dy 0.008 0.015 0.017 0.015 0.024 Dy 0.015 0.05 0.048 0.072 0.057 Ultra ICP 0.0065 0.0165 0.0185 0.0125 0.0265 Ultra ICP 0.0154 0.0504 0.0492 0.074 0.058 Chemex 0.006 0.014 0.015 0.010 0.024 Chemex 0.021 0.061 0.062 0.082 0.061 Shr 0.0064 0.0156 0.0161 0.0117 0.0229 Shr 0.0172 0.05 0.0494 0.0745 0.0588 Acme 0.02 0.02 0.02 0.03 Acme 0.02 0.06 0.05 0.08 0.06 Ultra XRF 0.005 0.0175 0.02 0.0145 0.0275 Ultra XRF 0.0155 0.05 0.0515 0.078 0.058 Phelps-Dodge 0.0082 0.0145 0.0149 0.0112 0.0186 Phelps-Dodge 0.021 0.061 0.058 0.088 0.07 Ultra/Dy 0.81 1.10 1.09 0.83 1.10 1.03 1.01 1.03 1.03 1.02 Fe Al Dy 13.1 45.8 49.6 51.6 49.1 Dy 13.3 4.31 1.61 1.55 2.17 Ultra ICP 13.4 46.9 52.3 53.3 50.9 Ultra ICP 15.4 4.96 1.86 1.78 2.61 Chemex 12.5 44.5 48.1 48.2 47.4 Chemex 14.5 4.50 1.73 1.60 2.34 Shr 12.8 44.6 49.0 50.0 47.7 Shr 13.9 4.24 1.56 1.52 2.20 Acme 12.8 48.4 51.1 52.2 51.4 Acme 15.01 5 1.87 1.87 2.97 Ultra XRF 13.2 46.9 51.9 53.1 50.2 Ultra XRF 15.7 4.9 1.87 1.79 2.54 Phelps-Dodge 14.9 47.5 51.3 52.0 49.6 Phelps-Dodge 14.1 3.91 1.41 1.41 2.15 Ultra/Dy 1.03 1.02 1.05 1.03 1.04 1.16 1.15 1.16 1.15 1.20
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2268-06 2270-08 2272-01 2272-04 2278-07 2268-06 2270-08 2272-01 2272-04 2278-07 Cr Dy 0.352 0.996 1.31 1.54 1.42 Ultra ICP 0.36 1.03 1.32 1.57 1.46 Chemex 0.380 0.942 1.28 1.47 1.36 Shr 0.421 1 1.32 1.56 1.44 Acme 0.48 0.97 1.29 1.5 1.41 Ultra XRF 0.389 1.03 1.35 1.59 1.44 Phelps-Dodge 0.62 1.71 2.18 2.52 2.24 Ultra/Dy 1.02 1.03 1.00 1.02 1.03 Notes: • Six-lab interlaboratory comparison on five Ambatovy drill core samples from Phelps Dodge. • Samples except Phelps Dodge’s were analysed in the latter half of October 2003.
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Table 15: Inter-laboratory comparison of assays results
Notes: • A6 is the certified reference material 610-1. The figures in the box are the certified values. • Samples were analysed in the first half of December 2003. Table 16: Certified reference standard analyses
Element Units Rep. 1 Rep. 2 Rep. 3 Rep. 4 Rep. 5 Rep. 6 Average STD Cert. Value Aluminum % wt 1.80 1.81 1.78 1.81 1.83 1.82 1.808 0.017 1.96 Cadmium % wt 0.001 0.001 0.001 0.001 0.001 0.001 0.001 0.000 Calcium % wt 0.095 0.096 1.05 0.093 0.093 0.09 0.253 0.391 Chromium % wt 1.81 1.86 1.84 1.79 1.82 1.85 1.828 0.026 1.84 Cobalt % wt 0.075 0.075 0.073 0.074 0.075 0.075 0.075 0.001 0.075 Copper % wt 0.004 0.003 0.003 0.004 0.003 0.003 0.003 0.001 Iron % wt 47.9 47.8 47.0 47.2 47.3 47.4 47.433 0.350 47.46 Magnesium % wt 1.86 1.87 1.83 1.87 1.89 1.87 1.865 0.020 1.86 Manganese % wt 0.618 0.615 0.616 0.606 0.613 0.612 0.613 0.004 0.581 Nickel % wt 1.52 1.52 1.51 1.53 1.53 1.56 1.528 0.017 1.48 Silicon % wt 3.11 3.12 3.13 3.13 3.14 3.13 3.127 0.010 3.16 Sulphur % wt 0.179 0.169 0.189 0.174 0.198 0.183 0.182 0.010 0.189 Titanium % wt 0.012 0.011 0.011 0.013 0.012 0.012 0.012 0.001 0.015 Zinc % wt 0.055 0.054 0.055 0.055 0.054 0.054 0.055 0.001 Notes: • Analysis of the certified reference standard 610-1 provided by IRSID, a French institute of iron metallurgical research. • The standard was analysed six times on 31 October 2003. Each replicate included lithium borate dissolution.
11.4.3 Quality Control The quality control plan was established in March 2004. The samples were analysed in batches of 20 to 25 including a duplicate of one of the samples in the batch, a quality control check sample, which is either an in-house standard or a certified standard. A reagent blank is also included. Table 17 shows a typical analysis report illustrating the insertion of duplicate, blank and standard samples in every 25-sample batch. Both the electronic and paper copies of the analysis reports are maintained.
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Table 17: Typical analytical report
Notes: • Date of analysis: 14 April 2004. • Report number: 200401621. • Method of analysis: ICP spectrometry. • All figures are in % weight.
11.4.4 Interlaboratory Cross-Checks Cross-checking of samples was carried out in the three labs. At the end of the feasibility study drilling program, 1,050 samples were cross-checked between the UltraTrace and DMSA labs, 979 samples between the Chemex and DMSA labs, and 858 samples were cross-checked among UltraTrace, Chemex and DMSA labs. Table 18 sets out the main analytical procedures conducted at each of the three laboratories selected for the assaying of the lateritic nickel samples. Each of the laboratories is ISO certified and carries out its operations under the rules and procedures set by the registration organisations. Table 18: Analytical procedures summary
Sample Analytical Standards and Laboratory Digestion Elements/Tests Changes and additions preparation methods duplicates Ni, Co, Fe, Mg Duplicates and Core samples are Mixed acids ICP Mn, Zn, Cu, Al, standards every None dried, crushed and S, Ca 12 samples UltraTrace pulverised to Sodium Blanks every 75 µm peroxide ICP Si, Cr, Ti, Mg 72–127 samples fusion 1 duplicate, 2 No sample Ni, Co, Fe, Mg, standards and 1 preparation Mixed acids ICP Mn, Zn, Cu, Al, None blank every 40 required; samples S, Ca Chemex pulverised by samples UltraTrace were Sodium ICP Si, Cr, Fe submitted in glass peroxide vials by DMSA fusion Gravimetric LOI at 1,000°C
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Sample Analytical Standards and Laboratory Digestion Elements/Tests Changes and additions preparation methods duplicates Beginning 13 April 2004, Samples pulverised 1 duplicate, 1 fusion time was changed Lithium Ni, Co, Fe, Mg, by UltraTrace were standard and 1 from 30 minutes to Dynatec borate ICP Mn, Zn, Cu, Al, dried at DMSA lab blank every 25 60 minutes; sample fusion S, Ca, Si, Cr, Ti before digestion samples weights were changed from 0.2 g to 0.15 g
11.5 2009 Round-Robin Laboratory Tests Round-robin tests involving four different laboratories were performed in 2009 to produce another set of standards. The samples were sent to SGS, CHEMEX, Sherritt Technology and UltraTrace laboratories. Samples were divided in four categories: high-grade nickel, low-grade nickel, LMS and aluminous ferralite. Standard samples were sent with control samples (Ambatovy Project, 2009c).
11.6 Validation Test of AMSA ICP Lab In 2009, the mine site analytical ICP laboratory became operational. AMSA conducted validation tests to guarantee that the mine lab could be used as the primary lab of the drilling sampling. The tests covered contamination, precision and accuracy. Two sets of samples were sent to Sherritt Technology – Fort Saskatchewan, to UltraTrace and to AMSA lab, including blanks, standard samples and duplicates. The results showed that contamination, precision and accuracy of AMSA ICP lab were adequate (Ambatovy Project, 2010c).
11.7 Ambatovy QAQC (2009 to 2017) Since 2009, Ambatovy joint venture implemented a QAQC procedure for the exploration program based on AMEC recommendation. Table 19 shows the internal QAQC sequence for 100 samples. Table 19: Internal QAQC
Type Percentage Standard 7% Primary 75% Twin sample 3% Fine blank 3% Coarse blank 3% Twin sample for external lab 4% Coarse duplicate 3% Fine duplicate 2%
For the samples sent to UltraTrace, the secondary external laboratory, the sequence is shown in Table 20. Table 20: Internal QAQC – UltraTrace
Type Percentage Fine blank 5% Fine duplicate 5% Primary 85% Standard 5%
Prior to 2016, the sample and QAQC validations were done with Microsoft Access and Excel software. In 2016, all validated data were migrated to Acquire, a proprietary geological database software. All QAQC validation is now done in Acquire. The standard, blank and duplicate analytical results are checked and if
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necessary, re-assays were done to avoid eventual issues. If results for standards fell outside the set limits, five samples before and five samples after the standard are re-analysed. Duplicates are treated similarly.
11.7.1 Standards Figure 18 and Figure 19 show Ni results from two standards between 2009 and 2017. The Ni assays are inside the limits, but the Ni assays of the standard shown in Figure 18 are consistently below the accepted standard value indicating a bias towards underestimation.
Ni Standard 2009-2017 1.5
1.45
1.4
1.35
1.3
1.25
1.2
1.15
1.1
1.05
1
2010 2012 2014 2017 2009 2009 2009 2009 2009 2009 2010 2010 2010 2011 2011 2011 2012 2012 2012 2013 2013 2013 2013 2013 2013 2014 2014 2014 2014 2015 2015 2016 2016 2016 2017 2017 2017 2017 Year assay standard value upper limit Lower limit
Figure 18: Internal standard 1 for nickel (%) from 2009 to 2017
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Ni Standard 2009-2017 1.05
1
0.95
0.9
0.85
0.8
0.75
0.7
0.65
0.6
2011 2013 2014 2017 2009 2009 2009 2009 2010 2010 2010 2010 2011 2011 2011 2012 2012 2012 2012 2013 2013 2013 2013 2014 2014 2014 2014 2015 2015 2016 2016 2017 2017 2017 2017 2009 Year assay standard value upper limit Lower limit
Figure 19: Internal standard 2 for nickel (%) from 2009 to 2017 For cobalt, the standard assays are rather variable but respect the limits (Figure 20); only 0.30% of assays are outside limits.
Standard 2009-2017 Co 0.13
0.125
0.12
0.115
0.11
0.105
0.1
0.095
0.09
0.085
0.08
2011 2012 2015 2015 2009 2009 2009 2010 2010 2010 2011 2011 2011 2012 2012 2012 2013 2013 2013 2013 2013 2013 2014 2014 2014 2014 2014 2014 2015 2016 2016 2017 2017 2017 2017 2017 2017 Year assay standard value upper limit Lower limit
Figure 20: Standard for cobalt (%) from 2009 to 2017
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11.7.2 Duplicates Coarse and fine duplicates are systematically integrated in the QAQC sequence, Figure 21 and Figure 22 show the Ni results. Note that when the nickel grade is low (<0.4% to 0.5%), there is considerable scatter outside the ±20% limit. After re-assays were done, there was no improvement in these results.
Ni Duplicate Coarse Duplicate 2009-2017 6.5 6 5.5 5 4.5 4 3.5 3 2.5 2 1.5 1 0.5 0 0 0.5 1 1.5 2 2.5 3 3.5 4 4.5 5 5.5 Ni orginal Coarse duplicate 10% limit 20% limit
Figure 21: Coarse duplicate for nickel (%) from 2009 to 2017
Ni Duplicate Fine Duplicate 2009-2017 6.5 6 5.5 5 4.5 4 3.5 3 2.5 2 1.5 1 0.5 0 0 0.5 1 1.5 2 2.5 3 3.5 4 4.5 5 5.5 Ni orginal Fine duplicate 10% limit 20% limit
Figure 22: Fine duplicate for nickel (%) from 2009 to 2017 For duplicate cobalt assays, the same issue arises at lower grade as is seen for nickel – considerable scatter outside the 20% limit (Figure 23). When an issue was noted, a re-analysis was performed to confirm the original sample grade.
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Co Duplicate Duplicate 2009-2017 1.9 1.8 1.7 1.6 1.5 1.4 1.3 1.2 1.1 1 0.9 0.8 0.7 0.6 0.5 0.4 0.3 0.2 0.1 0 0 0.1 0.2 0.3 0.4 0.5 0.6 0.7 0.8 0.9 1 1.1 1.2 1.3 1.4 1.5 1.6 Co orginal Duplicate 10% limit 20% limit
Figure 23: Coarse and fineduplicate for cobalt (%) from 2009 to 2017
11.7.3 Blanks Pink Granite is used to generate coarse and fine blanks to check for contamination during preparation or analysis process. Figure 24 and Figure 25 show the values. Less than 10% of the values are outside the allowable limit.
Ni Fine Blank 0.07
0.06
0.05
0.04
0.03
0.02
0.01
0
2009 2012 2013 2016 2017 2009 2009 2009 2009 2010 2010 2010 2010 2011 2011 2011 2011 2012 2012 2012 2012 2013 2013 2013 2013 2014 2014 2014 2014 2014 2015 2015 2016 2017 2017 2017 2017 2017 Year Ni upper limit
Figure 24: Fine blank for nickel from 2009 to 2017
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Ni Coarse Blank 0.07
0.06
0.05
0.04
0.03
0.02
0.01
0
2013 2013 2013 2009 2009 2009 2009 2010 2010 2010 2011 2011 2011 2011 2012 2012 2012 2012 2013 2013 2014 2014 2014 2014 2014 2015 2015 2016 2016 2017 2017 2017 2017 2017 Year Ni upper limit
Figure 25: Coarse blank for nickel from 2009 to 2017
11.7.4 Summary of 2009–2017 QAQC Results Table 21 shows the total numbers of failures for nickel between 2009 and 2017 in the QAQC process. Table 21: Failure from 2009 to 2017
Total Failure Percentage failed Blank 3,186 25 0.78% Standard 3,710 12 0.32% Duplicate 3,254 575 17.67% Ni >0.4% 2,030 143 7.04%
11.8 Inter-Laboratory Pulp Check Samples The external validation for pre-2014 analyses was discussed in the previous NI 43-101 Technical Report (Daigle et al., 2014). The external validation (pulp check samples at second laboratory) for nickel since the last Mineral Resource estimation in 2013 are presented in Figure 26 and Figure 27. Comparison between the data from 2014 and 2015 onwards indicates there were problems with the 2014 validations; when these are removed from the dataset, validation is considerably improved (Figure 27). Figure 28 shows the external validation for cobalt since 2014. Low-grade cobalt and nickel pulp check samples have high variability with scatter outside the ±20% limit. The variability noted in the duplicates with low-grade values (<0.4% Ni) will not have any impact on the Mineral Resource estimate due to the cut-off grade at 0.6% Ni.
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Figure 26: Internal assays vs external assays for nickel since 2014
Figure 27: Internal assays vs external assays for nickel since 2015
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external value (UltraT race) Internal vs External assays 2014-2017 0.8
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Figure 28: Internal assays vs external assays for cobalt since 2014
11.9 Qualified Person’s Opinion and Conclusions It is the Qualified Person’s opinion that security, sample collection, preparation and analytical procedures undertaken on the Ambatovy Project during the 2014–2017 drill programs are appropriate for the sample media and mineralisation type and conform to industry standards. It is the Qualified Person’s opinion that QAQC program undertaken during the 2014–2017 Ambatovy Project drill programs are appropriate and conform to industry standards. The 2014–2017 blank, certified reference material and duplicate sample results provide sufficient confidence in the 2014–2017 drill core assay values for their use in the estimation of CIM compliant mineral resources. The Qualified Person notes that the oxide analytical sums were checked and 189 samples (0.2% of the total samples) were excluded from the estimation because the oxide sum was below 30%. In addition, one drillhole was excluded due to no accurate coordinates and 226 drillholes were not used for this estimation due to pending external lab checks (2017 campaign).
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12 Data Verification
The Ambatovy project was visited by co-author Qualified Person, Dr Adrian Martinez Vargas, P.Geo., of CSA Global, for six days from 24 June to 1 July 2018. The purpose of this visit was to conduct a due diligence inspection of the Ambatovy mine site and the processing plant located at Toamasina, working with the Ambatovy team to review the Mineral Resource estimate and the informing data used, and to become familiarised with the local geology of the property and the operation in general. The mine site inspection was completed from 24 to 27 June 2018 and included a visit to both Analamay and Ambatovy Deposits, and also a visit to the mine site laboratory, the OPP and the mine pit located in the Ambatovy Deposit. The local geology was observed, including gabbro occurrence and the common distribution of limonites and saprolites in the lateritic profile (Figure 29). The ongoing drilling program at the Analamay Deposit was observed including sample collection, logging and documentation, and transportation. The sample preparation and storage facility and the local laboratory were visited, where the sample preparation, assaying and sample storage activities were observed (Figure 30). Density samples and density determinations were also observed (Figure 31). During this visit, the written protocols for sampling, assaying, sample QAQC, logging, database storage, and sample security, were reviewed and discussed with the local team.
Figure 29 Mining operations at the Ambatovy pit and exposure of the nickel laterite profile
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Figure 30 Drilling and sampling logging (upper row and middle left); sampling and logging facility (middle right), and sample storage facility (lower row)
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Figure 31 Density sample (left) and wet weight determination (right) The PAL Plant Site inspection was completed from28 June to 1 July 2018. This visit included a tour around the Processing Plant, a visit to the analytical laboratory located there, a visit to the port, and a visit to the TMF. The relevant Qualified Persons have reviewed the sample collection and analysis methodologies and are of the opinion that those methodologies are to current industry standards and permit a meaningful investigation ofthe mineralisation atthe Ambatovy Project for the purpose of resource estimation under CIM guidelines and provide the basis for the conclusions and recommendations reached in this Report. It is the opinion of CSA Global and the relevant Qualiifed Persons that the data made available to CSA Global are a reasonable and accurate representation of the Ambatovy Project and are of sufficient quality to provide the basis for the conclusions and recommendations reached in this Report.
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13 Mineral Processing and Metallurgical Testing
13.1 Summary A series of mineral processing and metallurgical testwork programs were conducted from 2003 to 2010. The results of the testwork programs were used to confirm the recovery of metal values from the ore, define the process flowsheet and process design criteria and to confirm the plant operating parameters for the initial ore feed. Initially, batch and continuous testwork was conducted in 2003 and 2004 using readily accessible ore samples to confirm the recovery of metal values from the ore samples and to define the flowsheet and design parameters. After 2004, a series of batch pressure leach tests were conducted to evaluate the amenability of a more diverse set of samples. Two OPP pilot test runs were done at the mine site, one in 2006 and one in 2008, to confirm the amenability of selected bulk ore samples to scrubbing and screening and to develop criteria which were used for the design of the commercial OPP. Ore slurry samples collected in the 2008 OPP pilot test run were used for batch and continuous metallurgical testwork conducted to confirm the amenability of ore projected as feed to the plant during the initial operating years. Finally, in 2011, batch testwork was conducted on samples collected from the low grade ore stockpile which was used as the initial plant feed to confirm the response and to provide guidance to the plant operators for the initial operating period.
13.2 2003 and 2004 Metallurgical Testwork Continuous and batch metallurgical testwork was conducted in 2003 and 2004 which was used to confirm the recovery of metal values and define the metallurgical plant process flowsheet and design parameters.
13.2.1 Sample Selection Sample material representing early and later years of operation, i.e. ferralite top (FT), ferralite bottom (FB), LMS and saprolite, in amounts reflecting relative proportions in the Ambatovy Deposit were selected for the continuous testwork program. For batch tests, sample materials were collected over a variety of principal bedrock types with a variety of laterite lithological types, and with reasonable geographic distribution.
13.2.2 Metallurgical Testwork After batch testwork to establish preliminary design criteria for continuous leaching, feed blends were made up from various representative lithological types. A continuous demonstration campaign, using optimum parameters as defined in preliminary batch testwork, was conducted. This campaign included simultaneous operation of the pressure leach, slurry neutralisation, countercurrent decantation and solution neutralisation unit operations. Solution produced during the continuous demonstration campaign was collected and then used as feed for continuous sulphide precipitation testwork. The sulphide precipitation testwork achieved recoveries of over 99% for nickel and cobalt from the PAL solution. The mixed sulphides produced were in turn used as feed to metal refining testwork which included batch pressure oxidative leaching, batch impurity removal, continuous solvent extraction, batch cobalt strip solution purification and metals recovery via batch hydrogen reduction. The testwork demonstrated a suitable refinery flowsheet for production of high-quality nickel and cobalt metal products.
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13.2.3 Ore Variability Testwork Samples of various ore types from Ambatovy West were collected from HQ core drilling in December 2003 for batch metallurgical testing by DMSA. The drillhole locations were selected to represent the laterite lithologies overlying dunite (the principal source rock), and a variety of other rock types including, magnetic dunite, peridotite, pyroxenite, olivine gabbro/dolerite and syenite. In addition, four holes locations were selected from each of the Analamay Deposit (North, Central and South) and Ambatovy Southeast, based on laterite lithology and analyses of the DMSA drill cores. Batch PAL and slurry neutralisation tests were conducted to evaluate the response of these ore samples. In addition, settling tests were performed on both feed and leach slurries, as well as the neutralisation slurries, to characterise the solid-liquid separation characteristics of the individual ore samples and the leach residues. Overall, the thickening, leaching, leach residue settling and leach slurry neutralisation characteristics of various feeds from Ambatovy West, Ambatovy Southeast and the Analamay Deposit, behaved similarly. These results, taken with the similarity of chemical compositions of the deposits indicate that the process plant designed on the continuous test results for Ambatovy West, will be capable of treating the ores from Ambatovy Southeast, Central and the Analamay Deposit.
13.3 Ore Preparation Pilot Plant Mineral Processing Campaigns OPP testwork was performed in a pilot plant at the Ambatovy mine site from 10 August to 14 September 2006. Ore from five locations was excavated from test pits and delivered to the pilot plant. Four of the test pits were located in the Ambatovy Deposit and one in the Analamay Deposit. The pilot plant consisted of two stages of scrubbing with vibrating screens after each scrubber to separate the undersize slurry from the oversize fraction. The undersize fraction from each stage of scrubbing was directed to a thickener; the oversize from the primary scrubber was directed to the secondary scrubber. The oversize from the secondary scrubber was collected as reject from the process. A second series of ore OPP testwork was conducted in 2008. Twenty-four test runs were completed on 12 distinct ore samples. The samples tested originated from six pits, two from each of the Ambatovy West, Ambatovy South East and Analamay orebodies. The pits were 10 m deep, with two samples processed from each pit, one sample from 0 m to 5 m and the second sample from 5 m to 10 m.
13.4 Pressure Leach Testwork Ore that was processed in the OPP pilot plant campaign was collected from three different zones in the Project deposit: Ambatovy West (AMB W), Ambatovy South East (AMB SE), and Analamay (ANA). These feed samples were each mined from a relatively narrow depth range, between 5 m and 10 m. Thirty-two drums of laterite ore slurry, produced in Madagascar as thickener underflow during the ore preparation pilot plant program in May and June 2008, were sent to Fort Saskatchewan. The material comprised about 1.7 t (dry basis) of ore.
13.4.1 Batch Testwork Batch pressure leach test-work was conducted with the three ore samples described above, as well as a fourth sample obtained from 0 m to 5 m in Ambatovy West. High nickel and cobalt extractions were obtained in the batch tests, consistently exceeding the PDC value of 95% for these elements.
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13.4.2 Continuous Testwork A continuous pilot plant program, seven days in duration, was conducted with the following unit operations: • Pressure acid leaching with Ambatovy West, Ambatovy South East, and Analamay Deposit ore feeds • Slurry neutralisation of the leach discharge • Countercurrent decantation( CCD) washing of the slurry neutralisation discharge • Raw liquor neutralisation with limestone • Tailings neutralisation with limestone and lime. Pressure Leach Results The target pressure leach operating conditions for most of the pilot plant campaign were the same as the process design criteria, including 75 minutes autoclave retention time at 260°C. Excellent nickel and cobalt extractions 95% or greater were achieved from all three of the ore types tested in the pilot plant campaign. Slurry Neutralisation The slurry neutralisation circuit consisted of three reaction tanks in series, operated at 90°C with a total retention time of 60 minutes with limestone slurry added into the first tank. Generally, there was little change in the concentrations of nickel, cobalt, aluminum, chromium, iron or manganese across the circuit in either the solids or solution phases, and the extractions of nickel and cobalt were not affected. The utilisation of limestone was approximately 97%. Countercurrent Decantation The CCD wash circuit included seven thickeners with counter current flows of neutralised slurry solids and overflow solutions. The circuit performance was excellent. Raw Liquor Neutralisation The raw liquor neutralisation circuit included four reaction tanks in series, operated at 85°C with a total retention time of between 84 minutes and 108 minutes. The discharge slurry was consistently maintained between pH 3.6 and pH 3.8, and iron was precipitated to less than 0.18 g/L in solution. The limestone utilisation throughout the pilot plant campaign was excellent. Tailings Neutralisation Continuous neutralisation of the CCD tailings with limestone and lime was conducted concurrently in three discrete one-day campaigns that were partially integrated with the operations listed above. Four reactors were operated in cascade at 75°C, with the first three simulating the commercial tanks (with 55 minutes of retention in total) and the final vessel simulating the tailings pipeline (60 minutes). Limestone slurry was added to Tank 1 to give a target pH of 6.0 in Tank 2, while slaked lime slurry was added to Tank 3 to give a target Tank 4 (“pipeline discharge”) pH of 8.0 and to precipitate manganese to less than 0.2 g/L. The limestone additions were typically150 –200% of the stoichiometric requirements. Aluminum, iron and chromium were typically precipitated to below 0.005 g/L in Tank 2. The addition of lime to meet the stoichiometric requirements for the target extents of manganese and magnesium precipitation was sufficient to precipitate manganese to below 0.2 g/L in Tank 3 and Tank 4.
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13.5 Low-Grade Ore Testwork At the mine site, mined low-grade ore was categorised as Code 4 for material between 0.8% Ni and 1.0% Ni and Code 5 for material between 0.6% Ni and 0.8% Ni and was placed in different stockpiles. Three samples were collected from each material type in December 2010 and testing was conducted in early 2011. All six samples exhibited favourable feed settling, leaching and leach slurry settling characteristics. Acceptable nickel and cobalt extractions, typically 95% or greater, were achieved under the process design conditions, and with considerably lower acid addition than in the process design.
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14 Mineral Resource Estimates
14.1 Introduction The Mineral Resource estimate of the Ambatovy and Analamay Deposits was prepared by Mr Manuel Domenech, Mine Geology Superintendent of the Ambatovy mine, under the supervision of Dr Adrian Martinez Vargas, Senior Consultant, P.Geo., and full-time employee of CSA Global, using all assays available by June 2017. Resource models for Ambatovy and Analamay Deposits were prepared using the same methodology. The process consisted of obtaining geological domains defined by surfaces and lithology domains defined at drillhole intervals, creating a block model and assigning geological domains and interpolating lithology proportions into blocks, interpolating the grade corresponding to each lithology, and assigning density to each lithology proportion into blocks. A single block may be composed by one or more lithology types, and its corresponding grade. It is assumed that certain lithologies can be mined selectively (e.g. gabbro and its weathering products can be sent to waste piles if its proportion in the block exceeds 10%). Lithologies with proportions under 10% of the block cannot be mined selectively, and for this reason its corresponding grade values were blended into lithologies with higher proportions in the block.
14.2 Informing Data
14.2.1 Drillhole Data Figure 32 and Figure 33 show the drillhole data used for Mineral Resource estimation for the Ambatovy and Analamay Deposits, with drillholes used in the previous Mineral Resource estimate (Daigle et al., 2014) shown in blue and new drillholes shown in red. Since 31 December 2013, 270 holes (13,758 m) were added in Ambatovy, representing a total of 1,744 drillholes (89,261 m). However, only 1658 drillholes were within the Deposit perimeter and used for interpolation. The drillhole spacing is mainly 50 m, except for Ambatovy Centre, where the drillhole spacing is 100 m and 200 m (Figure 32). Most of the new drilling is located in Ambatovy Southeast. The average depth of drillholes in this deposit is approximately 45 m. In Analamay, 158 drillholes (4,923 m) had been added since 31 December 2013, for a total of 777 drillholes (27,868 m) in this deposit. However, only 670 drillholes located with the deposit perimeter were used for interpolation. The average depth of these drillholes is 30 m and the most common drillhole spacing is 100 m (Figure 33).
14.2.2 Conversion from Laborde to UTM Coordinates Almost all the drillhole collars have been surveyed in the UTM coordinate system established at the mine site for the construction and development activities. Collar locations for some of the GENiM holes from the early 1970s could not be located and resurveyed in this high-precision survey system. For these 41 holes (which contain less than 5% of the total number of assays), their original collar coordinates, measured in the French colonial Laborde coordinate system, had to be transformed to UTM coordinates. A Laborde-to-UTM conversion was developed using a digitized version of the contour lines of topography that was prepared by GENiM and annotated in Laborde coordinates. The Laborde coordinates of the old topography map were translated, rotated and scaled to produce a version that matched, with a minimum squared error, the current topography. Since this procedure used thousands of data points from dozens of contour lines, it yielded a transformation that produced a more reliable result than previous attempts to establish a Laborde-to-UTM conversion. When the new coordinate transformation was used to convert the Laborde collar coordinates of the 41 GENiM holes, the discrepancies between the reported collar elevations and the ground elevation at the collar location were minor.
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14.2.3 Adjustments to Compensate for Bias in some Aluminum and Cobalt Assays Since 2004, the assay protocols have remained the same, and the ongoing QAQC program has confirmed the reliability of the assays through a consistent program that uses standards, blanks and duplicates (Section11 ). The assessment of the quality of the GENiM and Phelps Dodge assay programs, and its sample verification, was carried out with closely-spaced twin holes. Comparisons of statistical summaries, and of grade profiles in adjacent drillholes, confirmed that the assays from different drilling campaigns are consistent with each other, with no apparent bias except for two specific elements from the Phelps Dodge campaign of the 1990s: the cobalt and aluminum assays from the Phelps Dodge drillholes are systematically low by about 10%, on average. These cobalt and aluminum biases were also identified by QAQC studies done by Phelps Dodge in the late 1990s, and again by QAQC studies done by DMSA in 2004. With repeated confirmation of the low bias in the cobalt and aluminum assays of Phelps Dodge, these assay values were increased by 10%. With cobalt being one of the revenue-producing metals, the 10% increase in cobalt assays adds to the value of the resource; with aluminum being one of the deleterious metals that increases processing costs, the 10% increase in aluminum assays diminishes the value of the resource. These adjustments, which have been used in all the resource estimates since 2004, now affect less than a 15% of the samples in the assay database.
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Figure 32: Drillholes used in the current resource estimation of the Ambatovy Deposit Note: Drillholes used in the previous resource estimate (done in 2013) are shown in blue. New drillholes are shown in red. Source: AMSA, 2018
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Figure 33: Drillholes used in 2017 Mineral Resource estimation of the Analamay Deposit Note: Holes in blue were used in the 2013 estimates; holes in red are new to the 2017 estimates. Source: AMSA, 2018
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14.3 Geological Modelling Four main geological domains were identified: ferricrete, ferralite, saprolite and bedrock. Wireframes surfaces defining these domain contacts were created from polylines digitized in vertical sections with directions North-South and East-West, snapping points into drillhole intersects (Figure 34). Domain surfaces were used to flag drillholes and blocks with a domain code (Figure 35). However, the ferralite and ferricrete domains were merged into a single domain of limonites and used to interpolate grade and lithology proportions. The ferricretes domain is thin and with a limited number of samples. The ferricrete and ferralite domains were used separately to assign different densities. Interpolations were only in the limonites and saprolite domains. Four main lithology types were identified: ferralite and ferricrete, saprolite, bedrock, and gabbro. These were defined from lithology logged on drillhole samples and using geochemical information. Samples with over 7.5% Aluminum were reclassified as gabbro. Samples non-tagged as gabbro were redefined as saprolite if magnesium grade is over 1%, or to ferralite otherwise. The lithology types were then used to estimate lithologies proportions in the block and grade associated to each lithology per separate. At the Ambatovy Deposit there are three intrusive centers: a large one beneath the area known as Ambatovy West, and smaller ones beneath Ambatovy Central and Ambatovy South East (Figure 32). Ambatovy West is isolated from Ambatovy Centre and South East. Continuity exists between Ambatovy Centre and South East, and these two zones were combined for estimation.
Figure 34: Schematic vertical cross-section in Ambatovy with the different geological units
Figure 35: Schematic representationof Ambatovy block model with the geological codification
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14.4 Statistical Analysis and Grade Capping Drillhole data was composited to one metre intervals prior to statistical analysis and variography. De-clustering is not a requirement because there is no selective drilling in high-grade or low-grade areas. Anomalous high-grade values were truncated to a capping value (or top-cut; Table 22 and Table 23). Table 22 and Table 23 show a summary of the statistics. These tables show differences in statistics between gabbro and ferralite-saprolite in nickel, cobalt, aluminum, silicon, manganese, chromium and titanium. Table 22: Summary statistics of sample data used for resource estimation in the Ambatovy Deposit
Variable Ni Co Fe Al Mg Si Mn Cr Ti Ferralite-Saprolite Number of samples 42,891 41,321 43,036 43,036 42,494 26,525 26,713 26,590 26,553 Minimum value (%) 0.006 0.001 2.2 0.3 0.0018 0.0412 0.0097 0.0015 0.08 Maximum value (%) 5 2 67.4 23.9 19.8271 36.2786 10 10 10 Mean value (%) 1.03 0.09 45.59 4.27 0.33 2.55 0.72 1.51 0.94 Gabbro Number of samples 24,196 22,898 24,467 24,472 24,042 18,146 18,214 17,854 18,164 Minimum value (%) 0.001 0.001 0.1036 0.2534 0.0012 0.0051 0.0054 0.001 0.0404 Maximum value (%) 3.2683 2 54.8049 35 1 40.8239 10 9.62 7.77 Mean value (%) 0.37 0.05 28.07 11.54 0.18 8.07 0.45 0.57 1.98 Top cut value (%) 5 2 78 35 - - 10 10 10 Source: AMSA, 2018 Table 23: Summary statistics of sample data used for resource estimation in the Anamalay Deposit
Variable Ni Co Fe Al Mg Si Mn Cr Ti Ferralite-Saprolite Number of samples 12,821 11,978 12,858 12,858 12,816 9,822 10,027 10,068 9,604 Minimum value (%) 0.004 0.001 0 0 0.001 0.0379 0.01 0.01 0.0247 Maximum value (%) 2.6864 2 78 17.5 13.3 32.1 10 10 7.9 Mean value (%) 0.86 0.09 48.79 3.77 0.30 1.87 0.72 1.60 0.69 Gabbro Number of samples 5,466 4,927 5,527 5,527 5,482 4,691 4,723 4,722 4,668 Minimum value (%) 0.003 0.001 0 0 0.001 0.005 0.001 0.0015 0.0173 Maximum value (%) 1.4799 2 77.5 32.2 1 34 9.09 10 9.34 Mean value (%) 0.29 0.06 30.45 11.79 0.16 6.10 0.50 0.62 1.84 Top cut value (%) 5 2 78 35 - - 10 10 10 Source: AMSA, 2018
14.5 Variograms and Spatial Variability Experimental variograms were calculated and fitted to a variogram models for each one of the elements and lithology proportions. Examples ofvariograms are provided in Figure 36. The nuggets of the variogram models were deduced from experimental vertical variograms. The variogram models used for grade interpolation are shown in Appendix 1, and the variogram models used to interpolate lithology proportions are shown in Table 24. All models were fitted with spherical and one nugget structures. All the variables have greater continuity in the horizontal directions than in the vertical, except for calcium in the Analamay Deposit.
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Figure 36: Ni variogram inside ferralite wireframe for ferralite (above) and gabbro (below) Source: AMS
14.6 Estimation of Lithology Proportions Each block may be composed of a variable proportion of ferralite and ferricrete, saprolite, bedrock, and gabbro (or laterites produced by weathering of gabbroic rocks). Gabbro has high content of deleterious elements alumina and silica, and sub economic nickel grades (<0.6%). Laterites produced by alteration of gabbroic rocks can be visually identified in field by its pink and white colour, that contrast with the red, yellow, and brown-greenish colours of the limonites and saprolites and are selectively mined as waste when its volume is large enough. The bedrock is usually left in place and bedrock boulders are separated in the slurry preparation plant (OPP) before entering the metallurgical process. Limonites and saprolites are sent to the plant but have different metallurgical and physical properties. To capture these assumptions in the model the portion of gabbro, limonite (ferricrete and ferralite), saprolite and fresh rock were interpolated within the to geological domains of limonites and saprolites.
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Lithology proportions were interpolated using the same estimation parameters used to interpolate grade and the variogram models shown in Table 24. All proportions were adjusted afterk riging to ensure they added to 100% for each block. Low percentages of gabbro and limonites + saprolites were also adjusted to capture the fact that small proportions are unlikely to be mined selectively. If the estimated proportion of gabbro is less than 10% this material is reassigned to limonite or saprolites. If a block has less than 10% of limonite plus saprolite this material is reassigned to gabbro. Table 24: Variogram models used to interpolate proportions of ferralite and ferricrete, saprolite, bedrock and gabbro
Structure 1 Structure 2 Ratio Geological Ellipsoid Area Nugget Major/ Major/ domain bearing* Sill Range Sill Range Semi-major Minor Ambatovy Centre Limonites 120 0.04 0.12 16 0.08 417 1 1.7 and South East Saprolites 40 0.007 0.01 112 0.01 369 1 5.6 Limonites 160 0.06 0.06 31 0.08 561 1.2 4.6 Ambatovy West Saprolites 20 0.02 0.02 72 0.02 335 1.1 6 Limonites 120 0.08 0.05 57 0.09 335 1.8 4 Analamay Saprolites 70 0.09 0.04 105 0.04 277 1 8.8 * Variograms were defined along horizontal plane, without dipping angle
14.7 Grade Estimation The grades of 15 metals (nickel, cobalt, aluminum, calcium, chromium, copper, iron, potassium, magnesium, manganese, sodium, sulfur, silicon, titanium and zinc) were all estimated using ordinary kriging. The elements nickel, cobalt, aluminum, iron, magnesium, manganese, silicon and titanium were interpolated by geological domain (limonites and saprolites) and by lithology (limonite, saprolite, and gabbro). The grades of calcium, chromium, copper, potassium sodium, sulphur and zinc were estimated by deposit (Ambatovy and Analamay), combining limonites and saprolites into a single domain. Metal grades in the bedrock were not interpolated. Most metals follow a well-defined trend in the vertical direction. This trend is controlled by changes in chemical and physical properties of the different horizons of the weathering profile. Figure 37 shows the trend on nickel grades, with the nickel grades peaking near the base of the ferralite. Changes in grades within each geological domain is gradual. Changes in grade around the limit between ferralite and saprolite contact are usually sharp, especially for major elements, cobalt and nickel, and for that reason it was treated as a hard boundary for interpolation purpose; it was also considered as a hard boundary for the interpolation of proportions. The influence of the gradual trend was minimised by using flat search ellipses in the vertical direction and a limit in the number of samples per drillhole (Table 25). In addition to vertical trends there are also horizontal changes on grade distributions, mainly driven by the composition of the rock that originated the lateritic profile. Nickel grades tend to be higher in laterites originated from weathering of peridotites and decreases close to the border of the Deposits, where gabbro is more abundant. The influence of lateral trends was controlled by restricting the search distance horizontally (Table 25).
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Figure 37: Average nickel grade, calculated over a 10 m moving window, vs height above the base of the ferralite Source: Srivastava, 2010 Table 25 shows the search parameters used for interpolation. Three search passes were used for interpolation, the first pass includes the drilling pattern at 50 m, the second pass includes 150 m, and the last one includes the drilling pattern above 150 m. The second and third search passes were only used to estimate blocks not estimated in earlier passes, due to the lack of data. Table 25: Search parameters
Horizontal Vertical search Maximum sample Minimum Maximum Kriging Search type search ellipse ellipse by hole sample sample First pass Ellipsoid 75 10 4 16 50 Second pas Ellipsoid 200 10 4 16 50 Third pass Ellipsoid 1000 10 4 1 50
Blocks with magnesium grade over 1% were reclassified as saprolite, and blocks with magnesium grade under 1% as ferralite.
14.8 Density Estimation Density was assigned as a constant value per geological domain, as shown in Table 26. These density values were obtained as the averages of densities determined using a geochemical formula developed by Srivastava and described in the Technical Report Ni 43-101 of the Ambatovy Project, with effective dateof 1 September 2014 (Daigle et al., 2014). The formula is as follows:
Density = 0.6285-0.2160*log10(Ni)-0.0745*log10(Co)+0.2214*log10(Al2O3+TiO2)+0.2582*MgO/SiO2
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Table 26: Average density values assigned to geological domains and its corresponding lithologies
Geological domains (defined by wireframe surfaces) Lithology domain Ferricrete Ferralite Saprolite Ferricrete/ferralite 1.1 1.06 1.06 Saprolite 1.03 1.03 1.03 Gabbro 1.22 1.22 1.22 Bedrock 2.2 2.2 2.2
The average and standard deviations of densities derived from chemical composition were compared with those estimated using in situ physical bulk density determinations, as shown in Table 27 and Table 28. On average the geochemical densities are similar for Ambatovy and Analamay. Physical bulk density measurements show higher values and more variability than geochemical-derived density, and measurement errors are suspected. Table 27: Average densities by geological domain deduced from geochemical formula
Deposit Ferricrete Ferralite Saprolite Ambatovy 1.10 1.06 1.03 Average density Analamay 1.10 1.05 1.02 Ambatovy 0.09 0.10 0.09 Standard deviation Analamay 0.08 0.08 0.06 Ambatovy 1.48 1.50 1.47 Maximum Analamay 1.43 1.76 1.36 Ambatovy 0.82 0.74 0.75 Minimum Analamay 0.86 0.78 0.80
Table 28: Average densities by geological domain deduced from density samples collected in Ambatovy
Ferricrete Ferralite Saprolite In situ density 1.26 1.08 1.08 Standard deviation 0.28 0.2 0.24 Maximum 1.63 1.83 1.77 Minimum 0.63 0.34 0.51
Physical bulk density measured on gabbro samples also showed higher average and variance values than density obtained with geochemical formulas and the geochemical density was retained for resource estimation. Table 29 shows a reconciliation of tonnage calculated using the averaged densities shown in Table 26, using density interpolated in the resource model reported in 2014, and production tonnage until end of 2017. Tonnages calculated with averaged densities reconciles better than density interpolated. Table 29: Reconciliation of tonnage using density assigned as average vs density interpolated in the block
Tonnes (Mt) Using average density* 35.20 Using density interpolated* 32.60 OPP feed + stockpile 34.40 * Density deduced with geochemistry
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14.9 Block Model Validation Block model validation was completed per estimation domain and consisted of comparison of the average values of variables in drillhole and block model, visual validations in sectional view, and swath plots. In addition, global change of support was completed for nickel grades, using declustering of drilhole data with 200 m x 200 m x 4 m de-clustering cells. The block model of the Ambatovy deposit was validated with reconciliatory with production data from the regions that had already been mined since 2016. Figure 38 and Figure 39 show examples of sectional view validations completed using the 3D visualizer of Dassault Systèmes’® Surpac software. The validation consisted of confirming that grade estimates in block model are similar to the grades in nearby drillhole data.
Figure 38: Drillhole and block model section (east-west) with estimated nickel in Ambatovy
Figure 39: Drillhole and block model section (east-west) with estimated nickel in Analamay
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Swath plots were constructed to confirm that the fluctuations observed in the columns, rows and levels of the Block Model were consistent with the fluctuations observed in the original drillhole data (Figure 40 and Figure 41). Global change of support confirmed that no over-smoothing is affecting the estimate of nickel grades despite the large number of samples used for interpolation (Figure 40 and Figure 41, lower right side).
Figure 40: Swath plots (above and lower left) and global change of support (lower right) of kriged nickel grades in the geological domain of limonites+ferralites (and lithology domain/proportion of limonites+ferralites) of the Amabatovy Deposit
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Figure 41: Swath plots (above and lower left) and global change of support (lower right) of kriged nickel grades in the geological domain of limonites+ferralites (and lithology domain/proportion of limonites+ferralites) of the Analamay Deposit
14.10 Resource Classification Classification, or assigning a level of confidence to Mineral Resources, is undertaken in strict adherence to the “Definition Standards for Mineral Resources and Mineral Reserves” adopted by the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Council on 10 May 2014 (CIM Council, 2014). The classification of Mineral Resources into Measured, Indicated and Inferred categories was based on the confidence, quality and quantity of the informing data, the confidence in the geological interpretation of the deposit and the “reasonable prospects for economic extraction” of these resources. The search pass used to estimate the block was also used as reference to classify resources. Blocks estimated in the first search pass were assigned as candidates for Measured resources. These blocks were estimated with at least four drillholes located in the 50 m drillhole exploration grid and can be classified as Measured since the confidence on the resources in these blocks is high and sufficient for possible production planning purposes, after applying modifying factors. Blocks estimated in the second pass also make use of data from at least four drillholes, but blocks located in areas over 50 m and up to 200 m drillhole spacing were informed in this pass. Blocks estimated with the second search pass were assigned as candidates of Indicated Resources, since there is sufficient confidence in the grade and tonnage estimates to do mine planning. The slope of regression (SOR) was used as reference to classify saprolitic resources, the block was retained as candidate for Indicated saprolites if SOR ≥0.85. Grades on saprolites
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are more scattered, due to the low proportion of this rock, especially within the limonitic geological domain. Inferred category was assigned to blocks estimated in the third pass, that could be estimated with only one drillhole, and to blocks labeled as saprolites and with SOR <0.8. The final classification into Measured, Indicated, and Inferred resources was completed using a manual contouring approach to ensure a coherent Mineral Resource classification, without small patches of contrasting categories non-suitable for mine planning. Figure 42 and Figure 43 show in plan view examples of resource classification for Ambatovy and Analamay.
Figure 42: Left – three passes of kriging; Right – passes manually smoothed to use for resource classification
Figure 43: Left – three passes of kriging; Right – passes manually smoothed to use for resource classification
14.11 Mineral Resource Estimates
14.11.1 Reasonable Prospects of Economic Extraction The cut-off used for reporting is 0.45% Ni, which corresponds to a break-even economic cut-off calculated at a nickel price of US$14,000/t, 90% recovery, processing cost of US$47/t mined, and selling cost of US$2,185/t. This cut-off does not include revenue from cobalt content nor any penalties, such as extra acid consumption due to high concentration of Mg associated with saprolites. However, it allows the selection of material with reasonable prospects of economic extraction, as shown in Section 15 of this report and in Figure 52. The flat-lying nature of the Analamay and Ambatovy deposits, with mineral resources located right below the topographic surface, makes easy its extraction with open pit mining. The physical limits of these two deposits, defined at depth by the bedrock and laterally by gabbro formation is sufficient to constraint resources.
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14.11.2 Mineral Resource Reporting The Ambatovy and Analamay deposits Mineral Resources inclusive of Mineral Reserves, reported per ore types are shown in Table 30. Combined resources are presented in Section 1 Summary (Table 1). The Mineral Resources do not include any material that has been mined up to 30 June 2018 with the exception of material in stockpiles that have been set aside for future treatment. A fixed cut-off grade of 0.45% Ni has been applied to both the Ambatovy and Analamay Deposits in situ material. The stockpile inventory as at 30 June 2018 of 10.7 Mt at 0.81% Ni and 0.06% Co shown in Table 30. The tonnage and grade of stockpiles were estimated by tracking material moved from the pit to low-grade and high-grade stockpiles using truck counts and grade control data; a fleet management system (Wenco) has now been implemented to track of stockpile material. Stockpiled inventory is further discussed in Section15.5 and shown in Table 39. Table 30: Mineral Resource estimate1 for Ambatovy Project (inclusive of Mineral Reserves) above a reporting cut-off2 grade of 0.45 %Ni (with an effective date of 30 June 2018)
Material Classification4 Tonnage (Mt) Ni (%) 3 Co (%) 3 Ambatovy Measured 40.5 0.99 0.08 Ferralite Indicated 52.0 0.83 0.08 Inferred 20.4 0.74 0.08 Measured 2.6 1.37 0.05 Saprolite Indicated 14.3 1.18 0.05 Inferred 7.4 0.95 0.06 Analamay Measured 9.5 0.81 0.08 Ferralite Indicated 55.7 0.89 0.09 Inferred 33.7 0.84 0.09 Measured 0.0 0.78 0.05 Saprolite Indicated 7.7 1.20 0.07 Inferred 7.6 1.05 0.07 Stockpiles Total Measured 10.7 0.81 0.06 1. Numbers have been rounded to reflect the precision of a Mineral Resource estimate. 2. The reporting cut-off is calculated for an open pit with processing cost of US$47/t mined, selling cost of US$2,185/t, metallurgical recovery of 90% and a nickel price of US$14,000/t. Cobalt contribution and penalty elements are not considered in this calculation. These are Mineral Resources and not Reserves and as such, do not have demonstrated economic viability. 3. The average grade estimates reflect nickel and cobalt resources in situ, and do not include factors such as external dilution, mining losses and process recovery losses. 4. Resource classification as defined by the Canadian Institute of Mining, Metallurgy and Petroleum in their document “CIM Definition Standards for Mineral Resources and Mineral Reserves” of 10 May 2014.
14.11.3 Factors that may Affect the Mineral Resource As of the Effective Date,t he QP responsible for this section, Dr Martínez Vargas, is not aware of any known current environmental, permitting, legal, title, taxation, socio-economic, marketing or political factors that might materially affect these mineral resource estimates.
14.12 Comparison with the previous Mineral Resource Estimates The Mineral Resources presented in the previous Technical Report (Daigle, B. et.al., 2014) have an effective dateof 31 December 2013 and were prepared and supervised by R. Mohan Srivastava, using the same methodology used to generate a previous block model completed in 2009. These previous Mineral
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Resource models were developed in large part prior to the commencement of mining and in the absence of reconciliation data. The earlier Mineral Resource models were focused primarily on the limonite part of the resource and did not separately model the Gabbro dykes and saprolites. The 2018 Mineral Resource model is much better aligned to the mining and PAL constraints than the earlier 2013 model. In addition, more attention has been paid to modelling resources in saprolites. Other differences are: • The 2013 resource model grade variables were estimated with simple kriging with local trend (local average grade). The current model grade variables were estimated with ordinary kriging. • Density values in the 2013 resource model were interpolated using simple kriging and density values in sample intervals calculated from chemical data using formulas similar to the one presented in Section 14.8. In the current resource model densities were assigned to each lithology domain as a constant value. However, the density corresponding to each lithology was obtained as the average of densities calculated on sampling intervals from chemical data, using the formula presented in Section 14.8. • Mineral Resources reported in 2013 were at a 0.6% Ni cut-off. This cut-off was considered appropriate for nickel prices in 2013 but also corresponds to the cut-off historically used to report Mineral Resources for this property. The current Mineral Resources were estimated at a 0.45%Ni cut-off. This marginal cut-off was calculated witha five-year average nickel price and reference cost obtained from production data.The 0.6% Ni cut-off was found problematic to convert Mineral Resources to Mineral Reserves as it excludes a large number of blocks with actual economic extraction value, as shown in Figure 52. • Drilling in new areas allowed extension of the resource model on the Analamay Deposit, and infill drilling allowed increased confidence in the resources estimated in both Analamay and Ambatovy Deposits. The boundaries of the mineralisation differ slightly from one estimate to other in both Deposits. • The block size used in the 2013 model was 10 m x 10 m x 3 m, with only one grade content per block. The current model has 12.5 m x 12.5 m x 4 m minimum cell size and 25 m x 25 m x 4 m mining blocks, with grade variables corresponding to each lithology and lithology proportions, and it is optimised to allow selective separation of waste (gabbro) from mineralised material on a block-by-block basis. Differences in geological modeling and estimation approaches, and on mining assumptions, make the previous and current resource models non-comparable at a similar cut-off grade; for example, the current block model is more selective and produces lower tonnage but higher grade. However, a comparison of these two resource estimates is shown in Table 31 to provide a quantitative reference of the changes in grade and tonnage.
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Table 31 Comparison of 31 December 2013 and 30 June 2018 Mineral Resources
Deposit Year Category Tonnage (Mt) Ni (%) Co (%) Measured 68.3 0.9 0.08 2013 Indicated 62.1 0.8 0.07 Inferred 53.8 0.7 0.06 Ambatovy Measured 43.1 1.0 0.08 2018 Indicated 66.3 0.9 0.07 Inferred 27.8 0.8 0.07 Measured - - - 2013 Indicated 109.0 0.8 0.09 Inferred 21.3 0.8 0.08 Analamay Measured 9.5 0.8 0.08 2018 Indicated 63.4 0.9 0.09 Inferred 41.2 0.9 0.09 Notes: • Figures have been rounded and hence may not add up exactly to the given totals. • The Qualified Persons responsible for this section of the Technical Report have not done sufficient work to classify the 2013 historical Mineral Resource estimates as a current Mineral Resource estimates. Sherritt does not treat these historical Mineral Resource estimates as current Mineral Resource estimates. • 2013 resources are estimated at a 0.6% Ni cut-off. • 2018 resources are estimated at a 0.45% Ni cut-off. • Both 2013 and 2018 resources do not include any of the material that has been mined up to 30 June 2018. • Cobalt grade does not enter into the definition of the reporting cut-off grade.
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15 Mineral Reserve Estimate
15.1 Introduction In this Technical Report, the Mineral Reserve estimates for the Ambatovy Project contains forward- looking information. Material factors that could cause actual results to differ materially from the conclusions and estimates set out in this report include: (1) naturally occurring geological variability; (2) geological interpretations; (3) changes in the modifying adjustment factors as a result of geological variability (4) changes in the processing costs; (5) changes to the nickel and cobalt prices; and (6) changes to the throughput capacity/bottlenecks. The material factors, or assumptions, that were applied in drawing the conclusions, forecasts, and projections set forth in this Item are summarised in this, and other Items of this Technical Report. For this reason, readers should read this Item solely in the context of the full report, and after reading all other Items of this report. The Mineral Reserve estimate is based on the life of mine scheduled material quantities. The life of mine schedules that underpin the Mineral Reserve have been based on targeting Measured and Indicated Resources with appropriate modifying factors applied.
15.2 Modifying Factors The Mineral Resource presented in this Technical Report is underpinned by the updated 2018 Resource model – as this model is materially different than the previous 2013 resource model, new resource to PAL modifying factors had to be generated. The assumptions made in the conversion of the resource model data to material that can be considered for conversion to Mineral Reserves is outlined in Figure 44, Figure 45 and Figure 46. The key criteria outlined in these figures are: • If gabbro comprises more than 10% of a block, it is assumed it can be separately mined as waste (Figure 44) • If gabbro is 10% or less of a block, it is assumed it cannot be mined separately and is mined with the ore as dilution (Figure 45) • If the combined block proportion of ferralite and saprolite is 10% or less, it is assumed it cannot be mined as ore and is considered waste (Figure 46).
Figure 44: Resource block to mining block (gabbro >10%)
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Figure 45: Resource block to mining block (gabbro <= 10%)
Figure 46: Ferralite + Saprolite <= 10% The modifying factors are the adjustments to the material that has been defined as potential ore in the insitu resource required to estimate the resultant PAL feed (i.e. ROP) tonnes and quality.
15.2.1 Run of Mine (ROM) – Run of Preparation (ROP) Modifying Factors The preferred approach to defining the modifying factors is to reconcile each stage of the mining process and thereby estimate modifying factors for each stage of mining. The mine to mill stages that provide potential sampling points for input for Ambatovy can be seen in Figure 47.
Figure 47: Ambatovy process flow and sampling points
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IMC conducted a reconciliation of the recorded mining and milling data for 2016 and 2017 and reconciled this against both the 2018 resource model and the 2013 resource model. IMC found large variations year to year. In order to smooth out these variations, combined reconciliation data from 2016 and 2017 was used. In 2016/2017, IMC conducted a reconciliation between the 2013 resource model and production data. The earlier reconciliation work (which was based on a different and less robust mined dataset) found the historical production data to be inconsistent. It was concluded at that time that the most reliable sampling point for ROP data was the PAL sampling point (point 6 in Figure 47). IMC has used a modifying factor approach such that recovery is a % of total tonnes (e.g. 86% of tonnes recovered into ROP) whilst grade adjustments are fixed addition or subtraction (e.g. Ni resource grade - 0.04%). The flow of material against which reconciliation has been carried out is summarised in Figure 48. These same flow points are reflected in the column headings in Table 32. The starting point is the resource mined in the pit (A) which is simply the difference between the starting triangulated surface and the finish triangulated surface for the reconciliation period flagged against the resource model. The next measured point is (B) which is the stockpile delta – this simply the volume difference from survey of all ore grade stockpiles over the reconciliation period. This is a measure that does require an assumption as the volume of (B) is known from survey, however there is no measurement of density. The resource density has been assumed. The next measured point is (C) which is the OPP feed tonnes. This is measured on the belt that feeds the scrubber. The total tonnes to the OPP is known; however, it is not known what the split of tonnes is that comes from (A) directly or from (B). It is also unknown what proportion of the OPP feed is made up of sheeting material. The sheeting material is assumed to be all rejected at the OPP so does not impact the PAL feed; however, the estimation of quantity of sheeting material in the feed is important as: 1. Sheeting is not captured in the triangulation delta for (A); 2. Sheeting is captured in the triangulation delta for (B); and 3. The Resource recovery must factor in the various tonnage adjustments for sheeting to get the correct adjustments. The final measured point is (D) which is the material flowing out of the OPP (C) to the PAL. This is measured at a number of points; IMC has based the reconciliation on the data from the PAL feed after the bottom of the pipe after the PALtanks. This does not include any sheeting material as it has been rejected. The relevant mass balances for each step are summarised in Figure 48.
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B C D Stockpile Delta OPP Feed PAL Feed * Sourced from Stockpile Surveys *Sourced OPP Feed Belt Weight * Sourced from Site EOM Reports
OPP PAL Stockpile
Sheeting Mined Waste with Ore W Resource to Waste Sheeting Mined with Ore (DR)= A x 5% Pit Resource ROM Recovery = ((B + C) - DR) / A A OPP Rock Feed (SO) = (DR / (B + C)) x C Resource Removed Resource OPP Feed Recovery = D / (C - SO) * Sourced from Survey Resource to PAL Feed = ((B + C) - DR) / (D / (C - SO)) W = A x (1 - ROM Recovery )
Figure 48: Material flow for tonnage econciliationr Each of these equations can be solved linearly with the goal of calculating the ROM recovery (i.e. the unknown variable). One assumption that is required to enable this calculation is the amount of sheeting in the OPP feed. Historical measurements support an assumption that 15% of the Resource tonnes is added as sheeting – 10% of this sheeting is removed by selective mining, leaving 5% of sheeting fed to the OPP or stockpile. This is 5% of the insitu resource not 5% of the recovered resource. An example of the material flow is provided in Figure 49. IMC used the Gurobi Linear Programming (LP) engine to solve the equation; however, any LP solution could solve the equation, including goal seek and solver in Microsoft xcel E – the formula could also probably be solved against the one variable but that was more complicated than the simple Gurobi approach used by IMC. The reconciliation data and derivation of recovery is summarised in Table 32. The modifying factors for the grade elements was somewhat simpler than for the tonnage – the insitu grade of each element (with the exception of carbon) is known as is the grade in the PAL feed– accordingly the average delta is known. The modifying factors for the calculated tonnage adjustment and the grade delta applied to the 2018 Resource model are summarised in Table 33. The saprolite recovery has been estimated to be 50%. Very little data is available on this as the targeted material so far has been principally ferralite with saprolite generally entering the circuit as dilution. The 50% is considered conservative. Note that for the five-year plan, the mining recovery has been assumed to remain the same as the reconciliation data (86%) with OPP recovery of 91% of the resource material.
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From Year 6 onwards, it has been assumed that mining recovery will improve (through improved grade control and mine selectivity) to 95%, with OPP recovery remaining the same. The above gives an overall recovery of 79% over the 5-year period and 86% for the Life of Mine Plan (LOMP). The derivation of modifying factors has assumed that sheeting material is added to the insitu resource as 15% of the insitu resource and that 5% of this sheeting material is mined as ore to the OPP or stockpile, where it has then been assumed that 100% of that sheeting material is rejected at the OPP. The flow of the sheeting material is summarised in Figure 49.
Figure 49: Sheetingmaterial flowchart Table 32: Tonnage Resource – ROP modifying factors (data shown represents million tonnes in reconciliation period)
Resource Sheeting B C Sheeting Resource Resource A Resource D OPP mined Stockpile OPP feed OPP feed W ROM to PAL Resource OPP feed PAL recovery with ore Delta* total SO (Mt) recovery feed (Mt) RO feed D/RO (Mt) (Mt) (Mt) (%) (%) (Mt) (%) 2018 Model 12.53 0.63 1.52 9.92 0.54 9.38 8.56 1.71 86% 91% 79% 2018 Resource Recommended 12.53 0.63 1.52 9.92 0.54 9.38 8.56 1.71 86% 91% 79% 5-Year Plan 2018 Resource Recommended 12.53 0.63 1.52 10.95 0.55 10.40 9.46 0.69 95% 91% 86% LOM Plan * Stockpile Delta is the surveyed change in stockpile volume multiplied by assumed density.
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Table 33: 2018 Resource model to ROP (PAL feed) modifying factors
Recovery Recovery Material Ni Co Fe Al Mg Si Mn C (Year 1-5) (Year >5) Ferralite 79% 86% -0.04 0.00 -1.00 0.00 +0.70 +0.30 0.00 0.00 Saprolite 50% 50% -0.04 0.00 -1.00 0.00 +0.70 +0.30 0.00 0.00
The grade control model has been used where available to predict carbon in the ROM model; however, when not available a normative relationship between carbon and iron has been used to predict carbon. The reconciliation data above did not include carbon as no recorded data was available for carbon at the time of carrying out this work. In summary, the recommendations vis-à-vis the modifying factors (insitu – ROP conversion) are: • Mining recovery of 86% in the next five years increasing to 95% in Year 6 (through improved grade control and modelling systems). • Constant ore recovery through the OPP of 91%. • Ferralite insitu resource to ROP (PAL feed) tonnage recovery of 79% (86% mine recovery x 91% OPP recovery) compared to 89% in the 2017 work, increasing to 86% in Year 6 (95% mine recovery x 91% OPP recovery). • Saprolite insitu resource to ROP tonnage recovery of 50% the same as the 2017 work program. • Ni adjustment -0.04% compared to +.01% in 2017, based on reconciliation. • Mg adjustment is 0.7% compared to 0.41% in 2017 – this is consistent with the suggested impact of sheeting and saprolite transition zone. The sheeting material has an impact on the PAL feed as some sheeting is abraded in the OPP and reports to the PAL – this has an impact most notably on the Mg grade as sheeting is generally high in Mg.
• The Al2O3 adjustment was in large part ignored in the earlier work as it was known that the 2017 model incorrectly estimated the Al203 grades. In this work no adjustment has been applied to Al2O3 as it is
reasonable to assume that the insitu Al2O3 estimate is going to be more representative of the mined Al2O3 in moving forward with the modelling of the gabbro; however, it should be noted that the Al2O3 grade is lower in the updated Resource model.
15.2.2 Sheeting The impact of road sheeting on the OPP has been outlined in Item 0. Due to the high moisture content of the ore, any equipment with tyres on roadways will rapidly degrade the road to the extent it is unusable. In order to maintain trafficability of both primary roads and roads into the mining face, approximately half a metre of competent waste material is placed on top of geofabric. Various estimates of this road sheeting quantity have been made – both empirically and also through observation. The estimate is that the amount of road sheeting required is around 15% of the material excavated, irrespective if it is waste or ore. When the ore is mined, some effort is made by the excavator operator to remove the sheeting material and send it to the waste dump or to a sheeting stockpile for recycling for more sheeting. Estimates have been made through observation and measurement that 10% of the placed sheeting is selectively removed from the ore and sent to waste. The remaining 5% is sent to ore (note these ratios are of the proportion ofs heeting in total, meaning that in a 1,000 t block of insitu ore, 150 t of sheeting is
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added, 100 t of that is removed to waste which is 10% of the 1,000 t but is two-thirds or 66% of the added sheeting, one-third of added sheeting taken to OPP). As the sheeting is competent large waste material, it has been assumed that all sheeting that goes to the OPP is rejected at the OPP. It is not possible to know the grade of the sheeting material – accordingly, the reported ROM data does not include the tonnes or the grade of the sheeting. The tonnes are however taken into account when working out the mining cost.
15.2.3 Selection Based on Resource Classification and Domains The economic assessment of Ambatovy and the definition of Mineral Reserves has been based on only assigning value to Measured and Indicated Mineral Resources. The ultimate pit limit is defined laterally as either the lateral economic limit based on pseudo flow optimisation outputs, or by environmental/lease boundaries. In most instances the vertical extent of the pit is defined as the base of the saprolite/top of bedrock. This is irrespective of the economic viability of mining to the base of saprolite – for safety, it is not possible to leave small pillars of saprolite above the bedrock as they could potentially be geotechnically unstable. The overall impact of the additional mined material below the economic floor is not material to the estimate of the Mineral Reserve. The other consideration has been what should be done with mineralised material that is contained within the final pit limits but is classed as Inferred Mineral Resource. Whilst these zones of increased geological uncertainty are in all likelihood going to contain ore, which will most likely be Measured or Indicated once more drilling is complete, the grade and tonnes of that ore at this point in time cannot be stated with sufficient confidence to class the material as Measured or Indicated and therefore it cannot be converted under the rules and guidelines of CIM “Definition Standards for Mineral Resources and Mineral Reserves” (CIM Council, 2014) and NI 43-101 to Mineral Reserves. In the LOM plan which underpins this Technical Report, the Inferred Resources that are within the economic pit have been included in the scheduled quantities. The rationale for this is that it is more important in the LOM plan to understand the material movement requirements each year and the related equipment purchase requirements and infrastructure requirements. As there is a high level of confidence of converting the Inferred Resource to Measured or Indicated Resources well in advance of mining, the sensible course of action is to include this material in the planning process. Unfortunately, for reporting within the rules and guidelines of CIM Definition Standards andNI 43-101, no value can be assigned to these Inferred Mineral Resources in the reporting of Mineral Reserves. In order to meet the rules and guidelines of CIM Definition Standards and NI 43-101, a LOM schedule has been completed that is based on the detailed five-year tactical plan for the next five years of operation and then from Year 6 onward is based on a schedule and financial analysis that includes only Proven and Probable Mineral Reserve material derived from Measured and Indicated Mineral Resources. For transparency and completeness, Item 24 of this report discusses the Inferred Resources that are included in the LOM schedule that is the basis of the current site based mine planning.
15.2.4 Stockpile, Block Binning and Cut-Off Grade The stockpiling strategy is a key economic driver for the Ambatovy operations – if all ore above the economic cut-off grade were fed directly to the OPP, the value of the project would be reduced significantly and the production curve for nickel metal would be reduced in the near term.
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A minimum ore cut-off grade and a variable cut-off grade (i.e. between stockpiled and direct feed ore) is an essential part of the mining strategy at Ambatovy. The stockpiling strategy must be structured in such a way that stockpiling can control total revenue received and PAL throughput. The use of a block net revenue equivalence (which IMC has related back to what we call NiEq for each block) that takes into account the Ni and Co revenue along with the variable PAL costs for that block is the cornerstone of the cut-off grade/stockpiling strategy. The other less obvious aspect of the stockpiling strategy is the throughput constraint impact. This is primarily determined by the acid consumption of that block of ore through the process plant. Acid supply is one of the key constraints to throughput capacity. On the basis of the above, the stockpiling strategy enables selective stockpiling and recovery on the basis of both NiEq grade and acid consumption.
15.2.5 Block Binning and Block Allocation Introduction The allocation of blocks in the model as OPP direct feed, waste or stockpile is the mechanism by which the optimum variable cut-off grade strategy is implemented. This obviously also impacts the Mineral Reserve estimate. The block allocation (as OPP direct feed, waste or chosen stockpile) is an output from the LP scheduling process. The differentiation of blocks of material that can be potentially allocated as each material type is the driver behind the allocation. This process is called block binning. The block binning and subsequent block allocation approach presented in this report is a methodology that has evolved through many iterations of schedules and reviews of the operations, process drivers and most importantly the dynamic allocation of blocks to stockpile, waste and OPP direct feed in the LP scheduler. The following provides a brief overview of the approach. The essential elements in the block allocation have been to: • Consider all revenue streams (the nickel and cobalt revenue) in block allocation • Consider the net revenue contribution of each block (i.e. the revenue generated from the block less the cost of processing that block of ore) • Consider the consequential opportunity cost impacts of mining that block of ore, i.e. the impact the block has on the ability to produce nickel and cobalt metal (one key limiting factor in the process is acid generating capacity of the plant) • Consider the possibility to take into account other limiting factors (such as Mn/C, Al) • Align the block allocation with the optimised strategic dynamic block allocation • Ensure a simplified block allocation can be sensibly achieved on a day-to-day operating basis (i.e. the block allocation used in the strategic schedule can be implemented at the mine face) • Ensure a methodical approach to block allocation is implemented that matches the outcomes of the above. The outcome of numerous iterations to develop the best-balanced strategy is illustrated in Figure 50. This figure shows different colours for different bins of material based on NiEq grade and the acid consumption in the block (with acid consumption being the % of the ROP dry tonnes that is acid consumed – e.g. a block with 30% acid consumption means 300 kg of acid is consumed for each tonne of ROP).
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Figure 50: Block bins NiEq based In allocating blocks, the first step was to calculate the net revenue generated by each block. This was relatively straightforward as all of the revenue and process parameters are stored in the model through the ROM–ROP process. This net revenue per block could have simply been represented as $Rev/ROP tonne; however, it is best practice in nickel projects to convert the revenue to a meaningful baseline – thus the $Rev/ROP tonne calculated in the ROM–ROP algorithm was converted to an equivalent nickel grade that would generate the same revenue. This is called the NiEq grade. This is the Ni grade that would have resulted in that same total revenue when factored against the ROP tonnes in the block. As outlined above, there is also a significant impact on the block allocation from acid consumption. Various data distribution histograms were developed to look at the distribution of material in different acid zones and then comparing these to how the dynamic scheduler allocated the blocks – this culminated in the large-scale allocation as illustrated in Figure 50 and Table 34. This allocation was further reviewed by looking at bench plans for groupings of block allocation and looking at how the LP scheduler allocated blocks over incremental periods of the mine life. Based on this, it was found that larger super groups of block allocation can be defined. This culminated in the super groups used as the basis for stockpiling in the LOMP as illustrated in Figure 51 and Table 34.
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Figure 51: Consolidated bin allocation based on revenue/cost ratio vs ROP NiEq revenue Table 34: NiEq and acid consolidated bins
PAL NiEq % PAL NiEq % Acid minimum Acid maximum Consolidated bin minimum maximum (kg/t) (kg/t) - 0.40 - 250 - 0.40 250 350 Ultra Low Grade Ni Low Acid 1 - 0.40 350 450 - 0.40 450 600 0.40 0.70 - 250 0.40 0.70 250 350 Low Grade Ni Low Acid 2–4 0.40 0.70 350 450 0.40 0.70 450 600 0.70 0.90 - 250 0.70 0.90 250 350 Med Grade Ni Low Acid 5–6 0.70 0.90 350 450 0.70 0.90 450 600 0.90 100 - 250 0.90 100 250 350 High Grade Ni Low Acid 7–8 0.90 100 350 450 0.90 100 450 600 Ultra Low Grade Ni High Acid - 0.40 600 1,000 9 Low Grade Ni High Acid 0.40 0.70 600 1,000 10–12 Med Grade Ni High Acid 0.70 0.90 600 1,000 13–14 High Grade Ni High Acid 0.90 100 600 1,000 14–15
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15.2.6 Cut-Off Grade The Mineral Reserves have been based on the economic cut-off grade. The alternative to using the economic cut-off grade would be to apply a universal cut-off grade at an elevated grade. This is what has been done to date at AMSA in the grade control modelling – a minimum nickel grade of 0.6% has been defined as ore during the mining process. In addition to this grade control cut-off grade of 0.6% Ni, a number of previous Mineral Resource and Mineral Reserve estimates have been based on a 0.45% Ni cut-off grade. For laterite nickel projects, it is IMC’s experience that it is common for the cut-off grade used in selecting the plant feed to be materially higher than the marginal cut-off grade, which supports the 0.6% Ni cut-off grade. In order to provide a framework for the approach taken, particularly in the context of the transparency requirements of NI 43-101 technical reports, IMC undertook a review of the alternative cut-off grade options. The economic cut-off grade for each block of ore has been estimated based on the revenue/tonne from the nickel and cobalt in the block required to pay for the creation of saleablenickel and cobalt metal from that block of ore. In order to present the results, for each block in the resource the revenue generated per tonne (from nickel and cobalt) is divided by the cost to create the metal. Blocks with a ratio of >1 are economic. As the historical mining grade control model cut-off grade has been set as 0.6% Ni, the results of the economic analysis are presented in the context of the differences between the economically-based cut- off grade used in this Mineral Reserve estimate and the material that has historically been mined as ore. Figure 52 shows on one axis the nickel grade of blocks and on the other the ratio of revenue/cost: • The material that comprises the Mineral Reserve as stated in the Technical Report is the red and pink material to the right of the 1.0 x-axis revenue/cost ratio. • The material that is in the Mineral Reserve for both the economic cut-offgrade and the 0.6% Ni grade control cut-off is shaded as red (>=1 revenue ratio and >= 0.6% Ni). • The material that is included in the 0.6% Ni based grade control model but is not economic is shown as orange (0.15 Mt) (<1 revenue ratio and >= 0.6% Ni). • The material that is not included in the 0.6% Ni based grade control model but is economic is coloured pink (13.6 Mt) (>=1 revenue ratio and <0.6% Ni). • The material that is excluded by both approaches and is always waste is coloured green (<1 revenue ratio and < 0.6% Ni). • The pink material is the 13.6 Mt of material that has been included in the Mineral Reserve but without a change to the current grade control cut-off grade strategy would be considered waste. • The 0.15 Mt of orange material is not economic to mine at today’s prices and has been included in the grade control model as what amounts to internal dilution. The use of the economic cut-off grade approach is the most robust approach to cut-off grade for reporting of Mineral Reserves. In order for the operational performance to reflect the Mineral Reserve assumptions, the grade control strategy will be modified to reflect the economically-based cut-off grade rather than the historical 0.6% Ni cut-off grade. IMC understand this transition is underway and will be implemented with the updated five-year and LOM plan which underpin this Technical Report.
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Figure 52: Revenue/Cost ratio vsnickel grade
15.2.7 Moisture The moisture content has been difficult to incorporate in the resource model at Ambatovy; it is known that it is high and increases through the ferralite profile toward the saprolite roof, and it is also known that it peaks at around 50%. The high moisture is taken into account by the addition of sheeting and also by the division of material into P1 and P2 type material (with P2 material being a combination of the saprolite zone and the lower part of the ferralite zone that has high moisture, P1 being the upper horizon with lower moisture). For the purposes of estimation and planning, IMC has relied on average moisture measurement of 38% moisture in all mined material. IMC’s opinion is that this reflects the average moisture of mineralised material placed on trucks. It should be noted that all grade and resource estimation is based on dry tonnes, moisture is only of relevance for mining cost estimates.
15.3 Optimisation As outlined in Section 15.2.3 of this Technical Report, the use of optimisation tools, such as pseudo flow techniques, is only required, predominantly, to determine the lateral economic extents of the orebody – the vertical extent in almost all instances is defined by the base of saprolite. IMC utilised GEOVIA pit optimisation software to assist with the definition of the staged pit development sequence and also to help define the final pit limits. The inputs to the optimisation are summarised in Table 35.
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Table 35: Optimisation modelling inputs
Input Unit Value US$/t 15,041 Nickel price US$/lb 6.82 US$/t 55,631 Cobalt price US$/lb 25.23 Nickel selling cost US$/t 2,185 Cobalt selling cost US$/t 2,591 Processing cost fixed US$/t feed 42.76 Mining cost fixed US$/t mined 3.99 Processing cost reference US$/t feed 62.77 Cut-off grade Ni % 0.6 Wall slope degrees 27
The staged pit shells output from the optimisation for Ambatovy and Analamay are shown in Figure 53 and Figure 54. The primary purpose of the optimisation has been to define these pit development stages and therefore the optimisation was focused on relative economic benefit not absolute economic benefit. For many laterite projects, IMC do not run an optimisation but use a margin ranking approach similar to coal deposits. The reason this approach was not adopted for Ambatovy is the variability in the surface and floor slopes which results in the need to excavate bedrock to access parts of the orebody – an optimised shell approach takes into account the cost of this additional excavation, whereas an economic ranking would not. It is these parts of the resource (which are few) where the final Mineral Reserve was partly impacted by the outputs from the optimisation.
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Figure 53: Ambatovy optimised staged pit shells – Measured and Indicated only
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Figure 54: Analamay optimised staged pit shells – Measured and Indicated only The pit limits are in almost all instances defined by the mining lease constraints and the saprolite floor. The exceptions to this are areas where the saprolite is >27° or the ferralite is Inferred. These exceptions are illustrated in Figure 55 and Figure 56 for Ambatovy and Analamay respectively.
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Figure 55: Pit floor not saprolite floor – Ambatovy The areas of difference have been designated by letter in the figure as follows: A: Steeper than 27° and mining area less than 40 m. Not economic to push down. B: SAP Pinnacles flattened to allow for ramp access. C: SAP steeper than 27°. Bedrock benched for stability. D: High magnesium zone. Not economic. E: Mined-out area, too narrow to mine remaining ore. F: SAP steeper than 27°. Bedrock benched for stability. G: SAP steeper than 27°. Bedrock benched for stability. H: Area flattened to allow entry to main access ramp. I: Inferred material. J: Access ramp through bedrock to transition to lower South East area. K: SAP steeper than 27°. Bedrock benched for stability. L: SAP steeper than 27°. Bedrock benched for stability. M: SAP floor not used in 1023 design as it is currently near completion. The same data for Analamay is illustrated in Figure 56.
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Figure 56: Pit Floor Not Saprolite Floor - Analamay The areas of difference have been designated by letter in the figure as follows: A: Inferred material. B: Inferred material. C: Inferred material. D: Inferred material. E: Access width less than 40 m. Bedrock flattened. F: Inferred material. G: Inferred material.
15.4 Mineral Reserve Estimate – Insitu Resources The LOM ultimate pit shell that defined the material to be included in the LOM schedule has been based on mining the portion of the Measured and Indicated Mineral Resource that has been shown to be economically mineable. Within the final pit shell, there are Inferred Mineral Resources which have in part been included in the site-based scheduled quantities. These Inferred Resources are not included in the Mineral Reserve.
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Modifying factors have been applied to in-situ resources as outlined in Item 15.2. It is important to note that the ROM Mineral Reserve data has been based on applying some of the modifying factors, but not all, to the Mineral Resource – the sheeting is not included in the ROM data – it is removed through the OPP. Note that the tonnage of the sheeting has been taken into account for mining cost estimation. The economic viability of mining the ore has been based on the ROP adjusted material to define that material that ends up as feed to the PAL plant. The derivation of the ROP has been as defined in Item 0. The OPP (which is the main contributor to the change from ROM to ROP), in simple terms, involves separating out all ROM material that is greater than 0.8 mm in size – with the oversize material being taken to waste dumps or used for road sheeting. Some ore grade material is lost as part of this sizing process. The ROM Mineral Reserve is the portion of the Mineral Resource that has been shown to be the economically mineable part of a Measured or Indicated Mineral Resource as demonstrated by the LOM study that underpins this report; however, the economics are determined on the basis of the ROP characteristics of each block of ore. This life of mine study, on which the Mineral Reserve estimate has been based, has incorporated information on mining, processing, metallurgical, economic and otherrelevant factors to demonstrate, at the time of the work program, that economic extraction can be justified. The difficulty with reporting the Ambatovy Mineral Reserve, which is done in accordance with NI 43-101 rules and guidelines, is that the ROM Mineral Reserve, as outlined above, includes only some diluting material as these diluting materials are captured at the ROP/OPP stage of reporting (i.e. sheeting material). The ROP Mineral Reserve includes the impact of all diluting material and allowances for losses when the material is mined, however it also includes losses and upgrades/downgrades of material through the Ore Preparation Plant (i.e. the ROM–ROP factors). This is common to many laterite nickel projects. In order to ensure transparency, it is IMC’s practice to report Mineral Reserves for laterite nickel projects on the basis of the ROM material and also on the basis of the ROP material – with the focus on the ROP material. Accordingly, the Mineral Reserve (ROM) is summarised in Table 36 and the Mineral Reserve (ROP) is summarised in Table 37. The waste material that is mined in order to extract the Mineral Reserve material is summarised in Table 38. The waste has been reported on the basis of directly mined waste (which is called “waste insitu”) and waste that also includes the rehandle of sheeting material that is a necessary aspect of mining at Ambatovy (the total waste including the rehandled sheeting material is called “Waste (insitu+sheeting)”. The material that is the lost material (the difference between the Mineral Reserve (ROM) and the Mineral Reserve (ROP)) has not been reported in the Mineral Reserve statement. This material is primarily the reject material from the Ore Preparation Plant. Allowances for the cost of hauling and stockpiling the ROP reject material have been allocated in the mine operating cost model.
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Table 36: Mineral Reserve (nominal ROM – excludes sheeting)
Tonnage Ni Co Al Mg Ni metal Co metal Deposit Classification (Mt) (%) (%) (%) (%) (kt) (kt) Proven 35.3 1.01 0.08 4.69 0.60 355.3 29.2 Ambatovy Probable 47.5 0.92 0.07 4.72 1.16 435.2 34.5 Proven+Probable 82.8 0.96 0.08 4.71 0.92 790.4 63.7 Proven 8.3 0.81 0.08 4.08 0.31 67.3 6.7 Analamay Probable 55.1 0.94 0.09 3.90 0.92 516.3 47.4 Proven+Probable 63.4 0.92 0.09 3.92 0.84 583.6 54.1 Proven 43.6 0.97 0.08 4.57 0.55 422.6 35.9 All Deposits Probable 102.6 0.93 0.08 4.28 1.03 951.5 81.9 Mineral Reserve Proven+Probable 146.2 0.94 0.08 4.36 0.89 1,374.1 117.8 (ROM) Stockpiles Proven 8.1 0.81 0.06 6.64 1.82 65.5 4.0 Total Mineral Proven+Probable 154.3 0.93 0.08 4.48 0.94 1,439.6 121.8 Reserve (ROM)
Table 37: Mineral Reserve (ROP)
Tonnage Ni Co Al Mg Ni metal Co metal Deposit Classification (Mt) (%) (%) (%) (%) (kt) (kt) Proven 31.5 0.96 0.08 4.70 1.21 303.0 26.2 Ambatovy Probable 39.8 0.85 0.07 4.76 1.55 338.9 29.5 Proven+Probable 71.3 0.90 0.08 4.74 1.40 641.8 55.8 Proven 7.5 0.77 0.08 4.07 1.00 58.0 6.1 Analamay Probable 47.8 0.88 0.09 3.92 1.39 421.2 41.4 Proven+Probable 55.3 0.87 0.09 3.94 1.33 479.2 47.5 Proven 39.0 0.93 0.08 4.58 1.17 361.0 32.3 All Deposits Probable 87.6 0.87 0.08 4.30 1.46 760.1 71.0 Mineral Reserve Proven+Probable 126.6 0.89 0.08 4.39 1.37 1,121.1 103.3 (ROP) Stockpiles Proven 4.0 0.77 0.06 6.63 2.52 31.2 2.4 Total Mineral Proven+Probable 130.6 0.89 0.08 4.46 1.41 1,152.3 105.7 Reserve (ROP)
Table 38: Waste material within final itp
Deposit Classification Tonnage (Mt) Waste (insitu) 91.69 Ambatovy Waste (insitu+sheeting) 117.86 Waste (insitu) 60.47 Analamay Waste (insitu+sheeting) 79.05 Waste (insitu) 152.16 All Deposits Waste (insitu+sheeting) 196.91
15.5 Mineral Reserve Estimate – Existing Stockpile During the course of production to date, substantial low-grade (LG) and high-grade (HG) stockpiles have been developed. The status of these stockpiles as at 30 June 2018 is illustrated in Figure 57.
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Figure 57: Existing stockpiles Stockpiled material has in the most part been estimated based on truck counts and grade control data. More recently with the implementation of the Wenco system the tracking of stockpile material has been more rigorous. IMC were provided with an estimate of the stockpile material as of 30 June 2018. This is shown inTable 39. Table 39: Existing stockpile tonnes
Tonnages Ni Co Fe Al Mg Si Mn High grade 45,928 1.10 0.09 43.48 4.85 1.34 2.34 0.80 Low grade 10,691,458 0.81 0.06 39.80 6.65 1.82 3.81 0.64 Total 10,737,386 0.81 0.06 39.82 6.64 1.82 3.80 0.64
IMC has included this stockpile material in the Mineral Reserve. Conservative stockpile recovery factors of 75% of the LG are applied to the ROM Reserve (95% of the HG) and 50% of the LG to the ROP Reserve (91.9% of the HG), with similar figures included in the LOM plan schedule (see Table 45) with the material recovered towards the end of the mine life. The ROP grade has had the relevant grade adjustments applied.
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15.6 Mineral Reserve Statement The Mineral Reserves for Ambatovy are a subset of the Mineral Resource (i.e. the Mineral Resource is inclusive of the Mineral Reserve). The Mineral Reserve estimate based on the mining of the current insitu Mineral Resource is154.3 Mt ROM which is converted to 130.6 Mt ROP. The grade of the product in the Mineral Reserve that is fed in to the PAL plant is 0.89% Ni, 0.08% Co. This is the ROP grade. The modifying factors applied to the Mineral Resource have been summarised in Section 15.2. A proportion of the stockpiled material has been included in the Mineral Reserve.
15.6.1 QP Comments As of the Effective Date, the authorQP responsible for this section, Stewart Lewis, is of the opinion that as a producing property, the risks to which the mineral reserve estimates could be materially affected by mining, metallurgical, infrastructure, permitting and other relevant factors are negligible.
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16 Mining Methods
16.1 Introduction Ambatovy has been in production since 2013. The last update to the Mineral Reserve was reported in September 2014. Over the ensuing four-year period, the knowledge base of how to gain the highest value from the Ambatovy resource has increased substantially. One of the most fundamental changes in the mining approach is that the optimisation of the mining operations is undertaken on a holistic basis, i.e. the complete material flow from resource modelling through to saleable product is taken into account when developing mine plans.
16.2 Pit Limits The pit limits are well defined; generally, the base of the pits is the floor of the saprolite (the exception being if the slope is greater than 27° when the bedrock may form the pit floor), whilst the lateral extents are constrained either by lease/environmental boundaries or economic limits. The development of the ultimate pit limits is outlined in Section 15.3 of this report. The material within this pit limit defined the Mineral Reserve for Ambatovy.
16.3 Pit Design/Stages Definition The pit design work is undertaken in two phases, the first being the relatively straightforward final pit limit design (as outlined above), and the second phase being the design of staged pits within the LOM shell. The design of the staged pits was further subdivided into the five-year detailed design tactical pits and the Year 6 to LOM staged pits.
16.3.1 Five-Year Detailed Pit Designs Detailed pits were designed for Years 1 to 5 with input from the optimised shells, surface constraints such as haul road, infrastructure and environmental exclusion zones and the base of saprolite. The preliminary LOM schedule informed the design process as to which pits were to be targeted in the first five years, the material directly fed to the OPP that was recovered from stockpile, the material recovered from stockpile and the material sent to waste. The design parameters used for the pit designs are summarised in Table 40. Table 40: Pit design parameters
Pit Parameter Unit Value Ramp width (single/double) m 30 Ramp gradient % 8 Bench height m 4 Bottom pushback minimum size m 0 Standoff from haulroad m 10 Standoff from boundary m 30 Haulroad gradient % 8 Haulroad section m 300 Pushback minimum width m 30 Minimum switchback width m 60 Bench face angle (batter angle) m 45 Berm width m 8
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16.3.2 Tactical (Minesched) Five-Year Plan and Schedule Whilst the design stages for the five-year pits and preliminary schedule for the next five years of operation were based on the LP strategic scheduler. To develop a workable five-year mine development plan, a tactical schedule based on allocation of excavators to benches and detailed ramp design and dump plans is required. AMSA use the MineSched planning system for this tactical scheduling. The five-year tactical schedule used the LP schedule to guide the sequence and mining quantities. Mine production at Ambatovy remains constant for the remainder of 2018, 2019 and 2020. Total material movement is summarised in Figure 58; this shows the ramp-up in material movement from 2021 onwards. Several months of increased rehandle are required in 2021 to maintain PAL feed. Mining progression per pit design phase is shown in Figure 59. The Ambatovy deposit is favoured for mining up until completion of the North East Dam at Analamay in late 2020. Mining then ceases for some time in the Ambatovy pits and focuses on the high-grade areas in Analamay that were released for mining on completion of the dam. Completing the dam earlier would allow these pits to be mined sooner and may reduce the amount of rehandle required in 2021.
Figure 58: Five-year plan total movement by material type
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Figure 59: Five-year plan ore mined by pit
16.3.3 Phase 2 Life of Mine Design and Schedule Having completed the detailed five-year pit designs and tactical schedule, this phase of the scheduling locked in the five-year schedule and focused on the remaining LOM schedule using strategic LP scheduling (i.e. Years 6 to LOM). In the first pass, the phases were defined for Year 6–10 from pit design shells (without detailed ramps). From Year 11 onward, the staged pit development was controlled by consolidated optimisation shells that represented logical pit development steps. Through iterations of scheduling, design refinement and then re-scheduling, a final design/schedule iteration was chosen to define the LOM schedule. Apart from the design constraints, the other key constraints, as per the five-year plan, are the process constraints. It has been difficult to lock down all process constraints as the process plant is still going through improvements and de-bottlenecking programs. The problem was exacerbated somewhat by the fact that no definitive basis for constraints appears to exist. It is a somewhat circuitous process as it would appear that the process team take the outputs from the mine schedule and then put that data in a spreadsheet solution that then defines if a limit has been reached, they then send back a new set of constraints that has actually been informed from the constraints we imposed on the schedule that was input. In a perfect world, the constraint model used by the process team would be integrated as part of the mine scheduling system; however, it is most likely this solution is some time off as the process team are still developing an understanding of the drivers behind the process plant. Notwithstanding the above, a set of constraints has been agreed on as per Table 41.
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Table 41: LP LOM production onstraintsc
Year LP constraint Comment 2018 2019 2020 2021 2022 2023 2024+ Sourced from Maximum pipe “180509_Template_NI43-101 7.08 7.08 7.08 7.08 7.08 7.08 7.08 tonnes (Mt) inputs_a04 (Director Plant Operation) rev4.xlsx” 2018-2023 sourced “PAL Feed Rate PAL feed tonnes Inputs 2018 06+06_5YP Rev 2.502 5.407 5.817 5.791 5.990 5.850 5.895 (Mt) 01.xlsx”. 2024+ Limit from 2017 LOM Work 2018-2023 sourced “PAL Feed Rate PAL acid tonnes Inputs 2018 06+06_5YP Rev 1.75 1.909 2.001 2.109 2.134 2.131 2.131 (Mt) 01.xlsx”. 2024+ Limit from 2017 LOM Work Maximum slurry 12.37 12.37 12.37 12.37 12.37 12.37 12.37 Sourced from (Mm3) “180509_Template_NI43-101 Tailings dry inputs_a04 (Director Plant 10.8 10.8 10.8 10.8 10.8 10.8 10.8 tonnes (Mt) Operation) rev4.xlsx” Maximum 1.5 1.5 1.5 1.5 1.5 1.5 2.0 Determined by site stockpiling (Mt) Maximum 1 1 1 1 1 1 3 Determined by site reclaim (Mt) Maximum Ni 60 60 60 60 60 60 60 Limit from 2017 LOM Work metal (kt) Maximum Co 5.6 5.6 5.6 5.6 5.6 5.6 5.6 Limit from 2017 LOM Work metal (kt) Equivalent total Based on smoothing a LOM case 12,000 12,000 12,000 14,000 14,000 14,000 20,000 excavator hours that had no limitation. Equivalent total Based on smoothing a LOM case 120,000 120,000 120,000 120,000 120,000 120,000 120,000 truck hours that had no limitation.
16.3.4 Stockpile Strategy The stockpiling strategy is based on the stockpile parameters outlined in Section 15.2.4 of this report. Within the material classification, the strategic schedule (and for the five-year plan, the tactical schedule) allocates material dynamically to either go to one of the ore stockpiles or directly to the OPP. The LP strategic scheduler develops the stockpiling strategy that maximises the net present value of the project through balancing out the various processing constraints against the material that is available in each period either directly from the pit or recovered from a stockpile. The key driver of the stockpiling decision is the inherent value of the block and the PAL mass flow constraint, with acid demand limiting throughput in some years. IMC has endeavoured to ensure that the strategic scheduler is able to optimise stockpiling strategies without being constrained by equipment, i.e. IMC has erred on the side of overcapacity in the production fleet as the overarching requirement is to feed the PAL the optimum balance of feed. The stockpiles are shown for the LOM in Figure 60.
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Figure 60: Stockpile balance
16.4 Development Plan The mining development sequence (i.e. how the pit stages are sequenced) has been primarily driven by the optimised sequence developed sequence as shown in Figure 53 and Figure 54 in Section 15.3 of this report. Imposed on top of this optimum sequence has been timing constraints dictated by the construction of sediment containment structures in the Analamay area. This controls the earliest start-up time for Analamay. Together with the infrastructure constraint, IMC has imposed some practical constraints as to where mining is carried out from year to year. It would not be practical to bounce between Ambatovy and Analamay every few months so where possible, mining is focused in one area or the other. The final constraint imposed on the strategic schedule is that staged pits are forced to be mined out before moving to the next staged pit in the sequence. This stops the strategic scheduler from leaving the lower saprolite material in the floor. This development methodology forces this material to be cleaned up and taken to the relevant stockpile, or direct fed if appropriate.
16.5 Waste Disposal There is one main external waste dump for each of the Ambatovy and Analamay Deposits and two central low-grade ore stockpiles located adjacent to the OPP. The waste dumps for Ambatovy and Analamay are approximately 15 Mm3 in volume each, and the active stockpiles are approximately 40 Mm3 by volume. The majority of the waste and sub-economic mineralised material is dumped back into the mined-out pits. The footprint of these dumps and stockpiles are outlined in Figure 61 and Figure 62 (this does not include the in-pit waste dumping).
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The waste dumps are presently designed quite conservatively at a 15° overall slope. A work program is presently scheduled for some time in the next few years to optimise the dump designs, rationalise the dumping strategy (particularly as to how best to develop low grade and medium grade stockpiles within the dump footprint for later recovery). These changes to dump strategy will provide upside to the mine operations. At this time, final positioning and size of each waste dumps are still under study and need to be considered as work in progress.
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Figure 61: Ambatovy waste dump locations
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Figure 62: Analamay dumps and sediment containment
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16.6 Grade Control AMSA has various methods available for grade control. The focus of grade control is to align the grade control with the strategic plan. The difficulty in the AMSA grade control system is that: • The blocks that are ore and waste are based on an economic analysis – this takes into account the prevailing processing costs and the nickel and cobalt prices; obviously these all change on at least an annual basis • The stockpiling strategy changes dynamically from period to period. A mechanism is required to enable the site team to carry out block allocation with inputs from PAL factors, process costs factors and relative economics. In 2017, IMC developed such a system for AMSA that illustrates that it is possible to have a relatively simple system to flag blocks appropriately in the grade control model. This simple graphical user interface takes in a bench plan, production run or pit shells and then assigns grade bin allocations based on an interactive process that allows the appropriate balance to be found for that production run period by changing the allocation of bins on a short-term basis. It is beyond the scope of this Technical Report to go into the methodology in detail. What is important however is that IMC has made certain assumptions regarding cut-off grade and stockpiling strategies in the development of the Mineral Reserve. These assumptions are underpinned by a systematic approach to grade control block allocation that will facilitate the site teams achieving the modelled outcomes The grade control GUI includes fields for the daily target ROP production for that period (noting that the LP scheduler will have provided the optimum annual target for each year of operation via the quarterly five-year planning optimisation); this same LP schedule will have provided the target grade bin allocation. Once the selected blocks for the allocation have been chosen, a small grade control block model is created that as a bare minimum has the grades of each element that contributes to processing costs, constraints and revenue generation (along with X, Y and Z coordinates). The routine then works out how many ROP tonnes are contained within the grade control input model (based on the current ROM-ROP modifying factors) and thereby the number of operating days that the data represents based on the daily target OPP feed input. The daily acid consumption target is also input (noting that this should generally be the annual capacity divided by plant operating days). However, if the process plant run into problems with acid, the target may need to be reduced for short periods or maybe sometimes be increased. This target is not used in the base run; however, it is used to back calculate the OPP feed that would consume this acid. The schedule is then run for that period and the acid demand per day is calculated and presented on the screen along with other key drivers (Mn/C, Al, Mg). The goal is to match the PAL capacity and acid demand per day with the acid available whilst also meeting other constraints. If there is a mismatch, there are two options available: 1. Change grade bin allocation. 2. Change the daily OPP target feed.
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To assist in the selection of the appropriate mechanism to get the best outcome, a number of key parameters are provided in the interface, these being: • ROP OPP per day to meet acid – this is the tonnes of ROP that should be fed to the OPP to consume all the acid, the acid target having been based on the target acid per day tonnage field. It is worth noting that the economics of the mine are generally maximised if all of the acid is consumed as it is best to feed as much saprolite as possible due to its higher grade. The effect of maximising nickel grade through maximising saprolite feed results naturally in higher acid consumption, but also maximises project net present value. • The ROM OPP per day is shown for reference only. • The Ni metal per day, Co metal per day, NiEq virtual metal per day and the Normalised Revenue per day. An example grade control run is provided in Figure 63.
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Figure 63: Block allocation unr
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16.7 Mining Fleet The base case operating cost model developed by IMC is an owner-mining based model with no contractor support apart from construction of the sedimentcontainment dams which are considered capital projects. As it is a zero-based model and not based on supplier quotes, IMC has had to rely on AMSA supplied operating data together with in-house database of equipment. The running costs of the machines were sourced from the IMC cost database.
16.7.1 Load, Haul and Excavate IMC has assumed that the current operations using conventional truck shovel operations within the pits for the movement of ore, waste and road sheeting material will continue. A mixed fleet is used at Ambatovy, large Cat 6030 and 6020 excavators and Cat 777 haul trucks (90-t payload) move the bulk of ore and waste material. A smaller fleet of Cat 385 excavators (90-t, 5.4 m3 bucket) and Cat 745 (41-t payload) articulated dump trucks are used for residual ore mining, stockpiling reclamation, quarrying activities and road sheeting movement and to assist the large fleet when capacity is available. A high quantity of road sheeting material movement is required at Ambatovy, it is estimated to be in the order of 15% of total pit movement. The moist Ambatovy conditions require the laying of road sheeting at regular intervals (approximately every 50 m). The high moisture content of the laterite ore means haul trucks cannot be loaded on the lower bench to the excavator and are therefore often top loaded from the bench alongside the excavator. This setup will result in a lower excavator productivity due to the increased swing time for each pass.
16.7.2 Ancillary and Support Fleet A large cost at laterite nickel operations is the ancillary and support fleet. This fleet includes additional excavators, dozers, graders, wheel loaders, service trucks etc. The ancillary fleet is required to construct roads, sediment containment structures, mine road sheeting material (from the quarries) and carry out general clean-up operations around mining faces and to provide support to the primary excavation equipment. Additional dozers (D6s, D8s) are required for the laying of geofabric and road sheeting as well as land clearing and grubbing in preparation for pit mining. Approximately 840 ha of land is required to be cleared over the course of the mine life. It is assumed that all the required clearing will be completed in the proceeding seven years. Front-end loaders are required for OPP feed blending, coarse reject loading, road construction and stockpile reclamation (where necessary).
16.7.3 Mining Fleet Allocation (P1/P2 Fleet) Trucking at Ambatovy represents the largest cost for mining. An important consideration in forecasting operating costs has therefore been accurately forecasting truck fleet requirements to ensure trucking capacity does not limit feed to the OPP to meet PAL demand. This task is somewhat complicated at Ambatovy as two different fleets target different material types. The large fleet comprising Cat 6030/Cat 6020 Excavators loading Cat 777 trucks targets bulk mining and areas of lower moisture content, whilst a smaller fleet of Cat 390 or equivalent loading Cat 745articulated dump truck’s targets the areas that require more selective mining, high moisture areas and also carry out
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the mining of sheeting and quarrying as required. These two fleets have been designated the P1 and P2 fleet respectively. The large P1 fleet completes the majority of ore and waste mining but towards the bottom of the deposit the ore floor is interspersed with intrusions and difficult to navigate. This area is mined by the smaller P2 fleet but the point at which the P2 fleet replaces the P1 fleet is subjective. The P2 fleet mining in the floor of the pit is dependent on geology, operational requirements and/or fleet availability. The approach is premised on the fact that as mining gets closer to the top of the saprolite zone, ie toward the feralite/saprolite boundary, the digging becomes more complex, the moisture is higher and the material is therefore much more amenable to mining with the smaller P2 fleet. This has been managed in a dynamic scheduling system that works out the best allocation of equipment in the context of costs and balancing of medium-term fleet requirements.
16.8 Manning Mine operator staff is assumed to be on a two week on/one week off roster. Three panels are required to make up a 24-hour, seven-day per week operation. Staffing numbers allow for four weeks annual leave (20 days) and a sick leave/AWOL allowance of 2.5%. Tradesman staff are assumed to work day shift only, on fivea -day on/two-day off roster. Annual leave and sick leave/AWOL allowance is as above. In order to align IMC estimates with measured site values, tradesman manning numbers have been increased by a factor of 50% (1.5 actual man required, for each IMC tradesman calculated). IMC manning maintenance numbers are derived from the maintenance hour/operating hour ratio. The maintenance ratio is dependent on the expertise and efficiency of the maintenance personnel employed. A typical roster definition is shown in Figure 64.
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Figure 64: Labour roster defintion The organisation manning numbers are shown in in Table 42 for administration roles and Table 43 shows the operator numbers for the first 10 years of operation. It should be noted that the cost modelling system has two options for manning of equipment. One option is to only allocate operators to the fleet required to achieve production targets in each year of operation – this assumes operators can be re-allocated as required from year-to-year and results in less operators. The other option is to allocate operators to all fleet whether it is required or not. The latter option was used for the Ambatovy modelling.
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Table 42: Manpower summary – staff umbersn
Act ID Department Role Number Expatriate roles 1.1.2.1 Technical Mine Planning Coordinator 1 1.1.2.1 Technical Senior Planning Engineer 1 1.1.2.1 Technical Geology Coordinator 1 1.1.2.1 Technical Senior Geologists 2 1.1.2.2 Operations Surveyor 1 1.1.2.2 Operations Mine Manager 1 1.1.2.2 Operations Production Supervisor 1 1.1.2.3 Maintenance Logistics Coordinators 3 1.1.2.3 Maintenance Maintenance Supervisor 1 1.1.2.3 Maintenance Maintenance Manager 1 Local roles 1.1.3.1 Technical Technician Mine Planning 1 1.1.3.1 Technical Geology Assistants 4 1.1.3.2 Operations Senior Operation Engineer 1 1.1.3.2 Operations Blasting Technician 1 1.1.3.2 Operations Dispatch Technician 1 1.1.3.2 Operations Mining Technicians 2 1.1.3.2 Operations Auxiliary ADMs 2 1.1.3.2 Operations Survey Assistants 2 1.1.3.2 Operations Coordinator Mine Operation 2 1.1.3.3 Maintenance Stores Clerks 3 1.1.3.3 Maintenance Maintenance Planners 2
Table 43: Manpower summary – maximum and minimum Operations and Maintenance numbers
Operator type Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Total 395 432 423 401 423 469 544 532 471 564 Ancillary 41 41 41 39 38 38 38 38 38 38 Operators Dozer Operators 43 43 43 43 43 43 66 66 66 80 Drill Operators 7 7 7 7 7 7 7 7 7 7 Excavator 30 30 30 27 27 30 43 43 42 47 Operators Loader Operators 10 10 10 10 10 8 7 7 7 7 Tradesmen 148 159 150 144 159 176 218 197 158 226 Truck Operators 116 142 142 131 139 167 165 174 153 159
16.9 Mine Schedule The mine production schedule is a hybrid schedule incorporating Year 1–5 from the tactical schedule developed in Minesched and the remaining years in the LP scheduler. The scheduling development has been focussed on the mining operations not imposing a constraint on the PAL production rate. The constraints outlined in Table 41 were the core drivers of the schedule (those being the PAL constraints). In a perfect world, the 60kt of Nickel metal targeted would be produced each year; however, this is not possible in most years as the Mineral Resource mined is not capable of supporting that production– either
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through hitting an autoclave mass throughput constraint (ie the grade of Ni in the feed is not high enough to produce 60 kt of metal within the mass flow constraint) or hitting the limit of acid production capacity at the PAL. The balancing of these two constraints is largely the result of the inherent characteristics of the Mineral Resource whereby the ferralite zone is lower nickel and lower MgO and therefore consumes less acid (with MgO being the main acid consumer) whereas the saprolite zone has higher nickel but also higher MgO which consumes more acid. The obvious goal in development of the schedule is to at all times have a peak of both the autoclave mass capacity and optimised acid demand of the feed. This is achieved through selective stockpiling and recovery of the ore. In all periods of the mine life, the mass flow into the autoclaves has been maintained at the peak. It is not required due to presentation of ore from the pit, to always keep the acid demand at the peak. Figure 65 and Figure 66 illustrate this.
Figure 65: Ambatovy LOM PAL feed tonnes
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Figure 66: Ambatovy LOM PAL acid consumption
16.9.1 Scheduled Quantities The presentation of scheduled quantities for Ambatovy comprises many factors, including: • ROM mined • ROP component of the ROM mined • ROM and ROP direct feed to OPP • ROM and ROP to various stockpiles • ROM and ROP recovered from various stockpiles • Ni metal, Co metal and all other elements contained in the above • Sheeting material mined • Sheeting material to the OPP • Reject material from the OPP • Quarry material mined for sheeting and road construction • Waste mined. The mine life runs from January 2019 through to 2044. The mine life numbers are represented in Table 44 and Table 45 and Table 46 and Table 47. These tables for the LOM commence six months later than when the Mineral Reserves have been declared (which is 30 June 2018). The schedule is shown on a yearly basis for calendar year 2019 through to 2023 (the five-year plan) and then in five-year aggregates for the remainder of the mine life. Table 45 includes Inferred material within the five-year plan. Table 46 includes both Inferred material and stockpiled material with the relevant factors applied.
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Table 44: Waste mined
Total 2019 2020 2021 2022 2023 2024–2028 2029–2033 2034–2038 Remainder Mwt 186 4.72 3.71 8.02 8.07 5.42 48.44 51.17 33.97 22.20 Waste (ex-pit) Mdt 138 2.93 2.30 4.97 5.00 3.36 37.40 39.99 25.86 16.67 Mwt 47 1.44 1.57 1.70 1.87 1.70 13.58 12.52 8.27 4.80 Quarry (road sheeting) Mdt 43 1.36 1.49 1.61 1.78 1.62 12.22 11.26 7.44 4.32 Mwt 1 0.54 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 Quarry (other civils) Mdt 1 0.52 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 Mwt 48 1.98 1.57 1.70 1.87 1.70 13.58 12.52 8.27 4.80 Quarry (total) Mdt 44 1.88 1.49 1.61 1.78 1.62 12.22 11.26 7.44 4.32 Mwt 234 6.70 5.27 9.72 9.95 7.12 62.02 63.69 42.24 27.00 Waste (pit and quarry) Mdt 182 4.81 3.79 6.59 6.79 4.98 49.62 51.26 33.30 20.99
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Table 45: ROM data (excludes sheeting)
Total 2019 2020 2021 2022 2023 2024–2028 2029–2033 2034–2038 Remainder Total ROM Ore Mined Mdt 156 7.76 9.11 7.40 8.05 9.05 44.92 38.36 19.11 12.65 Ni % 0.95 1.08 1.07 1.02 1.00 1.03 1.00 0.90 0.90 0.77 Co % 0.08 0.09 0.09 0.08 0.09 0.08 0.09 0.08 0.07 0.07 Fe % 44.27 45.26 43.73 47.15 48.45 42.89 44.67 44.06 42.26 42.98 Al % 4.36 4.10 4.74 4.14 3.78 4.12 4.03 4.59 4.76 4.80 Mg % 0.90 0.73 0.62 0.55 0.44 1.15 1.10 0.83 0.87 1.04 Si % 3.47 2.82 2.60 2.28 2.20 4.26 3.51 3.37 4.34 4.30 Mn % 0.58 0.16 0.25 0.41 0.37 0.11 0.73 0.68 0.69 0.69 C % 0.08 0.03 0.04 0.07 0.05 0.03 0.10 0.10 0.10 0.10 Cr % 1.56 1.62 1.50 2.35 1.71 1.68 1.64 1.44 1.37 1.32 Mn:C (rop) % 6.75 5.90 5.96 5.77 6.67 3.72 7.34 6.94 7.26 7.19 ROM Ore to Stockpile Mdt 31 2.10 2.57 1.87 1.79 2.52 11.18 8.42 0.12 0.00 Ni % 0.81 1.07 1.06 0.98 1.00 1.11 0.69 0.66 0.52 0.57 Co % 0.06 0.10 0.09 0.07 0.08 0.09 0.05 0.05 0.04 0.04 Fe % 45.71 45.52 43.52 47.74 48.73 43.75 46.66 44.68 44.60 32.88 Al % 4.59 4.08 4.76 4.18 3.76 3.93 4.52 5.19 5.88 6.03 Mg % 0.56 0.63 0.61 0.45 0.39 1.00 0.54 0.49 0.36 2.62 Si % 2.80 2.81 2.72 2.09 2.05 3.85 2.70 2.97 2.22 11.01 Mn % 0.46 0.15 0.24 0.69 0.92 0.14 0.51 0.48 0.57 0.28 C % 0.09 0.03 0.04 0.14 0.14 0.08 0.10 0.10 0.10 0.08 Cr % 1.16 1.17 1.40 1.23 1.19 1.40 1.16 1.01 0.80 0.62 Mn:C (rop) % 4.96 5.87 6.23 4.84 6.40 1.74 5.09 4.85 5.81 3.36 ROM Stockpile Reclaim Mdt 42 0.63 0.18 1.13 0.55 0.55 1.12 4.35 15.36 18.11 Ni % 0.77 1.02 1.10 1.08 0.84 1.05 0.92 0.83 0.81 0.67 Co % 0.06 0.09 0.09 0.09 0.06 0.07 0.08 0.07 0.07 0.05 Fe % 43.11 43.40 43.07 45.68 47.30 45.34 42.59 39.55 41.46 45.04 Al % 5.10 4.25 4.11 4.32 4.60 3.99 4.91 5.53 5.24 5.02 Mg % 1.64 1.07 1.12 0.41 0.22 0.85 1.90 3.07 2.26 0.91 Si % 3.40 4.15 3.90 2.33 1.96 3.18 3.50 4.28 3.88 2.85 Mn % 0.50 0.15 0.13 0.19 0.17 0.08 0.38 0.57 0.58 0.49 C % 0.10 0.02 0.02 0.03 0.03 0.01 0.07 0.11 0.10 0.10 Cr % 2.37 0.00 0.00 0.00 0.00 0.00 2.84 4.07 3.19 1.63 Mn:C (rop) % 5.42 8.49 8.52 6.75 6.50 10.14 7.81 5.39 5.57 4.77
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Table 46: ROP tonnes (direct)
Total 2019 2020 2021 2022 2023 2024–2028 2029–2033 2034–2038 Remainder OPP ROP (direct) Mdt 111.6 5.66 6.54 5.53 6.26 6.53 28.51 25.76 16.19 10.67 Ni % 0.93 1.08 1.04 0.86 0.92 0.93 1.03 0.91 0.84 0.72 Co % 0.09 0.09 0.08 0.07 0.08 0.07 0.10 0.08 0.08 0.08 Fe % 43.04 45.16 42.63 38.97 44.49 39.28 44.07 43.68 41.99 43.00 Al % 4.25 4.10 4.61 3.43 3.48 3.86 3.90 4.44 4.80 4.83 Mg % 1.27 0.77 0.60 0.49 0.41 1.12 1.64 1.40 1.38 1.50 Si % 3.44 2.83 2.48 1.95 2.06 4.08 3.57 3.39 4.22 4.08 Mn % 0.60 0.17 0.25 0.27 0.19 0.09 0.81 0.75 0.71 0.71 C % 0.08 0.03 0.04 0.04 0.03 0.01 0.10 0.10 0.10 0.10 Cr % 1.63 1.78 1.50 2.27 1.71 1.65 1.81 1.56 1.37 1.32 Mn:C (rop) % 7.50 5.92 5.86 6.70 7.05 9.22 8.18 7.56 7.31 7.27 OPP ROP (from stockpile) Mdt 37.0 0.63 0.18 1.13 0.55 0.55 0.96 3.72 13.26 15.99 Ni % 0.75 1.02 1.10 1.08 0.84 1.05 0.91 0.83 0.80 0.64 Co % 0.06 0.09 0.09 0.09 0.06 0.07 0.08 0.07 0.07 0.05 Fe % 42.96 43.40 43.07 45.68 47.30 45.34 42.63 39.64 41.40 44.60 Al % 5.08 4.25 4.11 4.32 4.60 3.99 4.91 5.54 5.24 5.00 Mg % 1.89 1.07 1.12 0.41 0.22 0.85 1.90 3.06 2.43 1.41 Si % 3.40 4.15 3.90 2.33 1.96 3.18 3.47 4.23 3.84 2.93 Mn % 0.50 0.15 0.13 0.19 0.17 0.08 0.38 0.58 0.58 0.49 C % 0.13 0.02 0.02 0.03 0.03 0.01 0.27 0.28 0.19 0.05 Cr % 2.33 0.00 0.00 0.00 0.00 0.00 2.85 4.09 3.18 1.61 Mn:C (rop) % 2.76 8.49 8.52 6.75 6.50 10.14 1.41 2.07 3.65 1.32 Total OPP ROP Mdt 143.9 5.41 5.82 5.79 5.99 5.85 29.48 29.48 29.45 26.66 Ni % 0.89 1.03 1.03 1.01 0.94 0.96 1.03 0.90 0.82 0.67 Co % 0.08 0.09 0.09 0.08 0.08 0.08 0.10 0.08 0.07 0.06 Fe % 43.43 44.02 42.82 45.75 47.29 41.80 44.02 43.17 41.72 43.96 Al % 4.52 4.12 4.72 4.16 3.85 4.17 3.94 4.58 5.00 4.93 Mg % 1.60 1.49 1.33 1.25 1.13 1.87 1.65 1.61 1.86 1.44 Si % 3.54 3.24 2.87 2.63 2.52 4.61 3.57 3.49 4.05 3.39 Mn % 0.59 0.17 0.25 0.30 0.20 0.10 0.80 0.73 0.65 0.58 C % 0.09 0.03 0.04 0.04 0.03 0.01 0.10 0.10 0.10 0.10 Cr % 1.84 1.60 1.50 2.26 1.71 1.65 1.85 1.88 2.19 1.50 Mn:C (rop) % 6.94 5.98 5.86 6.70 7.02 9.19 8.13 7.27 6.49 5.75 *OPP ROP (Direct) only includes tonnes mined directly from the pit to the OPP. It does not include tonnes mined to the stockpile.
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Table 47: PAL feed
Total 2019 2020 2021 2022 2023 2024–2028 2029–2033 2034–2038 Remainder PAL Total Processed (from Pit & Stockpile) Mdt 143.9 5.41 5.82 5.79 5.99 5.85 29.48 29.48 29.45 26.66 Ni % 0.89 1.03 1.03 1.01 0.94 0.96 1.03 0.90 0.82 0.67 Co % 0.08 0.09 0.09 0.08 0.08 0.08 0.10 0.08 0.07 0.06 Fe % 43.43 44.02 42.82 45.75 47.29 41.80 44.02 43.17 41.72 43.96 Al % 4.52 4.12 4.72 4.16 3.85 4.17 3.94 4.58 5.00 4.93 Mg % 1.60 1.49 1.33 1.25 1.13 1.87 1.65 1.61 1.86 1.44 Si % 3.54 3.24 2.87 2.63 2.52 4.61 3.57 3.49 4.05 3.39 Mn % 0.59 0.17 0.25 0.30 0.20 0.10 0.80 0.73 0.65 0.58 C % 0.09 0.03 0.04 0.04 0.03 0.01 0.10 0.10 0.10 0.10 Cr % 1.84 1.60 1.50 2.26 1.71 1.65 1.85 1.88 2.19 1.50 Mn:C (rop) % 6.94 5.98 5.86 6.70 7.02 9.19 8.13 7.27 6.49 5.75 Slurry volume Mm3 261 10.00 10.50 10.50 10.50 10.50 53.89 53.34 53.54 48.21 PAL Acid Kt Kt 48144 1735 1882 1788 1759 2047 9508 9893 10623 8909 Ni metal t 1130231 49285 52920 51559 50024 49900 267923 235445 214992 158184 Co metal t 98899 4310 4308 4010 4311 3938 25399 20865 18173 13585 Rec Ni Constraint t 60000 60000 60000 60000 60000 300000 300000 300000 420000 Rec Co Constraint t 5700 5700 5700 5700 5700 28500 28500 28500 39900
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17 Recovery Methods
The process for the recovery of nickel and cobalt at the Ambatovy Project was designed by DMSA, which was acquired by Sherritt in 2007. Sherritt has extensive experience in the nickel industry and has been involved in the commercialisation of more than 15 nickel-cobalt projects on five continents. These include acid pressure leaching, ammonia pressure leaching, caron processing (reduction roast and ammonium carbonate leaching), atmospheric leaching, and hydrogen reduction. Sherritt’s extensive knowledge of nickel processing has been utilised to select the optimum process for treating the Ambatovy Project ore. A simplified overall flowchart of the OPP and Processing Plant is shown in Figure 67.
Figure 67: Ambatovy Project mine and process flowsheet
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17.1 Ore Preparation Plant The ore preparation circuit is designed for the processing of the limonite (ferralite) and LMS fractions of the orebody. Small amounts of saprolite material are fed to the ore preparation circuit, where the nickel- rich skins are scrubbed off of the competent high-magnesium boulders. The ROM ore is screened through a grizzly before being fed to two autogenous scrubbers in series. The first scrubber discharge material is screened, and undersize material is forwarded to the ore thickener. The remainder of the material is forwarded to the second scrubber for continued treatment and liberation of ore fines. Oversize material from the second scrubber (+8 mm) is rejected, and a middling stream is stockpiled for future treatment. The ore thickener allows preparation of consistent ore slurry pulp density for feed to the slurry pipeline. The ore slurry thickener underflow, at about 38–42% solids, is slightly diluted with water for feed to the pipeline. A 600 mm diameter slurry pipeline, approximately 220 km in length, transports the ore from the Property to the Processing Plant. Although large and long, the slurry pipeline operation is relatively simple in comparison with some other slurry pipelines which have been built in South America, where slopes are more extreme, multiple pump stations are often required, and multiple products are sometimes shipped in the same pipeline.
17.2 Leach and Sulphide Precipitation Plant All the steps of the leach and sulphide precipitation process have been successfully commercialised in other facilities and all equipment has been proven at the Ambatovy Project. Sherritt has significant experience in the design and operation of each step.
17.2.1 Ore Slurry Thickening Ore slurry from the pipeline is received in two thickeners which ensures a consistent feed slurry pulp density to the acid leaching circuit. Excess water in the ore slurry reports as thickener overflow and is used as part of the plant water supply.
17.2.2 Pressure Acid Leaching The PAL section consists of five parallel trains. The thickened ore slurry is fed to the pressure leach autoclaves through three stages of direct heating, with inter-stage pumping. The first two stages utilise flashed steam from the autoclave let down system, while the third stage utilises steam from the power plant/acid plant steam header. In the agitated horizontal autoclaves, the slurry is contacted with sulphuric acid at a temperature of approximately 260°C and is held for up to 75 minutes to complete the leaching of nickel and cobalt as well as precipitation of iron from solution. The leached slurry is discharged through three stages of flashing, which serve to cool and concentrate the slurry and generate steam for use in the preheat circuit and other areas of the Processing Plant.
17.2.3 Slurry Neutralisation and Solids Wash Circuit The cooled PAL discharge slurry from the autoclave is partially neutralised with limestone in the slurry neutralisation circuit prior to advancing to the six-stage, CCD wash circuit. The leach residue and other solids are discharged from the last thickener and sent to tailings neutralisation. The clarified raw leach liquor is drawn off from the first wash stage as overflow for further processing.
17.2.4 Raw Liquor Neutralisation The clarified leach solution from the CCD circuit is further neutralised in two parallel solution neutralisation circuits and precipitates ferric iron and other impurities, to produce a solution suitable for recovery of nickel and cobalt as a clean mixed sulphide intermediate product in the sulphide precipitation
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step. Additional heat supplied to the circuit from the ore leach flash circuit optimises the efficiency of operation. The solids are separated from the neutral solution, a portion of the solids are recycled to the neutralisation circuit as seed material, while the balance is sent to the CCD circuit for recovery of soluble and co-precipitated nickel and cobalt.
17.2.5 Sulphide Precipitation Product liquor from raw liquor neutralisation is mixed with nickel sulphide seed material and then contacted with fresh and recycled hydrogen sulphide in two parallel medium-temperature and pressure circuits. Reactions occurring within the circuits precipitate nickel and cobalt as mixed sulphide solids, leaving a solution that is largely free of nickel and cobalt. The product slurry is flashed to sub-atmospheric pressure, the flash gas is cooled and the steam condensed. The resulting gas, principally hydrogen sulphide, is recycled to the precipitation circuit, while the degassed slurry is forwarded to the thickener for recovery of the product mixed sulphide solids. The product solids are recovered and washed in a two- stage countercurrent wash circuit. The use of seed material minimises the scaling of reactor vessels and piping with mixed sulphides while the use of parallel circuits from feed vessels through to the thickeners allows continued operation of the Processing Plant while one circuit is being descaled. Seed material is prepared from a blend of fine precipitated sulphides and ground sulphides.
17.2.6 Tailings Handling The tailings from the CCD circuit are principally leach residue, with some gypsum and precipitated metal hydroxides. The tailings and sulphide precipitation discharge solution are further treated in a two-stage neutralisation circuit to remove remaining soluble metals such as aluminum, iron, nickel, cobalt, zinc, copper, and manganese. A portion of the magnesium is also precipitated. The resulting tailings slurry is forwarded to the TMF for permanent disposal. Due to local net precipitation and the overall plant water balance, a neutral water stream is decanted from the TMF; a portion of water is used in the Processing Plant while the balance is discarded via ocean disposal.
17.3 Nickel and Cobalt Refinery The mixed sulphides product from the leach plant are converted to metal in the refinery which is located at the Plant Site. All the steps of the refining process have been successfully commercialised in other facilities and equipment has been proven at the Ambatovy Project. Sherritt has significant experience in the design and operation of each step.
17.3.1 Sulphide Leach and Impurity Removal The mixed sulphides are leached in water using an overpressure of oxygen, resulting in a high strength nickel-cobalt sulphate solution with minor amounts of zinc, copper, and iron. The leaching process consists of two stages, with the unleached solids from the first stage progressing to the second stage. Second-stage leach residue is recycled to back to the second stage leach, and the iron-containing tailings bled from the circuit. This flow results in very high net extractions of nickel and cobalt while minimising autoclave volume and the risk associated with single-train operation. The combined leach solution is treated with ammonia to raise the pH and precipitate soluble iron, which is then separated from the solution with a thickener and pressure filters. The iron residue, which contains small amounts of co-precipitated nickel and cobalt, is recycled to the pressure acid leach stage of the leach plant. The iron-free solution is contacted with zinc sulphide to remove copper to trace levels. The resulting zinc-copper sulphide is separated from the solution and either sold or stockpiled. The solution is polish- filtered before being forwarded to the solvent extraction circuit.
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17.3.2 Solvent Extraction and Zinc Precipitation The zinc and cobalt are removed from the nickel solution in successive stages of solvent extraction with Ionquest 290 extractant. The zinc is removed first, with extraction at a pH of approximately 3.5. The zinc is loaded onto the organic extractant, which is separated from the zinc-free solution. The loaded organic is then stripped with dilute sulphuric acid. The zinc is precipitated from the high-strength zinc strip solution using hydrogen sulphide. A portion of the resulting zinc sulphide is recycled to the copper cementation stage, and the balance is sold. Cobalt is removed from the zinc-free solution via extraction into an organic phase at a pH of approximately 5.5. The loaded organic is scrubbed with dilute cobalt solution to remove entrained nickel solution and co-extracted nickel and is then stripped with dilute sulphuric acid. The high-strength cobalt strip solution is treated with zinc sulphide to remove trace amounts of copper, if needed, and then reports to the cobalt reduction area.
17.3.3 Nickel and Cobalt Reduction The separate nickel and cobalt solutions are treated by the proven hydrogen reduction process to produce pure metal powder products. The nickel and cobalt sulphate solutions are adjusted with addition of ammonia and ammonium sulphate, and then contacted with hydrogen at elevated temperature and pressure to precipitate the metals. The resulting slurries are flashed to atmospheric pressure and the metal powders are separated from the solution, washed, and dried. The powders are then sold as is, or briquetted and sintered in a reducing atmosphere furnace to produce high-quality nickel and cobalt briquettes.
17.3.4 End Solution Treatment The end solutions from nickel and cobalt reduction are treated with hydrogen sulphide to precipitate the trace levels of metals remaining in solution, which are then recycled as a mixed sulphide to the sulphide leach. Ammonium sulphate, a fertiliser, is then crystallized from the metals-free solution in a triple effect vacuum crystallization plant.
17.3.5 Utilities and Major Materials The mine and OPP use diesel-fired electrical generators set to produce power, while the Processing Plant uses coal-fired boilers to produce power and steam needed for the process. At full production coal consumption is estimated at 430,000 t per year. Water for the mine and OPP is sourced from the nearby Mangoro River and water for the Plant Site is sourced from the Ivondro River. At full production rates, the water requirement for the mine is estimated at approximately 4,000,000 t per year and the water requirement for the Plant Site is estimated at 8,000,000 t per year. Major bulk materials required for the process plant are sulphur and limestone. It is estimated that at full production the Processing Plant will require 1,700,000 t per year of limestone and 690,000 t per year of sulphur. Please refer to Section 18 of this report for further information on equipment, plant design and specifications.
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18 Project Infrastructure
18.1 Mine Site
18.1.1 Roads An 11 km access road has been built from the national road RN44. The access road is gravelled and is 9 m wide. It is maintained on a regular basis by a contract service crew. Several haul roads have been built to link the Ambatovy and Analamay pits to the OPP, to the waste dumps (Ambatovy and Analamay) and to the low-grade stockpile. The roads are gravelled and are 25 m wide.
18.1.2 Camps Three camps have been erected on the Property including the original exploration camp. A construction camp was built to lodge expatriate workers during the construction. Itcould accommodate up to 600 residents. It included a clinic, the training center (including the mine heavy and light vehicle driving simulator), a recreational building with related infrastructure and services. As the exploration and construction camps are built on top of the ore deposit, they will be dismantled in the future. A permanent operation camp has been constructed that can accommodate 100 residents. It includes a kitchen, dining room, and recreational room. Expatriates are lodged in this camp. National senior staff employees have the choice to live in the camp during their working days and commute during their time off or to live in Moramanga with their family and to travel every day from and to town. Potable water and waste water treatment plants have been installed just outside the operation camp area. An emergency fire storage reservoir has also been installed in the camp vicinity. A pipeline network links each main infrastructure in case of a fire.
18.1.3 Offices and Workshop Mine offices are provided for all administration and supervisory personnel for the mine and OPP. Space is provided for training employees, including inductions. The administration building is located next to the operation camp. The main part of the lower floor includes the warehouse facility. The washing, servicing and repair of the mining fleet are carried out in an area workshop with covered bays and a 6-t overhead travelling crane. This building also includes space for the following facilities: • Administration offices for mobile equipment maintenance personnel • Large truck workshop preventative maintenance bay • Tool store and parts storage area • Four containerised high bay workshops with interjoining office and parts storage • Four containerised workshop bays with 5-t overhead travelling cranes • Bulk lube storage and distribution with pumps and filtration to all workshops • Male toilets and washroom block, female toilet and washroom block, two locker room blocks and one shower block • Heavy equipment wash bay • Tyre storage, changing and repair workshop • Hydraulic hose fabrication workshop • Oil analysis laboratory • Two medium equipment workshop bays with 5-t overhead travelling crane • Light vehicle workshop.
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Apart from the main mobile equipment workshop, a facility for mechanical and electrical maintenance of the fixed equipment has been constructed.
18.1.4 Ore Preparation Plant The OPP is built approximately 1,000 m north of the mine infrastructure to minimise dust and noise at the infrastructure complex. The laboratory and slurry pumping station for feeding the slurry pipeline are located in the OPP area. The laboratory at the mine site is a stand-alone laboratory, which is supported by the main laboratory at the Processing Plant. It is linked into the data management system and is responsible for exploration samples, mine planning and production samples, samples in and around the OPP and water quality samples. The laboratory is located near the OPP.
18.1.5 Stockpiles The ROM stockpiles are located close to the ore preparation feed site. The stockpiles hold blended ore, ready to feed the OPP. Additional stockpile areas are located approximately 500 m south of the OPP feed site. The low-grade stockpile is located in a valley just south of the OPP. An under-drain has been installed at the bottom of the valley to allow groundwater base flows to continue to run down the valley without impacting the stability of the stockpiles.
18.1.6 Dumps The external waste dump for Ambatovy is located in a valley west of the Ambatovy West pit. An under- drain has been installed at the bottom of the valley. The external waste dump for the Analamay Deposit is located in a valley west of the Analamay Phase 1 pit. An under-drain has been installed at the bottom of the valley. Both external waste dumps are planned to be utilised until initial phases of the pits are completed, these will then be backfilled with the waste.
18.1.7 Mine Runoff Control Dams The area receives a large amount of rainfall. Mining, which is completed by open pit methods, therefore exposes disturbed and undisturbed lateritic materials to rainfall and runoff. This generates sedimentation, which must be controlled as part of the overall mining process. The Property is located along the surface water divide between the Mangoro and Vohitra rivers and contains the headwaters of six watersheds. The watersheds include the Sahaviara, Antsahalava, and Ankajathat that drain west towards the Mangoro River, and the Sahamarinana, Torotorofotsy, and Sakalava that drain east towards the Vohitra River. The slopes around the plateau consist of weathered granular ferricrete. A mine runoff control plan has been implemented to address the issue of sedimentation and potential environmental impact. A main component of the mine runoff control plan is the construction of runoff collection ponds downstream of the disturbed areas within the affected watersheds. Mine runoff is collected at these ponds to facilitate removal of sediments prior to discharge to the environment.
18.1.8 Pumping Station The Property’s water needs rely on Mangoro river water which is pumped to the OPP. The diesel high lift pumps transfer the untreated water to a lifting station located downstream of the Ambatovy North runoff collection pond. Electric pumps send the water to the OPP process water pond.
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Typically, all water used in the OPP (i.e. process water) reports to the ore slurry pipeline for transfer to the Processing Plant. Water users on the Property include the OPP, various process utilities, firewater and potable water.
18.1.9 Power Supply Nine diesel generators, each rated at 2.2 MW, have been installed for the requirements of the OPP and mine site in the power house located west of the OPP. In normal conditions, five generators operate with three on standby and one is under maintenance. The generators are presently designed to operate on light diesel. At the Property, the bulk power distribution is at 6.6 kV with large electrical requirements (in general 185 kW motors and larger) and 400 V for smaller electrical loads. Uninterrupted power supply is provided for the control systems where required. All necessary step-down transformers, distribution switchboards, motor control centres, small power and lighting installation and grounding are installed.
18.1.10 Pipeline Ore is slurried at the OPP located between the two ore bodies. The slurry is pumped from the OPP to the Processing Plant through a 600 mm diameter pipeline that is approximately 220 km in length.
18.2 Plant Site and Toamasina Area
18.2.1 Steam and Power A coal fired steam and power facility supplies power to the Plant Site and the construction and operations camps, and provides steam used in the Processing Plant. The steam and power facility consist of three fluidised bed boilers and three steam turbine generators, of which two boilers and two steam turbine generators are required during full production. The steam and power facility can export electrical power and high, medium and low-pressure steam to the Processing Plant. Each of the three turbine generators is capable of producing 45 MW of power. Peak electrical demand is estimated at 80 MW, with average demand of 65 MW at full production rate. The usable steam capacity of each powerhouse boiler is approximately 180 t/hr. Superheated high pressure steam from the power facility boilers is combined with superheated high-pressure steam from the acid plant waste heat boilers for use as process steam and turbine inlet steam. Majority of the superheated high-pressure steam is directed to the turbines for medium and low-pressure steam production and power generation. Letdown stations are also available for medium and low-pressure steam production should the turbines be unavailable. The remaining superheated high-pressure steam is cooled and consumed in the PAL circuit as a direct heating source. Medium-pressure steam extracted from the turbines at 3,600 kPa (abs) is combined with medium- pressure steam from the hydrogen plant waste heat boiler for use in various indirect heating processes in the refinery. Low-pressure steam is also extracted from the turbine at 400 kPa (abs) for direct and indirect use as process steam in the sulphide precipitation circuit, the metals refinery and boiler feed water preheating. The power facility receives demineralised water from the demineralised water plant, cooling water from the process water plant, and coal, diesel, limestone and sand from fuel storage and bulk materials storage yards.
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The power facility design is based on two de-aerators serving the operating boilers in the power plant and sulphuric acid plant. Ash generated by the power facility boilers is crushed and transported by truck to a disposal area. Thirty MW of supplemental diesel generators have been installed and will provide an additional source of electrical power during periods when the steam turbine generators are off line. The supplemental power supply can also be used for starting the steam and power facility.
18.2.2 Emergency Power Supply In case of a sudden plant-wide loss of electric power during Processing Plant operation, emergency diesel generators are provided as required for safety, and otherwise only to maintain the process equipment ready for quick resumption of production when normal power is restored. If required, the emergency power generators can also be used to start the steam and power facility. Ten emergency power diesel generators are on site, each generator has a generating capacity of 1.0 MW.
18.2.3 Water Supply The Ivondro River pump station provides all the necessary raw water requirements for the Processing Plant, taking into consideration the reclaim and tailings pond return waters available from the ore thickener and tailings pond, respectively. The untreated water is treated at the process water plant, to produce process water, demineralised water and potable water. A set of three pumps transfer water from the Ivondro River to the plant site. Diesel-fired generators that supply electricity to the pumps are also located at the pump station.
18.2.4 Petroleum Products, Supply and Storage The Processing Plant requires naphtha, diesel fuel and liquefiedpetroleum gas (LPG). Naphtha is delivered by dedicated pumps from the nearby oil terminal, owned and operated by a third party, to two-day tanks located at the plant site. An underground pipe connects the terminal with the day tanks. Diesel fuel is delivered to the plant site by trucks and unloaded to day tanks. From the day tanks, the fuel is delivered to the trucks and various vehicles by a fuel dispensing station. The naphtha and diesel fuel tanks are provided with flame arrestors and located in bounded areas sized for 110% of the storage capacity in the event of a leak. Hose reels and fire extinguishers are installed next to the bounded areas. LPG is delivered in trucks to bullets located on the Plant Site.
18.2.5 Bulk Materials Supply and Stockpiling Coal, sulphur and limestone arrive by ocean vessels and are unloaded at the port before being transported to the plant by train. At the plant, these bulk materials are discharged and then stockpiled by a stacker and reclaimed by front-end loaders, discharging into hoppers and feeding belt conveyors which deliver the materials to their respective end use. The stockpiles have a capacity of 60,000 t for coal, 75,000 t for sulphur and 130,000 t for limestone. These capacities were established based on daily consumption estimates and supply storage of 50 days for coal, 40 days for sulphur and 27 days for limestone. The rainwater runoff from the stockpiles area is contained in runoff ponds.
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18.2.6 Ammonia Anhydrous ammonia is imported as a liquid at minus 33°C in refrigerated tank ships and delivered to an off-site storage tank located near the oil terminal. The off-site storage tank facility is supplied electricity from the plant power facility, and it is also equipped with its own diesel fired electrical generator for backup power. The storage facility also includes a refrigeration system with redundant capacity, transfer pumps, water deluge system and control room. Pipelines connect the port ammonia unloading terminal to the storage tank and the storage tank to the on-site ammonia storage bullets.
18.2.7 Tailings Disposal The TMF has been designed according to the Equator Principles and to other more specific international standards, as set out by the Canadian Dam Association, International Commission on Large Dams, and the Mining Association of Canada. The tailings management area comprises two adjoining basins and covers an area of approximately 760 ha. Thirteen dams have been constructed to an elevation of 46 m which has sufficient capacity to sustain operations through to 2021. Stage 3 of the TMF construction plan has commenced to perform additional raises to provide sufficient storage capacity for the remaining mine life. Water management in the tailings basins is an essential consideration in such a high rainfall region. The water management plan involves containing supernatant and runoff so that it can be pumped back to the plant site, where a portion will be re-used in the process. The excess water will be decanted and pumped for discharge to the ocean in accordance with discharge parameters agreed for the Ambatovy Project. The effluent, which is discharged to the ocean, consists of neutralised process solutions that have been diluted by the high rainfall in the area. The design of the tailings pond provides for neutralisation of the tailings slurry with limestone and lime to remove iron, aluminum, nickel, cobalt, zinc, copper, chrome, the bulk of the manganese, and a portion of the magnesium prior to discharge to tailings storage and plans for dilution by rainfall. Initial outfall studies by marine experts and over five years of operation have both shown that this process effluent has no adverse effects on the local environment. The tailings area will be progressively reclaimed through the life of the Processing Plant. This will likely require a layer of fill on the tailings surface for stabilisation, prior to the establishment of vegetation. Topsoil and organics, stockpiled during development of the TMF, will be used for reclamation activities.
18.2.8 Air Strip An 883 m long and 20 m wide airstrip has been constructed that will allow private aircraft access to the Plant Site.
18.2.9 Port All material imports and exports are done through the port of Toamasina. For this, Mole B has been expanded to meet the Ambatovy Project requirements. The Mole B extension is 265 m long and 32 m wide with the addition of a mooring dolphin, projecting approximately 350 m beyond the north end of the extension. The structure is equipped with fenders and bollards suitable for 50,000 DWT class bulk carriers. A dedicated ammonia offloading terminal has been included in the Mole B expansion.
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18.2.10 Road and Rail A road link, approximately eight km long, from Toamasina to the Processing Plant, has been built for the construction and operation phases of the Ambatovy Project. The road is suitable for transporting Ambatovy Project loads, including the approximately 800-t loads during autoclave transportation. A rail spur, approximately 2 km long, has been constructed to connect the existing rail line, which has been double-tracked, to the plant site. This will facilitate movement of commodities by rail between the port and the Plant Site.
18.2.11 Camps and Housing A construction camp was erected in order to lodge workers during the construction. It accommodated up to 5,000 residents. It includes a clinic, a recreational building with related infrastructure and services. After the construction, this camp has been used to accommodate some of the operations work force as well as contractors. It has also been renovated to include facilities for staff and operations’ training programs. A permanent operations camp has been constructed to house expatriate operations personnel and can accommodate approximately 150 people. It includes a kitchen, dining room, swimming pool and recreational room. Residential expatriate employees are accommodated in the housing village which consists ofa combination of houses and duplexes. The village also includes a school, clinic, recreational room and swimming pool.
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19 Market Studies and Contracts
19.1 Market Studies
19.1.1 Nickel In recent years, the worldwide nickel market price experienced a continued decline as global production has exceeded demand with significant growth in low grade ferronickel, more commonly referred to as nickel pig iron (NPI). However, in the second half of 2017, nickel prices increased due to, among other factors, strong stainless steel production in China and optimism surrounding increased nickel demand from electric vehicles. Nickel prices on the LME were higher in 2017 than in 2016. The LME average cash settlement price for 2017 was US$4.72/lb, an 8% increase from the 2016 average of US$4.36/lb. Nickel opened 2017 at US$$4.63/lb and closed the year at US$5.47/lb and traded in a range between US$3.95/lb and US$5.82/lb. Nickel is a heavy silver-coloured metal whose principal economic value lies in its resistance to corrosion and oxidation and excellent strength and toughness at high temperatures. Nickel is used in the production of stainless steel, which accounts for approximately two thirds of worldwide nickel consumption. In 2016, the battery sector accounted for only 4.4% of total nickel demand, according to CRU International Limited (CRU), a leading provider of market analysis in the mining and metals industry. Nickel is also used in the production of industrial materials, including non-ferrous steels, alloy steels, plated goods, catalysts and chemicals. In 2017, approximately 88% of world primary nickel production was consumed in North America, Europe, Japan and China. Nickel demand is strongly influenced by world macro-economic conditions, which in turn influence the state of the world stainless steel industry, the single largest consumer of nickel. According to CRU, in 2017, Vale S.A., a Brazilian company, was the world’s largest producer of refined nickel. MMC Norilsk Nickel, a Russian company and Glencore, a Swiss company were the second and third largest producers, respectively. Production from the Ambatovy Joint Venture was 35,474 t or approximately 1.7% of annual world refined nickel production. Current world supply of refined nickel is estimated to be approximately 2.035 Mt/a. World nickel supply is broadly classified into primary and secondary nickel. Primary nickel is further subdivided into refined nickel (Class I) having a minimum nickel content of 99%, and charge nickel (Class II) having a nickel content of less than 99%. The main physical forms of Class I nickel are electrolytic nickel (cathode and rondelles), pellets, briquettes, granules and powder. Class II nickel includes ferronickel, nickel oxide sinter and utility nickel. Secondary nickel is the nickel contained in scrap metal, principally stainless steel scrap. World nickel supply has also been impacted by the growth of NPI in China. NPI is the lowest purity of what is considered refined nickel (as low as 2% nickel content) and is primarily used in China to make stainless steel. CRU estimates that NPI production in China was approximately 397,000 t of nickel equivalent in 2017 and an additional 185,000 t was produced in Indonesia. Total NPI production has been reported to have increased by approximately 117,000 t in 2017, making 2017 a new record year for world NPI production. Most major refined nickel producers supply nickel at grades ranging from 98.4% to 99.9% in purity. The Ambatovy Joint Venture’s sintered nickel briquettes, produced at a minimum of 99.8% purity, are well suited for stainless steel and alloy steel production and battery chemical applications, and are expected to continue to be sold to such industries.
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19.1.2 Cobalt Cobalt is a hard, lustrous, grey metal that is used in the production of high temperature, wear-resistant super alloys, catalysts, paint dryers, cemented carbides, magnetic alloys, pigments, rechargeable batteries and chemicals. The cobalt market is much smaller and more specialised than the nickel market. The cobalt market has been subject to significant price volatility due to the lack of a liquid terminal market. The LME introduced a 99.3% cobalt contract in February 2010 and in January 2017 announced that it was increasing the minimum purity to 99.8% to assist in contract adoption. Cobalt contract trading volumes were up 81% in 2017 over 2016 reflecting increased interest in the LME cobalt contract. The LME reported that 14,261 t of cobalt traded on the LME in 2017, representing approximately 52% of global refined metal production or 42% of total refined metal and chemical production. Other base metal contracts on the LME experience trading volumes of 50 or more times total production indicating that the LME cobalt contract is still in its infancy and remains a secondary pricing mechanism to the more widely accepted Metal Bulletin, as discussed below. Cobalt supply has evolved over the years from a reliance on unstable output associated with copper production in central Africa, to more diverse supply sources with material coming from a wider geographic area. Refined mainly as a by product of nickel and copper mining, approximately 64% of cobalt global production is processed through copper refining and 35% through nickel refining. The “copper belt” located in the Democratic Republic of the Congo (DRC) contains close to half of the world’s cobalt reserves. Australia, Cuba, Zambia, Madagascar, New Caledonia, Canada, Russia and Brazil hold most of the remainder. Cobalt production does not respond to cobalt demand. In the longer term, significant increases in supply are planned to be brought on stream from new large-scale international projects targeting copper production. The Ambatovy Joint Venture produces finished cobalt (briquettes and powder) at 99.9% purity, which exceeds the current LME specification. Based on data from CRU, worldwide supply of primary cobalt for 2018 is estimated to be approximately 131,572 t, an increase of approximately 13% from 2017. Ambatovy is among the leading suppliers of metallic cobalt to world markets. In 2017, cobalt was produced by 10 Cobalt Development Institute (CDI) member companies, with additional supplies coming from a variety of other companies. The non-CDI sources included individual companies such as Norilsk in Russia, as well as production from multiple refiners in China. Ambatovy’s operations supplied 3,053 t or approximately 2.7% of world primary cobalt in 2017. The relative importance of the different uses of cobalt has changed over the years, with demand for older, more established uses, such as pigment, magnets and carbides showing only modest, if any, growth over the period. Many of these traditional uses are strongly reliant on industrial growth for demand increases, so demand for these uses tends to rise and fall with global economic performance. Over the last decade, growth in the chemical sector (primarily in battery chemicals) has increased the demand for cobalt. The world’s reliance on global communications in the form of mobile phones and tablet technology has been a driving force for increased cobalt consumption. Strong recovery from the superalloy sector has also helped the market remain in relative balance. Over the long term, growth is expected in the rechargeable battery sector (hybrid and electric vehicle applications) and coal to liquid and gas to liquid catalyst sectors. The Metal Bulletin low-grade average cobalt price rose strongly during the year starting at US$14.65/lb and closing the year at US$36.00/lb. In 2017, low-grade average cobalt was quoted by the Metal Bulletin in a range between US$14.30/lb and US$37.00/lb, averaging US$26.53/lb (low-grade high/low year average), 125% higher than the average price for 2016 of US$11.78/lb. In 2017, the LME daily cash settlement price averaged US$25.39/pound with a low of US$14.97/lb and a high of US$34.25/lb. The LME price is considered the lowest openly traded market price for metallic cobalt meeting a minimum purity level of 99.3% with limited specifications for impurities.
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19.2 Contracts There are four primary operation and maintenance contracts and six primary commodity contracts considered as essential for the Ambatovy Project: • The Ambatovy Joint Venture has insourced the operation and maintenance of the Hydrogen Plant and the Air Separation Plant which produces oxygen and nitrogen (located at the Processing Plant) in September 2018. The O&M Contract with Air Liquide Grande Industrie Madagascar was terminated on 30 September 2018. • The Ambatovy Joint Venture has an agreement with Korea Plant Services & Engineering Co. Ltd for the operation and maintenance of the Power Plant and Water Treatment Plant. • The Ambatovy Joint Venture has an agreement with Madarail SA regarding the operation and maintenance of Ambatovy Project trains between the Plant Site and port. Each of the abovementioned contracts is in place and operational. The terms, rates and/or charges contained in such contracts are within industry norms. The Ambatovy Joint Venture has the following major commodity contracts with: • Sumitomo Corporation (Johannesburg) regarding the supply of steam coal • Sumitomo Corporation (Middle East) regarding the supply of limestone • Trammo AG regarding the supply of sulphur • Société Malgache des Petroles Vivo Energy SA (Vivo) regarding the supply of fuel (for AMSA and DMSA). The Ambatovy Joint Venture has agreed to sell to KORES, Sumitomo and Sherritt, and each of these buyers have agreed to purchase from the Ambatovy Joint Venture, the nickel metal production. In turn, KORES, Sumitomo and Sherritt have been authorised to undertake efforts to promote, solicit orders and negotiate sales of Ambatovy Project nickel. This agreement is for the life of the Ambatovy Project. An offtake agreement of this type is required for project eligibility under the LGIM. DMSA has entered into marketing and distribution agreements with Darton Commodities Limited (Darton) and Phoenixx International Resources LP (Phoenixx) under which Dartonand Phoenixx have agreed to act as the exclusive distributor of the Ambatovy Project’s cobalt products within their respective defined territories. Each of these contracts, as amended, is fully operational and contains terms, rates and charges which are within industry norms. It is envisaged that Sumitomo will take on increased responsibility for metal product distribution in mid-2019 as current contracts come to completion. DMSA has also entered into an agency agreement with International Raw Materials Ltd to be the exclusive agent for Ambatovy ammonium sulphate fertiliser, a by-product of the refining process. In addition to the contracts described above, DMSA has entered into several minor contracts with arm’s length third parties in relation to transportation, handling, sales, and materials and other services of this nature in accordance with the industry norms.
19.3 QP Comments As of the Effective Date, the QP responsible for this section, Glen Smith, has reviewed the market studies and contracts discussed in Sections 19.1 and is of the opinion that the results support the assumptions in the technical report. The assumptions are further supported by the Ambatovy Project’s production and sales results since entering production several years ago. The QP is also of the opinion that the contracts in place as discussed in Section 19.2 are within industry norms with respect to terms, rates or charges.
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20 Environmental Studies, Permitting, and Social or Community Impact
20.1 Summary of Environmental Studies and Challenges The principal environmental law in Madagascar is the Environmental Charter (law 90-033), Article 10 which requires an Environmental Impact Assessment (EIA) for investment projects as specified in Decree No. 99-954 as amended by Decree No. 2004-167 of 3 February 2004. Environmental and social performance of Ambatovy is regulated through a series of officially approved Environmental Management Plans (PDGESs) and subject to regular inspections coordinated by the Ministry of the Environment (Office National del’Environnement, ONE) with representatives from relevant government ministries. In terms of this, the company prepares an annual report that includes the results of all relevant monitoring programs. The terms of reference for the EIA were initially developed by DMSA in consultation with Golder Associates (Golder) and submitted to the ONE for review. After discussions and negotiations between Dynatec and ONE, the terms of reference were modified and approved by ONE. The final EIA was submitted in January 2006 (Dynatec Corporation, 2006) and was the subject of an extended evaluation process piloted by a multi-ministry technical committee. The evaluation included a series of public consultations involving communities, NGOs and national institutions. Studies conducted for the EIA confirmed the high biodiversity of the proposed mining site, most of which was covered in mid-altitude primary forest growing on a ferralitic substrate. The forest over the Deposits has a distinctive appearance, structure and floral composition and was described in the EIA as “azonal” forest or thicket to distinguish it from the surrounding “zonal” forest (Goodman & Raselimanana, 2010). The mine area also includes sensitive aquatic habitats (Rall et al., 2010). Subsequent scientific studies have confirmed the high biodiversity of the surrounding forest, which is now known to contain 12 species of lemurs, 38 species of other terrestrial mammals, 117 species of birds, 80 species of amphibians and 59 species of reptiles and over 1580 plant species (Goodman & Mass, 2010). Of these species, over 37 are listed as either critical or endangered by the International Union for Conservation of Nature (IUCN), including eight critically endangered species. A further 52 species were recently assessed in the IUCN Red List as critical (4) and/or endangered (48). The aquatic habitats support five endemic fish Evolutionary Significant Units (ESUs) that are currently being taxonomically described and have the potential to be new to science. The Ambatovy Joint Venture is treating these ESUs as if they had an IUCN Endangered status. To address this considerable challenge and meet regulator and lender requirements including International Finance Corporation (IFC) Standard 6 (IFC 2006), the Ambatovy Joint Venture operates a “no net loss” biodiversity strategy based on measures to avoid or minimise impacts and uses offsets to compensate for unavoidable biodiversity losses (Dickinson & Berner, 2010). The mine footprint has been kept to a minimum and is regulated by the mine’s environmental management plan (Ambatovy Project 2011a). Mitigation measures include paced directional clearing and species salvage (for both animals and plants), relocation and ex-situ preservation (for fish and frogs) and maintaining ecological connectivity (Mass et al., 2011). As part of the offsets program, two patches of azonal forest on the eastern margins of the ore bodies totalling approximately 300 ha have been set aside for biodiversity conservation (Ambatovy Project 2009a).
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20.2 Waste and Tailings Disposal, Site Monitoring and Water Management
20.2.1 Waste Disposal Malagasy environmental regulations require adequate measures to minimise and manage waste. The principal wastes generated by the Ambatovy Project include waste materials from construction, packaging and scrap, used oils, ash and other by-products, overburden from mining operations, tailings generated by the Processing Plant and domestic waste and sewage (Dynatec Corporation, 2006). The mine operates a waste collection scheme and is equipped with a waste sorting area and incinerator. Composting facilities have been developed for vegetable wastes. Reusable materials such as wood are collected and distributed to local communities. Timber from deforestation operations are distributed to local communities under supervision of the regional forestry authorities. Other products of forest clearing (topsoil, plant debris) are processed and stockpiled for future mine rehabilitation actions (Ambatovy Project 2011a).
20.3 Site Monitoring Regular surveys are conducted along the pipeline and the results are reported to the ONE. Management of the right-of-way, a road located above the buried pipeline that connects the mine and the plant site, is focused on control of erosion points. In 2012, air quality monitors were installed within the perimeter of the Property and Plant Site and now provide continuous monitoring of the ambient air. During 2013, the information received remotely from the new and pre-existing air monitoring equipment was integrated into a display showing real time air quality and meteorological data which is available in a number of key locations in the Plant Site. Preset alarms are triggered if an event that impacts air quality occurs. Water is extracted from the Mangoro River for the mine operations and from the Ivondro River for use in processing operations.
20.4 Water Management Water management is addressed in the specific management plans relating to the Ambatovy Project’s components (mine, pipeline, Plant Site, TMF and port extension). At the mine, sedimentation dams are constructed to prevent release of sediments from the mining area into local watercourses. A total of seven dams will be built during the LOM, five of which have been completed and a sixth is nearing completion. The dams are equipped with spillways as discharge structures. The dams are constructed to standards consistent with the requirements of the IFC. Regular monitoring of suspended solids and other water quality parameters is carried out downstream of the dams (Ambatovy Project, 2011a).
20.5 Tailings Disposal Owing to the Ambatovy Project’s design, which uses a pipeline to transport ore slurry to the Processing Plant, tailings are generated at the plant site. After nickel/cobalt extraction, the neutralised residues are pumped in slurry form to a TMF. The TMF has a spillway for the safe discharge of excess water in extreme storm events. However, water levels are controlled by decanting clear, mineralised water to an ocean outfall. This has a diffuser system extending 1.5 km out to sea and is designed to improve diffusion and mixing and to minimise impacts to ocean water quality. A connection to the Plant Site is also maintained, allowing for the recycling of water to the Plant Site for water conservation (Ambatovy Project, 2011c). The TMF will be developed over three distinct phases over the Ambatovy Project’s life and will ultimately hold an estimated 220 Mt of residues, occupying an area not exceeding 773 ha, including the largely undisturbed buffer zone, and built to a height of 72 m above mean sea level. Upon plant closure, the reclaiming tailings facility will be re-vegetated, and a trafficable surface will be created.
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20.6 Environmental Permits Environmental permit No. 47/06/MINENVEF/ONE/DG/PE was issued on 1 December 2006 to AMSA/DMSA conditional upon implementation of the ESMP (Ambatovy Project, 2006) and subject to an annual review process. The ESMP made provision for the development of specific management plans relating to project components (mine, pipeline, Plant Site, TMF and port extension), environmental thematic plans (air, water, soil, biodiversity) and other specific plans (village relocation, species specific plans, mine closure etc) and for the renewal of plans upon transition from the construction to operation phase of the Ambatovy Project. In addition to the environmental permit, the Ambatovy Project is obliged to obtain forest cutting permits for each parcel cleared. Permits are also required and have been obtained for water extraction and discharge, construction and the installation of a helipad and runway. In connection with the granting of the required authorisation to commence operations at its Processing Plant, the Ambatovy Joint Venture agreed work with the Malagasy government to establish a US$50 million environmental surety which could be accessed to correct any material breach of the applicable environmental laws, regulations and permits governing its environmental obligations, including closure obligations, which the Ambatovy Project failed to remedy. Since the environmental permit was granted, the Ambatovy Project has completed nine evaluations. Originally, two six-monthly visits by the Technical Evaluation Committee (CTE) and one annual report submitted to ONE were completed each year. Since the completion of construction, evaluations have been completed annually, and there continues to be a comprehensive annual report submission. Environmental performance at the mine has been found by the CTE to be satisfactory every year so far. In addition, the Ambatovy Joint Venture holds periodic ad hoc meetings with the regulator (ONE) so as to anticipate and address problems as they arise. An external Scientific Consultative Committee made up of national and international experts meets every second year to review the Ambatovy Project’s performance on biodiversity management. Environmental management at the Property is adaptive and consists of applying the “mitigation hierarchy” (which includes impact avoidance, minimisation and, where necessary, compensation or offsetting) informed by regular monitoring of the physical and biological environment. Physical environmental monitoring includes water quality (total suspended solids and other parameters), air quality (dust and other parameters) and meteorological monitoring. Biological monitoring includes monitoring the populations and health of affected lemurs (Mass et al., 2010), small mammals, birds, reptiles, amphibians and fish (e.g. Ambatovy Project, 2011b). Biological management actions of the mitigation hierarchy include defining clearly and minimising the mine footprint, slow directional clearing of forest (accompanied by the salvage and relocation of plant species of concern and the less mobile vertebrate animals) and the establishment and management of conservation zones or “offsets” (Ambatovy Project, 2007). The offsets include about 3,300 ha of forest surrounding the footprint, two set-aside parcels of “azonal” forest amounting to approximately 300 ha growing over part of the ore body and active support to regional conservation initiatives (Torotorofotsy wetland, Analamay-Mantadia Forest Corridor) (Ramahavalisoa et al., in prep).
20.7 Social and Community Requirements Malagasy environmental legislation requires consideration of social impacts and remedial actions to mitigate and compensate such impacts. Social aspects are addressed in a Social Development Plan, which primarily addresses resettlement. For the mine, only one household required a physical relocation. A total of 29 were farming on the mine footprint and required an economic relocation to an adjacent watershed called Ambolomaro. Outside the mine, at Berarano, one small hamlet of about 10 houses was shifted
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100 m to avoid interaction with pipeline construction. International guidelines on resettlement issued by the IFC and the World Bank have been followed. Upon securing the Property, a census of dwellings and agricultural holdings was made and agreements entered into with occupants not to extend their existing agricultural activities in return for compensation for the loss of a right to seek title to the occupied land. Many signatories have breached their agreements and there have been new unauthorised activities within the Property, some of which are in conservation forest. A process is underway to limit further expansion and control illegal occupants. Mine production is unaffected, but action is required to ensure compliance with environmental commitments andto maintain the integrity of the conservation zone. Environmental management actions are accompanied by social support and compensation measures including awareness raising, establishment of community-based management areas for forested zones outside the mine conservation forests, construction of a by-pass road around the mine and promoting sustainable methods of revenue generation i.e. ( intensified rice culture, fish farming etc), village relocation and compensation for lost access to resources. The comprehensive social baseline surveys were updated in mid-2012. The data collected from this baseline has served as the basis for social programming around the mine, particularly concerning livelihood activities.
20.8 Mine Closure The Ambatovy Joint Venture is committed to rehabilitation of the Property following exploitation. A Reclamation and Closure Plan has been prepared by consultants and is in place for all aspects of the project (Knight & Piésold, 2018). Mined areas will be reclaimed and rehabilitated as mining proceeds. The cost of mine site remediation and reclamation is estimated at US$100 million.
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21 Capital and Operating Costs
21.1 Capital Costs IMC was provided annual capital expenditure plans by AMSA. These are for items such as major haul road constructions, sediment containment dams, workshops and workshop relocation and other similar capital type items. The capital costs included over the next five years are US$68 million, US$70 million, US$52 million, US$45 million and US$23 million and then US$20 million per year for the subsequent five years and then around US$14 million untilY ear 18. Note the cost of mining fleet is not included in capital costs – the cost of equipment has been amortised in the hourly operating cost over the life of each item of equipment.
21.2 Operating Costs The following are the estimated uninflated operating costs of the Ambatovy Project from 2018 to 2022. Table 48: Uninflated annual operating costs (2018–2022)
Estimated cost Percent of
(US$’000) total Variable cost 264,044 45% Fixed cost 324,111 55% - Operating materials 112,181 19% - Consulting and contractors 63,183 11% - Manpower cost 82,163 14% - Logistics 15,021 3% - Others 51,563 9% Total Operating Cost 588,155 100% Note: Royalty payment is not included in the Operating costs above.
21.2.1 Variable Costs Sulphur, limestone and coal – these three commodities have the greatest impact on the plant site variable costs. The consumption of sulphur and limestone are driven by the impurities within the ore. The consumption of coal is impacted by rain which causes “wet coal” and causes more coal to be consumed. The variable costs are based upon projected consumption multipled by the following prices: Table 49: Commodity price assumptions used in the cash flow model (average price 2018–2022)
Commodity Price Sulphur US$118/t Coal US$92/t Limestone US$20/t Ammonia US$406/t Naphtha US$672/t
21.2.2 Fixed Costs Operating materials are driven by the maintenance needs of the equipment. This includes both planned and unplanned maintenance in mine and plant area.
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Manpower costs are driven by headcount, are based upon the expected number of expatriates and national employees on the Ambatovy Project’s payroll. These costs include estimated salaries and wages, benefits, travel and other allowances, as applicable, for all employees as part of their remuneration package. There will be a localisation of workforce from 2019 onward. The number of nationals is planned to increase in 2019 and is assumed to be stable until 2029, whereas the number of expatriates will be reduced significantly from 2019 onward. Contractors costs are AMSA contractor costs predominately consist of headcount to maintain the equipment. DMSA contractor costs mainly consist of operation and maintenance of facility and process consulting. Ambatovy is aiming to optimise consulting and contractor costs by in-sourcing. The top 50 contracts will be reviewed in depth to achieve the cost optimisation. Other fixed costs are associated with support departments which have tended to be quite stable. The following is a breakdown of estimated total operating costs by area from 2018 to 2047. Table 50: Total operating cost (2018–2047, uninflated)
US$ M Mine 1,512 PAL 6,256 Refinery 4,789 Administration and Marketing etc 1,157 Total Operating Cost 13,714 Note: Royalty payment is not included in the operating costs above.
21.3 Mine Operating Costs IMC has developed a zero-based cost model for the Ambatovy operations based on a combination of AMSA supplied data, in-house database information and IMC’s nickel laterite experience. The cost model is a hierarchical model which means that all costs are built up from the bottom of a tree with maintenance costs/operator costs/parts and ownership costs being at the bottom of the tree. The next step up the cost tree is each piece of equipment that these costs feed into (e.g. a Cat 777 truck will have the four cost feeders below it), then it steps up to the next level (e.g. Primary Equipment), then up to the next level (e.g. Mine Operations) and ultimately to the top level. The cost of mining fleet is amortised into the operating costs per hour.
21.4 Equipment Operating Costs Every equipment item has been costed separately based on labour cost, consumables, maintenance manpower and maintenance costs. An example of the detailed costs for the primary excavator is shown in Figure 68.
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Figure 68: Equipment Operating Cost – primary excavator
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21.5 Administration and Overhead Costs An outline of the administration operating cost centres is set out in Table 51. The camp running costs are approximately $10 per person per day. Table 51: Mining officedministration a cost centres
Operating cost category Average cost per year (US$ k) Total Admin Costs $3,725
21.6 Operating Cost Summary On the basis of the above inputs, the operating cost up to feeding the OPP has been estimated for each year of operation. Note the cost centre“ Processing” is the OPP cost and does not include any of the costs of the PAL. As outlined above, these costs are estimated down to the level of consumables on each item of equipment. In order to provide a summary of this data for the Technical Report, the operating costs for each year of the mine life are summarised in Figure 69 and Figure 70 graphically and in detail for the first 10 years of operations( Table 52).
Figure 69: Cost by cost centre
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Figure 70: Mining Operating Costs per year by category
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Table 52: Operating cost by year
Year Name 1 2 3 4 5 6 7 8 9 10 Total Total 118 119 113 109 115 116 135 129 109 145 Admin Admin 46.1 40.9 38.3 37.3 37.2 35.9 35.8 34.4 34.3 33.9 Admin:Oheads Oheads 41.7 36.5 33.9 32.9 32.9 31.5 31.4 30.0 29.9 29.5 Admin:Expat Expat 3.9 3.9 3.9 3.9 3.9 3.9 3.9 3.9 3.9 3.9 Admin:Local Local 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 Processing Processing 2.6 2.8 2.9 3.1 3.1 3.1 3.9 3.3 3.4 3.9 Processing:OPP OPP 2.6 2.8 2.9 3.1 3.1 3.1 3.9 3.3 3.4 3.9 Mining_Ops Mining_Ops 59.9 67.7 64.0 60.5 67.6 68.5 84.4 81.4 61.4 95.5 Mining_Ops:Mine Mine 59.2 67.0 63.2 59.8 66.9 67.8 83.7 80.6 60.7 94.7 Mining_Ops:Mine:Mine_Support Mine_Support 4.9 5.0 5.0 4.8 4.8 4.9 5.0 5.0 4.9 5.0 Mining_Ops:Mine:Primary Primary 51.5 60.9 53.8 52.5 59.8 62.4 77.7 75.1 54.4 89.3 Mining_Ops:Mine:Primary:Ore_Direct Ore_Direct 27.5 34.3 23.7 25.2 31.2 36.1 31.2 30.4 23.4 28.7 Mining_Ops:Mine:Primary:Waste Waste 13.0 12.4 21.6 19.6 16.1 22.3 38.8 33.2 26.8 44.8 Mining_Ops:Mine:Primary:Ore_Stockpile Ore_Stockpile 11.0 14.2 8.6 7.7 12.5 4.0 7.7 11.5 4.3 15.8 Mining_Ops:Mine:Stockpile_Reclaim Stockpile_Reclaim 2.3 0.6 3.9 2.0 1.9 0.0 0.5 0.0 0.9 0.0 Mining_Ops:Mine:Mine_Maintenance Mine_Maintenance 0.4 0.4 0.5 0.5 0.5 0.4 0.5 0.5 0.5 0.4 Mining_Ops:Grade_Control Grade_Control 0.8 0.8 0.8 0.8 0.8 0.8 0.8 0.8 0.8 0.8 Enviro Enviro 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Enviro:Sed_Dam Sed_Dam 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Enviro:Rehab Rehab 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Enviro:Clearing Clearing 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Small_Fleet Small_Fleet 9.5 7.2 7.4 8.2 7.6 8.1 10.9 9.6 9.6 11.8 Small_Fleet:Quarry_Roads_Sheeting Quarry_Roads_Sheeting 6.5 7.2 7.4 8.2 7.6 8.1 10.9 9.6 9.6 11.8 Small_Fleet:Quarry_Civil Quarry_Civil 3.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
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22 Economic Analysis
As Sherritt is a“ producing issuer”, as defined in NI 43-101, the Ambatovy Project is in production and this Technical Report does not include a material expansion of current production, an economic analysis for the Ambatovy Project is not a requirement for this Technical Report.
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23 Adjacent Properties
The Authors are unaware of any significant exploration activity or results on immediately adjacent third- party mineral properties.
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24 Other Relevant Data and Information
24.1 Production Targets In this Report, the total material presented in the LOM schedule has been derived from the operational LOM plan. This LOM plan includes Inferred Resource material that is within the final pit – with the final pit having been developed on the basis of only Measured and Indicated Resources. The reasons that the LOM plan used these Inferred Resources include: • The cost and uncertainty in converting the Inferred to Indicated Mineral Resources is relatively low; however, there is no reason that AMSA should carry out the drilling required for conversion until shortly before mining in those areas is undertaken. • The 2018 Mineral Resource model has used a different approach to classification than the previous model. This has resulted in some areas that are in the short-term planning cycle (and some areas presently being mined) being changed from Indicated to Inferred. As this material is being mined and has been grade control drilled subsequent to completion of the 2018 Mineral Resource, it is reasonable to treat the material as a production target. None of these scheduled Inferred Mineral Resources have been included in the Mineral Reserve statement.
24.2 Political Environment Mining investment in Madagascar is regulated by the Mining Code and the LGIM. The Mining Code, which was updated in 1999, covers all aspects of mining except those specifically dealt with in the LGIM. The Mining Code sets out the conditions for both exploration and exploitation permits which must be applied for sequentially. The exploitation permit requires an environmental assessment that is issued by environmental regulatory authorities. The Ambatovy Joint Venture is in possession of all exploration and mining exploitation permits required for the Ambatovy Project. The LGIM, which was enacted in 2002 and developed with the support and assistance of the World Bank, establishes the legal framework for developing and operating large-scale resource projects in the country and provides the equivalent of a stability agreement for at least 25 years. The LGIM guarantees that the terms of a permit will not be changed after it has been granted, provides for legal stability, and provides investment incentives for qualifying projects. In 2007, the Ambatovy Project received notification from the Malagasy government of the Ambatovy Project’s eligibility certification under the LGIM. The Ambatovy Joint Venture continues to monitor the political climate in Madagascar and to engage in ongoing communication with representatives of the national, regional and local governments as well as multilateral institutions, key embassies and key Malagasy stakeholders. To date, there have been no material disruptions in activities at the Ambatovy Project in conjunction with any political uncertainties in Madagascar. The Ambatovy Joint Venture has active communication with relevant Ministers and officials of the Malagasy government and continues its engagement with multilateral institutions, key embassies and key Malagasy stakeholders.
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25 Interpretation and Conclusions
Based on the current identified Mineral Resources and Mineral Reserves and the assumed prices and parameters, the authors of this Technical Report have concluded that profitable operations can be sustained for approximately 26 years at the Ambatovy Project. Mining activities commenced in 2010 and the first finished nickel and cobalt briquettes were produced from the refinery in September 2012. Since January 2014, the Ambatovy Project has been officially in commercial production. This update to Mineral Reserve for Ambatovy reflects an evolution in the understanding of the resource, the mining strategies and the process drivers. The Mineral Resource model and the Mineral Reserve model will continue to evolve as the knowledge base expands in how to get the most economic benefit out of a complex mining and processing environment. The key for the success of the project is to continue to improve strategies that are informed by what the process plant is capable of processing on a short-term basis, and then have a systematic approach to implementing appropriate grade control and stockpiling strategies such that the appropriate feed is maintained to the process plant. Aligned to this is the requirement for the mine operations to have adequate mining fleet with spare capacity such that there is not a situation where they are unable to supply the PAL with appropriate feed. Like many operations around the world, the biggest challenge will be to convert the strategic optimised strategy into short term tactical strategies that result in alignment of short-term and long-term strategies. In the context of the above, the assumptions that underpin the mining operations in this Technical Report has allowed for a five-year period in which the operations can continue to improve. IMC’s opinion is that the Mineral Reserve presented in this Technical Report is a reasonable estimate on the basis of information available to IMC at this time.
25.1 Upside Potential CSA Global considers the following points to represent upside potential to the Ambatovy Project’s financial assessment: • Conversion of Inferred Resources to higher classifications and subsequent inclusion in mine planning. • Discovery and evaluation of new Mineral Resources and Mineral Reserves. • Nickel or cobalt grade may locally be higher than modelled once mining takes place, since the grades from high-grade drill intercepts were smoothed during the geostatistical interpolation process. The continuity of high-grade intersections is unknown but may offer flexibility and opportunities during mining. • Expansion of existing deposits both laterally and vertically. Recommended actions and opportunities for improving project value are outlined in Section 26.
25.2 Downside Risk As with almost all mining ventures, there are risks that can affect the outcome of the Ambatovy Project. The major risk areas identified in this study are: • Lack of control over external drivers such as nickel or cobalt price and exchange rates • Poor control of mining dilution and loss during excavation activities
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• Not achieving the operating costs, productivities and other assumptions made in this study • Possible pit wall stability due to conditions not anticipated or modelled.
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26 Recommendations
The review of data and documents for this Technical Report leads to the recommendations set forth below. The recommendations are not phased, and each recommendation is not contingent on positive results from the other recommendations. All recommendations have been communicated previously to the Ambatovy Joint Venture; many of them are already being implemented or investigated. Some already have budgeted funds allocated for their implementation in the near future. Costing is not presentedas it is considered the recommendations could be completed as part of the ongoing mine operations.
26.1 In-Pit Screening of Coarse Saprolite The attrition tests confirmed that the nickel is mostly concentrated at the surface of the saprolite rocks, and little energy is required to wear the softest and weathered layer of the rocks. The distribution of nickel increased in the finer particles of the saprolite samples after grinding, suggesting that nickel was removed from the surface of the coarse particles. Thus, the use of mobile in-pit screening technologies would allow recovery of the fine saprolite with high nickel content.
26.2 Ground Penetrating Radar When mining the basal ferralite, the low-Mg saprolite and, possibly, the saprolite, ground penetrating radar (GPR) may improve production planning. In particular, it is likely to be able to provide reliable images of boulder zones, of bedrock pinnacles and of troughs in the bedrock. Previous field trials of GPR at the Ambatovy Project suggest that it might also be able to predict the locations of the high-Al gabbroic units that are generally going to be waste or, if ore, might need to be stockpiled as a high aluminum ore that requires special treatment or blending. A field trial of GPR technology was planned in 2014 to improve ore/waste discrimination but it is not known if this has taken place.
26.3 Mining The Ambatovy Joint Venture should evaluate the potential to convert lower classified grade material into Mineral Reserves. Geotechnical investigation and analysis for highwall designs greater than 50 m in height should be conducted. Consideration to residual geological structures and potential intersection with the groundwater table should be taken into effect for determining design stability and for factors of safety.
26.4 Hydrogeology An investigation should be undertaken that would lead to a better understanding of groundwater and its interaction with the highwall pit slope design and geotechnical stability.
26.5 Geology The resource model presented in this Technical Report introduced the use of gabbro proportions to account for internal block waste material that may be mined selectively. CSA Global recommends updating the grade control and reconciliation procedures to track and investigate variations on lithology proportions. CSA Global also recommends mapping of gabbro dykes as a standard grade control procedure, since it may help in predicting variations on gabbro proportions.
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27 References
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Appendix 1: Variogram Models
Geological Lithology Ellipsoid Structure 1 Structure 2 Ratio Area Elements Nugget domain domain bearing Sill Range Sill Range Major/Semi-Major Major/Minor Ni 140 0.02 0.08 27 0.06 528 1.6 3.4 Co 90 0.002 0.0034 60 - - 1.0 2.0 Fe 110 12.75 14.00 151 26.90 597 1.3 4.0 Al 140 1.51 1.73 66 1.11 457 1.7 2.2 Limonite Mg 120 0.10 0.03 36 0.16 335 1.4 7.0 Si 100 1.78 1.73 68 7.29 613 1.1 4.3 Cr 90 0.13 0.37 120 0.17 500 1.2 5.1 Mn 110 0.06 0.07 120 0.10 316 1.2 6.2 Ferricrete- Ti 140 0.17 0.24 195 - - 1.0 2.8 ferralite Ni 140 0.02 0.03 54 0.03 376 1.1 2.0 Co 90 0.0007 0.0004 103 0.002 526 1.7 3.0 Fe 110 15.14 33.13 63 57.31 717 1.9 2.6 Ambatovy Al 140 2.80 3.67 34 1.85 548 1.0 2.3 Centre and Gabbro Mg 120 0.00 0.01 96 0.01 636 1.1 15.3 South East Si 100 2.04 15.70 72 30.00 778 1.8 4.1 Cr 90 0.04 0.06 93 0.19 699 1.8 2.9 Mn 90 0.03 0.01 58 0.07 599 1.9 3.1 Ti 140 0.30 0.17 51 0.40 300 1.0 2.3 Ni 120 0.15 0.09 59 0.17 249 1.5 2.0 Co 110 0.00008 0.0038 53 - - 1.0 4.0 Fe 100 25.90 81.00 50 30.70 354 1.0 5.1 Al 100 2.30 3.70 60 3.00 293 1.4 3.3 Limonite- Saprolite Mg 110 10.00 15.10 66 - - 1.0 1.9 Saprolite Si 100 10.00 13.70 79 12.00 263 1.4 3.8 Cr 140 0.30 0.35 44 0.14 234 1.0 2.0 Mn 170 0.08 0.03 65 0.11 702 1.1 1.0 Ti 90 0.15 0.14 88 0.17 285 1.0 3.4
Geological Lithology Ellipsoid Structure 1 Structure 2 Ratio Area Elements Nugget domain domain bearing Sill Range Sill Range Major/Semi-Major Major/Minor Ni 170 0.03 0.07 38 0.07 421 1.0 4.7 Co 60 0.002 0.002 64 0.0015 216 1.4 4.8 Fe 40 8.99 12.80 64 48.20 722 1.1 8.2 Al 40 1.98 1.85 137 2.27 513 1.1 4.8 Limonite Mg 160 0.15 0.01 94 0.24 469 1.1 7.5 Si 160 1.67 19.20 621 - - 1.2 6.9 Cr 170 0.15 0.32 70 0.31 384 1.1 6.8 Mn 50 0.07 0.07 40 0.09 185 1.0 4.7 Ferricrete- Ti 130 0.20 0.11 106 0.12 500 1.4 2.9 ferralite Ni 170 0.015 0.04 83 0.04 440 1.9 2.9 Co 160 0.00089 0.0002 150 0.0027 577 1.5 3.4 Fe 160 12.60 58.10 116 65.50 478 1.2 3.9 Al 170 2.99 2.22 67 3.15 222 1.0 3.8 Ambatovy Gabbro Mg 170 0.01 0.02 131 0.01 727 1.0 5.1 West Si 160 1.11 31.00 121 38.00 528 1.1 7.6 Cr 170 0.06 0.08 188 0.12 524 1.1 3.6 Mn 90 0.04 0.02 89 0.07 534 1.2 3.3 Ti 80 0.24 0.44 89 0.21 383 1.0 2.6 Ni 0 0.11 0.13 73 0.19 443 1.3 3.5 Co 110 0.00027 0.0008 41 0.0005 326 2.5 5.5 Fe 70 25.80 70.20 96 37.90 503 1.3 5.3 Al 30 1.99 4.40 83 1.85 400 1.0 2.9 Limonite- Saprolite Mg 80 9.05 9.14 103 12.40 164 1.9 4.1 Saprolite Si 150 6.76 22.80 107 15.40 414 1.0 6.6 Cr 160 0.41 0.69 254 - - 1.2 8.0 Mn 40 0.02 0.05 78 0.04 492 1.3 11.0 Ti 70 0.19 0.06 153 0.15 452 1.0 10.0 Ca 40 0.07 0.05 70 0.19 470 1.0 7.6 Cu 0 0.00 0.00 106 0.00 613 1.0 4.8 K 100 0.05 0.03 263 0.13 881 1.0 13.7 Ambatovy Global Global Na 80 0.05 0.03 146 0.20 729 1.0 1.0 S 0 0.01 0.00 83 0.01 310 2.0 2.2 Zn 0 0.00 0.01 91 - - 1.0 7.4
Ellipsoid Structure 1 Structure 2 Ratio Area Geological unit Material type Elements Nugget bearing Sill Range Sill Range Major/Semi-Major Major/Minor Ni 120 0.02 0.048 139 0.05 412 1 8 Co 120 0.002 0.0022 103 0.0018 219 1.15 8.8 Fe 120 7.8 19.6 100 19.9 574 1.1 4.6 Al 50 1.61 0.63 100 2.14 443 1.27 5.77 Limonite Mg 30 0.011 0.054 113 0.11 439 1.5 14 Si 30 0.2 4.04 110 4.2 482 1.1 9.6 Cr 20 0.04 1.49 191 - - 1 8.16 Mn 120 0.062 0.011 107 0.073 336 1.5 10.5 Ferricrete- Ti 120 0.11 0.12 188 0.09 465 2.9 8.9 ferralite Ni 120 0.0057 0.008 51 0.0029 444 2.2 4.76 Co 70 0.0018 0.0012 97 0.004 540 1.23 3 Fe 120 15.3 32.3 89 50.7 477 1.27 3.89 Al 120 2.53 2.85 44 4.62 402 1.29 6.47 Analamay Gabbro Mg 30 0 0.015 91 0.012 541 1.9 11.3 Si 30 1.93 18.4 161 16.3 494 1.06 6.66 Cr 150 0.034 0.216 346 - - 1.65 3.7 Mn 140 0.025 0.04 112 0.14 416 1.78 5.48 Ti 150 0.25 0.58 209 - - 1 3.9 Ni 120 0.074 0.23 435 - - 1.4 13.5 Co 120 0.0013 0.0021 277 - - 1.6 11.8 Fe 20 58.9 54.8 148 50.7 302 1 5.76 Al 130 2.24 4.1 86 2.65 670 2.1 4.9 Limonite- Saprolite Mg 30 11.3 11.8 127 - - 1 2.4 Saprolite Si 30 9.7 23.8 211 - - 1 7.1 Cr 120 0.34 0.84 346 - - 1.1 21.3 Mn 120 0.06 0.11 368 - - 1 20 Ti 120 0.08 0.19 136 0.12 295 2.15 10.6 Ca 40 0.21 0.04 147 0.055 635 1 1 Cu 100 0.032 0.32 147 - - 1 8.9 K 80 0.001 0.009 612 - - 1 25.5 Ambatovy Global Global Na 50 0 0.025 508 - - 1 14.1 S 80 0.001 0.013 231 - - 1 10.25 Zn 150 0.000052 0.00031 83 0.00009 400 1.5 7.4