Technical Report

Project N TechnicalNI 43-101 Report RiverIron OreTrucking Mary Mary River, Baffin Island- o . 165926 Mary River, Baffin Island, Canada Mary River Iron Ore Trucking NI 43-101 Technical Report

Prepared by: AMEC Americas Ltd. Prepared for: Baffinland George H. Wahl, P. Geo Iron Mines Corporation Rene Gharapetian, P. Eng. February 2010 James E. Jackson, P. Eng Project No. 165926 Vikram Khera, P. Eng VOLUME 1 Gregory G. Wortman, P. Eng REPORT Effective Date, January 13, 2011

IMPORTANT NOTICE

This report was prepared as a National Instrument 43-101 Technical Re- port for Baffinland Iron Mines Corporation (BIM) by AMEC Americas Lim- ited (AMEC). The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in AMEC’s services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by (BIM) subject to the terms and conditions of its contract with AMEC. This contract permits (BIM) to file this report as a Technical Report with Cana- dian Securities Regulatory Authorities pursuant to National Instrument 43- 101, Standards of Disclosure for Mineral Projects. Except for the purposes legislated under provincial securities law, any other uses of this report by any third party is at that party’s sole risk.

Baffinland Iron Mines Corporation Mary River Baffin Island Mary River Iron Ore Trucking NI 43-101 Technical Report

TABLE OF CONTENTS

1.0 EXECUTIVE SUMMARY ...... 1-1 1.1 Introduction and Scope of Work ...... 1-1 1.2 Property Description and Location ...... 1-2 1.3 Exploration Licenses ...... 1-3 1.4 Accessibility, Climate, Local Resources, Infrastructure and Physiography ...... 1-3 1.5 History ...... 1-7 1.6 Geology and Mineral Resource Methodology ...... 1-8 1.7 Mine Engineering ...... 1-14 1.8 Mineral Processing...... 1-16 1.9 Geotechnical Investigations ...... 1-17 1.10 Mary River – Crushing, Screening and Stockpiling ...... 1-19 1.11 Shipping ...... 1-21 1.12 Infrastructure ...... 1-22 1.13 Port Layout ...... 1-25 1.14 Environmental ...... 1-26 1.15 Closure ...... 1-30 1.16 Capital Costs ...... 1-32 1.17 Operating Costs ...... 1-34 1.18 Project Execution Plan and Schedule ...... 1-35 1.19 Marketing ...... 1-35 1.20 Economic Analysis BIM to provide ...... 1-35 1.21 Risks and Opportunities ...... 1-36 1.22 Conclusions and Recommendations ...... 1-38

2.0 INTRODUCTION ...... 2-1 2.1 Terms of Reference ...... 2-2 2.2 Sources of Information ...... 2-2 2.3 Site Visits ...... 2-2 2.4 Scope of Facilities ...... 2-3

3.0 RELIANCE ON OTHERS ...... 3-1 3.1 Consultants and Contributors ...... 3-1 3.2 Qualifications of Consultants ...... 3-2

4.0 PROPERTY DESCRIPTION AND LOCATION ...... 4-1 4.1 Property Locations ...... 4-1 4.2 Mineral Titles ...... 4-1 4.3 Royalty Agreements and Encumbrances...... 4-7 4.4 Environmental Liabilities ...... 4-7 4.5 Existing Permits ...... 4-7 4.6 Jurisdiction ...... 4-7

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...... 5-1 5.1 Access ...... 5-1

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5.2 Climate ...... 5-1 5.3 Physiography ...... 5-2 5.4 Local Resources and Infrastructure ...... 5-3

6.0 HISTORY ...... 6-1 6.1 Ownership ...... 6-1 6.2 Past Exploration ...... 6-1 6.3 Work in Progress ...... 6-5

7.0 GEOLOGICAL SETTING ...... 7-1 7.1 Regional Geology...... 7-1 7.2 Local Geology ...... 7-5 7.3 Geology of the Mary River Iron Deposits ...... 7-6

8.0 DEPOSIT TYPES ...... 8-1

9.0 MINERALIZATION ...... 9-1

10.0 EXPLORATION ...... 10-1 10.1 History of Relevant Exploration Work ...... 10-1 10.2 Exploration Work by WGM ...... 10-1 10.3 Exploration Work by BIM (2004 to 2010) ...... 10-2 11.0 DRILLING ...... 11-1 12.0 SAMPLING METHOD AND APPROACH ...... 12-1 12.1 WGM – 1964/65 ...... 12-1 12.2 BIM 2004 to 2009 Exploration Campaigns ...... 12-2 12.3 Drill Core Handling Procedures ...... 12-2 12.4 Metallurgical Sampling Procedure ...... 12-3 13.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY ...... 13-1 13.1 Sample Preparation and Assaying ...... 13-1 13.2 Quality Controls and Quality Assurance ...... 13-4 14.0 DATA VERIFICATION ...... 14-1 14.1 WGM 1964 and 1965 Assay Verification ...... 14-1 14.2 BIM 2004 to 2009 Database and Assay Verification ...... 14-2 14.3 SGS Internal Blank, Duplicate Pulp and Standard Results ...... 14-3 14.4 Quantile Plots ...... 14-3 14.5 Twinned Drill Holes ...... 14-3 14.6 External BIM Duplicate Samples ...... 14-5 15.0 ADJACENT PROPERTIES ...... 15-1 16.0 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 16-1 16.1 Historical Testwork ...... 16-1 16.2 2003-2005 Testwork ...... 16-3 16.3 2006-2007 SGA Testwork ...... 16-8 16.4 2008 to 2010 Testwork ...... 16-12 16.5 Summary and Conclusions ...... 16-21

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16.6 Mineral Processing...... 16-22

17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ...... 17-3 17.1 Mineral Resources ...... 17-3 17.1.1 Resource Database ...... 17-3 17.2 Solid Modeling ...... 17-5 17.3 Compositing ...... 17-6 17.3.1 Statistics ...... 17-7 17.4 Resource Estimation Methodology ...... 17-18 17.4.1 Variography ...... 17-19 17.4.2 Estimation Methodology ...... 17-19 17.5 Mineral Resource Classification ...... 17-21 17.6 Validation of Block Model ...... 17-22 17.6.1 Assessment of Reasonable Prospects of Economic Extraction ...... 17-26 17.6.2 Mineral Resource Cut Off Grades ...... 17-28 17.6.3 Mineral Resource Statement ...... 17-29 17.7 Mining and Trucking Based Mineral Reserves ...... 17-31 17.7.1 Mineral Reserve Statement ...... 17-33 18.0 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES ...... 18-1 18.1 Proposed Mining Operation ...... 18-1 18.1.1 Pit Optimization ...... 18-1 18.1.2 Pit Design ...... 18-2 18.1.3 Proposed Crushing and Screening Plant ...... 18-4 18.1.4 Milne Inlet – Planned Transport, Stockpiling, Reclaiming and Ship Loading ... 18-4 18.1.5 Proposed Production Schedule ...... 18-5 18.2 Waste Rock Management ...... 18-9 18.2.1 Shift Roster and Required personnel ...... 18-9 18.3 Equipment Selection ...... 18-10 18.4 Infrastructure ...... 18-12 18.5 Environment ...... 18-12 18.6 Capital Cost Estimate...... 18-12 18.7 Operating Cost Estimate ...... 18-13 18.8 Markets ...... 18-14 18.9 Taxation ...... 18-18 18.10 Financial Analysis ...... 18-19 18.11 Basis of Financial Analysis ...... 18-20 18.11.1 Results of Base Case Financial Analysis ...... 18-21 18.11.2 Base Case Sensitivity Analysis ...... 18-23 18.11.3 Alternate Case Financial Analysis ...... 18-24 18.12 Risk and Opportunity Analysis ...... 18-26 19.0 OTHER RELEVANT DATA AND INFORMATION ...... 19-1 21.0 INTERPRETATIONS AND CONCLUSIONS ...... 21-1 22.0 REFERENCES ...... 22-1 22.1 Geology and Mineral Resources ...... 22-1 22.2 Metallurgy ...... 22-3 22.3 Geotechnical References ...... 22-3 23.0 DATE AND SIGNATURES ...... 23-2

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LIST OF TABLES

Table 1-1: Deposit 1, 2 & 3 Mineral Resources ...... 1-11 Table 1-2: Combined Deposit 1, 2 & 3 Mineral Resources ...... 1-11 Table 1-3: Mineral Reserves within the Trucking Study Designed Pit ...... 1-12 Table 1-4: Capital Cost Estimate Summary ...... 1-19 Table 1-5: Operating Cost Estimate Summary ...... 1-20 Table 1-6: Base Case Financial Analysis Summary ...... 1-21 Table 1-7: Alternate Case Financial Analysis Summary ...... 1-21 Table 4-1: Mary River Mineral Lease Status...... 4-1 Table 4-2: Nunavut Tunngavik Inc. Exploration Areas ...... 4-2 Table 4-3: BIM McOuat and Glacier Lake Staked Claims ...... 4-3 Table 4-4: BIM Turner, Cockburn and Rowley Claims Staked ...... 4-4 Table 11-1: Database Mary River Deposit No. 1 ...... 11-1 Table 11-2: Database Mary River Deposit No. 2 ...... 11-1 Table 11-3: Database Mary River Deposit No. 3 ...... 11-2 Table 11-4: Database Mary River Deposit No. 4 ...... 11-2 Table 11-5: Drill Core Recoveries ...... 11-3 Table 13-1: Detection Limits and Mean Grades of Deleterious Elements ...... 13-3 Table 14-1: WGM Comparative Analysis ...... 14-1 Table 14-2: Comparison of Twin Drill Holes S-2 and MR1-05-67...... 14-4 Table 14-3: Comparison of Twin Drill Holes S-5 and MR1-05-63...... 14-4 Table 14-4: Comparison of Twin Drill Holes S-6 and MR1-05-60...... 14-5 Table 16-1: Summary of Metso Testwork ...... 16-10 Table 16-2: Grade Differences Between Lump and Fines ...... 16-11 Table 16-3: 0.5 mm Size Fraction Results ...... 16-12 Table 16-4: Metallurgical Testwork by Deposit ...... 16-13 Table 16-5: Lump Iron Ore Physical and Chemical Attributes-Crusher Samples ...... 16-15 Table 16-6: Fine Iron Ore Physical and Chemical Attributes-Crusher Samples ...... 16-16 Table 16-7: Lump Trial Cargo Physical and Chemical Attributes – Receiving Port...... 16-17 Table 16-8: Fine Ore Trial Cargo Physical and Chemical Attributes – Receiving Port...... 16-18 Table 16-9: Fine Ore Lab-based Sinter Testwork ...... 16-20 Table 17-1: Exploration Database Used for Resource Estimation Mary River Deposit #1 ...... 17-4 Table 17-2: Database Used for Resource Estimation Mary River Deposit #2 ...... 17-4 Table 17-3: Database Used for Resource Estimation Mary River Deposit #3 ...... 17-4 Table 17-4: Deposit #1 Zone Solids Used to Constrain Grade Interpolation ...... 17-6 Table 17-5: Balance of Composite Length Analysis ...... 17-7 Table 17-6: Comp_100 Statistics ...... 17-8 Table 17-7: Comp_200 Statistics ...... 17-9 Table 17-8: Comp_300 Statistics ...... 17-10 Table 17-9: Comp_400 Statistics ...... 17-11 Table 17-10: Comp_600 Statistics ...... 17-12 Table 17-11: Comp_700 Statistics ...... 17-13 Table 17-12: Comp_1300 Statistics ...... 17-14 Table 17-13: Comp_1400 Statistics ...... 17-15 Table 17-14: Statistics Report Deposit #2 Upper 200 Zone ...... 17-16 Table 17-15: Statistics Report Deposit #3 Main Zone ...... 17-17

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Table 17-16: Major Axis Variogram Results Deposit No. 1 – North Limb ...... 17-19 Table 17-17: Deposit No. 1 Search Ellipse Parameters ...... 17-20 Table 17-18: Deposit No. 2 & 3 Search Ellipse Parameters ...... 17-21 Table 17-19: Comparison of Inverse Distance Squared and Ordinary Kriging Interpolated Grades for Deposit No. 1 ...... 17-22 Table 17-20: Optimisation Parameters ...... 17-28 Table 17-21: Deposit 1, 2 & 3 Mineral Resources ...... 17-30 Table 17-22: Combined Deposit 1, 2 & 3 Mineral Resources ...... 17-30 Table 17-23: Optimization Parameters ...... 17-32 Table 17-24: Mining Recovery and Mining Dilution ...... 17-32 Table 17-25: Summary of the Pit Slope Angle used in the Optimization ...... 17-33 Table 17-26: Mineral Reserves within the Designed Pit for Trucking Study ...... 17-33 Table 18-1: Summary of the Pit Optimization Result ...... 18-1 Table 18-2: Summary of the RoM Production Schedule ...... 18-7 Table 18-3: Pit Rock Movement Schedule during the Life of Mine ...... 18-8 Table 18-4: Mining and Ancillary Fleet Requirement ...... 18-11 Table 18-5: Cost Summary ...... 18-13 Table 18-6: Operating Cost Summary ...... 18-14 Table 18-7: Iron Ore Consumption by Country ...... 18-16 Table 18-8: Capital Cost Estimate Summary ...... 18-19 Table 18-9: Operating Cost Estimate Summary ...... 18-20 Table 18-10: Financial Summary ...... 18-20 Table 18-11: Revenue ...... 18-21 Table 18-12: Base Case Financial Analysis Summary ...... 18-21 Table 18-13: Financial Model ...... 18-22 Table 18-14: Summary of the Pre-Tax Annual Cash Flow ...... 18-23 Table 18-15: Owner Operated ($ 000) ...... 18-24 Table 18-16: Alternate Case, Summary of Contractor Operator Capital Costs ...... 18-25 Table 18-17: Alternate Case, Summary of Contractor Operator Operating Costs ...... 18-25 Table 18-18: Alternate Case Financial Analysis Summary ...... 18-26 Table 21-1: Drilling Budget for 2011 ...... 21-6 Table 21-2: Overall Budget (CAD Million) ...... 21-6

LIST OF FIGURES

Figure 4-1: Property Location Map Regional ...... 4-5 Figure 4-2: Deposit and Property Location Map ...... 4-6 Figure 4-3: Review Process Diagram ...... 4-10 Figure 5-1: Overall Mine Site Plan ...... 5-6 Figure 5-2: Milne Inlet Overall Site Plan ...... 5-9 Figure 7-1: Regional Geology Map of Baffin Island (Jackson et al, 2000) ...... 7-2 Figure 7-2: Geology of Northern Baffin Island ...... 7-4 Figure 7-3: Local Geology Map of Mary River Deposits ...... 7-8 Figure 7-4: Deposit No 1 Geology Map ...... 7-9 Figure 7-5: Typical Section of Deposit No. 1 ...... 7-10 Figure 7-6: Geology of Deposit Nos. 2 and 3 ...... 7-11

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Figure 7-7: Section of Deposit No. 2 ...... 7-12 Figure 7-8: Section of Deposit No. 3 ...... 7-14 Figure 7-9: Geology of Deposit No. 4 ...... 7-15 Figure 14-1: Fe% HARD Duplicates ...... 14-5 Figure 14-2: Fe% Duplicates...... 14-6 Figure 14-3: Fe% Duplicates >40% ...... 14-6 Figure 14-4: Fe% HARD Duplicates ...... 14-7 Figure 14-5: Fe% Duplicates...... 14-7 Figure 14-6: Fe% Duplicates <40% ...... 14-8 Figure 14-7: SiO2% HARD Duplicates ...... 14-8 Figure 14-8: SiO2% Duplicates ...... 14-9 Figure 14-9: SiO2% Duplicates <15% ...... 14-9 Figure 14-10: AI203% HARD Duplicates ...... 14-10 Figure 14-11: AI203% Duplicates ...... 14-10 Figure 14-12: Al2O3% Duplicates <10% ...... 14-11 Figure 14-13: MgO% HARD Duplicates ...... 14-11 Figure 14-14: MgO% Duplicates ...... 14-12 Figure 14-15: MgO% Duplicates ...... 14-12 Figure 14-16: P% HARD Duplicates ...... 14-13 Figure 14-17: P% Duplicates ...... 14-13 Figure 14-18: Duplicates <0.8% ...... 14-14 Figure 14-19: S% HARD Duplicates ...... 14-14 Figure 14-20: % Duplicates ...... 14-15 Figure 14-21: S% Duplicates <4% ...... 14-15 Figure 14-22: Mn% HARD Duplicates ...... 14-16 Figure 14-23: Mn% Duplicates ...... 14-16 Figure 14-24: Mn% Duplicates <4% ...... 14-17 Figure 16-1: Grain Size Distribution of Sinter Feeds and Concentrates ...... 16-19 Figure 16-2: Material Handling Flowsheet ...... 16-23 Figure 17-1: Swath Plot North 100 Zone Fe% ...... 17-23 Figure 17-2: Swath Plot North 100 FEO% ...... 17-23 Figure 17-3: Swath Plot North 100 Zone Al2O3% ...... 17-24 Figure 17-4: Swath Plot North 100 MgO% ...... 17-24 Figure 17-5: Swath Plot North 100 SiO2% ...... 17-25 Figure 17-6: Swath Plot North 100 Mn% ...... 17-25 Figure 17-7: Swath Plot North 100 P% ...... 17-26 Figure 18-1: General Layout of the Designed Final Pit, Waste Dump. Haul Road and Crusher Location ...... 18-3 Figure 18-2: Scheduled Fe, P, and S Grade per Period ...... 18-8 Figure 18-3: World Production of Iron Ore and Crude Steel ...... 18-14 Figure 18-4: Iron Ore Price Forecast ...... 18-18

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C ONTENTS

1.0 SUMMARY ...... 1-1 1.1 Location and Access ...... 1-1 1.2 Mineral Tenure, Surface Rights, and Royalties ...... 1-2 1.3 Permits and the Environment ...... 1-2 1.4 Geology and Mineralization ...... 1-3 1.5 History and Exploration ...... 1-4 1.6 Drilling ...... 1-5 1.7 Analyses and Quality Assurance/Quality Control ...... 1-5 1.8 Data Verification ...... 1-7 1.9 Metallurgical Testwork ...... 1-7 1.10 Mineral Resources ...... 1-8 1.11 Mineral Resource Statement ...... 1-10 1.12 Mineral Reserve Estimation...... 1-10 1.13 Mineral Reserve Statement ...... 1-12 1.14 Mine Plan ...... 1-12 1.15 Planned Operations ...... 1-13 1.16 Waste Rock ...... 1-15 1.17 Infrastructure ...... 1-15 1.18 Cost Estimates ...... 1-18 1.19 Marketing ...... 1-18 1.20 Taxation ...... 1-19 1.21 Financial Analysis ...... 1-21 1.22 Conclusions ...... 1-22 1.23 Recommendations ...... 1-22

T ABLES

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1.0 SUMMARY

AMEC Americas Limited (AMEC) was commissioned by Baffinland Iron Mines Corporation (BIM) to prepare an independent Qualified Person’s Review and NI 43- 101 Technical Report (the Report) on a feasibility study update for the wholly-owned Mary River Iron Ore Project (the Project) located in Baffinland, Nunavut. The objective of the engineering for the feasibility study for this project was to determine the technical viability of a trucking option for the Project and to develop conceptual engineering and the design basis together with the preliminary project execution plan and schedule and a project cost estimate in the order of ±15%.

The project contains the Deposit Nos. 1, 2, 3, 4, and 5 deposits, and the Deposit No’s 6, 7, 8 and 9 prospects.

BIM will be using the Report in support of the BIM Press Release “Baffinland Announces Economics for the Mary River Direct Shipping Road Haulage Project Option” dated January 13, 2011.

At the time of commencement of the feasibility study, BIM was a listed and publicly- traded Canadian entity. In September 2010, a hostile takeover was launched by Nunavut Iron Ore Acquisition Inc; the terms of the offer were rejected by BIM. In November 2010, BIM announced a support agreement with ArcelorMittal S.A, whereby Arcelor would acquire all of the outstanding common shares of BIM. On 14 January 2011, the offer for BIM became a joint offer from Nunavut and Arcelor. As at 7 February 2011, the joint venture partners had 90% control of BIM common share stock and were moving to compulsory acquisition. As of the date of this report, BIM remained a public company.

1.1 Location and Access

The Project area is located on the northern half of Baffin Island on National Topographic Systems (NTS) Sheet 37 G/5 at Latitude 71° and Longitude 79° approximately 160-km south of Mittimatalik (Pond Inlet), 270 km southeast of Nanisivik, 300-km north of Hall Beach, and 1000-km northwest of Iqaluit, the capital of the Nunavut Territory. The Universal Transverse Mercator (UTM) location of Deposit No. 1 is, 7915000N, 563000E [North American Datum (NAD) 1983 Zone 17).

Access to the Mary River project is by fixed wing aircraft using a 1,830-m gravel airstrip of which 1,463 m are currently useable. Access is also available by float or ski plane on nearby Sheardown Lake. Iqaluit has daily scheduled flights and Pond Inlet is serviced by an ATR turboprop aircraft six days a week. BIM currently operates a

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three-day a week regular charter service from Iqaluit to the site using a nine-passenger Dornier during the summer field season.

1.2 Mineral Tenure, Surface Rights, and Royalties

The mineral properties of BIM consist of three mining leases over the Mary River Deposit No. 1 (Lease 2484), Deposit Nos. 2 and 3 (Lease 2485), and Deposit No. 4 (Lease 2483). These leases cover a total area of 1,593 ha. BIM retains the lease rights to the four deposits. Deposits Nos. 5 to 9 are located on mineral claims totalling 53,700 ha that have been staked by BIM. BIM also holds an Exploration Agreement signed with the Nunavut Tunngavik Inc for an additional 18,120 ha proximal to the current three mining leases. The total property (Leases/Mineral Claims/Exploration Area) covers an area of approximately 73,413 ha over nine separate deposits.

The leases are valid for a 21-year period and will be required to be renewed on or before August 27, 2013. In order to obtain mineral leases under the Canadian Mining Regulations, legal surveys of the leases were required and completed by a qualified surveyor.

The mining leases predate the May 25, 1993, Nunavut Land Claims Agreement (NLCA), but are surrounded by Inuit-owned land (IOL), either designated as surface and subsurface rights (around mineral leases 2484 and 2485) or surface rights only (around mineral lease 2483). The mineral leases are administered by Indian and Northern Affairs Canada (INAC) under the Canadian Mining Regulations. Gaining access to land for which the Inuit have surface ownership is through the regional Qikiqtani Inuit Association in Iqaluit.

There are no royalties, back-in rights, payments or other obligations pertaining to the Mary River property, other than taxation pertaining to any mining property being developed in Nunavut.

1.3 Permits and the Environment

A number of permits have been obtained to carry out exploration activities. All exploration activities to date have been undertaken in accordance with the appropriate Nunavut or Canadian regulations.

Various permits, licences, and approvals will be required to be issued for construction and operation on successful completion of the review processes. Key permits to be obtained will include archaeological and paleontological permits, Type A Water

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Licence, Fisheries Act authorization applications, approvals or exemptions under the Navigable Waters Protection Act, and a licence for explosives manufacture.

As part of base line studies, BIM and its third-party consultants have completed vegetation surveys, terrestrial and marine mammal surveys, migratory bird studies, and surveys of marine and lacustrine/riverine species.

BIM filed is Draft Environmental Impact Statement in January 2011. It is expected that the environmental assessment process with be completed by the end of 2012.

1.4 Geology and Mineralization

The Mary River deposits are interpreted to be examples of Algoma-type iron formations. Global examples of this deposit type include Kudremuk (India), Cerro Bolivar (Venezuela), Carajás (Brazil), part of Krivoy Rog (Russia).

Within the Project, iron ore mineralization is developed in metasedimentary rocks of the Mary River Group, part of the Committee Belt, an assemblage of granite- greenstone terrains, rift basin sediments and volcanic rocks which lie within the northern Churchill Province and extend from southwest of Baker Lake for over 2000 km to north-western Greenland.

The Mary River Group consists of a diverse assemblage of metasedimentary and metavolcanic rocks, preserved in narrow, folded greenstone belts. The greenstone belts generally show a lower basal sequence of varied metavolcanic rocks, overlain by middle metasedimentary-metavolcanic sequences including iron formation, succeeded by an upper group of metavolcanic and metapelitic clastic sedimentary units with high- level metamorphism.

The Mary River deposits consist of a number of lensoid bodies which differ in iron, silica, and sulphur content, and in the proportions of their main oxide minerals, hematite, magnetite, and specularite. These chemical and mineralogical differences may reflect primary compositional variations of the exhalative sediments which formed the iron formation horizons, and secondary hypogene changes during metamorphism, as well as supergene alteration near unconformities.

Deposit No. 1, on which the trucking-option feasibility study was completed, is currently the largest defined iron deposit in the Mary River area. Deposit No. 1 can be divided into an approximately 1,300 m-long northern portion (North Limb) which strikes at 041° and dips at -77° to the southeast, and an approximately 700 m-long southern portion (South Limb) which strikes at 316º and dips at -65° to the northeast. The high- grade lower 100 zone oxide iron formation forms a tabular, 105 m to as much as 150

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m-thick body with chlorite–actinolite schist and/or garnetiferous amphibolite at the hanging wall and quartz-mica schist and quartz-feldspar-mica gneiss at the footwall. Bands of chlorite-actinolite schist with garnet and/or magnetite, banded oxide facies iron formation, and staurolite-cordierite-mica schist are rarely interlayered within the iron deposit particularly with the high-grade magnetite of the north limb. These waste zones are laterally continuous and average 1 m to 15 m in thickness and separate the lower, middle and upper zones. High-grade hematite-dominated iron formation predominates along the south limb and core of the synformal structure, while specularite occurs adjacent to the site where the north limb is disrupted by a north- northwest-trending fault. The iron formation along the south limb forms an assemblage of at least 290 m in thickness. The assemblage consists of two major sequences comprising a lower high-grade iron formation band (up to 120 m in true thickness) overlain by a mixed zone of variably alternating high-grade iron formation, banded oxide iron formation, chlorite-actinolite schist, amphibolite and cordierite- staurolite-mica schist layers 2 m to 18 m thick.

The high grade iron formation zone at Deposit No. 2 crops out over an elevation of 610 m to 670 m, strikes east-northeast–west-southwest to east–west and is characterized by sub-vertical dips to the south and south-southeast. Deposit No. 2 appears to be more complex than Deposit No. 1, in terms of continuity of high-grade iron formation and continuity of thickness.

Deposit No. 3 is situated on the crest and lower slope of a ridge 670 m south of Deposit No. 2. The deposit occurs at an elevation of 490 m to 530 m, and consists of high-grade hematite and specularite iron formation.

High-grade iron formation of Deposit No. 4 crops out on a low ridge 27 km northwest of Deposit No. 1. Exposures of high-grade magnetite and specularite iron formation occur as a series of elongated lenses or bands, 5 m to 75 m wide, which crop out intermittently over a strike length of 2,800 m within the mine lease boundary

Deposit No. 5 consists of hematite-enriched iron formation. The largest outcrop can be traced over an approximate 700 m strike length and an apparent exposed width of 70 m.

1.5 History and Exploration

Prior to BIM’s project involvement, work was completed by British Ungava Explorations Limited (Brunex) and Baffinland Iron Mines Ltd (BIML). Work completed included access construction works, airstrip construction, camp construction, geophysical surveys, geological mapping, core drilling, metallurgical tests, a mineral resource

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estimate, and numerous engineering studies and project reviews. This work is referred to as legacy data.

BIM commenced work in 2004, and has conducted re-establishment of the exploration camps, geological mapping, surveying, a magnetic airborne geophysical survey, channel sampling, bulk sampling, metallurgical testwork, core drilling in support of mineral resource and mineral reserve estimates, geotechnical evaluations, port site evaluations, and proposed rail corridors, and engineering studies.

A scoping study was performed on Deposit No. 1 in 2006 and a feasibility study completed on the same deposit in 2008. The key assumptions in both studies were that lump ore would be transported by rail to a port at Steensby Inlet, and then shipped to overseas smelters. Under the assumptions in the studies, positive project economics were returned. Both studies are considered historic, and information within the studies should not be relied upon, as the mineral resource and mineral reserve estimates used to support the studies have been superseded by the mineral resource and mineral reserve estimates in this Report.

In 2010, an updated feasibility study was commissioned to investigate the economics of a trucking operation from Deposit No 1 at the Project site to Milne Inlet, stockpiling of ore at a port facility, and then shipping of the ore to overseas smelters during the open water season.

1.6 Drilling

The BIM drill database consists of exploration drill data collected by in the 1960s, as well as six successive drill campaigns by BIM. Drilling on Deposit No’s. 1, 2, 3 and 4 total 640 core drill holes (41,298 m).

Core was logged for geological and geotechnical parameters. Drill collar locations were surveyed. Down-hole surveys for the BIM programs were performed using a Maxibor instrument.

Core from the legacy and BIM programs was typically halved. Sample intervals were 2m for BIM programs. Legacy drill data could have sample intervals from 3–6 m.

1.7 Analyses and Quality Assurance/Quality Control

Several primary assay laboratories have been used for routine analyses over the Project history. These include Technical Services Laboratories (TSL) and Superior Laboratories (SL) for the legacy sampling programs, and SGS Laboratories (SGS) for the BIM programs.

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Samples of half core from the legacy drill programs were crushed to -6 mm in a steel- plate crusher, a 300 g split of the crushed sample was pulverized to -10 mesh (1.7 mm), and 5 g assayed for soluble iron, silica and sulphur by wet chemical methods. For an accurate identification of ore types, a 100 g split of the -10 mesh material was used for standard Davis Tube analysis, which was the main analytical procedure used to classify the samples based on their hematite-magnetite content.

Initially, for the BIM core processing, preparation of samples included a primary crush of the sample to -43 mm maximum. This was followed by a secondary crush to 34 mm maximum. Samples were then screened at 6.3 mm, weighed and recombined. Samples were then riffle split to 500 g for tertiary grinding and pulverizing. A tertiary crush of the 500 g split to 3 mm completed the crushing preparation. . In 2005, the secondary crush to -34 mm was eliminated as analysis of screened fractions completed by SGA in Germany showed little or no variance. Analytical determinations consisted of:

H2O: Moisture content was determined at 105° on -3mm material by heat and weight loss. Fe: Total iron by potassium dichromate titration FeO: Non-oxidizing acid leach, followed by potassium dichromate titration Cl: UV spectrophotometry S: Total sulphur by Leco carbon sulphur analyser

Whole rock analysis determinations consisted of assays for SiO2, Al2O3, P2O5, Na2O, TiO2, V2O5, CaO, MgO, K2O, MnO, using borate fusion followed by X-ray fluorescence. Elemental detection limits for all elements except Na2O and K2O was 0.01%; Na2O and K2O was 0.05%. Loss on Ignition (LOI) was determined by high-temperature weight loss. Minor and trace element determinations were performed using inducted coupled plasma (ICP) spectrometry, following acid digestion. A suite of ICP analyses were also evaluated to see if any elements were potentially problematic.

There is limited currently-available information on any quality assurance/quality control (QA/QC) programs for the legacy data.

Approximately 5% of the 2004, 2005, 2006, 2007, and 2009 BIM assays were sent to an external referee laboratory, SGA in Germany, for duplicate pulp analysis. SGS’s internal quality control procedures include duplicate samples, spiked blanks, spiked replicates, reagent/instrument blanks, preparation control samples, certified reference material (CRM) analysis, and instrument control samples; these data were used to provide the QA/QC data for the BIM drill programs.

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1.8 Data Verification

Review of the legacy data in the database consisted of check assaying of a small population of samples by sending these samples to external laboratories. The results

suggested that the legacy Fe and SiO2 grades for the seven samples could be reasonably reproduced at external laboratories.

To check the quality of assaying by TSL, a set of length-weighted composite samples

of minus 10-mesh rejects were re-assayed by TSL and SL for Fe, P, SiO2, Mn, Al2O3, CaO, MgO, S, TiO2 and LOI. The TSL and SL composite assays showed generally good agreement for most elements with TSL reporting generally slightly higher values, and phosphorous having the largest variation.

Data verification completed by G. Wahl in support of mineral resource estimation consisted of data entry validation, analysis of available external duplicate data, analysis of twinned holes, checking the results of SGS’s internal quality control, checking the quality of core logging and sampling protocols, comparisons of drill campaign results via quantile plots, and validating of drill hole collar locations. No areas of bias or quality concern were identified.

1.9 Metallurgical Testwork

Initial metallurgical testwork programs included reduction-disintegration tests, reducibility and swelling tests, and preliminary pellet tests. Under BIM, testwork performed included chemical analysis, size analysis and tumble tests, decrepitation tests and evaluation of the dynamic and static low-temperature disintegration behaviour, evaluation of the reducibility, the reduction index, and the reduction-under- load behaviour, evaluation of crushing behaviour, porosity and density measurements, tumbler and abrasion tests, crushing pilot plant tests and sinter tests. In addition, high- temperature properties such as softening, melting and dipping of the iron oxides were evaluated using a REAS test. Three bulk sample cargoes (2-lump and 1-fine iron ore) totalling 113,000 tonnes were utilised in blast furnaces in Europe with outstanding results.

Testwork results confirmed the high quality of Mary River lump ore. Chemical analysis showed high Fe contents for Mary River lump ore and low gangue contents. The contents of alkalis, manganese, titanium and vanadium are generally low. Phosphorus values may need to be controlled via blending.

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1.10 Mineral Resources

The exploration database for Deposit No. 1 consisted of 156 drill holes, totalling 30,170m and 220 channel samples totalling 1,090 m. Legacy drill holes in Deposit No. 1 were removed for resource estimation purposes where those holes were replaced by current BIM drilling. Because of the location of channel samples on very limited exposure and difficulty in obtaining continuous and representative samples, the channel sample data was also excluded.

A total of eight drill holes (1,314 m) and 22 channel samples totalling 130 m were included in the Deposit No. 2 database.

A total of 11 drill holes (2,553 m) and eight channel samples totalling distance of 82 m were included in the Deposit No. 3 database. As the 2007 channel samples on Deposit 3 contained a complete deleterious dataset and were largely repeats of the 1960s era channel sampling, only the 2007 channel samples were used to support resource estimation.

A LiDAR topographic coverage at 30 cm resolution was used to delimit Deposit No. 1.

Wireframes were interpreted for lithological units on approximately 75 m-spaced drill- hole sections at Deposit No. 1. An assay boundary of ~55% Fe was used to define hanging- and footwall solid boundaries; however, where the massive iron formation graded along strike into a banded iron formation, a lower cut-off was used to allow grade interpolation parameters to define gradational extent of lower-grade mineralization along strike between drill sections. A lower zone and upper zone solid were defined for Deposit No. 2. Deposit No. 2 is narrower and bifurcates into two separate zones at its eastern extent. A single main zone was interpreted for Deposit No. 3.

Samples were composited to 5 m composites for statistics and grade interpolation. Remaining composite intervals with a minimum length of at least 3.25 m were also included within these composite files.

Estimation domains were defined following evaluation of statistical distributions.

The model for Deposit No. 1 used a block size of 15 m x 15 m x 15 m, whereas that for Deposit Nos. 2 and 3 used 25 m x 25 m x 10 m blocks. Model block grade attributes comprised Fe%, SiO2%, Al2O3%, FeO%, Mag%, MgO%, S%, P%, Mn%, Na2O%, K2O%, and Lump%. Additional block attributes included resource classification, apparent density, distance to nearest sample, rock code and average distance to nearest sample.

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Density data consisted of 387 metallurgical samples. Waste rock was assigned an assumed density of 2.65 g/cc. Overburden was assigned a density of 2.24 g/cc. For the ore zones an iron grade-based regression formula where apparent density = (0.0288* Fe%) + 2.65 was estimated for each block.

Ordinary kriging was used to inform block grades used for resource estimation for the Deposit No 1 block model while inverse distance squared was used to interpolate grades to the Deposit Nos. 2 and 3 block model.

Local block grade estimates were validated by comparing block grades to drill-hole grades on a section-by-section and elevation-by-elevation basis. The model was further validated by comparing the ordinary kriged grades versus the inverse distance squared estimate. The model was considered to appropriately reflect the raw data.

The following criteria were used to pre-classify blocks into classification categories:

Measured Mineral Resources: Block informed by a minimum of three samples, within an interpreted mineralized solid, and within 45 m of the nearest informing sample. Indicated Mineral Resources: Block informed by a minimum of three samples, within an interpreted mineralized solid and greater than 45 m and less than 120 m of the nearest informing sample. Inferred Mineral Resources: Block informed by a minimum of three samples, within an interpreted mineralized solid and greater than 120 m and less than 450 m of the nearest informing sample.

Reasonable prospects of economic extraction were assessed by applying preliminary economic constraints within an open pit shell. Parameters used to constrain the Mineral Resources to assess reasonable prospects of economic extraction were based on the rail option evaluated during 2008. Although the trucking study that supports Mineral Reserves supersedes that estimate, and the rail study is not current, the derivation of the cost parameters was considered appropriate to support the open pit optimization for Mineral Resources.

The optimization and final/ultimate pit shell selection for Deposit 1 was carried out by using a 59% Fe cut-off grade (CoG) for orebody 100 and 60% Fe for the remaining orebodies. This was to have consistency in generating the nested pit shell with the previous studies. However, in order to produce a direct shipping product with an average in-situ grade of +66.0% Fe, a mining CoG of 59% Fe was selected to estimate the mineral inventory inside the selected pit shells for Deposit No. 1.

The diluted iron grade product in Deposit Nos. 2 and 3 was selected at +64% Fe and the CoG was 60.5% Fe. This was achieved by applying 2.5% dilution at zero percent

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Fe in the optimization. As a result, the average insitu grade of the total mineral inside the selected ultimate pit shell for Deposits 2 and 3 is 65.8% Fe.

1.11 Mineral Resource Statement

Mineral resources for Deposit Nos. 1, 2, and 3 are shown in Table 1-1 and Table 1-2. The tables include estimated deleterious element grades and are inclusive of Mineral Reserves. Mineral resources that are not mineral reserves do not have demonstrated economic viability. The effective date of the estimate is 30 December, 2010.

1.12 Mineral Reserve Estimation

Mineral Reserve estimation for Deposit No. 1 was carried out using Whittle and Gems software (Lerchs–Grossmann (LG) based packages) to develop a series of nested pits for a portion of the Measured and Indicated mineral resources, considering open pit mining and the initial topography. In order to design the open pit, and for practical reasons, a larger shell, containing 120 Mt of mineralization was selected, and a 60 Mt pit was designed within the larger shell.

Access and in-pit ramps were developed at 33 m and at a 10% gradient. A mining CoG of 59% Fe and 60% Fe was used in the optimization process for the Upper Orebody, and the Middle and Lower orebodies respectively, to simulate a direct shipping product at an average >65 % Fe.

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Table 1-1: Deposit 1, 2 & 3 Mineral Resources Deposit No Resource Category Million Tonnes Al2O3 % Fe % FeO % Mn % P % S % SiO2 %

1 Measured 207 1.2 66.3 13.8 0.16 0.05 0.19 2.7 1 Indicated 211 1.0 66.3 15.7 0.21 0.05 0.26 2.5 1 Measured & Indicated 418 1.1 66.3 14.8 0.19 0.05 0.23 2.6 1 Inferred 213 0.9 66.9 22.9 0.40 0.05 0.43 1.9 2 & 3 Indicated 26 1.0 65.0 2.0 0.02 0.02 0.02 5.2 2 & 3 Inferred 336 1.1 65.9 3.7 0.10 0.03 0.01 2.4

Note 1 : Deposit No 1 Mineral Resources are Above 59% Fe Cut off and within Pit Shell Note 2 : Deposit 2 & 3 Mineral Resources Above 60.5% Fe Cut off and within Pit Shell

Table 1-2: Combined Deposit 1, 2 & 3 Mineral Resources

Deposit No Resource Category Million Tonnes Al2O3 % Fe % FeO % Mn % P % S % SiO2 % 1, 2 & 3 Measured 207 1.2 66.3 13.8 0.16 0.05 0.19 2.7 1, 2 & 3 Indicated 237 1.0 66.2 14.2 0.2 0.05 0.23 2.8 1, 2 & 3 Measured & Indicated 444 1.1 66.2 14.1 0.2 0.05 0.22 2.8 1, 2 & 3 Inferred 549 1.0 66.3 11.1 0.2 0.04 0.17 2.2 Note 1 : Deposit No 1 Mineral Resources are Above 59% Fe Cut off and within Pit Shell Note 2 : Deposit 2 & 3 Mineral Resources Above 60.5% Fe Cut off and within Pit Shell

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For scheduling purpose additional mining dilution and mining recovery factors were assigned to each orebody according to the orebody geometry. The lower and middle orebodies are solid and consist of some areas of internal waste. A one to two percent mining dilution at zero grade was applied to the in situ iron grade. The Upper orebody is smaller than the other orebodies with more internal waste, sharp transition between ore and waste rock, and irregular orebody geometry. Higher mining dilution and loss of 5% (at zero iron grades) were applied to calculate the diluted mineral grade in this area. Dilution was not considered for the in situ grade of the deleterious elements.

Geotechnical parameters were used to design pit slopes. The surveyed topography surface (LiDAR Surface) was used for mineral reserve estimation. This surface was adjusted to reflect the excavated rock during the bulk sample operation.

1.13 Mineral Reserve Statement

The mineral reserves for the Mary River Project are tabulated in Table 1-3. Mineral reserves have an effective date of 30 December 2010.

Table 1-3: Mineral Reserves within the Trucking Study Designed Pit

Mt Fe% P% S% Mn% SiO2% Al2O3% FeO% Proven 33.7 66.1 0.023 0.069 0.12 1.89 1.37 9.3 Probable 27.1 66.3 0.027 0.061 0.14 1.69 1.13 7.8 Total 60.7 66.2 0.025 0.065 0.13 1.79 1.26 8.6

AMEC notes that the 2008 feasibility study, which was based on an assumption of using rail transport, evaluated a larger area of mineralization than the current study. There remains significant upside Project potential in Deposit No. 1 if this mineralization can be integrated into the current study.

1.14 Mine Plan

Dates discussed in this section of the Technical Report are for illustrative purposes only, as a decision to proceed with mine construction still requires regulatory approval and approval by BIM management

The proposed trucking option mine plan envisages 3 Mt/a of run-of-mine (RoM) production of a direct-shipping product with an average grade of >66.0% Fe. An

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average lump – fine ratio of 75% - 25% was estimated by BIM for the final product based on test results.

The pre-production period is projected to start in the first half of 2013 reaching the full 3 Mt/a production rate in the second half of 2013.

During the pre-production period, a minimum of 300 kt of ore is planned to be crushed and shipped. In spite of the scheduled 300 kt for crushing and shipping, the mining schedule provides 600 kt of ore during the pre-production period. This allows for flexibility in the mining operation if it be required to crush and ship, up to an additional 300 kt of ore.

The selected CoG strategy was to maintain the continuity of the orebody on each bench during the mining operation, and to provide a direct shipping product at an average grade of >66% Fe in the mining schedule. An insitu CoG of 57% Fe was used for the Lower Orebody and a 50% CoG for the Middle and Upper orebodies due to limited tonnage (219 kt) of the mineralized rock in the 50%-57% Fe with minimal effect on the overall RoM grade.

In Year 9, 64 kt of mineralized rock (Lower Orebody) in the 57-62% category was considered as part of the low-grade stockpile due to the initial setup for the mining schedule, and was not included in the RoM. This was to maintain an average grade of 66%Fe in Year 9. It is likely that all or part of this tonnage will be included in the RoM during the mining operation with in-pit blending and/or utilizing the crushed ore stockpile.

The stripping ratio in the preproduction period is 1.2 and then it is decreased to about 0.4 in the first five years. In Years 6 to 8 the stripping ratio is 0.2. This is due to longer haul distances in this period and to maintain a constant truck fleet during the mining operation. The stripping ratio is then increased to 0.3 until Year 15 and then it is decreased to 0.2 in the final five year period. The total average stripping ratio is 0.3.

The objective of the mine scheduling was to maximize the iron content in the RoM and to avoid sharp fluctuation of the iron and deleterious elements grades. Product specification for the deleterious elements was not considered as a constraint in the schedule. The duration of the project, excluding the six months of pre-mining period, is 20 years.

1.15 Planned Operations

Mining is envisaged as a conventional truck and shovel operation. Trucks will feed a crushing and screening plant that is designed to operate 300 d/a, with an operating

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availability of 75%. The primary objective of the plant is to maximize the production of lump product (-25/+7 mm), while at the same time, keeping fines product (-7 mm) at a minimum. The plant design throughput is planned at 600 t/h with the product streams consisting of 75% lump product and 25% fine product.

Crushed ore will be transported by haul trucks 100 km from Mary River to Milne Inlet, taking four to five hours for a round trip, depending on the time of the year. Due to the short and variable shipping season (70–100 d/a) the total annual production (3.0 Mt) will be stockpiled at the port.

The Port of Milne is planned to be a Panamax iron ore-loading port for approximately 100 days per year during the open water season, from approximately July 15th to October 15th. During the season it is expected to load 40 to 45 Handymax and Panama size ore vessels.

The mining fleet estimation was carried out by using the prepared mining schedule and estimating the haul distances from the open pit to the crusher and the waste dump. The 55t dump trucks and 4.75m3 backhoe excavators are selected for the mining operation at Mary River. Smaller equipment may was also considered for the relatively small mining operation however, due to the mined rock hardness and abrasiveness it was decided to select the above mentioned equipment for rigidness and to have the available excess capacity.

Based on the estimates; eight trucks, two backhoe excavators, and two production drills capable of drilling up to 7” blast hole are required for the mining operation. Supporting equipment such as secondary loader, track and wheel dozers, and water cart, together with secondary ancillary equipment such as fuel truck are also included in the required mine fleet estimates.

The work regime for the mining operation consists of 365 days per annum, assuming two shifts a day, and 12 hours per shift. The approach for this working regime was based on an allowance of six weeks trucking delay from Mary River to Milne Inlet, and three weeks of weather delay during the winter season. There are no planned shutdowns. Major overhaul and maintenance of the equipment will be planned during the 65 days of trucking-delay period; however, personnel will be available on site should limited mining operation be required for waste stripping.

A total of 161 personnel are required in the pre-production period and 177 personnel in Year 1 and beyond.

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1.16 Waste Rock

Waste and mineralized rock were categorized into potentially acid-generating (PAG) rock and clean waste. The PAG dump requires proper drainage in order to collect the potential run-off. Detailed studies were not carried out to quantify the PAG rock within the mining limit therefore a preliminary estimate of 20% of the total waste was considered to be PAG rock. Mineralized rock below the CoG was included in this category and a separate pile was not assumed for the low-grade mineralized rock.

The waste dump capacity is estimated by using a swell factor of 40% for all rock types. The non-PAG dump capacity is 7.4 Mm3 and for the PAG dump (including the low- grade stockpile) the required dump volume is 1.9 Mm3.

1.17 Infrastructure

Proposed infrastructure includes:

Mary River Mine Site

Open pit mine Access and site roads Temporary construction facilities Construction camp Contractor offices Quarry and borrow sites Temporary fuel storage Aggregate crusher and stockpiles Concrete batching plant Power generators Maintenance shops Warehouses/stores Laydown areas Communication systems Waste dump Ore crushing and screening facilities Material handling system Ore stockpiling/reclaim facilities

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Truck loading and weigh scale facilities Offices and administration Permanent camp Assay laboratory Communication systems Heavy equipment fleet parking Airstrip and associated facilities Bulk fuel storage and distribution Explosive manufacture and storage (Contract) Maintenance shops Water supply (fresh water / fire water / industrial water) Water treatment facility(sewage treatment and sludge incineration) Power generation Emergency power Waste management facilities (solid waste and recycling)

Tote Road Mary River to Milne Inlet

Bridges / culverts Quarry and borrow sites Refuge stations Potable and raw water supply Sewage handling disposal during construction / operation Temporary fuel storage

Milne Inlet Port Site

Access and site roads Temporary construction facilities Construction camp Contractor offices

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Quarry and borrow sites Temporary fuel storage Aggregate crusher and stockpiles Concrete batching plant Power generators Maintenance shops Warehouses/stores Laydown areas Freight dock Communication systems Truck receiving facilities Ore stockpiling Reclaim and materials handling facilities Ship loading facilities Floating ore dock Over winter berthing facilities Freight receiving and laydown area Mooring buoys Navigation lights and buoys Dredging Offices and administration Permanent camp Communication systems Airstrip and terminal facilities Bulk fuel receiving, storage and distribution Water supply (fresh water / fire water / industrial water) Water treatment facility (sewage treatment and sludge incineration) Power generation Emergency power Waste management facilities (solid waste and recycling)

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Maintenance shops Parking

1.18 Cost Estimates

The capital and operating costs were estimated accordingly based on the required mining equipment, the annual utilized hours, mine consumables per annum, and the required technical staff, workshop personnel, and mine operators. Recent quotes from equipment manufacturers and suppliers were used for the cost estimates for the major mining and ancillary equipment.

The initial capital cost for the mining equipment is CAD$20 M and the sustaining capital cost for the project is CAD$17.5 M. This cost is based on equipment budget prices at the Port of Montreal, and includes the initial spare parts. Contingency and shipping cost from Port of Montreal to Milne Inlet are added in the overall financial model of the project. In addition to these costs, CAD$10.8 M of the estimated operating cost during the preproduction period in Year 2013 will be considered as the project capital cost. Capital costs are summarized in Table 1-4Table 1-4.

The total mine operating cost during the 20 years project life (excluding the preproduction period) is CAD$493 M or CAD$6.43/t of mined rock (CAD$8.21/t of ore). Shipping cost of the consumables (excluding the fuel/lube from the port of Montreal) and the contingency on the operating cost will be included in the project financial model. The annual cost of the mine personnel (62%) and the fuel cost (15%) are the two major components of the total mine operating cost. Operating costs are summarized in Table 1-5.

1.19 Marketing

Testwork has confirmed the high quality and attractive attributes of the Mary River iron ores. Current market conditions have seen an almost 700% increase in the price of iron ore in the last 10 years. It is expected that over the next 10 year that price will stabilise at some level with new production coming to market.

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Figure 1-1: Iron Ore Price Forecast

Iron Ore Price Forecast - Atlantic Basin

300.0

250.0

200.0

per dmtu) per 150.0 ¢

100.0 Price (US Price

50.0

0.0 1998 1999 2000 2001 2002 2003 2004 2005 2006 2007 2008 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 Year

Lump Iron Ore Price Forecast Fine Iron Ore Forecast

Source: AME, CRU

Despite the price increases, most new production has been from near-mine Brownfield development and much of the new production is finer and finer grained iron ores that inhibit sinter productivity. Over the next 10 years, most of the new production will be pellet feed (< 150μm) material that can only be used to make a pellet. In addition increasing energy cost wills make quality lump and fine iron ore sought-after products.

Baffinland’s Mary River lump and fine iron ore have been clearly demonstrated a high quality premium iron ores that will be valued blast furnace feeds and may very well attract an additional price premium to existing forecasted iron ore prices.

1.20 Taxation

Taxation in Nunavut consists of Federal and Nunavut Income Tax. In addition, mining income is subject to the Nunavut Mining Royalty, a graduated tax from 5 % to 14% based upon net income. It is treated the same as income tax excepting for interest on debt which is not deductible for calculation of the Royalty.

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Table 1-4: Capital Cost Estimate Summary Mary River Iron Ore - Trucking Option Project Cost Summary

Project Area Description Cost (CAD x'000) Direct Costs Mining 59,470 Crushing & Screening Plant 23,350 Stockpiling & Shiploadout 53,770 Infrastructure at Mary River (Mine) 122,244 Infrastructure at Milne Inlet (Port) 96,378 Ore Dock Facilities 58,216 Tote Road 85,268 Total Direct Costs 498,696

Indirect Costs Project Indirect Costs 142,797 Owner's Costs 8,970 Royalties (Gravel) 12,540 Total Indirect Costs 164,307 Subtotal Direct and Indirect Costs 663,003 Total Contingency 77,280 TOTAL PROJECT COST 740,283

Closure Costs 27,000 Sustaining Capital 146,057

Table 1-5: Operating Cost Estimate Summary

CAD/DMT Mining 8.21 Processing 2.80 Road haul 8.85 Stock Piling 1.45 Ship Loading 3.50 Camp 2.06 Air Transportation 1.27 G&A 0.79 Total 28.93 Ocean Shipping* Excluded

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1.21 Financial Analysis

The results of the economic analysis represent forward-looking information that are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.

BIM completed the financial modeling. AMEC reviewed the pre-tax financial modeling and conducted the sensitivity analysis. Tax calculations were reviewed for reasonableness; however, AMEC does not provide expert advice on taxation matters.

The financial analysis does not include certain other costs that would be associated with executing the Trucking Project Option including any financing costs, corporate G&A and the socio-economic costs associated with the Inuit Impact and Benefits Agreement, including financial participation.

The base case financial data is assumes an Owner-mining scenario. Financial analysis of the base case showed the pre-tax project NPV at 8% to be CAD$1,407 million, the after-tax NPV at 8% to be CAD$1,030 million and the internal rate of return (IRR) after tax to be 30.6%. The payback period is 2.6 years. A summary of the cashflow analysis is included in Table 1-6.

An alternate financial analysis (Table 1-7) was performed to assess the economics of a mining operation that was performed under contract. Financial analysis of the alternate case showed the pre-tax project NPV at 8% to be CAD$795 million, the after- tax NPV at 8% to be CAD$577 and the internal rate of return (IRR) to be 22.3%. The payback period is 3.6 years.

Table 1-6: Base Case Financial Analysis Summary

Owner Operated ($ millions) After-tax internal rate of return 30.6% Project after-tax cash flow $3,070 Project pre-tax net present value 8% $1,407 Project after-tax net present value 8% $1,030 Capital cost of the Project $740 Operating cost (per tonne) $29 Payback period (in years) 2.6

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Table 1-7: Alternate Case Financial Analysis Summary

Contractor Operated ($ millions) After-tax internal rate of return 22.3% Project after-tax cash flow $1,957 Project pre-tax net present value 8% $795 Project after-tax net present value 8% $577 Capital cost of the Project $605 Operating cost (per tonne) $63 Payback period (in years) 3.6

1.22 Conclusions

In the opinion of the QPs, the Project that is outlined in this Report has achieved its objectives in that a deposit that could support mine development has been identified. A decision to proceed with development will require appropriate permits, and approval by both relevant statutory authorities and BIM’s board.

1.23 Recommendations

Recommendations include further infill and regional exploration and geotechnical drilling. Total drilling is 6000 m at a total cost of $6 million. Additional environmental, geotechnical, geological, and design engineering and construction planning activities are also planned for a grand total cost of $41 million.

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C ONTENTS

2.0 INTRODUCTION ...... 2-1 2.1 Qualified Persons ...... 2-2 2.2 Information Sources ...... 2-2 2.3 Site Visits ...... 2-3

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2.0 Introduction

AMEC Americas Ltd (AMEC) was retained in July, 2010 by Baffinland Iron Mines Corporation (BIM) to carry out a Feasibility Study for the Mary River Iron Ore Trucking Project (the project). The objective of the engineering for the feasibility study for this project is to determine the technical viability of the project and to develop conceptual engineering and the design basis together with the preliminary project execution plan and schedule and a project cost estimate in the order of ±15%. This report (the Report) summarizes the results of the Feasibility Study.

BIM will be using the Report in support of the BIM Press Release “Baffinland Announces Economics for the Mary River Direct Shipping Road Haulage Project Option” dated January 13, 2011.

The overall Mary River Project is a world-class direct ship iron ore deposit located on Baffin Island in the high arctic of Nunavut, Canada. In 2008, a development proposal was produced and evaluated in a Definitive Feasibility Study by Aker Solutions (AKER 2008 FS) that included an 18 Mt/a, crushing and screening operation with transportation via rail to a deep water port and year round shipping.

To expedite the Project, an accelerated development concept based on trucking the ore from Mary River to Milne Inlet, with shipping to customers only during the open water season at a capacity of approximately 3 Mt/a was investigated through conceptual studies by both AMEC and BIM. The results of those studies were encouraging and Baffinland elected to progress the concept further by carrying out a feasibility study.

The Mary River deposit is located in the high arctic with very challenging logistics due to the extreme cold climate and short open water shipping season. Special emphasis was also required on the protection of the sensitive northern environment both during construction and operation. Engineering also considered optimization of equipment and modularization / pre-fabrication, wherever possible, with an emphasis on rapid construction techniques.

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This Report includes an update to the mineral resource estimate and mining engineering used for the 2008 feasibility study prepared by Aker Kvaerner. The mineral resource estimate includes drill and assay results from an exploration drilling program completed in 2009. The mineral resources from Deposit No. 2 and Deposit No. 3 were included for the purposes of providing a complete inventory of Mary River mineral resources. The mineral resource estimate is intended to be used by BIM to further the development of their properties by providing an updated estimate of mineral resources in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) resource classification system.

This mineral resource estimate was commissioned by BIM in January 2010 and provides the basis for this Report.

2.1 Qualified Persons

The following are the Qualified Persons (QPs) as defined in National Instrument 43-101, Standards of Disclosure for Mineral Projects, for portions of the Report:

George Wahl, P. Geo., Independent Consultant to the Issuer

Rene Gharapetian, P. Eng., Independent Consultant to the Issuer

Vikram Khera, P. Eng., AMEC, Financial Analyst

James Jackson, P. Eng., Specialist Engineer – Civil, AMEC.

Gregory Wortman, P. Eng., AMEC, Technical Director, Process

George Wahl is responsible for the preparation of Sections 4 to 15 and 17 of this Report.

Rene Gharapetian is responsible for preparation of mining related portions of Sections 17, 18 and Section 21,

Vikram Khera is responsible for preparation of the Financial Analysis included in Section 18.

James Jackson is responsible for input to Sections 5 and 18 of the report.

Gregory Wortman is responsible for preparation of the balance of the Report.

2.2 Information Sources

Information that supports the Report has been obtained from BIM and third-party consultants as follows:

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Information used to support mineral resource estimation was obtained from BIM officials and technical staff for the major part. Such information was obtained from company documents or obtained from BIM staff via personal communication. Information regarding product marketing and taxation applicable to the Project was provided by BIM. Some historical data was obtained from early Watts, Griffis and McOuat reports as cited in the References section of this Report. Regional and local geological information was obtained from reports of the Geological Survey of Canada. Information was also sought from trade journals and the technical literature. AMEC has also used information contained in previous technical reports on the Project.

Reference documents are cited in the text as appropriate and summarized in Section 22 of this Report.

AMEC provided the remaining project management, engineering design and cost estimating services for the Feasibility Study.

2.3 Site Visits

Site visits were completed by George H. Wahl, P. Geo for a period of 1 to 2 weeks in each of the past five years. The most recent visit was June 13 to 19, 2010.

The scope of the site visit by George H. Wahl was to examine drill core and outcrops from each deposit, review core logging, sampling, surveying, drilling, quality control, database management procedures and assist with maintaining ore quality during the bulk sample program.

R. Gharapetian visited the site for three days in June 2006 for general site familiarization.

James Jackson visited the site in July, 2010 for evaluation of the road routing between Mary River and Milne Inlet, and sourcing borrow materials for the same.

The QPs who visited site are not aware of any material scientific and technical changes to the information on the Project between the dates of the site visit and Report signature date.

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C ONTENTS

3.0 RELIANCE ON OTHER EXPERTS ...... 3-1 3.1 Consultants and Contributors ...... 3-1 3.2 Mineral Tenure ...... 3-1 3.3 Surface Rights ...... 3-1 3.4 Royalties ...... 3-1 3.5 Permitting...... 3-2 3.6 Environmental ...... 3-2 3.7 Geotechnical Studies ...... 3-2 3.8 Shipping Costs and Logistics ...... 3-2 3.9 Marketing ...... 3-2 3.10 Taxation ...... 3-3 3.11 Contract Mining, Crushing and Screening, Road Haulage Design and Unit Price Estimating ...... 3-3 3.12 Port and Dock Design Information and Unit Pricing ...... 3-3

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3.0 RELIANCE ON OTHER EXPERTS 3.1 Consultants and Contributors

The QPs, as authors of this Report, state that they are qualified persons for those areas identified in the relevant “Certificate of Qualified Person” attached to this Report. The QPs have relied on, and believe there is a reasonable basis for this reliance, upon the following reports of other experts, which provided information regarding mineral rights, surface rights, and environmental status in sections of this Report as noted below.

3.2 Mineral Tenure

The QPs have not reviewed the mineral tenure, nor independently verified the legal status, ownership of the Project area, underlying property agreements or permits. AMEC has fully relied upon, and disclaims responsibility for, information derived from legal experts as provided by BIM.

This information is used in Section 4.2.

3.3 Surface Rights

The QPs have fully relied upon information supplied by BIM’s staff and experts retained by BIM for information relating to the status of the current Surface Rights.

This information is used in Section 4.

3.4 Royalties

The QPs have fully relied upon information supplied by BIM’s staff and experts retained by BIM for information relating to the royalty status for the Project.

This information is used in Section 18.

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3.5 Permitting

The QPs have fully relied upon information supplied by BIM’s staff and experts retained by BIM for information relating to the status of the current permitting for the Project.

This information is used in Section 4.

3.6 Environmental

The QPs have fully relied upon information supplied by BIM’s staff and experts retained by BIM for information relating to the environmental status of the Project.

This information is used in Section 4.

3.7 Geotechnical Studies

The QPs have fully relied upon information supplied by BIM’s staff and experts retained by BIM for information relating to geotechnical conditions expected in the open pit and waste dump design.

This information is used in Sections 17 and 18.

3.8 Shipping Costs and Logistics

The QPs have fully relied upon information supplied by BIM’s staff and experts retained by BIM for information relating to logistical concerns and shipping costs

This information is used in Section 18.

3.9 Marketing

The QPs have fully relied upon information supplied by BIM’s staff and experts retained by BIM for information relating to marketing for product from the Project as follows:

AME Mineral Economics Group, 2010, (Confidential Report).

CRU Group 2010, (Confidential Report).

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This information is used in Section 18.

3.10 Taxation

The QPs have fully relied upon information supplied by BIM’s staff for information relating to taxation applicable to the Project

This information is used in Section 18.

3.11 Contract Mining, Crushing and Screening, Road Haulage Design and Unit Price Estimating

The QPs have fully relied upon information supplied by BIM’s staff and experts retained by BIM for information relating to design information and unit pricing for evaluation of the design/supply/operate contract mining, crushing screening and road haulage option for the Project as follows:

Nuna Logistics Limited, Court Smith, P. Eng., Executive Vice President, “Baffinland Trucking Scenario Feasibility Study – Nuna Cost Estimates, Mary River Project, Baffinland Iron Mines Corporation, December 20, 2010.

This information is used in Section 18.

3.12 Port and Dock Design Information and Unit Pricing

The QPs have fully relied upon information supplied by BIM’s staff and experts retained by BIM for information relating to design information and unit pricing for evaluation of the design/supply/operate contract port stockpiling/reclaiming and shiploading option for the Project as follows:

Ocean Construction Inc., Steve Cote, “Modular Wharf for the Milne Inlet Project, Preliminary Budget Pricing”, November 10, 2010.

Logistec Stevedoring Inc. Alain Pilotte, “Acquisition and operation costs estimate; Baffinland’s Iron Ore shipping 2013”, December 16, 2010.

This information is used in Section 18.

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C ONTENTS

4.0 PROPERTY DESCRIPTION AND LOCATION ...... 4-1 4.1 Property Locations ...... 4-1 4.2 Mineral Titles ...... 4-1 4.2.1 Mineral Leases ...... 4-1 4.2.2 Exploration Areas ...... 4-2 4.2.3 Mineral Claims ...... 4-2 4.3 Royalty Agreements and Encumbrances ...... 4-7 4.4 Environmental Liabilities ...... 4-7 4.5 Existing Permits ...... 4-7 4.6 Jurisdiction ...... 4-7 4.7 Permits Required to Support Development ...... 4-8 4.8 Environment ...... 4-9 4.8.1 Baseline Studies ...... 4-9 4.8.2 Environmental Impact Statement ...... 4-9 4.8.3 Company Policies ...... 4-12 4.8.4 Closure Plan ...... 4-12 4.9 Socio-Economics ...... 4-12 4.10 Comment on Section 4 ...... 4-13

T ABLES

Table 4-1: Mary River Mineral Lease Status ...... 4-1 Table 4-2: Nunavut Tunngavik Inc. Exploration Areas ...... 4-2 Table 4-3: BIM McOuat and Glacier Lake Staked Claims ...... 4-3 Table 4-4: BIM Turner, Cockburn and Rowley Claims Staked ...... 4-4

F IGURES

Figure 4-1: Property Location Map Regional ...... 4-5 Figure 4-2: Deposit and Property Location Map ...... 4-6 Figure 4-3: Review Process Diagram ...... 4-10

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4.0 Property Description and Location

4.1 Property Locations

The Project area is located on the northern half of Baffin Island on NTS Sheet 37 G/5 at Latitude 71 and Longitude 79 approximately 160-km south of Mittimatalik (Pond Inlet), 270-km southeast of Nanisivik, 300-km north of Hall Beach, and 1000-km northwest of Iqaluit, the capital of the Nunavut Territory. The Universal Transverse Mercator (UTM) location of Deposit No.1 is, 7915000N, 563000E (NAD 1983 Zone 17) (Figure 4-1).

4.2 Mineral Titles

4.2.1 Mineral Leases

The mineral properties of BIM (Figure 4-2) consist of three mineral leases over the Mary River Deposit No. 1 (Lease 2484), Deposit Nos. 2 and 3 (Lease 2485), and Deposit No. 4 (Lease 2483). These leases cover a total area of 1593.4 ha (Table 4-1). The leases are valid for a 21-yr period and subject to renewal on August 27, 2013. In order to obtain mineral leases under the Canadian Mining Regulations, legal surveys of the leases were required and completed.

Table 4-1: Mary River Mineral Lease Status

Lease Issue Date Expiry Date Area (ha) Deposits (No.)

2483 1992-08-27 2013-08-27 129.73 4

2484 1992-08-27 2013-08-27 557.20 1

2485 1992-08-27 2013-08-27 906.43 2, 3

The mining leases predate the May 25, 1993, Nunavut Land Claims Agreement (NLCA), but are surrounded by Inuit-owned land (IOL) (Figure 4-2), either designated as surface and subsurface rights (around mineral leases 2484 and 2485) or surface only (around mineral lease 2483). The mineral leases are administered by Indian and Northern Affairs Canada (INAC) under the Canadian Mining Regulations. Gaining access to land for which the Inuit have surface ownership is through the regional Qikiqtani Inuit Association in Iqaluit.

The original leases were surveyed in 1971 by the Federal Land Surveyor and leases were issued in 1971, by the Northwest Territories Government, renewed in 1992 and

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are in good standing to August 27, 2013. Pursuant to the incorporation of BIM, the successor company to the earlier BIM, the leases were legally transferred to the former on March 4, 2005.

Survey markers of the lease boundaries comprising steel pins and rock cairns were relocated by BIM personnel in the summer of 2007 using a Trimble GPS to confirm the previously defined latitude and longitude coordinates and the current NAD (1983) Zone 17 UTM datum. The BIM surveyed UTM points indicate that Deposit Nos. 1 and 2 and the western extent of Deposit No. 3 are all well contained within the lease boundaries. BIM has indicated that legal surveys are filed with the Legal Surveys Division of the Department of Energy Mines and Resources in Ottawa. After the initial 21-yr period, the leases were renewed in 1992 for a second 21-yr period and in 2013 can be renewed for another 21 year period.

4.2.2 Exploration Areas

In 2008, BIM obtained the right to earn a 100% interest in two exploration areas through an Exploration Agreement with the Nunavut Tunngavik Inc. one exploration area totalling 16,695 hectares surrounding mineral leases 2484 and 2485 and the second exploration area totalling 1,425 hectares situated to the east of mineral lease 2483. The two exploration agreement areas are in good standing for a period of five years and expire in 2013. Agreement areas are summarized in Table 4-2 and shown in Figure 4-2

Table 4-2: Nunavut Tunngavik Inc. Exploration Areas

Block Issue Date Expiry Date Area (ha) Deposits (No.)

1 2008-05-01 2013-05-01 16,695 3 Possible Eastern Extension 2 2008-05-01 2013-05-01 1425 5 Portions of Deposit 18,120

BIM’s mining leases are not subject to the Exploration Agreement and subsequent Joint Venture agreements should Baffinland choose to continue with the Exploration Agreement and meet the expenditure and annual payment requirements.

4.2.3 Mineral Claims

In October 2008, Baffinland staked 18 mineral claims totalling 12,956 hectares known as the McOuat block which encircle the leases covering Deposit No 4 and extend to the southeast to cover Deposit 5. See Table 4-3 and Figure 4-2.

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In 2010, BIM staked four additional blocks of claims: the Glacier Lake Block consisting of twenty claims covering Deposit No 6 the Turner River Block consisting of nine claims covering Deposit No 7 the North Cockburn River Block consisting of eight claims covering Deposit No 8 the Rowley River Block consisting of 4 claims covering Deposit No 9

Claim information is summarized in Table 4-3 and Table 4-4 and locations are shown in Figure 4-2.

Table 4-3: BIM McOuat and Glacier Lake Staked Claims Claim Claim Number Name Block Hectares Acquired Expiry K12448 BIM 18 McOuat Lake 1,045 2008-10-28 2010-10-28 K12447 BIM 17 McOuat Lake 1,045 2008-10-28 2010-10-28 K12446 BIM 16 McOuat Lake 1.045 2008-10-28 2010-10-28 K12445 BIM 15 McOuat Lake 1,045 2008-10-28 2010-10-28 K12444 BIM 14 McOuat Lake 1,045 2008-10-28 2010-10-28 K12443 BIM 13 McOuat Lake 1,045 2008-10-28 2010-10-28 K12442 BIM 12 McOuat Lake 269 2008-10-28 2010-10-28 K12441 BIM 11 McOuat Lake 856 2008-10-28 2010-10-28 K12440 BIM 10 McOuat Lake 1,045 2008-10-28 2010-10-28 K12439 BIM 9 McOuat Lake 1,045 2008-10-28 2010-10-28 K12438 BIM 8 McOuat Lake 1,045 2008-10-28 2010-10-28 K12437 BIM 7 McOuat Lake 126 2008-10-28 2010-10-28 K12436 BIM 6 McOuat Lake 104 2008-10-28 2010-10-28 K12435 BIM 5 McOuat Lake 66 2008-10-28 2010-10-28 K12434 BIM 4 McOuat Lake 484 2008-10-28 2010-10-28 K12433 BIM 3 McOuat Lake 388 2008-10-28 2010-10-28 K12432 BIM 2 McOuat Lake 522 2008-10-28 2010-10-28 K12431 BIM 1 McOuat Lake 731 2008-10-28 2010-10-28 K13811 M 1 Glacier Lakes 543 2010-07-12 2012-07-12 K13812 M 2 Glacier Lakes 1,015 2010-07-12 2012-07-12 K13813 M 3 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13814 M 4 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13815 M 5 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13816 M 6 Glacier Lakes 522 2010-07-12 2012-07-12 K13817 M 7 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13818 M 8 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13819 M 9 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13820 M 10 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13821 M 11 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13822 M 12 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13823 M 13 Glacier Lakes 522 2010-07-12 2012-07-12

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Claim Claim Number Name Block Hectares Acquired Expiry K13824 M 14 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13825 M 15 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13826 M 16 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13827 M 17 Glacier Lakes 522 2010-07-12 2012-07-12 K13828 M 18 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13829 M 19 Glacier Lakes 1,045 2010-07-12 2012-07-12 K13830 M 20 Glacier Lakes 1,045 2010-07-12 2012-07-12 Note: Claims K12431 through K12448 inclusive were renewed in late 2010; review of the assessment report and update of the INAC land cadastral has not yet been completed.

Table 4-4: BIM Turner, Cockburn and Rowley Claims Staked Claim Claim Number Name Block Hectares Acquired Expiry K14313 Turner-1 Turner River 1,045.1 2010-09-08 2012/09/12 K14314 Turner-2 Turner River 1,045.1 2010-09-08 2012/09/12 K14315 Turner-3 Turner River 1,045.1 2010-09-08 2012/09/12 K14316 Turner-4 Turner River 1,045.1 2010-09-08 2012/09/12 K14317 Turner-5 Turner River 1,045.1 2010-09-08 2012/09/12 K14318 Turner-6 Turner River 1,045.1 2010-09-08 2012/09/12 K14319 Turner-7 Turner River 1,045.1 2010-09-08 2012/09/12 K14320 Turner-8 Turner River 1,045.1 2010-09-08 2012/09/12 K14321 Turner-9 Turner River 1,045.1 2010-09-08 2012/09/12 K14301 NCR-1 North Cockburn 1,045.1 2010-09-08 2012/09/12 K14302 NCR-2 North Cockburn 1,045.1 2010-09-08 2012/09/12 K14303 NCR-3 North Cockburn 1,045.1 2010-09-08 2012/09/12 K41304 NCR-4 North Cockburn 1,045.1 2010-09-08 2012/09/12 K14305 NCR-5 North Cockburn 1,045.1 2010-09-08 2012/09/12 K14306 NCR-6 North Cockburn 1,045.1 2010-09-08 2012/09/12 K13407 NCR-7 North Cockburn 1,045.1 2010-09-08 2012/09/12 K13408 NCR-8 North Cockburn 1,045.1 2010-09-08 2012/09/12 K13409 NRR-1 North Rowley River 1,045.1 2010-09-08 2012/09/12 K14310 NRR-2 North Rowley River 1,045.1 2010-09-08 2012/09/12 K14311 NRR-3 North Rowley River 1,045.1 2010-09-08 2012/09/12 K14312 NRR-4 North Rowley River 1,045.1 2010-09-08 2012/09/12

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Figure 4-1: Property Location Map Regional

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Figure 4-2: Deposit and Property Location Map

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4.3 Royalty Agreements and Encumbrances

BIM is the exclusive owner of the mineral rights within the three mining leases that host the Mary River deposits 1, 2, 3 and 4. There are no royalties, back-in rights, payments or other obligations pertaining to the Mary River property, other than taxation pertaining to any mining property being developed in Nunavut.

4.4 Environmental Liabilities

Given that the Mary River site was explored in the 1960s, it is possible that some environmental liabilities exist in the form of petroleum contamination from general site activities, and from an original camp landfill. The potential environmental liability from these operations is thought to be minimal, based on an assessment of the area carried out in 2006. Surface clean up of materials and equipment remaining from exploration in the 1960s has been undertaken by BIM since 2004 including removal of steel drums and equipment from the Project site. Progressive remediation has been completed for all exploration work and work associated with the bulk sampling program.

4.5 Existing Permits

A number of permits have been obtained to carry out exploration activities. All exploration activities to date have been undertaken in accordance with the appropriate Nunavut or Canadian regulations.

A list of granted permits provided by BIM include Water License (2BB-MR0710, NWB), Authorization S.35(2) of Fisheries Act (NU-06-0084, DFO), S.5(1) Navigable Waters Protection Act (8200-09-10415 8200-09-10425, 8200-09-10414, 8200-09- 10424,Transport Canada, Inuit Land Use Lease and Aggregate Concession (Q10L3C010, QIA), Land Use Permit (N2007F0004, N2006C0036, INAC) and Quarry Permit (2010QP0088, INAC).

4.6 Jurisdiction

The Mary River iron ore deposits are located in the Territory of Nunavut, which was created in 1999 through the Nunavut Land Claims Agreement. The leases predate this inception date, but responsibilities for land use and rights have devolved to the Territorial Government and its agencies.

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Lands in the Nunavut Territory are classified as Crown Lands, Inuit Owned Lands (IOL) (surface rights) and IOL (surface and subsurface rights) and Commissioners lands. The Mary River leases of BIM are situated on land held by Crown mineral leases, but adjoin IOL held by the Inuit through the Qikiqtani Inuit Association (QIA).

The Territorial Land Use Regulations (TLUR) regulates surface activities related to mineral exploration and mining. The Canadian Mining Regulations regulate subsurface mineral explorations, exercised through INAC. BIM’s mineral leases have surface rights. Subject to completion of the environmental assessment process, BIM will extend or convert part or all of the existing commercial (surface) lease with the Qikiqtani Inuit Association for development and infrastructure facilities. BIM has recently filed a Draft Environmental Assessment Statement with the Nunavut Impact Review Board.

4.7 Permits Required to Support Development

Various permits, licences, and approvals will be required to be issued for construction and operation on successful completion of the review processes.

Archaeological and paleontological permits are required from the Department of Culture, Language, Elders and Youth (CLEY) for survey and mitigation of archaeological and paleontological sites prior to development.

Land tenure through long-term leases and shorter-term land use permits will be required from the QIA to access Inuit-owned lands that surround the Mary River Mine Site and from INAC.

Other key approvals include a Type A Water Licence from the Nunavut Water Board (NWB), Fisheries Act authorization applications with the Department of Fisheries and Oceans (DFO), approvals or exemptions under the Navigable Waters Protection Act administered by Transport Canada Navigable Waters Protection Program (TC-NWPP), and a licence for explosives manufacture from Natural Resources Canada (NRCan) under the Explosives Act.

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4.8 Environment

4.8.1 Baseline Studies

No plant species considered to be “rare” in Canada were found to occur in the survey locations. Terrestrial mammals in the region include barren-ground caribou of the North Baffin herd, wolf, arctic and red fox, ermine, arctic hare, and lemmings. Marine mammals are found in abundance in the region, including polar bears, narwhals, beluga whales, and bowhead whales, several species of seals, and walrus. North Baffin caribou are currently present at low densities and their numbers seem to vary in accordance with a 60- to 70-year cycle. Caribou are expected to remain at low numbers for the next couple of decades. There is evidence that caribou occur throughout the entire region. While some populations of caribou migrate between preferred habitats in summer and winter, North Baffin caribou appear to be non- migratory and are likely to be found relatively equally in many locations throughout the Project area.

Air quality and noise levels in the Project area are typical of remote environments. Freshwater quality measurements in the Mary River area indicate naturally elevated concentrations of dissolved oxygen, turbidity, aluminum, and iron. Some average values for pH, as well as cadmium and mercury in fresh water are greater than levels recommended by the guidelines of Canadian Council of Ministers of the Environment.

4.8.2 Environmental Impact Statement

BIM has recently filed a Draft Environmental Assessment Statement with the Nunavut Impact Review Board. The Process is tabulated in Figure 4-3.

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Figure 4-3: Review Process Diagram

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4.8.3 Company Policies

BIM is committed to protection of the health and safety of employees and the environment, and to ongoing community involvement and participation in the Project. The company embraces the principle of Social Responsibility as outlined by the emerging voluntary standard of the International Standards Organization, Guidance for Social Responsibility. The Project will be carried out in conformance with applicable Nunavut and Canadian laws, regulatory requirements, agreements, permits, and licences. In addition, on conclusion of the EIS process, BIM will complete an Inuit Impact and Benefits Agreement (IIBA) under negotiation with the Qikiqtani Inuit Association.

4.8.4 Closure Plan

The closure plan is preliminary but was developed in accordance with the regulatory framework established by the Inuit, Federal and Territorial governments as applicable to the feasibility phase of design.

The overall intent of the Closure Plan is to restore the Project areas to a near-natural condition, that are self-sustaining and to mitigate impacts from mining activities. In such a condition, the Project areas will provide wildlife and aquatic habitat. All creeks and rivers will return to pre-development conditions. A new lake will be created from the open pit.

Final closure of the Mary River Project will generally include removal and disposal of all equipment and materials to the Mine Site landfill or the materials and equipment will be sealifted to an approved facility located offsite. The open pit will be allowed to naturally re-flood developing a pit lake following closure. The waste rock stockpile will be re-contoured as necessary and will be allowed to naturally re-vegetate. Areas disturbed by project activities will be recon-toured as necessary to restore natural drainage and will naturally re-vegetate.

A post-closure monitoring program compliant with the applicable guidelines and regulations will be implemented to ensure the reclamation measures remain effective and continue to provide a high level of protection for the public and the environment.

A total of US$27 M has been allocated in the financial analysis for closure.

4.9 Socio-Economics

The five communities of northern Baffin Island in the immediate vicinity of the Mary River Project, listed alphabetically, include Arctic Bay (280 km), Clyde River (415 km),

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Hall Bay (192 km), Igloolik (155 km), and Pond Inlet (160 km). Each of these communities has historical socio-economic and ecosystem ties to the Project area.

Harvesting from the land (hunting, trapping and fishing) is a key livelihood component for many residents of North Baffin. Supplementing the in-kind income generated through harvest activities, residents earn money through employment and various social transfers. Other income generating activities include arts and crafts, carving, prints, tapestries, and wall hangings. Residents have expressed enthusiasm for wage- based work, even when this means working in remote locations away from the community.

4.10 Comment on Section 4

In the opinion of the QPs, the information discussed in this section supports the declaration of Mineral Resources and Mineral Reserves, based on the following:

Information from BIM supports that the mining tenure held is valid and is sufficient to support declaration of Mineral Resources and Mineral Reserves Surface rights are held either by the Crown, or by the Inuit as Inuit Owned Lands, and Commissioners Lands. Project development will be possible on the Crown- owned lands with appropriate permits. Surface rights access to the Inuit Owned Lands will require negotiation with the appropriate parties. There are no royalties payable to any third party. Permits obtained by the company to date to undertake exploration activities are sufficient to ensure that activities are conducted within the regulatory framework required by the Nunavut Government Additional permits will be required for Project development and operation, as summarized in this Section. At the effective date of this report, environmental liabilities are limited to those expected for an exploration-stage project that has been under consideration since the 1970s, and includes potential contamination from historic exploration site activities and a landfill. There are no historical environmental liabilities from BIM drill and bulk sample programs.. The current state of knowledge on environmental and permit status for the Project supports the declaration of Mineral Resources.

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C ONTENTS

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...... 5-1 5.1 Access ...... 5-1 5.2 Climate ...... 5-1 5.3 Physiography ...... 5-2 5.4 Local Resources and Infrastructure ...... 5-3 5.4.1 Existing Infrastructure ...... 5-3 5.4.2 Proposed Infrastructure ...... 5-4 5.4.3 Access ...... 5-10 5.4.4 Port ...... 5-10 5.4.5 Power...... 5-10 5.4.6 Water ...... 5-11 5.4.7 Communications ...... 5-11 5.5 Waste ...... 5-11 5.6 Comment on Section 5 ...... 5-11

F IGURES

Figure 5-1: Overall Mine Site Plan ...... 5-6 Figure 5-2: Milne Inlet Overall Site Plan ...... 5-9

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5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Access

BIM currently operates a three-day a week regular charter service from Iqaluit to the site using a nine-passenger Dornier during the summer field season.

Milne Inlet, the closest navigable water, is located 100 km to the northwest of Nuluujaak Mountain, the site of the Mary River Deposit No. 1. A gravel road (the tote road) constructed in the 1960s (suitable for use as a winter road) connects the Mary River deposits with Milne Inlet. There is also an airstrip at Milne Inlet.

Major movements of supplies and equipment would most likely originate from Montreal and be brought to Milne Inlet by ship, and then transported to Mary River by the tote road. During the 2007 field season BIM rehabilitated the Milne–Mary River tote road and airstrip in preparation for the extraction of a bulk sample during 2007–2008. The tote road from Mary River to Milne Inlet is the proposed route for the trucking option feasibility study that is the subject of this Report.

Steensby Inlet, located about 130 km south-southeast of Mary River, also provides a navigable access route and is being proposed as the port site from which Mary River ore will eventually be exported via rail to the port and year-round marine shipment. This inlet provides a longer ice-free season, but at this time, there is currently no road route or port facility available for use through this entry.

5.2 Climate

Frost-free conditions are short and are from late June to late August. There is continuous daylight from approximately May 5 to August 7. The months of July and August bring maritime influences and are usually the wettest (snow may still occur). Fog increases at this time due to arrival of moist air from southern Canada.

During September to November, temperature and the number of daylight hours start to decrease, and by mid-October the mean daily temperature is well below 0ºC. The highest amount of snowfall typically occurs during this period. A condition called

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“Arctic white out” often occurs during this time, where diffuse white clouds blend into the white snow-covered landscape, reducing visibility and increasing the likeliness of disorientation. This condition can also occur in April and May.

It is expected that mining, crushing and screening, ore transportation and stockpiling will be possible year-round. These operations are designed based on 300 operating days per year. Port access for delivery of materials and supplies and for product shipping will be constrained to ice-free periods, normally sixty to ninety days between about July 15 and October 15. These operations are based on an average of 75 available operating days.

5.3 Physiography

The Mary River iron deposits are situated close to the Central Borden Fault Zone, a major tectonic (structural discontinuity) and morphological feature which separates topographically lower land to the southwest from a higher plateau to the northeast which is deeply dissected by river valleys.

The northeast valley wall forms a strong linear scarp rising to the plateau, forming flat uplands at 450 m to 500 m that are deeply dissected by southward-flowing drainage valleys. The Mary River situated between Deposit Nos. 1 and 2 falls from an elevation of 530 m to 300 m elevation in a little over 4 km. The rivers drain south to Angajurjualuk Lake, or north to Tay Sound and Paquet Bay. The high land is characterized by a short snow-free period, and by ice caps and lakes with perennial ice cover.

The airstrip and old campsite at Sheardown Lake are situated on the south side of the Central Borden Fault at an elevation of approximately 180 m, while the Mary River Deposit No. 1 on Nuluujaak Mountain (698 m elevation) is on the northeast side of the Central Borden Fault.

A broad area in the fault valley about 1-km southwest of Deposit No. 1 at 180 m to 200 m elevation provides ample well-drained, relatively flat spaces for any contemplated future construction sites.

The Project lies within the zone of continuous permafrost, with an active layer thickness of up to 2 m and a total permafrost depth of about 500 m.

The extremely cold temperatures of the region, combined with permafrost ground conditions result in a short period of runoff that typically occurs from June to September. All rivers and creeks, with perhaps the exception of the very largest systems, freeze completely solid during the winter months. The runoff coefficient is very high, due to the combination of low temperatures, low infiltration, and minimal

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vegetative cover, and correspondingly, surface water is abundant, and the region is dotted with thousands of small lakes and streams.

The vegetation of northern Baffin Island contains fewer species and typically less ground coverage compared with more southerly tundra environments. Vegetation communities include upland areas recently emerged from under glacial ice; mixed tundra on lower slopes; heath tundra on drier slopes and sheltered banks; tundra polygons on expansive lacustrine or glacio-fluvial deposits; wetlands, and riparian associations.

Terrestrial wildlife in the region is comprised of the following eight species: caribou, wolf, arctic fox, red fox, ermine, Brown and Pearyland collared lemmings, and arctic hare. Notable bird use in the area consists of some flyover of migratory birds (particularly geese) moving to and from Bylot Island, and an abundance of raptors and loons. Plenty of raptor nesting habitat exists throughout the region, and the Peregrine Falcon (subspecies tundris), which have been recovering from near extinction in the late 1960s and were upgraded from being „threatened‟ to being a „species of special concern‟ in 1992, are abundant throughout the region.

Arctic char (Salvelinus alpinus) is the most abundant and widely distributed fish species in the lakes, rivers, and streams of Baffin Island.

During August and September, narwhal, bowhead whale, ringed seal, bearded seal, and harp seal are present within the waters of Milne Inlet and Eclipse Sound. Beluga and killer whales may also occasionally occur in those waters during this time.

5.4 Local Resources and Infrastructure

5.4.1 Existing Infrastructure

The only existing local resources available at the site are freshwater, sand, gravel and coarse stone. There is no timber of any kind. A potential hydroelectric source is situated near Separation Lake/ Rowley River, approximately 60 km to the northeast of Steensby Inlet, reportedly capable of 17.6 MW production. Wind power generation is also a potential source of energy.

The infrastructure currently available is a simple port facility at Milne Inlet, the Milne- Mary River tote road and the airstrips at Mary River and Milne Inlet. A 100 person Weatherhaven BIM camp at Mary River and a 55 person camp owned by Nuna at Milne Inlet were constructed to support the bulk sampling program. A 40 person camp has also been constructed at Steensby to support geotechnical and environmental assessments of the proposed future rail route. Additional capacity includes a 60

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person tent camp at Mary River and a 10 person tent camp at Milne. The 1462 m airstrip near the Mary River site is in use and the airstrip at Milne Inlet (approximately 800 m) are usable, but inadequate for large aircraft.

As a result of the bulk sample program, an inventory of equipment remains on site. This is comprised of BIM-owned equipment as well as equipment leased from Nuna. BIM equipment includes a 320 excavator, D5 and D4 dozers, Cat Telehandler, and Cat 950 loader. Explosives magazines and detonator magazines owned by Dyno are also located at Mary River. Nuna equipment on site includes fuel tank farms, fire tuck, tire truck, 16H and 14H graders, water truck, D7 and D8 dozers, (3) 773E haul trucks, 930 loader, 966 loader, (2) 980H loaders, 988H loader, fuel truck, radial stackers, crushing plant and (2) 345B excavators, ambulance, and (6) TA800 Kenworth trucks with (6) haul pups.

The surrounding communities of Iqaluit, Clyde River, Igloolik and Pond Inlet provide a potential labour pool for operations. Some Inuit have received training in basic trades and skills while working on the now closed DEW line and Nanisivik operations.

5.4.2 Proposed Infrastructure

Proposed infrastructure includes:

Mary River Mine Site (See Figure 5-1, 165926-6000-121-GAD-0100 – OVERALL MINE SITE PLAN)

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Waste dump Ore crushing and screening facilities Material handling system Ore stockpiling/reclaim facilities Truck loading and weigh scale facilities Offices and administration Permanent camp Assay laboratory Communication systems Heavy equipment fleet parking Airstrip and associated facilities Bulk fuel storage and distribution Explosive manufacture and storage (Contract) Maintenance shops Water supply (fresh water / fire water / industrial water) Water treatment facility(sewage treatment and sludge incineration) Power generation Emergency power Waste management facilities (solid waste and recycling)

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Figure 5-1: Overall Mine Site Plan

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Tote Road Mary River to Milne Inlet

Bridges / culverts Quarry and borrow sites Refuge stations Potable and raw water supply Sewage handling disposal during construction / operation Temporary fuel storage

Milne Inlet Port Site(See Figure 5-2, 165926-7000-121-GAD-0100 –MILNE INLET OVERALL SITE PLAN)

Access and site roads Temporary construction facilities Construction camp Contractor offices Quarry and borrow sites Temporary fuel storage Aggregate crusher and stockpiles Concrete batching plant Power generators Maintenance shops Warehouses/stores Laydown areas Freight dock Communication systems Truck receiving facilities Ore stockpiling Reclaim and materials handling facilities Ship loading facilities Floating ore dock Over winter berthing facilities Freight receiving and laydown area

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Mooring buoys Navigation lights and buoys Dredging Offices and administration Permanent camp Communication systems Airstrip and terminal facilities Bulk fuel receiving, storage and distribution Water supply (fresh water / fire water / industrial water) Water treatment facility (sewage treatment and sludge incineration) Power generation Emergency power Waste management facilities (solid waste and recycling) Maintenance shops Parking

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Figure 5-2: Milne Inlet Overall Site Plan

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5.4.3 Access

The proposed mine can only be accessed year-round by airplane. As part of the design, a runway capable of landing jet aircraft (Boeing 737-200) or turboprop aircraft (L382 Hercules) transport aircraft was considered. The runway will be located well away from permanent facilities on relatively flat topography to minimise the earthfill requirements and environmental impact. A modular construction terminal building will house the required security systems, and is capable of processing the passenger load of the Boeing 737.

5.4.4 Port

Fednav, a leading Canadian ship-owning and operating group, was selected by BIM to provide the shipping solutions for the Project. Fednav anticipates that it will do this by providing a combination of their Supramax (55,000 dwt) vessels and market Panamax (70,000 dwt) vessels. These vessels will arrive at Milne Inlet at first open water, on or about July 15th, ice permitting, and vessels will continue to arrive and be loaded until open water season ends, on or about October 31st. The ore dock will receive about 44 vessels during the 100 days open water season; about 14 per month.

Port operation practice and procedures will be similar to those at other large Canadian iron loading ports such as Sept Iles and Port Cartier

The ore dock proposed for use at Milne Inlet is a proprietary floating dock design by Group Ocean Inc. It will be used during the open water period only and will be removed to a secure protected storage area during freeze up for over-wintering.

5.4.5 Power

The estimated power demands for the mine, process, stockpile and infrastructure at Mary River are -5.25 MW. The current plan is to have four diesel generator sets, three operating and one spare to allow for maintenance, and an additional unit for emergency power supply.

The estimated power demand for the port is 5,304 KW. The power plan assumes four diesel generator sets, three operating and one spare to allow for maintenance, and an additional unit for emergency power supply.

Power station buildings will be of structural steel construction and each will have an overhead maintenance crane.

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5.4.6 Water

Camp Lake was identified as the fresh water source for Mary River. Milne Inlet has two different sources depending on the season. Philips Creek has been identified as the summer source. As this source freezes in winter, Thirty Two Kilometre Lake will be used as the winter source, with water trucked to Milne Inlet. This is practical as staffing at Milne is much reduced during winter as there is no shipping activity.

5.4.7 Communications

An existing two way radio communication system will be upgraded with additional base station and hand held units and antennas. for Mary River, Haulage and Milne Inlet communications. VHF radio and satellite telephone will be used for external communications..

5.5 Waste

Careful consideration was given to the choice of all ancillary facilities including the incinerator, Primary warehouse water treatment and sewage treatment, and glycol boiler building. Most of these facilities are to be housed in buildings, and some, such as the landfill area (for Mary River mine site only), must be remotely located from other permanent workplace facilities for reasons of health, safety and compliance with government regulations. There is no landfill at Milne Inlet port site; all landfilled items will be sent by truck to the Mary River site.

5.6 Comment on Section 5

In the opinion of the QPs:

Current access methods support exploration-level work programs. During planned mine development, additional access methods will be required. This will encompass marine transportation to and from Milne inlet i including sea lift transportation for equipment, materials and supplies, Panamax ore carrier transportation for product, and tanker transportation for fuel. . Upgrading of the existing Mary River airstrip to accommodate 737 Combi aircraft, and relocation of the Milne inlet airstrip for turboprop type aircraft will be required. The existing gravel road between Mary River minesite and Milne Inlet will be widened and upgraded to an all-weather access road.

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There is sufficient suitable land available within the mineral leases for any future mine infrastructure. No processing facilities such as a mill or tailings storage facilities are required under the envisaged mine plan. Project development plans have considered the availability of staff, power, water, infrastructure and communications facilities to suit the planned development requirements.

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C ONTENTS

6.0 HISTORY ...... 6-1 6.1 Ownership...... 6-1 6.2 Past Exploration ...... 6-1 6.3 Work in Progress ...... 6-5

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6.0 HISTORY

6.1 Ownership

In 1962 a total of 119 mining claims in four blocks were staked and an additional two prospecting permits over adjacent areas covering some 1470 km2 were acquired, followed by a further six claim groups. Three of the claim groups covered the areas now known as mining lease 2484 (Deposit No. 1); mining lease 2485 (Deposits Nos. 2, 3) and mining lease 2483 (Deposit No. 4). The claims were held by British Ungava Explorations Limited (Brunex), the original stakeholders.

A private company, Baffinland Iron Mines Ltd (BIML) had been established in 1963, organized by the financial participants and prospectors of the Brunex group, to hold the Mary River claims and leases, and to conduct the development of the prospects.

In 1963, an exploration program was undertaken by contract to the Toronto geological consulting firm of Watts Griffis and McOuat (WGM) of which Murray Watts was one of the founding principles. This program lasted through 1963 to 1966, with most of the field work carried out in the summers of 1964 and 1965.

In 1966, on completion of Brunex’s obligations, the three claim groups were transferred to BIML and taken to lease in 1971. At that time, management of the company (BIML) was taken over by Hudson Bay Mining & Smelting (HBM&S) supported by WGM. HBM&S was one on the founding investors in the Brunex syndicate.

In 2004, management of the private company was transferred to the current management group and through the reverse takeover of Glimmer Resources Inc., Baffinland Iron Mines Corporation was created Private placement and public share financings were completed to support initial funded operations and administration in 2004-2009. Stock ownership is now widely distributed.

6.2 Past Exploration

Exploration work by WGM between 1963 and 1966 (although no field work in 1966) included geophysical surveys, geological mapping, and diamond drilling of 30 holes (attempted and completed), with a total length of 3319 m at Deposit No. 1. This work required construction of a 100-km tote road to Milne Inlet, and construction of gravel airstrips near the Mary River base camp, at Milne Inlet and at Katiktok Lake, 40-km northwest of the base camp and near Deposit No. 4.

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Drilling was terminated after a total of 143 Mt grading 67.3% Fe had been identified by WGM in the central 1500 m of the Deposit No. 1, which has a total length of about 3800 m, based on outcrop and magnetic anomalies. A conceptual 300 Mt of potential ore defined by sporadic outcrops and some channel sampling in four other deposits was also recognized. The WGM tonnage estimates are provided for historical context only and are no longer valid due to subsequent infill and extension drilling by BIM.

This early work on the geology of the Mary River Area and the iron deposits, as well as the diamond drilling, resource estimate and metallurgical testing are discussed in more detail later in the following reports (WGM, 1963; WGM, 1964; WGM, 1965 and WGM, 1966 1 and 2).

Between 1966 and 2002, when the reformed BIM assumed control, no field work took place apart from the required land surveys completed in 1966 and filed in 1971. From time to time some metallurgical work and re-examinations of project economics were undertaken by Watts Griffis & McOuat (WGM) and HBM&S. Details of work and studies from this period are reported in the report by Strathcona Mineral Services (Von Guttenberg, R and Farquarson, G., 2003) and in the Aker Kvaerner Scoping Study (Edwards, F.A. et al, 2006).

Commencing in 2002, the newly formed BIM group reassembled and reviewed all available data and made plans for the 2004 summer season. In 2004, equipment was moved to site, a new camp was established, and some 2813 m of drilling in 15 holes were completed. This program increased the explored length of the Deposit No. 1 to over 2700 m, confirmed its vertical extent to 400 m (as measured from the crest of the ridge), and set the true width to be in excess of 400 m at the widest section. A new north-eastward trending limb was discovered above the hanging wall.

Twenty large samples of drill core collected from the 2004 drill program were sent to a laboratory in Germany for testing of the physical, geochemical and metallurgical properties of the potential ore. Results of this testwork are discussed in later sections of this report.

Exploration continued in 2005, with the completion of 34 more diamond drill holes totalling 8393 m. Work was advanced on mapping, and surveying. Additional metallurgical samples were collected from drill core and forwarded to Germany. The first exploration hole was drilled into Deposit No. 2.

A resource estimate was completed in 2006 as part of a scoping study completed by Aker Kvaerner. The resource estimate incorporated the WGM and BIM’s 2004 and 2005 drill results. Exploration work in 2006 was comprised of drilling on Deposit Nos. 1, 2 and 3. A total of 4135 m in 22 holes were drilled on Deposit No. 1 while 1172 m

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were drilled in 7 holes on Deposit No. 2 and 636 m were drilled in 3 holes on Deposit No. 3. Channel sampling in 2006 comprised of 12 samples or 67 m on Deposit No. 1.

Metallurgical mapping of Deposit No. 1 continued with a combined total of 209 samples forwarded to Germany by the end of 2006. A 9,338 metre drilling program in 2007 focused on geotechnical data collection and exploration/Geomechanical data collection. Drilling was also undertaken at the Milne Inlet and Steensby port sites, as well as along the Steensby railway alignment, intended to support project permitting and basic engineering, which was a new focus of the 2007 drilling program.

In the Deposit No. 1 2007 program, a total of 4,392 metres of drilling which includes both geotechnical and exploration drilling was completed. This drill program was comprised of 3 bulk sample pilot test holes on the north limb, a further 3 north limb infill holes, 3 south limb infill holes, 10 fold nose holes and 3 geotechnical-specific holes. An additional 88 core samples from 2006 Deposit No. 1 drilling were submitted for metallurgical testwork.

A total of eight additional holes on 450-metre step outs were drilled on Deposit No. 3 in 2007 for a total of 1,918 metres.

In February 2008, the Company announced the execution of a commercial lease of surface lands from the Qikiqtani Inuit Association (“QIA”) covering approximately 10,567 hectares of surface lands on Baffin Island, Nunavut. The lease permits the Company to use the subject area for the purposes of the exploration of minerals, the undertaking of engineering, geotechnical and environmental studies in support of the planning for a potential major mining development and to complete a bulk sampling program. The lease is valid from August 1, 2007 until October 31, 2009 and was renewed for an additional one year by the Company, subject to the terms and conditions of the lease.

In February 2008, the Company announced an updated mineral resource and mineral reserve statement for the Mary River Property completed by Aker Kvaerner E&C. The DFS is a detailed study of the technical and economic feasibility of Deposit No. 1 and is posted on Sedar. The DFS based on transporting ore by rail to Steensby Inlet indicates that, based on the shipment of 18 million tonnes of ore per year to the European market, there was sufficient tonnage to sustain a mine life of over 20 years.

In July 2008, the Company revised its bulk samples targets from up to 250,000 tonnes to a range of 120,000 to 150,000 tonnes. This target was revised primarily due to the advent of an earlier than anticipated spring melt combined with severe rainfall in the later part of June resulting in a need for extensive construction and maintenance work along the full length of the 100-kilometre tote road to the shipping site at Milne Inlet.

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Haulage of the bulk sample on this road at the time had been interrupted for approximately 9 weeks and resumed in late July 2008.

By November 2008, three shipments of the bulk sample program iron ore arrived in Europe. The first trial cargo of lump iron ore was shipped to the port of Vlissingen in the Netherlands for ThyssenKrupp. The trial cargo weighed approximately 54,464 tonnes by draft survey at the discharge port and the lump iron ore graded more than 68% iron with low moisture content of 1.28% and low levels of deleterious elements. The undersize (less than 6.3 mm) portion of the cargo was 3.4% of the cargo weight.

The second trial cargo of lump iron ore discharged approximately 31,050 tonnes at the port of Bremen, Germany for ArcelorMittal Bremen. This trial cargo also graded more than 68% iron with low moisture of 1.26% and low deleterious elements. The undersize portion of the cargo was 4.2%.

The third and final trial cargo was discharged at the port of Vlissingen in the Netherlands for ThyssenKrupp. This trial cargo weighed approximately 27,701 tonnes and graded more than 66% iron with moisture content of 3.32%.

The bulk sample open pit was selected to mine and ship lump iron ore that was representative metallurgically of what would be shipped in the initial 10 to 15 years of production. The three trial cargos of Mary River iron ore were consumed in the blast furnaces of ArcelorMittal and ThyssenKrupp in 2009.

For the year 2008, a total of approximately 12,411 metres of drilling were completed, with 5,071 metres of exploration and infill drilling on Deposit No. 1 in 27 holes, and 7,340 metres of geotechnical drilling in 442 holes. The focus of the program was to provide information and to improve confidence in the reserves and resources in the upper 250 metres of Deposit No. 1. More than half of the drill holes were drilled flat (less than 20 degrees) on the steep upper slopes of the Deposit to provide this information. For the 2008 metallurgical testwork program, a further 77 metallurgical samples from 2007 drill core and channel sampling were completed. The geotechnical drilling included 315 holes along the railway corridor, 64 holes at the Steensby port location, 33 holes at the quarry and tunnel locations along the railway corridor, and 30 infrastructure holes at the Mary River mine site location.

In October 2008, 18 mineral claims totalling approximately 32,016 acres (12,956 hectares) were staked. The claims surround Deposit No. 4 and cover prospective stratigraphy to the southeast.

In November of 2008, New-Sense Geophysics Ltd released results from a high resolution magnetic airborne geophysical survey. The survey was based on 200m spaced regional traverses with 100m infill lines surrounding the known deposit areas. A total of 6,184.9 line kilometres of survey were flown. The survey outlined additional

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stratigraphic targets in the vicinity of Deposit 4 as well as targets to the east and west of Deposit 3.

In 2009, enriched iron oxide mineralization was identified on the recently staked claims southeast of Deposit No 4 forming what is being referred to as Deposit No. 5.

During the summer of 2009, BIM drilled another 2317m on the south limb of Deposit No 1 to better define the south-eastern extent of mineralization. The previously interpreted extent of mineralization was confirmed with decreased in thickness at depth and a gradational increase banded iron formation with depth.

6.3 Work in Progress

In January 2010, the Company approved its 2010 exploration and development plans, which included a budget of $30 million in expenditures: (i) to fund a 2,200 metre drill program on Deposit No.4 and 5, (ii) complete updated resource estimation on Deposit No.1, (iii) continue to fund progressive reclamation, (iv) maintain regulatory compliance, (v) preserve core assets at site, (vi) advance the EIS. A decision to commence the trucking option feasibility study was made in mid 2010.

Work in the 2010 field season included diamond drilling on Deposits No 1, 2, 3, 4 and 5. Regional exploration led to the discovery of Deposits 6, 7, 8 and 9.

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C ONTENTS

7.0 GEOLOGICAL SETTING ...... 7-2 7.1 Regional Geology ...... 7-2 7.2 Local Geology ...... 7-6 7.3 Geology of the Mary River Iron Deposits ...... 7-7 7.3.1 Geology of Deposit No. 1 ...... 7-7 7.3.2 Geology of Deposit No. 2 ...... 7-11 7.3.3 Geology of Deposit No. 3 ...... 7-14 7.3.4 Geology of Deposit No. 4 ...... 7-15 7.3.5 Geology of Deposit No. 5, 6, 7, 8 and 9 ...... 7-16

F IGURES

Figure 7-1: Regional Geology Map of Baffin Island (Jackson et al, 2000) ...... 7-3 Figure 7-2: Geology of Northern Baffin Island ...... 7-5 Figure 7-3: Local Geology Map of Mary River Deposits...... 7-9 Figure 7-4: Deposit No 1 Geology Map ...... 7-10 Figure 7-5: Typical Section of Deposit No. 1 ...... 7-11 Figure 7-6: Geology of Deposit Nos. 2 and 3 ...... 7-12 Figure 7-7: Section of Deposit No. 2 ...... 7-13 Figure 7-8: Section of Deposit No. 3 ...... 7-15 Figure 7-9: Geology of Deposit No. 4 ...... 7-16

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7.0 GEOLOGICAL SETTING

7.1 Regional Geology

The Mary River Group is part of the Committee Belt, an assemblage of granite- greenstone terrains, rift basin sediments and volcanic rocks which lie within the northern Churchill Province and extend from southwest of Baker Lake for over 2000 km to north-western Greenland (Jackson and Berman, 2000). The Committee Belt is joined to the south by the approximately 1.9- to 1.8-Ga Baffin Orogen (see Figure 7-1). The Committee Belt has been divided into major assemblages, which include the following.

Archean-age banded granite migmatites and three or more phases of gneissic granitic intrusions, traversed by deformed amphibolite dikes. Ages 3.7 to 2.85 Ga, These are unconformably overlain by the Mary River Group. The units are strongly metamorphosed. Late-Archean Mary River Group; a diverse assemblage of metasedimentary and metavolcanic rocks, preserved in narrow, folded greenstone belts. Ages 2.76 to 2.72 Ga. Belts generally show a lower sequence of varied metavolcanics, overlain by metasedimentary-metavolcanic sequences including iron formation, succeeded by an upper group of metavolcanic and metapelitic clastic sedimentary units with high-level metamorphism. Paleoprotorozoic Piling Group; metasedimentary/metavolcanic sequence including quartzite, marble, sulphidic iron formation, black schists, mafic metavolcanics. Ages 1.9 to 1.8 Ga with medium-level metamorphism. Mesoproterozoic Bylot Super group, in the Borden Rift Basin; siliciclastic and carbonate sedimentary rocks, some mafic volcanic units. Age 1.27 Ga with low- level metamorphic facies. Early Palaeozoic Cambro-Ordovician (Turner Cliffs-Ship Point Formation); unmetamorphosed clastic and carbonate sedimentary rocks, locally preserved in north-westerly-trending grabens. Age 400 to 500 million years.

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Figure 7-1: Regional Geology Map of Baffin Island (Jackson et al, 2000)

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Mary River Group greenstone belts form isolated remnants in a roughly 65-km-long stratigraphically coeval volcanic series, which has been highly contorted, deeply eroded and forms a discontinuous series of volcanogenic units within older gneiss and migmatite formations (Figure 7-2). The belts consist most commonly of a lower sequence of felsic to mafic metavolcanic rocks and an upper sequence of turbiditic pelite-greywacke. The stratigraphic position of iron formation, quartzite, and conglomerate, minor marble and volcanic breccia may vary on a regional scale (Jackson and Berman, 2000). The group is well developed and may be up to 4000 m in total thickness.

The iron formations, of varied types and quality, occur discontinuously within the metasedimentary mid-section units of the Mary River Group. Their outcrop patterns are interrupted and truncated by original local sedimentary facies variations and by post-depositional folding, faulting and erosion. Only in a few locations, such as Mary River, do they reach thickness, quality and continuity sufficient to be of potential economic interest.

The regionally-extensive Franklin dolerite dike swarm generally follows the direction of major northwest-trending fault sets.

These major fault systems persist for hundreds of kilometres and are marked by fault- line valleys and scarps, and show evidence of very large vertical and horizontal displacements. One of these, the Central Borden Fault Zone, passes about 1 km to the south of the Mary River iron deposits. The fault separates the Mary River Group rock (~2.75-Ga age) on the northeast from the early Palaeozoic formations (Turner Cliffs and Ship Point formations) to the southwest.

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Figure 7-2: Geology of Northern Baffin Island

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7.2 Local Geology

The local geology is presented on Geological Survey of Canada Map 1451A “Geology of Icebound Lake, District of Franklin, 1:250 000 (Jackson G.D., Morgan, W.C., and Davidson, A. 1978). Closed-file maps of the geology at 1:50 000 scale are presented by Geological Survey of Canada, Geology of NTS Sheets 37G/5 and 37G/6 (1965) by G.D. Jackson. The geology of the Mary River area has been synthesized by BIM’s senior geologist, T Iannelli (Iannelli, T.R., 2005).

High-grade iron deposits were discovered in 1962 within a deformed granite- greenstone terrain at Mary River on northern Baffin Island, about 160 km south of Pond Inlet. Initial fieldwork in the 1962 to 1965 period outlined the presence of four exposed deposits (Deposit Nos. 1, 2, 3, and 4) of high-grade hematite-magnetite mineralization hosted within extensive belts of banded iron formation (Figure 7-3). Deposit No’s 1 to 3 occur within a single 30-km2 area, while Deposit No. 4 is situated 27 km to the northwest. The sedimentary-volcanic succession in which the iron formations are developed was designated as the Mary River Group by Jackson (1966).

The complexly folded Mary River Group greenstone belt, in the Mary River area, has a strike length of about 65 km and an inferred width of up to 6 km (Jackson et. al., 1978a). Rocks of the Mary River Group in the Project area included banded oxide and silicate facies iron formation, high-grade iron formation (hematite, magnetite and mixed hematite-magnetite-specularite varieties), mixed meta-sedimentary rocks (quartzite, metaconglomerate, metapelites, metagreywacke and related derived cordierite- staurolite-garnet-mica to quartz-feldspar-biotite schists and gneisses) (WGM, 1964, 1965; Jackson et. al., 1978; Jackson, 2000).

The iron formation-bearing assemblage, as outlined at Deposit No. 1, is stratigraphically underlain (to the west) by quartz-feldspar-mica gneiss (i.e., “quartz augen gneiss”) with minor interleaved bands of quartz-mica schist and quartzite, and overlain (to the east) by chlorite-actinolite schist and garnetiferous amphibolite. Thin bands of chlorite-actinolite schist, staurolite-garnet-mica schist, amphibolite, and banded iron formation occur interlayered within the high-grade iron formation across the strike width of the high-grade iron deposits (WGM, 1964, 1965). The iron formation assemblage may be as much as 400-m thick (Iannelli, 2005).

The Mary River iron ore deposits represent secondary enrichment of Algoma-type iron formation (Gross, 1996). The deposits are characterized by zones of massive, layered to brecciated hematite and magnetite, variably intermixed with banded oxide to silicate facies iron formation. They are spatially associated with a major northwest-trending discontinuity - the Central Borden Fault Zone. This crustal scale structure extends more than 200-km northwest from Angajurjualak Lake to Milne Inlet and southern Borden Peninsula (Figure 7-3). Field investigations by Duke (2010) suggest evidence

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for post-collisional extensional unroofing of the Hudsonian metamorphic terrain in north central Baffin, with keels of the Mary River Group bordering flow folded high grade gneissic domes. Duke and McCleod (2009) describe the cores of domes as interfolded Mesoarchean and Neoarchean granite-greenstone metamorphic assemblages refolded during the Hudsonian. Near contacts with the Mary River supracrustals, the gneiss is interleaved with melt sheets of Hudsonian augen granite. On the contact with the Mary River lithologies these augen granite sheets exhibit strong mylonitic foliation.

Duke (2010) indicates that the high-strain mylonitic boundary conditions between the domal gneisses and keels of the Mary River metavolcanics and metasediments are attributed to extensional unroofing of the Hudsonian gneisses, facilitated by domal rises and suprastructural sinks. The flow-folding within the domes is related to Hudsonian collision which ended around 1850 Ma. The development of mylonitic fabrics post-date collision, but predate the injection of non-mylonitized pegmatite bodies that probably date about 1,780 Ma.

Duke (2010) suggests a potential role for the mylonitic schist boundary in the secondary enrichment of the Mary River ores described in Section 8 - Deposit Types.

7.3 Geology of the Mary River Iron Deposits

7.3.1 Geology of Deposit No. 1

Deposit No. 1 is currently the largest defined iron deposit in the Mary River area. The deposit has a total strike length, as defined by outcrop and magnetic anomalies of about 3,800 m. Outcrops of high-grade iron oxides consisting of hematite and magnetite in various proportions and of specularite are exposed along the margin and crest of Nuluujaak Mountain at elevations ranging from 250 to 700 m, over a strike length of 2,500 m. The deposit is drilled off to the south; however, magnetics indicate the continuation of the iron formation for about 750 m to the north (WGM, 1964, 1965; Figure 7-3 and Figure 7-4).

Deposit No. 1 can be divided into an approximately 1,300 m-long northern portion (North Limb) which strikes at 041° and dips at -77° to the southeast, and an approximately 700m long southern portion (South Limb) which strikes at 316º and dips at -65° to the northeast (Figure 7-4). Minor extensions to the northeast and southeast vary locally in strike and dip. The limbs occupy the flanks of a steep north-easterly plunging syncline. Four stratigraphic lenses of high-grade mineralization are located within the fold nose. These include from footwall to hanging wall the 100, 200, 300 and 400 zones (Figure 7-4 and Figure 7-5).

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The high-grade lower 100 zone oxide iron formation forms a tabular, 105 m to over 150 m-thick body with chlorite–actinolite schist and/or garnetiferous amphibolite at the hanging wall and quartz-mica schist and quartz-feldspar-mica gneiss at the footwall. Bands of chlorite-actinolite schist with garnet and/or magnetite, banded oxide facies iron formation, and staurolite-cordierite-mica schist are rarely interlayered within the iron deposit particularly with the high-grade magnetite of the north limb. These waste zones are laterally continuous and average 1 to 15 m in thickness and separate the lower, middle and upper zones. High-grade hematite-dominated iron formation predominates along the south limb and core of the synformal structure, while specularite occurs adjacent to the site where the north limb is disrupted by a north- northwest trending fault.

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Figure 7-3: Local Geology Map of Mary River Deposits

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Figure 7-4: Deposit No 1 Geology Map

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Figure 7-5: Typical Section of Deposit No. 1

The iron formation along the south limb forms an assemblage of at least 290 m in thickness. The assemblage consists of two major sequences comprising a lower high- grade iron formation band (up to 120 m in true thickness) overlain by a mixed zone of variably alternating high-grade iron formation, banded oxide iron formation, chlorite- actinolite schist, amphibolite and cordierite-staurolite-mica schist layers 2- to 18-m thick. The lower high-grade iron formation band forms the surface outcrop along the crest of Nuluujaak Mountain.

7.3.2 Geology of Deposit No. 2

Deposit No. 2 outcrops on a ridge 2.6-km east of Deposit No. 1 (Figure 7-3). The deposit consists of dark steel- to blue-grey weathered high-grade specularite iron formation outcrops up to 40 m in width and 90 m in length and can be traced for over 500 m along strike. The deposit strikes west-southwest and grades into a south to

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south-southeast dipping up to 100-m-wide belt of banded oxide facies iron formation (Figure 7-6 and Figure 7-7). The banded iron formation belt trends westward for 900 m, where it is truncated by an interpreted northwest-trending fault. The high grade iron formation zone at Deposit No. 2 outcrops over an elevation of 610 to 670 m, strikes east-northeast – west-southwest to east – west and is characterized by sub-vertical dips to the south and south-southeast. Mapping, drilling results and correlation with the Mary River Group assemblage at Deposit No. 1, suggest that high-grade specularite iron formation at Deposit No. 2:

Stratigraphically overlies light grey pink-grey-brown weathered quartz-feldspar- mica gneiss and quartz-sericite to chlorite-amphibole-mica schist to the north and northwest. Is overlain by a sequence of green-grey-brown to dark green weathered chlorite- amphibole-mica schist with subordinate serpentinized mafic intrusive bands and banded oxide facies iron formation, in turn overlain by banded oxide facies iron formation (with minor high-grade magnetite plus specularite iron formation; upper banded iron formation zone) which is in turn overlain by chlorite-mica-quartz- feldspar schist (to the south and south-southeast).

Figure 7-6: Geology of Deposit Nos. 2 and 3

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Figure 7-7: Section of Deposit No. 2

The deposit has been drilled across a strike length of 500 m, extending east-northeast. An expanded strike length of 850 m can be tentatively interpreted based on occurrences of high-grade float towards the west and the extension of the high-grade zone eastwards along strike to the contact with the northeast-trending fault zone (at section line 567,000E) that separates Deposit Nos. 2 and 3.

Deposit No. 2 appears to be more complex than Deposit No. 1, in terms of continuity of high grade iron formation and continuity of thickness. The high-grade zone reaches a maximum (true) thickness of 90 m in drill hole MR2-06-82 and thins to 22 m within drill hole MR2-06-88 (300 m to the east-northeast). To the west-southwest, the high-grade zone consists of up to four bands of high-grade specularite to hematite iron formation interleaved with banded oxide facies iron formation across an interval of 79 m in (true) width in drill hole MR2-06-99. A zone of high grade magnetite to specularite iron formation, 17 m in true thickness, was encountered in the upper banded iron formation zone within drill hole MR2-06-85; the total strike length and depth extent remains to be determined. Typical analytical results, from 2006 drilling through the high-grade zone,

include grades of 66.3% Fe, 3.7% SiO2, 0.014% P, 0.01% S and 0.7% Al2O3 across a drill intersection of 119.8 m in drill hole MR2-06-82, and 66.2% Fe, 3.0% SiO2, 0.025% P, 0.01% S and 1.4% Al2O3 across a drill intersection of 44.4 m in drill hole MR2 06- 85.

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7.3.3 Geology of Deposit No. 3

As a result of the 2007 drill program, the historic nomenclature “Deposit Nos. 3, 3A and 3B” centered around groups of widespread outcrops along the same stratigraphy have been renamed to refer to a single zone, Deposit No. 3.

Deposit No. 3 is situated on the crest and lower slope of a ridge 670-m south of Deposit No. 2 (see Figure 7-6 and Figure 7-8). The deposit consists of high-grade hematite and specularite iron formation that occur at an elevation of 490 to 530 m. The iron formation which hosts Deposit No. 3 follows an east-northeast trending airborne magnetic anomaly which extends continuously approximately 8km. High- grade iron formation outcrops and float associated with a belt of Mary River Group metasedimentary and metavolcanic rocks can be traced along strike in outcrop and in aeromagnetic anomaly patterns intermittently for more than 15 km east-northeast and northeast to the Glacier Lake area (Baffinland Iron Mines Ltd., 1964, 1965; Jackson et. al., 1978a).

There is no outcrop evidence to suggest the continuation of high-grade iron formation northwards of Deposit No. 3 to connect to Deposit No. 2, as had been inferred from previous mapping (Jackson, 2000, 2006; Jackson et. al., 1978a). A single hole drilled in 2006 (Hole MR3-06-100; drilled due west at an inclination of 50º), located 320-m north-northeast of the western most outcrop at Deposit No. 3 and south of Deposit No. 2, failed to intersect iron formation. Instead, the drill hole intersected interleaved bands of granitic to granodioritic gneiss and chlorite-mica schist. The complex structural setting, across this area, suggests that continuity of the high-grade belt between the two deposits is not likely. However, additional drilling and detailed ground magnetometer surveys are planned to fully explore the nature of the north-northwest trending magnetic anomaly that lays midway between the two deposits and northwest of the site drilled in 2006.

Foliation in gneiss and schist and foliation plus relict banding in high-grade iron formation at Deposit No. 3 strikes east-west and at its western extent dips -88º to the north, while at its eastern extent the dip decreases to -67º towards the north. Drilling at Deposit No. 3 has shown that at the footwall contact, high-grade specularite iron formation is in contact with chlorite-amphibole to mica-sericite-chlorite schist which in turn is in contact with a thick sequence of quartz-feldspar-mica gneiss. Comparison with the correlative stratigraphic sequence noted at Deposit No. 2, suggests that the Mary River Group succession is overturned at Deposit No. 3. Thus, the high-grade iron formation assemblage is stratigraphically overlain (down slope) by silicate and/or banded oxide facies iron formation with minor chlorite-amphibole schist, which is in turn overlain by amphibolite (Jackson, 2000, 2006; Baffinland Iron Mines Ltd., 1964, 1965).

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Figure 7-8: Section of Deposit No. 3

Drilling, during the 2006 season in drill hole MR3-06-108, resulted in an incomplete drill intersection comprised almost entirely of remarkably homogeneous high-grade specularite iron formation that measured 169.8 m in length grading 65.8% Fe, 1.8%

SiO2, 0.03% P, <0.01% S and 1.1% Al2O3. This drill intersection represents a high- grade iron formation zone with an estimated true thickness of approximately 140 m (Figure 7-8). Drilling in 2007 on 450-m spaced drill sections indicated a strike length of at least 2,450 m of high grade iron formation, with the high-grade zone pinching out in thickness towards the east. At its western extent, however, the zone is interpreted to be terminated by a southwest-trending fault which is also interpreted to limit the eastern extension of Deposit No. 2 (Figure 7-6). Potential exists beyond the eastern- most extent of Deposit No. 3 for the discovery of additional targets of high-grade iron formation. Equally, potential exists beyond the western-most extent of Deposit No. 2 for discovery of additional targets of high grade iron formation.

7.3.4 Geology of Deposit No. 4

High-grade iron formation of Deposit No. 4 outcrops on a low ridge 27-km northwest of Deposit No. 1 and 3 km west of the Central Borden Fault zone (see Figure 7-3 and Figure 7-9). In this area, exposures of high-grade magnetite and specularite iron

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formation occur as a series of elongated lenses or bands, 5- to 75-m wide, which outcrop intermittently over a strike length of 2,800 m within the mine lease boundary. The exposures outcrop at a maximum elevation of 308 m and cover a combined surface exposure area of approximately 90,000 m2. The average grade of the deposit area based on 196 m of trench samples taken by WGM was determined to be 66.8%

Fe, 2.1% SiO2, 0.l08% P, 0.02% S and 0.7% Al2O3 (Baffinland Iron Mines Ltd 1965; von Guttenberg and Farquharson, 2003).

Figure 7-9: Geology of Deposit No. 4

7.3.5 Geology of Deposit No. 5, 6, 7, 8 and 9

The airborne magnetic survey completed in 2008 outlined a banded iron formation extending to the southeast, hosting Deposit No 5 approximately 5 km southeast of Deposit No 4. The deposit is comprised of hematite enriched iron formation. The largest outcrop can be traced for an approximate 700m in strike length and an exposed width of 70m.

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The deposit is comprised of hematite enriched iron formation. The zone can be traced in sporadic outcrops and with airborne magnetic for almost six kilometres and the average grade of twenty composite samples was approximately 67% iron.

As part of the 2010 regional exploration program 1970’s era aeromagnetic maps were used to identify a further 4 regional targets which resulted in additional significant discoveries of high grade massive and enriched magnetite and hematite deposits.

Deposit No 6, also known as Glacier Lake, is comprised of a zone of high grade massive magnetite, and enriched banded iron formation, exposed for 1km on surface. A coincident with an airborne magnetic high extends along a 5km east-northeast trending limb as well as along a 15km northwest-trending limb. The deposit is located 25km east-northeast of the Mary River camp. The exposed mineralization is up to 170m in width. A ground magnetic survey indicated that high grade magnetite is also exposed 12km southeast of the original discovery site.

Deposit No 7, also known as Turner River, is located 1 km north of Glacier Lake on the south side of Turner River. Mineralization is comprised of massive magnetite, mixed magnetite hematite and specular hematite. The main zone is exposed over a strike length of 700m and extends over a width of 100m. A second high grade zone 1.3km to the southeast is comprised of a 5-10m thick zone approximately 300m in strike length.

Deposit No 8 known as North Cockburn River, is in an area 95km east of Mary River and extends along the north drainage basin of the Cockburn River system. Exposed mineralization is comprised of high grade, coarse, hard, specular hematite bounded by coarse magnetite to mixed magnetite-hematite iron formation. The exposed high grade zone extends 400 m in strike across a width of 50-60 m. A second high grade zone 500m to the east, is comprised of an exposed zone of high grade 600m along strike and 40-50m in width.

Deposit No 9 known as Rowley River, is approximately 20-25 km east of Deposit 8 and 115 to 120 km southeast of Deposit No 1. Aeromagnetic surveys indicate a strong magnetic anomaly that extends 15km. Exposures consist of high grade and enriched magnetite mineralization 10’s of metres in thickness occupying several stratigraphic intervals.

The knowledge of the geological setting within Deposit No. 1 is sufficient to support a Mineral Resource and Mineral Reserve estimate. The geological data and interpretations available for Deposit No’s 2 and 3, support estimation of Mineral Resources.

Interpretation of the geology is currently underway on Deposit No’s 4 and 5. The intent is to include these in a Mineral Resource estimate in 2011.

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The Deposit Nos. 6, 7, 8, and 9 are at an earlier stage of exploration, and the geological setting is currently insufficiently understood to support estimation of Mineral Resources.

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C ONTENTS

8.0 DEPOSIT TYPES ...... 8-1

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8.0 Deposit Types

The Mary River deposits can be described by the depositional model proposed for Algoma-type iron formation. The characteristic stratigraphy for Algoma-type iron formations comprise of a lower succession of typically intermediate volcanics interbedded with acid volcanics, volcaniclastic sediments, and their sedimentary derivatives. This succession is followed by banded iron formation comprised of alternating cherty bands and iron oxides or bedded or banded pure oxides ranging from metres to hundreds of metres in thickness, depending on depths of the original basins. This succession in turn is overlain by a sequence of volcanics and volcaniclastic sediments often intruded by ultramafics and granitic intrusive bodies.

Mary River Group greenstone belts across northern Baffin Island consist of variable amounts of banded iron formation (BIF), which are associated with metasedimentary, metavolcanic and metaigneous rocks (Jackson et. al., 1978a, b, c; Jackson and Morgan, 1978). The banded oxide and silicate facies iron formations comprise the lateral equivalents of the high-grade Mary River deposits, which are almost entirely composed of iron oxides, suggesting a significant enrichment of iron through secondary processes.

Duke (2010) suggests growing evidence for a secondary Fe enrichment mechanism associated with widespread retrograded mylonitic detachments along the base of the high grade and enriched BIF across north central Baffin Island in places such as the Mary, Turner, Cockburn and Rowley River areas.

These detachment faults are marked by a chlorite retrograded mylonitic structural contact between underlying ductile rising domes of polydeformed mafic-felsic gneisses and overlying supracrustal sinking keels which host the Mary River iron deposits (Duke and McCleod, 2009). These detachments are thought to be post regional folding of the Hudsonian orogeny around 1,850 Ma (Duke, 2010).

The high grade massive and enriched iron oxide ores occur where Mary River BIF border directly on detachment schist that separates the mylonitic roof of infrastructural domes from the high strained keels of Mary River Group. At detachment schist contacts, magnetite BIF shows evidence of undergoing extreme pressure solution resulting in the highly efficient leaching of quartz from magnetite. In large deposits, the primary banding defining early isoclinal folding is completely destroyed by metamorphic differentiation and recrystallization, resulting in >100m thicknesses of massive magnetite/hematite. On the flanks of the high grade iron deposits, Fe- enrichment is marked by various degrees of silica loss due to continued silica removal. Several regional examples of exposed giant quartz veins occurring along detachment schist boundaries adjacent to high grade magnetite zones account for the removal of silica from the banded iron formations. Duke (2010) suggests that the hydrothermal

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removal of silica was likely facilitated by dewatering of the nearby footwall serpentinized komatiites.

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C ONTENTS

9.0 MINERALIZATION ...... 9-1

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9.0 MINERALIZATION

The Mary River deposits are comprised of a number of iron formations which have been enriched and altered to varying degrees. Original banded iron formations comprised of alternating layers of magnetite and hematite, are preserved in several locations along strike of existing high grade deposits. The regional metamorphism and folding associated with the Hudsonian Orogeny resulted in significant crustal thickening. This period also resulted in zones of weak to very efficient leaching of silica. Subsequent hypogene and metamorphic events led to the alteration of magnetite to hematite and specular hematite. Surface outcrops differ in iron, silica and sulphur content and in the proportions of their main oxide minerals – hematite, magnetite and specularite.

The microscopic examination of drill core specimens by R.A. Blais (1964) suggests that magnetite was the stable mineral of the iron formations during metamorphism (amphibolite facies), and was replaced to various degrees by hematite (martitization), resulting in a range of hematite-magnetite (martite) compositions. This alteration process occurred during a late stage of metamorphism, under conditions of localized stress and high oxygen pressure, and may have been associated with folding and faulting of the tabular high-grade magnetite iron formation.

The strong structural influence on channelling of oxidizing solutions is supported by the alignment of lineations produced by tubular pores in the hard high-grade hematite iron formation (i.e., parallel to the axial plane and plunge of the synform) and the concentration of high-grade hematite ore at the core of the fold.

Euhedral specularite crystals in magnetite rock indicate that specularite did not form by a process similar to martitization, but through the possible metamorphic replacement of magnetite. It is also possible that the specularite and other hematite mineralization formed as a primary rather than secondary mineral. However, what caused the crystallization of flaky, course-grained specularite as compared to fine-grained massive hematite is not fully understood. Field evidence suggests that higher fluid or vapour pressures in zones of shearing may have facilitated the growth of specularite.

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C ONTENTS

10.0 EXPLORATION ...... 10-1 10.1 History of Relevant Exploration Work ...... 10-1 10.2 Exploration Work by WGM ...... 10-1 10.3 Exploration Work by BIM (2004 to 2010) ...... 10-1

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10.0 Exploration

10.1 History of Relevant Exploration Work

The exploration history of the Mary River iron deposits can be divided in the work completed during the initial discovery years from 1962 to 1965, mainly by WGM and the work programs carried out since 2004 by BIM. Since the revival of interest in the Project, BIM has completed seven drill campaigns in 2004, 2005, 2006, 2007, 2008, 2009 and 2010 on Deposit No.’s 1, 2, 3, 4 and 5. The scope of this report covers work completed up until the end of the 2009 field season for Deposit 1 and up until the end of the 2007 field season for Deposit No’s 2 and 3. Assays for 2010 were still pending. Regional exploration work commenced in 2010 with the discovery of 4 new deposits of high grade mineralization, Deposit No’s 6, 7, 8 and 9.

10.2 Exploration Work by WGM

Exploration by WGM was comprised of magnetometer surveys, geological mapping, channel sampling of all four initial deposit areas and diamond drilling of Deposit No. 1. Additional engineering work included metallurgical testwork, tote road construction between Milne Inlet and the Mary River Camp, construction of three airstrips to support exploration, topographical surveys, hydrographic surveys, engineering studies of facilities and services, terminal/ship loading studies, ocean shipping studies, an investigation into sources of hydroelectric power, and soil testing for road/rail construction.

A total of 26 holes with a combined length of 3,319 m were drilled at Deposit No. 1 in 1964 and 1965. Only 15 holes were completed as planned, with the remaining 11 holes either lost (9) or could not be completed (2). Drill-hole assay data for the WGM

sampling program was limited to Fe, SiO2, and S. WGM also collected channel samples representing 1,030 m. of sample returning high grades (>60% Fe) on Deposit No’s 1, 2, 3 and 4. WGM also completed magnetometer surveys indicating that magnetic anomalies were consistent with high-grade outcrops, suggesting significant potential strike extent for each of the deposit areas.

10.3 Exploration Work by BIM (2004 to 2010)

Drilling by BIM in 2004 was focused on extending the northern and southern extents of mineralization previously defined by WGM drilling on Deposit No. 1. A total of 2,814 m of drilling was completed. In 2004, a total of 68 composite samples (60- to 120-kg)

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were collected from the remaining drill core and forwarded for metallurgical testwork at SGA in Germany. Each composite represented approximately a 14- to 16-m drill core interval. The drill program was successful in extending the northern and southern extents of Deposit No. 1.

The first drill hole (MR2-04-42) was drilled into Deposit No. 2 intersecting approximately 85 m true width of high-grade iron formation indicating that further drilling on Deposit No. 2 was warranted. A further seven metallurgical samples were collected from core and two metallurgical samples from outcrop in Deposit No. 2 and also forwarded to SGA in Germany for testwork.

In 2005, 34 holes were completed for a total of 8,073 m. The drill program was largely aimed at testing the depth extent of Deposit No. 1 to delineate the footwall contact. Drilling in Deposit No. 1 was successful in indicating that the surface widths encountered by WGM extended to approximately 600 m in depth with no indication of pinching out at depth. An additional 135 metallurgical composites collected from Deposit No 1 drill core were dispatched to SGA in Germany for metallurgical testwork. Another 15 metallurgical samples from surface outcrops on Deposit No. 1 were also forwarded to SGA in Germany.

An aerial topographic survey was completed by Eagle Mapping Services to produce a digital terrain map of the area covering Deposit Nos. 1, 2, and 3.

According to BIM, the 2006 drilling program was comprised of 7,067 m of drilling which included exploration, infill drilling, as well as geotechnical drilling for pit walls and infrastructure.

Deposit No 1 drilling in 2006 was comprised 22 drill holes (4,137 m). Thirteen holes were focused on the north limb, one on the south limb, four on the fold nose and the remaining four holes were comprised of geotechnical pit wall holes.

Exploration in 2006 also commenced on Deposit Nos. 2 and 3. Seven holes were drilled at Deposit No. 2 on sections spaced 100-m apart. At Deposit No. 3, 3 holes were drilled. One of these three holes, MR3-06-108, intersected a drilled width of 169.8 m of massive high-grade specular hematite, suggesting a true thickness of at least 140 m at the western extent. An additional 17 composited core samples from Deposit 2 and an additional 10 composited core samples from Deposit 3 were submitted for metallurgical testwork.

In 2007, BIM completed a total of 9,338 m of drilling which included exploration and infill drilling as well as geotechnical drilling of pit walls, potential infrastructure locations, proposed port facilities and planned railway alignments.

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Exploration drilling in 2007 and 2008 on Deposit No. 1 was targeted towards drilling the upper 250 m in elevation, in order to provide metallurgical material, upgrade interred resources and to obtain a more complete deleterious database for the upper portion of deposit; the portion which would provide the raw material for the first years of mine life. This program was only partially completed in 2008 and continued in 2009.

In the Deposit No 1 2007 program, BIM reports the completion of a total 4,392 m of drilling which includes both geotechnical and exploration drilling. This drill program was comprised of 3 bulk sample pilot test holes on the north limb, a further 3 north limb infill holes, 3 south limb infill holes, 10 fold nose holes and 3 geotechnical-specific holes. An additional 87 core samples from 2006 were submitted for metallurgical testwork.

A total of eight additional holes were drilled on Deposit No. 3 in 2007 for a total of 1,918 m on 450-m step outs. A further eight channel samples for a total of 82 m were collected in 2007. The drill program was successful in defining an inferred resource over a 2,450-m strike length.

In 2008, a total of 5,071 m were drilled in 27 infill holes. This program was targeted towards improving the confidence level of the upper 250m of the deposit. Drill results indicated a more localized control on deleterious elements and also indicated in some locations a greater thickness of near surface high grade iron formation than had been previously interpreted.

For the 2008 metallurgical testwork program, a further 54 metallurgical samples composited from the 2007 Deposit No 1 drill core were sent to SGA as well as another 23 composited 2007 core and outcrop samples for Deposit No 3.

During 2008, New-Sense Geophysics was contracted to complete a regional airborne magnetic survey covering all of the Baffinland leases and intervening areas. The survey was comprised of 6,185 line kilometres using a helicopter mounted high sensitivity Cesium magnetometer at 200 and more detailed 100m line spacing (Yukovenko, 2008).

In 2009, BIM continued drilling with a focus on the further defining the extent of mineralization within the south limb of Deposit No 1 in order to confirm the optimal potential crusher location. Drilling was comprised of 13 drill holes for a total of 2,317m. Drill results indicated a flattening of the dip towards the north and a gradational transition to a thinner banded iron formation at depth. Drilling also resulted in a thickening of the high grade mineralization near surface over previous interpretations.

In 2010, 17 drill holes comprised of 3,170m and drilled on 300m spaced sections were targeted on the strike extent of Deposit No 4. NQ holes ranging in depth from 104 -

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299m were drilled at a -50 degree dip towards the north. Drilling at Deposit No. 5 comprised 2,669 meters in 20 holes across an approximate 5 kilometre strike length of the deposit. Drilling at Deposit Nos. 2 and 3 were to better define geometry, while drilling at Deposit No. 1 was in support of the environmental assessment process, specifically to fill in potential gaps in Acid Rock Drainage testwork. In total, there was 6,802 metres of drilling in 45 drill holes. Results were still pending at the time of this report and therefore are not included herein.

Regional exploration traverses coincident with airborne magnetic anomalies in 2010 identified a further four deposits (6, 7, 8 and 9).

Exploration results from BIM were successful in that they allowed for the definition of continuous zones of high-grade mineralization for Deposit Nos. 1, 2 and 3, and the subsequent estimation of mineral resources for these deposits. Exploration was also successful in defining further mineralization at Deposits 4 and 5. Regional work also led to the discovery of a further four Deposits 6, 7, 8 and 9.

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C ONTENTS

11.0 DRILLING ...... 11-1

T ABLES

Table 11-1: Database Mary River Deposit No. 1 ...... 11-1 Table 11-2: Database Mary River Deposit No. 2 ...... 11-1 Table 11-3: Database Mary River Deposit No. 3 ...... 11-2 Table 11-4: Database Mary River Deposit No. 4 ...... 11-2 Table 11-5: Drill Core Recoveries ...... 11-3

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11.0 Drilling

The BIM exploration database is comprised of WGM exploration data completed in the 1960s, as well as 6 successive years of exploration by BIM. Table 11-1, Table 11-2, Table 11-3 and Table 11-4 summarize the exploration database completed on Deposit Nos. 1, 2, 3 and 4 by the end of 2009. A complete list of 2010 drilling is pending and is therefore not included in the following tables.

Table 11-1: Database Mary River Deposit No. 1

Database Mary River Deposit No. 1 Year Company Type No. of Collars No. of Metres 1964 WGM DDH 26 3 318 2004 Baffinland DDH 12 2 350 2005 Baffinland DDH 34 8 393 2006 Baffinland DDH 22 4 136 2007 Baffinland DDH 22 4 492 2008 Baffinland DDH 27 5 071 2009 Baffinland DDH 13 2 317 Total 156 30 077

1963 WGM CHA 208 1 023 2006 Baffinland CHA 12 67

Table 11-2: Database Mary River Deposit No. 2

Database Mary River Deposit No. 2 Year Company Type No. of Collars No. of Metres 2004 Baffinland DDH 1 122 2006 Baffinland DDH 7 1193 Total 8 1315

1963 WGM CHA 22 130

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Table 11-3: Database Mary River Deposit No. 3

Database Mary River Deposit No. 3

Year Company Type No. of Collars No. of Metres 2006 Baffinland DDH 3 636 2007 Baffinland DDH 8 1 918

1963 WGM CHA 19 97 2007 Baffinland CHA 8 82

Table 11-4: Database Mary River Deposit No. 4

Database Mary River Deposit No.4

Year Company Type No of Collars No of Metres 1963 WGM CHA 32 191

(DDH:Diamond drill core; CHA: channel sampling.)

Drilling was completed using mainly HQ (core diameter 63.5 mm), in order to improve core recovery and to provide sufficient sample for BIM to metallurgically map Deposit No. 1. All hole collars were surveyed on an ongoing basis and down hole surveys are completed using a Maxibor instrument which collects dip and strike data of the entire length of the hole. Where drilling became difficult, holes were extended using NQ equipment (core diameter 47.6 mm). Drilling was completed using Longyear L-38 and LM-30 modular rigs.

Drill core recoveries generally ranged between 92 to 99% for the BIM drilling and were found to be appropriate for resource estimation (Table 11-5). Occasional intervals of poor recovery were a result of broken or fragmentary hematite or magnetite generally related to apparent zones of faulting or shearing, friable, porous hematite within the fold axis; or fractured or brecciated waste rock above the hanging wall.

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Table 11-5: Drill Core Recoveries

Drill Core Recoveries Year Company Recovery % 2004 Baffinland 99 2005 Baffinland 96 2006 Baffinland 95 2007 Baffinland 92 2008 Baffinland 98 2009 Baffinland 98

1964 WGM 88 1965 WGM 70

The quantity and quality of the sampling and drilling procedures are sufficient to support a Mineral Resource and Mineral Reserve estimate.

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C ONTENTS

12.0 SAMPLING METHOD AND APPROACH ...... 12-1 12.1 WGM – 1964/65 ...... 12-1 12.2 BIM 2004 to 2009 Exploration Campaigns ...... 12-2 12.3 Drill Core Handling Procedures ...... 12-2 12.4 Metallurgical Sampling Procedure ...... 12-3

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12.0 SAMPLING METHOD AND APPROACH

12.1 WGM – 1964/65

Surface chip and trench samples in the 1964 to 1965 WGM programs were taken on sections perpendicular to the strike of the deposits. The sample weight was approximately 0.5 lb/ft (745 g/m).

Deposit No. 1 was channel sampled at approximately 150-m intervals and also on some intermediate sections for a total of 208 samples and a total sample length of 1,034m.

Deposit No. 2 was channel sampled at approximately 50-m intervals where outcrops permitted over a strike length of 335 m for a total of 22 samples and a total sample length of 130m.

Deposit No. 3 was channel sampled in only two general areas of outcrops spaced approximately 1.5-km apart for a total of 19 samples and a total sample length of 96m.

The surface samples at Mary River provide a limited representation of the composition and distribution of iron oxides in the underlying deposits. Surface sampling is limited by outcrop exposure, which only provides a limited exposure of the true thickness of the cross strike extent of the iron formation. Glacial debris mixed together with frost- heaved bedrock make it difficult to discern outcrop from float. Outcrops vary from either massive bedrock to highly fractured rock.

Core from the drilling programs at Deposit No. 1 in 1964/65, with visible iron oxides, was split longitudinally with one-half stored at Mary River and the other half shipped for assaying to Technical Service Laboratory (TSL) in Toronto, or to Warnock Hersey Company Limited in Montreal. Initially, core containing little waste material was sampled and assayed in 6-m lengths, later the sample length was reduced to 3 m and different types of iron ore and bands of waste were sampled and assayed separately. A total of 701 samples with a combined length of 2,728 m from 25 drill holes were assayed, with lengths of individual samples varying from 0.3 to 16.8 m, with an average of 3.9 m. The core samples show no bias and reflect the entire ore zone thickness. (The above notes are adapted from von Guttenberg and Farquarson, 2003.)

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12.2 BIM 2004 to 2009 Exploration Campaigns

The sampling method employed in 2004, 2005, 2006, 2007, 2008 and 2009 by BIM was core drilling which was predominantly contracted to Boart Longyear Inc. Drilling on Deposit No. 1 was targeted towards eventually achieving a drill density of 75 m between drill holes along strike and down plunge. Drilling on Deposit No. 2 is currently drilled at 100 m spacing, while drilling on Deposit No. 3 is limited to 450-m spacing between drill sections.

A limited amount of surface channel sampling was completed on Deposit No. 1 in 2006. A total of 48 channel samples covering 48 m were sent to SGS for assay. Channel samples were cut with a diamond saw and sample volume per metre roughly equated that of HQ diameter core.

An additional 11 channel composites weighing approximately 200 kg each were sent for metallurgical testwork on Deposit No. 1 and another 2 approximately 200-kg channel sample composites were sent for metallurgical testwork on Deposit No. 2. Channel sampling on Deposit No. 3 in 2007 was comprised of eight channel samples for a total of 82 m.

12.3 Drill Core Handling Procedures

The HQ-sized drill core was logged and 2-m sample lengths were marked for sampling and assaying. Sample lengths were less than 2 m when intervals were encountering changes in lithology. The core was cut using a diamond saw. Samples were collected not only for ore grade material but also for internal waste, hanging wall and footwall, as well as adjacent banded iron formation, in order to assess the grade impacts of waste deleterious elements and dilution. Remaining core samples were stored in core boxes on site for future reference or for metallurgical testwork. Sample mass for HQ sized core ranged from 11 to 15 kg, while sample mass for NQ sized core ranged from 6 to 7 kg.

Drill core logging captured major and minor rock types, core recovery, rock quality distribution, presence of sulphides and structure. Geological logging of drill core was completed on laptops to reduce transcription errors. Core was photographed prior to sampling.

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All samples were air-shipped in sealed containers to SGS laboratories in Sudbury. No evidence of tampering of samples was encountered. Sample assay values forwarded to different laboratories returned similar values.

12.4 Metallurgical Sampling Procedure

BIM surpasses industry standards of many iron ore development projects in Canada and in Brazil in the matter of the degree of deposit-wide metallurgical mapping in preparation for economic analysis. Metallurgical mapping of the deposit is an effective tool used by some major iron ore mining companies to reduce project risk by providing a full suite of metallurgical blast furnace characteristics for the entire deposit. Approximately 209 drill core samples and a further 17 samples were collected from surface outcrops in the 2004 and 2005 exploration programs. During 2006, a further 87 core samples were selected from Deposit No 1, 17 core samples were selected from Deposit No 2 and 10 core samples were selected from Deposit No 3. Another 53 near surface core samples on Deposit No. 1, 18 core samples from Deposit No. 3 and 6 surface samples from Deposit No. 3 followed in 2007. This type of deposit wide metallurgical testwork allows for an assessment of the anticipated range of blast furnace characteristics and facilitates significantly improved quality control of a blended product.

These samples were comprised of approximately 10-m intervals of core, weighing approximately 200 kg each, and were assayed for a full suite of deleterious elements, as well as size fraction analysis, mechanical characteristics (physical strength, wear and impact tests, density) and blast furnace characteristics such as reduction under load, reducibility, porosity, disintegration, abrasiveness, tumble strength and decrepitation.

All collection, splitting, and bagging of core samples were carried out by company personnel, with the company and personnel varying depending on the date of the drill program. No factors were identified with the drilling programs that could affect the reliability of the sample data used for Mineral Resource and Mineral Reserve estimation.

The sampling methods are acceptable, meet industry-standard practice, and are adequate for Mineral Resource estimation purposes.

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C ONTENTS

13.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY ...... 13-1 13.1 Sample Preparation and Assaying ...... 13-1 13.1.1 WGM (1964/1965) ...... 13-1 13.1.2 BIM 2004 to 2009 ...... 13-2 13.1.3 Assay Procedures ...... 13-2 13.1.4 Whole Rock Analysis ...... 13-2 13.2 Quality Controls and Quality Assurance ...... 13-4

T ABLES

Table 13-1: Detection Limits and Mean Grades of Deleterious Elements ...... 13-3

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13.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY

13.1 Sample Preparation and Assaying

13.1.1 WGM (1964/1965)

In 1964, assays were completed on chip and trench samples from the Mary River deposits for total iron, phosphorous and sulphur by standard wet chemical procedures. Silica, alumina, manganese, and titanium were determined by spectro-chemical methods with an accuracy of ±5%.

Samples of half core were crushed to -6 mm in a steel-plate crusher, a 300-g split of the crushed sample was pulverized to -10 mesh (1.7 mm), and 5 g assayed for soluble iron, silica and sulphur by wet chemical methods. For an accurate identification of ore types, a 100-g split of the -10 mesh material was used for standard Davis Tube analysis, which was the main analytical procedure used to classify the samples based on their hematite-magnetite content.

To chemically characterize the iron ore types separated by core logging and Davis Tube tests, 45 composite samples from 10 holes (S-2, -4, -5, -7, -7A, -8, -9, -10, and -

11) drilled in 1964 were analyzed for soluble Fe and for SiO2, Al2O3, MnO, TiO2, CaO, MgO, S, P, and LOI). Composites (10) of these samples were analyzed for 50 elements by spectrography.

Based on composite samples and assays from the 1964/65 drilling programs by spectrography and whole-rock analysis, (Technical Services Laboratories, Superior

Laboratories), other elements or oxides, such as Al2O3, Cu, Pb, Zn, Mn, Ti, V, Na and K all occur in amounts which would not affect the marketability of the iron ore.

The analytical work showed that iron levels and accessory elements generally fell within acceptable limits for lump iron ore in all composites, except waste-rock units. Sulphur in five composites exceeded the normal limit of 0.05%, and four magnetite-rich composites carried phosphorus in excess of the normally accepted level of 0.04%.

Comparative assaying was the only quality control measure discussed in the reports, from 1963 to 1965, describing the assaying of core and surface samples. Original laboratory reports for that period are no longer available. The laboratories used for assaying of core, surface and metallurgical samples in 1963 to 1965 were not certified which was not unusual at that time.

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The areas of concern related to the analytical work from that period, which may influence the quality of the resource, are the accuracy of the sulphur assays and the determination of the magnetite content by the Davis Tube magnetic separator.

The above descriptions of WGM sample preparation and assaying were adapted from von Guttenberg and Farquarson (2003).

In general, the sampling methodology, sample preparation and analytical procedures were adequate for the definition of the Mary River mineralization.

13.1.2 BIM 2004 to 2009

Sample preparation of split core was completed at SGS laboratories in Sudbury, Ontario. Prepared pulps were then forwarded to SGS in Lakefield for assay. Preparation of samples included a primary crush of the sample to -43 mm maximum. This was followed by a secondary crush to 34 mm maximum. Samples were then screened at 6.3 mm, weighed and recombined. Samples were then riffle split to 500 g for tertiary grinding and pulverizing. A tertiary crush of the 500-g split to 3 mm completed the crushing preparation. In 2005, the secondary crush to -34 mm was eliminated as analysis of screened fractions completed by SGA in Germany showed little or no variance. The elimination of the secondary crush also provides a more representative assessment of the lump ore percentage. A 500 gram riffle split sample was then pulverized to 150 mesh.

13.1.3 Assay Procedures

Analytical Determinations

H2O Moisture content was determined at 105° on -3mm material by heat and weight loss.

Fe Total iron by potassium dichromate titration FeO Non-oxidizing acid leach, followed by potassium dichromate titration Cl UV spectrophotometry S Total sulphur by Leco carbon sulphur analyser 13.1.4 Whole Rock Analysis

SiO2, Al2O3, P2O5, Na2O, TiO2, V2O5, CaO, MgO, K2O, MnO. Method: Borate fusion followed by X-ray fluorescence

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Detection: All elements 0.01%, except Na2O and K2O (0.05%). Loss on Ignition (LOI) is determined by high temperature weight loss.

Minor and trace element determinations were made by inducted coupled plasma spectrometry, using acid digestion. A suite of ICP analyses were also evaluated to see if any elements were potentially problematic. Table 13-1 of raw assay data from the 2004 and 2005 drill programs indicate that a bulk of these deleterious elements are either below the detection limit or proximal to the detection limit, suggesting that they are not present in sufficient quantities to negatively affect ore quality.

Elements and oxides that do present themselves at elevated quantities locally within

the Mary River deposits include MnO, MgO, P2O5, S and Al2O3.

Table 13-1: Detection Limits and Mean Grades of Deleterious Elements

Grade Parameter Units Detection Limit Mean As g/t <40 BD Ba g/t <.5 5.84 Bi g/t <60 BD Cd g/t <10 BD Co g/t <10 25.16 Cr g/t <4 18.12 Cu g/t <0.5 15.1 Mo g/t <10 BD Sn g/t <20 BD Tl g/t <30 BD Zn g/t <50 BD

Na2O % <0.002 0.007

TiO2 % <0.01 0.04

V2O5 % <0.01 BD CaO % <0.01 0.4

K2O % <0.003 0.01

BD = Below Detection Limit

For the purposes of resource estimation, MnO% and P2O5% in the assay certificates were recalculated by G. Wahl to provide Mn% and P% values in the BIM database.

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13.2 Quality Controls and Quality Assurance

SGS is accredited by the Standards Council of Canada (SCC) for specific mineral test listed on the scope of accreditation to the ISO/IEC 17025 standard. ISO/IEC 17025 addresses both the quality management system and the technical aspects of operating a testing laboratory.

Quality control procedures by SGS include duplicate samples, replicates, reagent/instrument blanks, preparation control samples, certified reference material analysis and instrument control samples.

Security was achieved by BIM restricting access to samples on-site, rigorous sample tracking, tamper-proof packaging and secure shipment routes.

After analysis, results were sent electronically by SGS to BIM offices in Toronto. Signed certificates were later forwarded by mail. Pulps were retained in storage at SGS for future reference as required. Core is stored at site.

The quality of the iron and deleterious element analytical data are sufficiently reliable to support Mineral Resource and Mineral Reserve estimation. Sample preparation, analysis, and security are generally performed in accordance with industry practices and standards.

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C ONTENTS

14.0 DATA VERIFICATION ...... 14-1 14.1 WGM 1964 and 1965 Assay Verification ...... 14-1 14.2 BIM 2004 to 2009 Database and Assay Verification ...... 14-2 14.3 SGS Internal Blank, Duplicate Pulp and Standard Results ...... 14-3 14.4 Quantile Plots ...... 14-3 14.5 Twinned Drill Holes ...... 14-3 14.6 External BIM Duplicate Samples ...... 14-5

T ABLES

Table 14-1: WGM Comparative Analysis ...... 14-1 Table 14-2: Comparison of Twin Drill Holes S-2 and MR1-05-67 ...... 14-4 Table 14-3: Comparison of Twin Drill Holes S-5 and MR1-05-63 ...... 14-4 Table 14-4: Comparison of Twin Drill Holes S-6 and MR1-05-60 ...... 14-5

F IGURES

Figure 14-1: Fe% HARD Duplicates ...... 14-5 Figure 14-2: Fe% Duplicates ...... 14-6 Figure 14-3: Fe% Duplicates >40% ...... 14-6 Figure 14-4: Fe% HARD Duplicates ...... 14-7 Figure 14-5: Fe% Duplicates ...... 14-7 Figure 14-6: Fe% Duplicates <40% ...... 14-8 Figure 14-7: SiO2% HARD Duplicates ...... 14-8 Figure 14-8: SiO2% Duplicates ...... 14-9 Figure 14-9: SiO2% Duplicates <15% ...... 14-9 Figure 14-10: AI203% HARD Duplicates ...... 14-10 Figure 14-11: AI203% Duplicates ...... 14-10 Figure 14-12: Al2O3% Duplicates <10% ...... 14-11 Figure 14-13: MgO% HARD Duplicates ...... 14-11 Figure 14-14: MgO% Duplicates ...... 14-12 Figure 14-15: MgO% Duplicates ...... 14-12 Figure 14-16: P% HARD Duplicates...... 14-13 Figure 14-17: P% Duplicates ...... 14-13 Figure 14-18: Duplicates <0.8% ...... 14-14 Figure 14-19: S% HARD Duplicates...... 14-14 Figure 14-20: % Duplicates ...... 14-15 Figure 14-21: S% Duplicates <4% ...... 14-15 Figure 14-22: Mn% HARD Duplicates ...... 14-16 Figure 14-23: Mn% Duplicates ...... 14-16 Figure 14-24: Mn% Duplicates <4% ...... 14-17

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14.0 DATA VERIFICATION

14.1 WGM 1964 and 1965 Assay Verification

Only limited assay and database verification work was completed by WGM during their exploration programs.

Previous validation of the WGM assay database was comprised of sending a small population of samples to external laboratories. Table 14-1 includes seven results from this WGM duplicate testwork program. Although too small a population to reliably reflect the quality of the WGM assay database, the results suggest that the WGM Fe

and SiO2 grades for the seven samples could be reasonably reproduced at external laboratories.

Table 14-1: WGM Comparative Analysis

Sample Tech. Serv. Lab Lakefield Steep Rock Superior Labs

No. Fe (total) SiO2 Fe (total) SiO2 Fe (sol.) SiO2 Fe (sol.) SiO2 14 69.97 0.13 68.65 0.78 70.68 0.24 69.46 0.51 24 70.17 0.20 68.88 0.07 35 70.87 0.54 69.67 0.88 42 71.26 0.09 70.18 0.30 70.05 0.20 71.30 0.20 47 69.48 0.22 68.82 0.52 54 71.93 0.15 70.69 0.24 72.17 0.21 71.42 0.31 57 70.06 0.11 69.33 0.46

To check the quality of assaying by Technical Services Laboratories (TSL), a set of length-weighted composite samples of minus 10-mesh rejects were re-assayed by

TSL and Superior Labs (SL) for Fe, P, SiO2, Mn, Al2O3, CaO, MgO, S, TiO2 and LOI.

The TSL and SL composite assays showed generally good agreement for most elements with TSL reporting generally slightly higher values, and phosphorous having the largest variation.

The iron assays compared reasonably well, and some negative bias seen in the

composite averages for SiO2, and S had a significant amount of scatter in the individual assays. Of more concern was the lack of standards which would allow an assessment of the accuracy of the assay methods. This particularly affects sulphur, which can occur at levels critical for the marketability of the iron ore, although its absolute concentrations are low. There is also concern regarding the phosphorous precision and accuracy.

In order to address the shortcomings of the 1960s era deleterious dataset, BIM commenced a drill program in 2005 to entirely replace this dataset.

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14.2 BIM 2004 to 2009 Database and Assay Verification

Data verification completed by G. Wahl was comprised of data entry validation, analysis of available external duplicate data, analysis of twinned holes, checking the results of SGS’s internal quality control, checking the quality of core logging and sampling protocols, comparisons of drill campaign results via quantile plots, and validating of drill hole collar locations.

The drilling/sample database used for resource estimation was obtained from BIM as MS excel files that were imported into Surpac Mining software. These data files were examined both visually and statistically. Any issues that arose during data import were raised with BIM personnel, corrected and rechecked.

A variety of adjustments were made as data was imported into Surpac. Collar file coordinates were cross-checked with recent drill-hole collar survey files provided by BIM. WGM drill-hole location maps were imported and cross-referenced with the most recent survey data as a secondary check. Drill-hole collar elevations were also cross- referenced against the most recent LiDAR topographic survey.

Approximately 5% of the electronic drill-hole assay database was checked against original assay records for data entry errors. Although assay certificates for the WGM drilling were no longer available, original assay tables were used to check data entry. Drill logs and drill core from the WGM drill campaign were no longer available for verification purposes. During data input, assay intervals were checked and corrected for overlapping samples, and data entry errors.

SGS completed the assay work in BIM’s 2004, 2005, 2006, 2007, 2008 and 2009 drill programs. SGS is accredited by the SCC for specific mineral tests listed on the scope of accreditation to the ISO/IEC 17025 standard. ISO/IEC 17025 addresses both the quality management system and the technical aspects of operating a testing laboratory. SGS documentation provided by BIM indicates that SGS also participates in round robin reference material certification programs.

SGS’s internal quality control procedures include duplicate samples, spiked blanks, spiked replicates, reagent/instrument blanks, preparation control samples, certified reference material analysis, and instrument control samples.

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14.3 SGS Internal Blank, Duplicate Pulp and Standard Results

Duplicate pulp assay results taken by SGS as part of their in-house quality assurance/quality control (QA/QC) program were collected for the BIM drill program and were reviewed by G Wahl. Results of the SGS duplicate pulp testwork for all major grade and deleterious grade attributes were evaluated for each year. No contamination issues were noted with blanks. SGS’s internal standard results were within two standard deviations of the population mean with no anomalous values. Regressions of duplicates indicated good correlation coefficients with the only issues pertaining to those samples assayed near their detection limit which is expected and not an issue of concern.

14.4 Quantile Plots

In order to verify the quality of the WGM assay database, a quantile-quantile plot was generated for the WGM and BIM 5-m composited datasets. The results indicated a reasonable correlation for percentage of iron.

The percentage of sulphur quantile plot indicates that the BIM drill campaign encountered a significant population of higher grade sulphur results. It is possible that the difference is based on localized geological variations, based on where these samples were taken. The percentage of sulphur data anomalies may also be the result of laboratory errors or differences in analytical methodologies. The problematic WGM sulphur results reflect the Strathcona (2003) observation that metallurgical testwork in 1971 highlighted similar sulphur grade uncertainties. Because of the problematic S data and missing oxide assays from the WGM era drilling, recent BIM drill campaigns have replaced WGM near surface drilling with a more comprehensive and reliable dataset.

14.5 Twinned Drill Holes

Three sets of WGM holes were twinned by BIM, in order to further assess the quality of the WGM database (see Table 14-2 to Table 14-4). In order to extract comparable data from the variable sample length dataset, the assay data which fell within the interpreted iron formation solids was composited into 2-m lengths. Extracted grade

attributes common to both datasets included Fe%, SiO2% and S%. Table 14-2, comparing WGM hole S-2 and BIM MR1-05-67, indicates that there was a significantly

high variability in Fe% and SiO2% and S% minimum and maximum grades, resulting in

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a larger discrepancy in average Fe% and SiO2% grades. These outliers impact the overall mean grade of the iron formation interval.

All twinned intervals reflect a slightly lower average Fe% grade in the WGM holes. As

well, SiO2% and S% mean grades are consistently higher in the WGM drilled intervals. The variance in sulphur content over a 58-m interval in the second twin in Table 14-3 cannot be explained. The S% bias is also not consistent with the sulphur quantile plot results. It is not clear whether this is a result of contamination in the WGM drilling, lower core recovery, actual geological features or as a result of sampling or laboratory error.

Table 14-2: Comparison of Twin Drill Holes S-2 and MR1-05-67

WGM S-2 BIM MR1-05-67

Fe% SiO2% S% Fe% SiO2% S% Average 66.09 1.84 0.04 68.55 0.73 0.03 Minimum 8.44 0.18 0.00 64.31 0.13 0.01 Maximum 70.72 33.03 0.13 70.62 3.46 0.37 25th Percentile 68.40 0.27 0.02 68.18 0.28 0.01 50th Percentile 68.77 0.45 0.03 68.96 0.49 0.02 75th Percentile 69.54 0.81 0.04 69.50 1.01 0.04 Tri Mean 68.90 0.51 0.03 68.88 0.59 0.02

Comprised of 29 composites or 58 m of continuous mineralization.

Table 14-3: Comparison of Twin Drill Holes S-5 and MR1-05-63

WGM S-5 BIM MR1-05-63

Fe% SiO2% S% Fe% SiO2% S% Average 68.09 0.53 0.07 68.62 0.41 0.01 Minimum 64.40 0.30 0.02 66.74 0.21 0.01 Maximum 69.12 1.19 0.11 69.40 1.41 0.02 25th Percentile 67.78 0.40 0.04 68.37 0.27 0.01 50th Percentile 68.63 0.44 0.07 68.74 0.31 0.01 75th Percentile 68.80 0.48 0.08 69.05 0.39 0.01 Tri Mean 68.09 0.53 0.07 68.62 0.41 0.01

Comprised of 29 composites or 58 m of continuous mineralization.

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Table 14-4: Comparison of Twin Drill Holes S-6 and MR1-05-60

WGM S-6 BIM MR1-05-60

Fe% SiO2% S% Fe% SiO2% S% Average 67.57 0.97 0.03 68.30 0.80 0.02 Minimum 64.94 0.25 0.01 61.99 0.14 0.01 Maximum 68.90 2.08 0.05 70.29 3.89 0.08 25th Percentile 66.92 0.61 0.02 67.62 0.27 0.01 50th Percentile 67.72 0.64 0.03 69.18 0.32 0.02 75th Percentile 68.31 1.51 0.04 69.58 0.59 0.04 Tri Mean 67.65 0.92 0.03 68.79 0.39 0.02

Comprised of 23 composites or 46m of continuous mineralization.

14.6 External BIM Duplicate Samples Approximately 5% of the 2004, 2005, 2006, 2007 and 2009 assays were sent to an external referee laboratory SGA in Germany, for duplicate pulp analysis. Regressions for each element assayed for each year were run to assess for any persistent or incremental laboratory error. All regressions returned good coefficients of correlation. Figure 14-1 to Figure 14-24 indicate good inter-laboratory repeatability of the latest 2004-2007 assay duplicates. SGS Lakefield however indicates a slight and consistent bias in higher Mn grades over SGA. As well the correlation of S between the two labs is low. The 2009 laboratory duplicates were not yet available for this report.

Figure 14-1: Fe% HARD Duplicates

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Figure 14-2: Fe% Duplicates

Figure 14-3: Fe% Duplicates >40%

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Figure 14-4: Fe% HARD Duplicates

Figure 14-5: Fe% Duplicates

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Figure 14-6: Fe% Duplicates <40%

Figure 14-7: SiO2% HARD Duplicates

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Figure 14-8: SiO2% Duplicates

Figure 14-9: SiO2% Duplicates <15%

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Figure 14-10: AI203% HARD Duplicates

Figure 14-11: AI203% Duplicates

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Figure 14-12: Al2O3% Duplicates <10%

Figure 14-13: MgO% HARD Duplicates

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Figure 14-14: MgO% Duplicates

Figure 14-15: MgO% Duplicates

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Figure 14-16: P% HARD Duplicates

Figure 14-17: P% Duplicates

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Figure 14-18: Duplicates <0.8%

Figure 14-19: S% HARD Duplicates

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Figure 14-20: % Duplicates

Figure 14-21: S% Duplicates <4%

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Figure 14-22: Mn% HARD Duplicates

Figure 14-23: Mn% Duplicates

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Figure 14-24: Mn% Duplicates <4%

In 2007, BIM contracted SGA of Germany to create six standards representing a range

of Fe, SiO2, Al2O3, FeO, P and S grades. The standards were generated from Mary River metallurgical samples, in order to provide reference material consistent with the Mary River ore types. Standards were inserted at a rate of 1 in 25 within the BIM sample stream. Standard results from the 2007 and 2008 exploration program were analyzed and plotted for various elements. A total of 60 plots were generated by G. Wahl to assess for laboratory drift. No significant assay quality issues were identified other than the normal level of minor sample mix-ups.

In summary, the analyses of the various QA/QC data by G. Wahl indicate that the current BIM drill-hole assay database is appropriate for resource estimation. In order to address the problematic sulphur assays from WGM and build a more extensive deleterious database appropriate for future mine planning, G. Wahl recommended in 2005 that the near-surface WGM drill-hole database be replaced with new drilling. This program of drilling the upper 250-m elevation of Deposit No. 1 continued in 2008 and was largely complete in 2009.

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C ONTENTS

15.0 ADJACENT PROPERTIES ...... 15-1

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15.0 ADJACENT PROPERTIES

In 2008 BHP Billiton Ltd. has allowed its adjacent land holdings to Deposit No 4 to lapse. These have been subsequently staked by BIM.

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C ONTENTS

16.0 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 16-1 16.1 Historical Testwork ...... 16-1 16.1.1 1964 – Ontario Research Foundation ...... 16-1 16.1.2 1965 – Bulk Sample ...... 16-2 16.1.3 1971 – University of Minnesota Mineral Resources Research Centre ...... 16-2 16.2 2003-2005 Testwork ...... 16-3 16.2.1 General ...... 16-3 16.2.2 2004 Program ...... 16-4 16.2.3 Sampling Procedure ...... 16-4 16.2.4 Test Program ...... 16-4 16.2.5 Testwork Results ...... 16-6 16.3 2006-2007 SGA Testwork ...... 16-8 16.3.1 REAS Testing ...... 16-8 16.3.2 Crushing Testwork ...... 16-9 16.3.3 Reassessment of the Lump to Fines Ratio ...... 16-10 16.3.4 Lump Ore Drop Tests ...... 16-10 16.3.5 Chemical Distribution – Lump and Fines...... 16-11 16.3.6 Screen Fraction Metal Analysis ...... 16-11 16.3.7 ProMet ...... 16-12 16.4 2008 to 2010 Testwork ...... 16-12 16.4.1 Lump Testwork ...... 16-13 16.4.2 Bulk Sample Program Summary ...... 16-13 16.4.3 Lump Trial Cargoes ...... 16-17 16.4.4 Fine Ore Trial Cargo ...... 16-18 16.4.5 Laboratory-based Sinter Testwork ...... 16-19 16.4.6 Future Metallurgical Testwork ...... 16-21 16.5 Summary and Conclusions...... 16-21 16.6 Mineral Processing ...... 16-22

T ABLES

Table 16-1: Summary of Metso Testwork ...... 16-10 Table 16-2: Grade Differences Between Lump and Fines ...... 16-11 Table 16-3: 0.5 mm Size Fraction Results ...... 16-11 Table 16-4: Metallurgical Testwork by Deposit ...... 16-13 Table 16-5: Lump Iron Ore Physical and Chemical Attributes-Crusher Samples ...... 16-15 Table 16-6: Fine Iron Ore Physical and Chemical Attributes-Crusher Samples ...... 16-16 Table 16-7: Lump Trial Cargo Physical and Chemical Attributes – Receiving Port ...... 16-17 Table 16-8: Fine Ore Trial Cargo Physical and Chemical Attributes – Receiving Port ...... 16-18 Table 16-9: Fine Ore Lab-based Sinter Testwork ...... 16-20

F IGURES

Figure 16-1: Grain Size Distribution of Sinter Feeds and Concentrates ...... 16-19 Figure 16-2: Material Handling Flowsheet ...... 16-23

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16.0 MINERAL PROCESSING AND METALLURGICAL TESTING

The main classes of iron ore materials in trade are lump, fines, and concentrate. Lump is defined as material of size between 6.3 and 35 mm (occasionally up to 45 mm), and can be charged directly to the blast furnace. Lump ores typically command a premium in iron ore trade and markets.

Fines are of size between 1 and 6.3 mm. This material must be sintered (heated to partial fusion and aggregation) before charging. Concentrate is material less than 1 mm in size and is used generally to produce iron ore pellets, in which the material is aggregated with a neutral mineral binder, rolled into pellets (balls about 20 mm in diameter), and fused to yield a high crushing-strength furnace feed.

16.1 Historical Testwork

Information in this subsection is summarized from von Guttenberg and Farquarson (2003).

16.1.1 1964 – Ontario Research Foundation

In 1964, three 1 t samples were collected at surface and shipped for processing to the Ontario Research Foundation in Weston, Ontario (Ontario Research Foundation, 1964).

The three high-grade (69.7% top 70.9% iron) samples used for the tests consisted of hematite, hematite with 10% to 15% magnetite, and magnetite with 3% to 5% hematite, with low levels of silica, manganese, sulphur and phosphorous, and a pore space of 5% to 10% in the hematite-rich samples.

The samples were described as dense and tough (specific gravities of 5.1 to 5.2) which made crushing difficult, but limited the amount of fines produced by crushing, and they had relatively high compressive strengths and good abrasion resistance. Reduction-disintegration tests showed the hematite-rich samples were suitable for blast furnace feed, while the magnetite sample was considered as potential ore for the open hearth process, due to its slower rate of reduction and the excessive amounts of fines produced in the blast furnace during testing.

The chemical composition of the three samples compared favourably with iron ore concentrates and pellets from North America and Brazil, which generally had less iron and more silica than the Mary River samples.

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16.1.2 1965 – Bulk Sample

The WGM report summarizing the 1965 field program stated that 80 t of high-grade hematite and 97 t of high-grade magnetite were recovered from outcrops and placed in empty fuel drums. On the return flights for the movement of supplies from Montreal in the spring of 1965, 15 t of hematite and 27 t of magnetite were flown out with the balance left in drums at Milne Inlet; where they remain today.

Unfortunately, no record remains (in the documentation researched by Strathcona) of what was done with that portion of the 1965 bulk sample that was transported by aircraft to Montreal.

16.1.3 1971 – University of Minnesota Mineral Resources Research Centre

In 1971, a metallurgical test program on Mary River core samples was conducted at the University of Minnesota under the guidance of H.E. (Buzz) Neal, a well-known iron ore specialist with substantial experience in the discovery and development of the large iron ore deposits in Labrador and northeastern Quebec. The test program had been initiated by HBM&S and the results incorporated in the updated economic reviews of the Mary River Project undertaken by HBM&S in 1972-73. The testwork focused on reducibility and swelling tests using Mary River crude ore and pellets (H.E. Neal and Associates, 1971). The core was from nine different core holes spread over a strike length of the deposit of about 1 km. Testing was completed on 15 samples, each weighing in the 5 kg to 9 kg range.

Testing utilized both the Burghardt and Gakushin methods for measuring reducibility. Results were generally parallel, although the Gakushin test showed lower reduction results of high magnetite samples. Hematite- (specularite) rich samples showed excellent reducibilities, but unacceptable levels of swelling and fusion. Samples with intermediate (15% to 29%) magnetite showed good to excellent reducibility and low swelling, and samples with 54% to 94% magnetite showed generally good reducibility (with the exception of one sample). The swelling, measured by tests designed for pellets, was unacceptable in two of the samples. It is important to recognize that the swelling tests are designed for pellets and are not currently used for lump ores. Only three of the 15 samples showed good to excellent reducibility, little swelling, and a low sulphur contents; another 21 specularite-rich and one magnetite-rich were also low in sulphur, but slightly outside the required ranges for reducibility and swelling.

The 1971 tests only partly confirmed the results from the 1964 work at the Ontario Research Foundation, which had indicated that hematite ore would be suitable as blast furnace feed while the magnetite-dominant samples showed unacceptable metallurgical characteristics.

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Preliminary pellet tests at the University of Minnesota indicated that excellent pellets could be produced from high-grade hematite and magnetite ores, and a superior sinter product was produced from the ore.

16.2 2003-2005 Testwork

16.2.1 General

In their recommendations, Strathcona recommended that “Any future metallurgical tests should use samples weighing some 200 kg each, equivalent to 14 m of HQ-size core (or 8 m of PQ-size core) of high-grade iron ore. Testing should be done at recognized institutions by standard procedures as specified by the ISO or by Japanese Industrial Standards (JIS).”

On reviewing the historical results, Strathcona and BIML (later to be BIM) initiated discussions SGA for a recommended testwork program. SGA is well known throughout world iron ore circles as a premier laboratory for analysis, testwork and pilot studies in this field.

SGA’s testwork was performed in accordance with ISO standard procedures and included a chemical analysis, size analysis and tumble test, decrepitation test and evaluation of the dynamic and static low-temperature disintegration behaviour, the reducibility, the reduction index, and the reduction-under-load behaviour. In addition, high-temperature properties such as softening, melting and dipping of the iron oxides were evaluated using a REAS test.

It was agreed that two samples collected by Strathcona at site in the summer of 2003, each weighing about 30 kg, would be sent to SGA for initial program design. All of the above tests, except the tumble test (for lack of sufficient material) and the REAS test, were completed. One sample was hematite-rich and the other magnetite-rich.

SGA confirmed that the hematite sample had the physical and chemical characteristics of a high-quality lump ore and compared well to lump ores from Brazil and Australia. The magnetite-rich sample was very high grade, 71.6% iron, which is at the stoichiometric limit for iron in pure magnetite. SGA commented at the time that they had not encountered such sample material before. The sample contained an excessive level of phosphorous (0.085%), but the most telling comment was that “the reducibility of this ore is, however, very low and this ore will not be accepted for blast furnace operation”. It would, when crushed, be accepted as a good quality sinter feed, providing phosphorus could be controlled.

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16.2.2 2004 Program

To further the metallurgical understanding of the potential ores, the plans for the first season of drilling by BIM included provision for the collection of a number of samples of a large enough mass (80 kg to 140 kg) for a full battery of metallurgical tests. A contract was arranged with SGA to process the samples at the end of the drilling season and to report the results in early 2005.

The basic conceptual plan for the Project was founded on the assumption that the Project would have minimal processing of ores on site. Therefore, exploration would focus on suitable material for direct shipping ores, allowing for only crushing, blending and selective mining to determine product classes and quality.

16.2.3 Sampling Procedure

Based on the geological logs and the available assay results, 20 zones were selected in seven of the completed drill holes to provide a range of magnetite ratios, and included materials of varying visible sulphide mineral content and phosphorous content. The samples were taken from the archive material (split) cores remaining after the other split had been sent to assay. The sampled zones varied between 11 and 42 m in length, with an average of 27 m, and comprised a total of 267 samples, each between 10 kg and 12 kg in weight. A total of 2.8 t of samples were sent to Germany for testing.

16.2.4 Test Program

SGA was instructed to composite the samples for each zone and to perform a broad battery of tests on the resulting mixed material lots. The test program included the following chemical, physical and metallurgical tests (SGA, 2005).

Crushing Test

The test material is crushed and then divided by screening into nine size fractions, from 0.0 to +40 mm, and fractions are weighed, to give a profile of the crushing characteristics. Of particular importance is the total of fractions over 6.3 mm and the total weight of fractions under 6.3 mm. This is the lump:fines ratio. A ratio of over 80% lump is desirable, but material yielding over 60% is acceptable.

Chemical Analysis

Samples are analyzed according to a protocol outlined in Section 13 on sample preparation, analysis and security, necessary for the correlation of physical, chemical and metallurgical results.

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Bulk Density

Using standard methods, the specific gravity/apparent density is measured. This is important for stockpile planning and railcar specification.

Porosity

Porosity is measured by the gas compression test. The porosity is important in furnace behaviour: more porous materials are more easily reduced.

Tumbler and Abrasion Test, ISO 3271

A quantity of -40- to +10-mm material is placed in a drum and rotated 200 times, screened at 6.3 and 0.05 mm, and the fractions are weighed. The Tumbler Index is expressed as the percentage of material >6.3 mm and the Abrasion Index is the percentage of material <0.05 mm. This is a measure of the susceptibility to breakage during handling and transportation. Over 80% in the +6.3 mm and under 10% in the -0.05 mm classes are desirable.

Degradation Test (also called Disintegration Test), ISO 4691

A quantity of material is heated at 500ºC in a high-temperature reduction furnace. The resulting material is screened and the fraction of material smaller than 3.1 mm is taken as the index. During the test, hematite is converted to magnetite and some bound water is released. The volume changes consequent on these reactions can impose internal strains and breakage in the material which, on a larger scale, can affect furnace behaviour.

Decrepitation Testing, ISO 8371

A quantity of material in the <25- to +20-mm range is heated at 700ºC for 30 minutes. The resulting material is screened and the position of <5 mm is expressed as the Decrepitation Index. This is a measure of bursting and parting caused by combined water and volatile minerals. This can also affect behaviour in the blast furnace. Values in good materials are between 1% and 10%.

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Reduction Tests, ISO 4695 and ISO 7215

A quantity of material is heated in the reduction furnace at 900ºC to 1,000ºC. This removes the oxygen from the sample and as the sample mass is continuously weighed during the test the loss of weight with time (dR/dt) is a measure of how easy/difficult it is to reduce the material in the furnace. Typical final reduction values range from 30% to 60%.

Low Temperature Reduction/Degradation Tests (RDI), ISO 4696-1 and 4696-2

Material is subjected to physical strength testing.

Reduction Under Load (RUL) Test, ISO 7992

Material is heated in a retort at high temperature under strongly reducing conditions, while a pressure ram exerts a constant known load on the charge. The test measures reducibility, shrinkage of the test charge, and reduction gas permeation of the charge.

REAS Test

This is a proprietary test of SGA and is not an ISO test. The test is carried out in a high-temperature reduction furnace and is directed at measuring reducibility, softening, melting and dripping characteristics. REAS is a German acronym for these words.

Sintering Test

This is a test series directed at the sintering properties on screen fines (<6.3 mm). These tests are intended to show the coke consumption and physical properties of the resultant sinter product.

Technical details of the testwork performed are outlined by Misra, (1998).

16.2.5 Testwork Results

SGA (2005) reported the following results

Crushing Behaviour

The ratio of lump (>6,3 mm) to lump plus fines (<6.3 mm) was between 60% and 80% (mean: 72%), which is favourable. Lump commands a higher sale price, and the higher the ratio, the better.

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Chemical Analysis

All samples demonstrated superior levels of iron content. Contents of alkalis, Ti and V were favourably low in all samples. Two samples showed unacceptably high phosphorous levels. The latter is not necessarily a handicap in lump class ores, but indicates that selective mining and blending may be required to address this issue. The phosphorous problem will probably require selective mining and blending.

Bulk Densities

Bulk densities lay between 4.66 and 5.11, reflecting the generally high proportions of the two dominant iron minerals in the test material.

Porosities

Porosities varied widely between 0.8% and 21%, volume voids/total volume. The mean was 9.26%. The low-porosity materials were the samples of hard, dense magnetite sectors of the deposit. The porosity correlated with reducibility and reduction data. The pores allowed the reducing gases to permeate and reduce the material much more readily.

Tumbler and Abrasion Tests

Tumbler and abrasion tests were generally very favourable, with only two samples (2 and 13) giving unacceptable values. The range was 72% to 90% (mean 77.56%), leaving aside the two low samples which has values of 46.9% and 41.9%. These were softer, hematite-rich materials. For the tumbler test, values over 80% are preferred, but values over 60% are acceptable. Abrasion rest values showed the inverse of these tests, with values generally in the range of 3.8% to 22.3%, with the exception of the two anomalous samples cited above. These values are acceptable.

Degradation Test

Degradation testing in both dynamic and static test procedures were all within acceptable limits, except for three samples, Nos. 2, 13 and 20, which showed a clear tendency to disintegration under stress. RDI results are generally favourable.

Decrepitation Test

Decrepitation test results were favourably low for all samples, with a maximum DI of 1.72% <6.3 mm in composite sample No 1.

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Reduction Tests

Reducibilities were generally within acceptable ranges. Generally, magnetite-rich samples are harder to reduce, especially if they are of the hard brittle variety. The reduction results are generally good in low-magnetite samples, but some composites high in magnetite (Nos. 8, 2, 16 and 17) showed very good reducibility. Medium magnetite composites show reducibilities in the range of acidic blast furnace pellets, meaning they would be acceptable furnace feed.

RUL Test

Generally, the RUL test showed low differential pressures, which is a favourable indication. An anomalously high differential was found in only one sample. Sample Nos. 18 and 19 showed very slow rates of reduction under load.

REAS Tests

All samples reached melting and dripping stages. The REAS tests also confirmed the reducibility results found in other tests above.

Sinter Tests

Low- and medium-magnetite composites showed good sintering qualities for their fines. High-magnetite material was more difficult to sinter (i.e., the process was less productive). Sinter quality was generally comparable to reference materials.

16.3 2006-2007 SGA Testwork

Additional lump ore tests on the 2005 samples from Mary River Deposit 1 were performed in 2006 and 2007. Additional testwork included a crushing pilot plant test on two samples by Metso. At the same time, SGA did a reassessment summary of the lump to fines ratio for Deposit Nos. 1 and 2, and drop tests to evaluate the breakage of lump ore in handling. Chemical analysis of the lump and fine fractions after screening on 6.3 mm for 135 samples, and screen fraction metal analyses on 25 samples including 12 composites and 13 channel samples were also done. ProMet Engineers of Australia were hired to give an independent overview of the test work completed.

16.3.1 REAS Testing

SGA reported for the 2007 REAS testwork results that “The experimental softening and melting behaviour was found to be consistent across the range of mineral types

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within the Mary River Deposit No. 1, while different characteristics of single lump samples show variations from the general trends. However, melting and dripping of metallic iron was reached for all samples, even for the samples with high magnetite content, despite of their low reducibility during indirect reduction.”

Aker Kvaerner (2008) agreed that the REAS tests confirm the findings for reducibility behaviour of the Mary River Lump ore, and that it was important to establish that for all samples tested, melting and dipping conditions for metallic iron could be reached.

16.3.2 Crushing Testwork

Crushing pilot plant tests were conducted by Metso Minerals at their Mineral Research and Test Center in Milwaukee WI, on two 2.5 t each samples of magnetite and hematite ore. The crushing was done in two stages, first in a jaw and then a cone crusher. The lump to fines ratio for the magnetite sample after cone crushing was projected to be over 88% lump in production and for the hematite sample it was to be over 80% lump (-31, +6.3 mm).

The results show that both samples were very competent. However the results were not used in feasibility-level studies because the Metso report noted that the small diameter crusher that was used produced a coarser product than would be made in a larger commercial crusher. Results were adjusted for pilot plant correlation. Metso’s test results are summarized in the Table 16-1. The material is abrasive and will cause primary liner life to be in the range of 5 Mt to 7.5 Mt and secondary crusher liner life to be 2.5 Mt to 5 Mt according to ProMet’s interpretation.

The crushing work indices are relatively normal except that the magnetite tested was softer than expected, but in the pilot plant this was confirmed with the jaw crusher power draw being 12 kW for magnetite and 36 kW for hematite and the cone crusher drawing 33 kW for magnetite and 58 kW for hematite.

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Table 16-1: Summary of Metso Testwork

Item Magnetite Hematite Remarks CWI Average, kWh/T 6.0 13.9 Magnetite is surprisingly low CWI (max), kWh/T 7.9 19.7 Wide variability CWI (min), kWh/t 1.4 3.2 Wide variability Abrasion Index, grams 0.75 1.06 Classified as Abrasive French Abrasion, g/t 2210 2684 500 g of 3M x 5M for 5 min. Crushability Index 47 39 % Passing 10 mesh Jaw Crusher, kW 12 36 Operating draw, not peak Cone Crusher, kW 33 58 At 1.25inch CCS Cone Crusher, STPH 158 196 Also at 1.25 CCS Cone Crusher, kWh/t 0.19 0.27 Based on throughput

16.3.3 Reassessment of the Lump to Fines Ratio

SGA reported that for Deposit No. 1, the lump ore ratio median value was observed to be over 80% plus 6.3 mm. For the limited work on Deposit No. 2, the median value was over 77%. Channel samples from the surface in Deposit No. 1 returned a lump ratio median value of 88%. Due to certain areas where the lump ore ratio is known to be low, SGA concluded that the value for the operation will be approximately 75%. This will be managed by exercising grade and quality control in the mining operation by setting aside or avoiding the minor soft ore zones.

ProMet noted that the average lump:fines ratio was indicated by the test work to be between 2.5 and 4:1, or between 71% and 80% plus 6.3 mm after crushing. This agrees well with SGA’s conclusion. Drop tests confirmed the competence of these ores.

16.3.4 Lump Ore Drop Tests

Two bulk ore samples were received at SGA, one each of magnetite and hematite. The hematite sample weighed 1.32 t and the magnetite 1.46 t. Each was crushed to 100% passing 40 mm and sub-sampled to give over 500 kg of each ore type. Screening revealed that the crushed hematite was 83% plus 6.3 mm and the magnetite was 92% lump (over 6.3 mm).

The samples were then dropped four times in batches of about 20 kg from a height of 15 m onto a steel plate. The resulting size distributions were measured and found to be 73% for the hematite and just under 80% for the magnetite.

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It is seen that a reduction of about 10% in the lump ore will occur with severe conditioning. Assuming that the ratio of hematite to magnetite will be about 50/50, the average lump ratio will be about 76.5% in the operating plant.

16.3.5 Chemical Distribution – Lump and Fines

All 135 samples, in batches 3 and 4, were screened and assayed for metals and impurities in the plus and minus 6.3 mm size fractions on each of the large 100 kg composite samples. This was done to show the difference in grade between the lump fraction (+ 6.3 mm) and the fines fraction (- 6.3 mm). Results are included as Table 16- 2.

Table 16-2: Grade Differences between Lump and Fines

Value % Fe % SiO2 % P % S Lump Vs fines grade, avg. diff. - 0.14 - 0.14 0.007 0.056

Iron and silica on average are slightly lower in the fines while phosphorus and sulphur are a little higher on average. But considering that high sulphur values generally are found with the dense and higher magnetite content, these differences are so small that the larger issue will be control of the mined material to meet shipping specifications, by excluding off-spec material from the plant feed if that is required.

Due to the variability in the difference between lump and fines where sometimes the fines can be lower or higher in grade than the lump, it is concluded that no significant deleterious effect is or will be created by the conditioning of the ores in handling.

16.3.6 Screen Fraction Metal Analysis

Metal analysis was performed on a range of different screen-size fractions, ranging, in millimetres, as shown:

+20, 20 x 10, 10 x 6.3, 6.3 x 3.15, 3.15 x 0.5 and 0.5 x 0

The minus 0.5 mm fraction is included as Table 16-3, as an example and is based on average data for 24 samples.

Table 16-3: 0.5 mm Size Fraction Results

PRODUCT WT % % Fe %SiO2 % P % S Composite 100 68.1 1.3 0.046 0.343 Minus 0.5 mm 3.5 67.0 1.5 0.077 0.687

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This shows that there is a concentration of silica, phosphorus and sulphur in the minus 0.5 mm fines, but that this fraction is only 3.5% of the feed weight on average. Similarly, the iron content of these ultra fines is about 1% lower than the feed ore. This confirms the observation made above where there was a minor concentration of phosphorus and sulphur in the minus 6.3 mm fines and very little difference in the silica values.

16.3.7 ProMet

ProMet’s specific mandate was to review the test work program for completeness and the company was provided with documentation relating to the Mary River test work, mineral resources, and the proposed process plant. ProMet essentially confirmed the conclusions reached by SGA, and recommended that:

Conditioning tests should be carried out on large bulk samples to correctly determine the lump/fines ratio. Conduct TUNRA-style test work on required bin slopes. Metso had done five small tests and determined slopes varying between 37º and 46º for the resultant piles; more tests on larger piles should be done. Perform an options study to find ideas that will save operating costs and ensure that all client requirements are met. Continue reduction testing of ore samples until there is confidence that the ore quality will be good in the actual operation. Perform additional CWI, UCS, and ore abrasion testing, to allow a good assessment of operating costs. Metso did not provide any UCS results and because the samples tested by Metso may not have been representative.

16.4 2008 to 2010 Testwork

Metallurgical testwork continued at SGA with additional lump testwork, laboratory based sinter pot tests, and analysis of the consumption of three cargoes shipped to two steel companies in 2008.

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16.4.1 Lump Testwork

Lump testwork was completed on both composited drill core and surface channel samples (Table 16-4). Each lump test sample was composited to an average minimum 100 kg to 200 kg sample size. The majority of the work was focused on Deposit No. 1.

Table 16-4: Metallurgical Testwork by Deposit

Channel Core Core Only Total Total 2004 2005 2006 2007 2008 2009 Totals Deposit No. 1 15 372 68 135 88 54 19 8 372

Deposit No. 2 2 23 7 0 16 0 0 0 23

Deposit No. 3 7 26 0 0 10 16 0 0 26

Deposit No. 4 3 0 0 0 0 0 0 0 0 Total 448 27 421 75 135 114 70 19 8 421

In 2010, lump testwork on the 2009 drill core was more limited in scope, effectively filling in data gaps in Deposit No. 1’s metallurgical database. Drill core testing is completed at SGA in Germany and the metallurgical testwork is completed on the previous years drill core.

16.4.2 Bulk Sample Program Summary

In 2008, BIM shipped 113,217 t of iron ore in three cargoes. The first cargo was 54,464.6 t of lump iron ore shipped to ThyssenKrupp Steel in September 2008. The second cargo was 31,050.7 ts of lump iron ore shipped to ArcelorMittal Bremen in October 2008 and the final cargo was 27,701.5 t of fine iron ore shipped to ThyssenKrupp, also in October 2008. Due to the economic crisis, the cargoes were not consumed until 2009.

BIM choose two areas, a hematite-dominant and magnetite-dominant area to mine and blend the two ore-types. These areas were chosen so that the blended product was metallurgically representative of the average material to be mined in the first ten years of production. In both the trial cargo and flat drill holes drilled under the two planned open pits, the concentration of deleterious elements, specifically sulphur and phosphorous were significantly lower in tenor than predicted by the block model. In addition, the expected concentration of magnetite-dominant mineralization in both pit areas was significantly less than predicted in the block model. Based upon this result, and additional infill drilling, the high sulphur zones within Deposit No.1 appear to be

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more restrictive than is indicated in the block model. However, additional drilling would be required to improve confidence in this observation.

Crusher samples were taken at timed intervals during the bulk sample. The timed samples for both the lump and fine iron ore were composited and analysed. The lump iron ore results exhibited excellent consistency throughout. The fine iron ore exhibited moderate consistency affected by the fault gouge (Table 16-5 and Table 16-6).

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Table 16-5: Lump Iron Ore Physical and Chemical Attributes-Crusher Samples

Crusher Samples Sample Sample Sample Sample Sample Sample Sample Lump Ore Composite 1 Composite 2 Composite 3 Composite 4 Composite 5 Average 1-5 Cargo #1 Sample moisture % 0.78 0.93 0.95 0.96 1.11 0.946 1.16 Sample weight kg 258.99 217.88 218.14 235.87 231.26 232.4 166.1 Chemical Analysis Fe tot % 68.31 67.89 68.02 68.74 68.63 68.3 68.08 Magnetite % 16.8 12 18.1 13.2 18.7 15.8 16.2 FeO % 5.32 4.05 6.1 3.89 4.82 4.84 5.01 SiO2 % 0.73 0.91 1.21 0.53 0.65 0.81 0.98 Al2O3 % 0.56 0.6 0.75 0.49 0.47 0.57 0.60 CaO % 0.012 0.019 0.042 0.008 0.015 0.019 0.018 MgO % 0.06 0.085 0.1 0.04 0.06 0.07 0.142 P % 0.012 0.01 0.011 0.012 0.009 0.011 0.009 S % 0.031 0.029 0.024 0.02 0.03 0.027 0.026 Na2O % 0.012 0.007 0.014 0.007 0.01 0.01 0.016 K2O % 0.012 0.015 0.02 0.013 0.015 0.015 0.015 Mn % 0.242 0.123 0.162 0.11 0.132 0.154 0.137 L.O.I. % 1.1 1.45 1.04 0.81 0.97 1.074 1.19 Size distribution > 40,0 mm Sum. % 0.0 0.0 0.0 0.0 0.0 0.0 0.0 > 31,5 mm Sum. % 0.0 0.0 0.0 0.0 0.0 0.0 0.0 > 25,0 mm Sum. % 11.7 14.1 14.3 14.6 17.2 14.4 16.8 > 20,0 mm Sum. % 39.6 49.4 45.5 39.7 47.9 44.4 54.8 > 18,0 mm Sum. % 53.1 62.1 58.4 50.4 59.3 56.7 67.0 > 16,0 mm Sum. % 61.1 69.4 65.7 57.4 65.9 63.9 71.1 > 12,5 mm Sum. % 78.8 83.8 81.7 74.5 81.6 80.1 83.4 > 10,0 mm Sum. % 87.8 90.3 89.4 85.3 90.1 88.6 89.5 > 6,3 mm Sum. % 96.1 95.6 96.1 95.0 97.0 96.0 95.0 > 5,0 mm Sum. % 97.1 96.8 96.7 95.7 97.3 96.7 95.8 > 1,0 mm Sum. % 98.4 97.2 97.4 96.9 97.9 97.6 97.0 > 0,0 mm Sum. % 100.0 100.0 100.0 100.0 100.0 100.0 100.0

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Table 16-6: Fine Iron Ore Physical and Chemical Attributes-Crusher Samples

Crusher Samples Sample Sample Sample Sample Sample Sample Sample Fine Ore Composite 1 Composite 2 Composite 3 Composite 4 Composite 5 Average 1-5 Cargo #3 Sample moisture % 0.63 0.89 0.47 0.3 1.31 0.720 3.32 Sample weight kg ~225 ~185 ~185 ~205 ~190 198 ~200 Chemical Analysis Fe tot % 66.83 66.71 65.81 67.88 68.13 67.07 66.645 Magnetit % FeO % 5.17 3.73 5.72 4.02 5.21 4.77 N/A SiO2 % 2.11 1.99 3.19 1.22 1.07 1.92 2.46 Al2O3 % 0.81 0.83 1.37 0.75 0.56 0.86 0.95 CaO % 0.04 0.03 0.033 0.023 0.026 0.030 0.04 MgO % 0.11 0.16 0.145 0.06 0.055 0.106 0.11 P % 0.015 0.011 0.014 0.014 0.01 0.013 0.014 S % 0.036 0.036 0.032 0.026 0.037 0.033 0.032 Na2O % 0.033 0.028 0.034 0.02 0.016 0.026 N/A K2O % 0.048 0.039 0.052 0.028 0.022 0.038 N/A Mn % 0.25 0.13 0.185 0.13 0.133 0.17 0.15 L.O.I. % 1.41 1.72 1.37 1.04 1.21 1.35 0.93 Size distribution >25,0 mm Sum. % ------>20,0 mm Sum. % ------>18,0 mm Sum. % ------>16,0 mm Sum. % ------>12,5 mm Sum. % 0.0 0.6 0.4 0.3 1.4 0.5 >10,0 mm Sum. % 0.1 2.5 1.4 2.1 3.2 1.9 > 8,0 mm Sum. % 1.1 7.0 4.9 6.5 7.6 5.4 > 6,3 mm Sum. % 12.2 19.2 15.0 19.8 20.7 17.4 20.9 > 5,0 mm Sum. % 29.3 37.1 29.2 37.4 37.6 34.1 > 4,0 mm Sum. % 41.4 49.2 40.3 49.1 49.3 45.9 > 3,15 mm Sum. % 51.9 59.2 49.6 59.2 58.7 55.7 58.4 > 2,5 mm Sum. % 59.4 65.8 56.1 65.3 64.1 62.1 > 1,0 mm Sum. % 78.2 81.9 74.6 81.9 81.9 79.7 79.7 > 0,5 mm Sum. % 83.3 85.9 79.8 85.8 85.0 84.0 > 0,315 mm Sum. % 86.1 87.8 82.9 87.8 86.9 86.3 88.2 > 0,2 mm Sum. % 88.1 89.4 85.5 89.3 88.6 88.2 > 0,16 mm Sum. % ------> 0,1 mm Sum. % 91.4 91.9 89.4 91.8 91.1 91.1 92.6 > 0,063 mm Sum. % 93.2 93.4 91.4 93.4 92.6 92.8 96.9 > 0,04 mm Sum. % 94.5 94.6 93.2 94.8 93.9 94.2 > 0,025 mm Sum. % 95.4 95.4 94.3 95.6 94.5 95.0 > 0,0 mm Sum. % 100.0 100.0 100.0 100.0 100.0 100.0 100.0

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16.4.3 Lump Trial Cargoes

Two lump cargoes were shipped to Europe. A summary of the chemical and physical attributes is tabled in Table 16-7.

Table 16-7: Lump Trial Cargo Physical and Chemical Attributes – Receiving Port

Lump Trial Cargoes

Screen Analysis Cargo #1 Cargo #2 Chemistry Cargo #1 Cargo #2 Plus 31.5 0.1% 0.7% Fe (total) 68.015% 68.635% 16 mm -31.5 mm 22.9% Al2O3 0.64% 0.47% 71.5% 16 - 25 mm 49.1% 12.5 mm - 16 mm 10.1% 8.9% CaO 0.02% N/A 10 mm - 12.5 mm 9.3% 7.1% MgO 0.10% 0.10% 8 mm - 10 mm 3.5% 4.7% MnO 0.175% 0.18% 6.3 mm - 8 mm 2.1% 2.4% P2O5 0.020% 0.018% minus 6.3 mm 3.4% 4.2% SiO2 0.92% 0.56% Moisture Content 1.28% 1.26% S 0.026% 0.027% Metallurgy P 0.008% 0.008% RUL [dR/dt]40 %/Min 0.92 Mn 0.13% 0.14% RI % 41.9 FeO 4.5% 4.4% Strength [ISO 13930} % 65.3 LOI (det) 0.80% 0.66% Tumble Strength % 93.9 V N/A Abrasion Index % 3.8 Ni N/A

Consumption of the two lump ore cargoes occurred in mid-2009. At ThyssenKrupp, the lump ore averaged 15% or 250 kg of Mary River lump ore per tonne of hot metal. Productivity was maintained and increased slightly in a low to moderate productivity environment. ThyssenKrupp used 100% coke with no injectants during the consumption of the Mary River lump ore. Conclusions reached were that the Mary River lump iron ore was a high quality ore that showed excellent handling characteristics.

There was a small size distribution with 72% of the ore in the +16mm size fraction and 91% +10mm fraction. Despite no screening at the port site, undersize material averaged less than 4% for the two cargoes. In addition, the material sat on the port site for some nine months before being consumed; however, the increase in undersize was well under 1%. In normal sales contracts for lump iron ore, allowable undersize (- 6.3mm) is generally 12% to 14%, before price penalties are applied. As ores are screened at 6.3 mm to prevent undersize from being charged to the blast furnace, use of Mary River appears to give the customer ~10% more lump.

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At ArcelorMittal Bremen, similar physical attributes were measured and blast furnace technical staff stated that handling and transport characteristics were excellent. Approximately one-third of lump cargo was consumed in late 2008 prior to the downturn caused by the economic crisis. Results were very positive so in 2009, ArcelorMittal Bremen technicians planned a more aggressive test.

Initial iron ore burden mix was approximately 60% sinter and 40% pellets; Mary River lump was added to the burden mix as a direct substitute for pellets. The lump ore was added to the mix at a rate of 100 kg/t of hot metal increasing 100 kg’s every 12 hours until the rate was 400 kg/t of hot metal. Productivity increased with an average daily hot metal production of 6,800 t versus an average productivity of 5,900 t/d. Furnace behaviour was excellent and results were exceptional in a high productivity environment as Mary River lump iron ore was an effective direct substitute for pellets.

16.4.4 Fine Ore Trial Cargo

BIM shipped 27,702 t of fine iron ore to ThyssenKrupp which was consumed in early 2009 as part of two sinter strands with each strand approximately 150,000 t. The fine iron ore cargo attributes are tabled in Table 16-8.

Table 16-8: Fine Ore Trial Cargo Physical and Chemical Attributes – Receiving Port Screen Analysis Mutually Determined % Chemistry Analysis Plus 6.3 20.9% Fe (total) 66.635% 3.15 mm - 6.3 mm 37.5% Al2O3 0.94% 1 mm - 3.15 mm 21.3% CaO N/A 0.315 mm - 1 mm 8.5% MgO N/A 0.125 mm - 0.315 mm 4.4% MnO 0.19% 0.050 mm - 0.125 mm 4.3% P2O5 0.027% minus 0.050 mm 3.1% SiO2 2.45% Moisture Content 3.32% S 0.032% Tumble Strength N/A P 0.012% Mn 0.15% FeO 4.6% LOI (det) 1.01%

To date screen chemical analysis of drill core composites of the +6.3 mm and -6.3 mm fractions have seen no statistically significant difference in the average chemistry. In the trial cargo mining, a narrow metre to two metre-wide fault gouge was encountered. In a normal environment, this gouge would have had no impact on crushing and screening. However, the small crushing plant combined with the earlier than normal spring freshette resulted in BIM replacing the 6.3 mm square (¼ inch by ¼ inch) screen with a 57.2 mm by 6.3 mm (2 ¼ inch by ¼ inch) slotted screen. This resulted in almost

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21% oversize material in the trial cargo. The gouge also resulted in an increase in the amount of silica, alumina, and minor increases in alkalis versus what was predicted from drilling.

The average grain size (MS-50) of the fine trial cargo was 3.5 mm, an excellent attribute. The grain size distribution can be noted for the fine iron ores used in the sinter testwork by ThyssenKrupp (Figure 16-1).

Figure 16-1: Grain Size Distribution of Sinter Feeds and Concentrates

100 Robe River

Carajas SF 90 MBR SF

80 Gülbs GFM Carol Lake conc.

70 Itabira SF

BHP

60 Mary River

50 Screenings[%] 40

0 0,01 0,1 1 10 Granulation [mm]

Mary River fine iron ore amounted to 9.6% of the sinter feed mix. Despite the relatively small amount of ore within the sinter mixture, there was a 4.6% increase in productivity in the making of sinter, from 36.6 t/m2/d to 38.3 t/m2/d. Sinter quality and metallurgical attributes remained high and fuel (coke) consumption was reduced.

16.4.5 Laboratory-based Sinter Testwork

BIM also completed laboratory-based sinter pot testwork to compare and contrast the consumption with the bulk sample with laboratory tests. The laboratory-based sinter testwork was completed on a reference sinter mix very similar to what was used in the Thyssen-Krupp sinter testwork. The reference sinter mixture included a blend of the following fine iron ore and concentrates:

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Carajás sinter fines (Brazil) Itabira sinter fines (Brazil) Carol Lake concentrates (Canada) Robe River sinter fines (Australia)

MBR-SiO2 sinter fines (Brazil) Guelbs concentrates (Mauritania) Blast furnace return fines

Sinter plants add variable amounts of return fines, mill scale, and flue dust to the sinter mix to recycle this material in the sinter mix.

BIM fine iron ore was added to the sinter mixture at 10%, 30%, 50% of the mixture and a sinter pot test was completed on 100% Mary River fine iron ore. BIM used the composited crusher samples for the sinter mixture which was the same as the material consumed in the sinter plant testwork.

There was insufficient sample material to complete a 100% sinter pot test with crusher sample material. The 100% Mary River sinter fines testwork used material sampled from the remaining fine iron ore stockpile. This material had higher silica and alumina and lower iron content (65.3% Fe; 2.4% SiO2; 1.5% Al2O3) than seen in the crusher sampled material (66.6% Fe; 2.4% SiO2; 1.0% Al2O3). Other chemistry was similar in tenor.

Laboratory-based testwork confirmed the excellent results as achieved in the sinter plant (Table 16-9). At a 10% amount of Mary River fines, the productivity increase was 4.9%, approximately the same as seen in the sinter plant results. At 100% Mary River fine iron ore, the productivity gain was 39.3% greater the standard reference mixture.

Table 16-9: Fine Ore Lab-based Sinter Testwork Mary River Fines Productivity Result Productivity (%) (tonnes/metres2/24 Hrs.) Increase (%) Reference (0%) 34.9 - 10% 36.6 4.9% 30% 37.2 6.6% 50% 43.2 23.8% 100% 48.6 39.3%

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16.4.6 Future Metallurgical Testwork

Additional lump testwork continues at the Mary River iron ore deposits but on a more reduced pace. The outstanding results achieved in both the lump and fine iron ore testwork and comparable lab-based has allowed the Company to be more selective.

Lump testwork will continue to be completed on both surface channel samples and on drill core. The metallurgical database at Deposit No. 1 is more than adequate to support mine planning and the blending of lump and fine iron ore sale products. No further work is needed at Deposit No. 1.

16.5 Summary and Conclusions

Results of the testwork are:

Most of the testwork was done in analyzing the chemical, physical and metallurgical properties of lump ore samples from drill core and surface samples from Mary River Deposit No. 1 The testwork confirmed the high quality of Mary River lump ore. Due to the unique character of the deposit with ore types from almost pure hematite at low FeO contents to almost pure magnetite at high FeO contents with the majority of in- between mineralization, the individual test results differ in a rather large range. Accordingly extensive test work is necessary to enable the development of a proper mine planning system. Chemical analysis shows high Fe contents for Mary River lump ore and low gangue contents. The contents of alkalis, manganese, titanium and vanadium are generally low. The phosphorus content of the Deposit No. 1 ore will be favourably low, despite some high phosphorus assays on some samples containing high magnetite values. Phosphorus in the shipped products would therefore need to be controlled by proper mine planning and material blending. Some samples were assayed with high sulphur content. For the shipped product, sulphur content well below 0.2% can be expected. High sulphur in the product is more acceptable in for lump ores than for fines. The lump ore ratio of the direct shipping material will be high and is expected to be in a range over 75% (plus 6.3 mm).

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The tumble strength and abrasion values of the lump ore is expected to be on a well-accepted level in the mine and material handling areas of the plant. The metallurgical testing resulted in favourable reducibility for lump ore for the low magnetite content samples. As expected, reducibility is lower for the higher magnetite-content samples. However, it can be stated that for a number of high magnetite content samples, R40s around 0.5%/min according to ISO 4695 and 0.8%/min according to ISO 7992, respectively, were evaluated, being in a similar range as acidic blast furnace pellets. Most of the samples revealed favourable, and part of the samples excellent, results in disintegration testing. Decrepitation tendency was found to be low for all samples tested. Favourably low differential pressures were found in the reduction-under-load tests. The REAS tests confirmed the findings for reducibility behaviour of the Mary River lump ore. In particular, the dense pure magnetite samples reached only low reduction indices during indirect reduction. For all samples, melting and dripping could be reached, as well as for high magnetite content ores. As a conclusion it can be stated that Mary River ores will be of high quality. The high lump ratio of around 75% is especially advantageous.

16.6 Mineral Processing

The capacity of the proposed mining, comminution and handling facilities are considered modest for iron ore operations. The subarctic location will necessitate thorough weatherproofing, but otherwise the crushers, screens and other processing and transfer equipment are conventional in their selections.

The process flow sheet (See Figure 16-2) consists of sequential crushing by jaw and cone crushers in closed-circuit with vibrating screens. The plant will deliver two products, lumps and fines, to stockpiles for blending prior to transfer to the port for ocean transport.

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Figure 16-2: Material Handling Flowsheet

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Lump iron ore is expected to be the main product from the Project and will be about 75% of the total production by weight. This product will have high iron grade by world standard, and will have variable impurity content, depending on the mine plan and schedule. The lump product has been found suitable for blending as direct feedstock for blast furnace iron making. The minority fines product would also be saleable depending on its grade and market conditions.

The proportion of lump generated from laboratory crushing of core samples was high compared to other iron ores. This trend is likely to be confirmed from mined samples. Physical tests will be required on ore samples that represent mine deliveries to define the crushing and screening properties for detailed plant engineering and economic analysis of future production.

The process plant for the Mary River Project has been designed with the maximizing of lump product as a primary objective. It is a two-stage crushing and screening plant which produces two products, a -25 +6 mm lump product and a -7 mm fine product, each of which is then conveyed to storage piles ahead of the load out facility for road haul to Milne Inlet.

Further work during the crushing and screening of the material during the bulk sampling program indicates that the physical attributes of the Mary River lump and fines are outstanding. The lack of undersize (~4% less than 6.3 mm) based upon screening at the discharge port indicates superior physical attributes that may allow tighter physical specifications of a lump ore sale product. It is important to note that the small portion of undersize was achieved without discharge port screening.

Material above this screen size is recycled to the secondary crusher. All material less than 7 mm will be shipped for sinter plant feed at the smelter, where it will be blended with other fine products.

Comment on Section 16

In the opinion of the QPs, the metallurgical test work conducted to date supports the declaration of Mineral Resources based on the following:

The metallurgical testwork completed on the Project has been appropriate to establish a process route that is applicable to the mineralization types Tests were performed on samples that were representative of the mineralization for the purposes of establishing an optimal conceptual process flowsheet The process route proposed uses conventional technology

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Metallurgical mapping of the deposit has established clear parameters for likely deleterious elements. Such elements can be appropriately controlled during mine planning and ore control Recovery factors from the tests are appropriate to the mineralization types and selected process route based on the available testwork data.

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C ONTENTS

T ABLES

Table 17-1: Exploration Database Used for Resource Estimation Mary River Deposit #1 ...... 17-4 Table 17-2: Database Used for Resource Estimation Mary River Deposit #2 ...... 17-4 Table 17-3: Database Used for Resource Estimation Mary River Deposit #3 ...... 17-4 Table 17-4: Deposit #1 Zone Solids Used to Constrain Grade Interpolation ...... 17-6 Table 17-5: Balance of Composite Length Analysis ...... 17-7 Table 17-6: Comp_100 Statistics ...... 17-8 Table 17-7: Comp_200 Statistics ...... 17-9 Table 17-8: Comp_300 Statistics ...... 17-10 Table 17-9: Comp_400 Statistics ...... 17-11 Table 17-10: Comp_600 Statistics ...... 17-12 Table 17-11: Comp_700 Statistics ...... 17-13 Table 17-12: Comp_1300 Statistics ...... 17-14 Table 17-13: Comp_1400 Statistics ...... 17-15 Table 17-14: Statistics Report Deposit #2 Upper 200 Zone ...... 17-16 Table 17-15: Statistics Report Deposit #3 Main Zone ...... 17-17 Table 17-16: Major Axis Variogram Results Deposit No. 1 – North Limb ...... 17-19 Table 17-17: Deposit No. 1 Search Ellipse Parameters ...... 17-20 Table 17-18: Deposit No. 2 & 3 Search Ellipse Parameters ...... 17-21 Table 17-19: Comparison of Inverse Distance Squared and Ordinary Kriging Interpolated Grades for Deposit No. 1 ...... 17-22 Table 17-20: Optimisation Parameters ...... 17-28 Table 17-21: Deposit 1, 2 & 3 Mineral Resources ...... 17-30 Table 17-22: Combined Deposit 1, 2 & 3 Mineral Resources ...... 17-30 Table 17-23: Optimization Parameters ...... 17-32 Table 17-24: Mining Recovery and Mining Dilution ...... 17-32 Table 17-25: Summary of the Pit Slope Angle used in the Optimization ...... 17-33 Table 17-26: Mineral Reserves within the Designed Pit for Trucking Study ...... 17-33

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F IGURES

Figure 17-1: Swath Plot North 100 Zone Fe% ...... 17-23 Figure 17-2: Swath Plot North 100 FEO% ...... 17-23 Figure 17-3: Swath Plot North 100 Zone Al2O3% ...... 17-24 Figure 17-4: Swath Plot North 100 MgO% ...... 17-24 Figure 17-5: Swath Plot North 100 SiO2% ...... 17-25 Figure 17-6: Swath Plot North 100 Mn% ...... 17-25 Figure 17-7: Swath Plot North 100 P% ...... 17-26

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17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

17.1 Mineral Resources

The resource estimate for Mary River was completed by Resource Geologist, G. Wahl P. Geo., who is a QP as defined by NI 43-101. The effective date of this resource estimate is Dec 30th, 2010.

This section provides details in terms of key assumptions, parameters and methods used to estimate the mineral resources together with G Wahl’s opinion as to their merits and possible limitations. Deposit Nos. 1, 2 and 3 are included in the following description of mineral resources. The resource estimate of Deposit No.1 has been updated with drill results from 2009, while the resource estimates from Deposits No. 2 and 3 have been restated for the purposes of providing a complete inventory of Mary River Mineral Resources within a single report.

This AMEC Trucking FS only considers upgrading a limited and initial 60Mt from these mineral resources into reserves as directed by BIM. The current assumption by BIM as the basis for this Study is that after the initial 60MT have been mined, the balance of mineral resources will then be removed by rail transport to the coast as per the Aker FS (2008). The pit shells used to report Mineral Resources in this report are based on rail costs associated with the 18Mtpa throughput rate defined in the Aker FS (2008).

The database used to estimate Mineral Resources was audited by G Wahl and the geological interpretations and grades were found by G Wahl to be appropriate for resource estimation. Methodologies used by BIM to generate the database were reviewed by G Wahl and were generally found to be appropriate.

All solid modeling, geostatistics and resource estimation work was completed in Surpac Version 6.1.

17.1.1 Resource Database

The exploration database provided to G Wahl by BIM for Deposit No. 1 was comprised of a total of 156 drill holes, totalling 30,170m and 220 channel samples totalling 1,090m (see Table 17-1). WGM holes in Deposit No. 1 were removed for resource estimation purposes where those holes were replaced by current BIM drilling. Because of the location of channel samples on very limited exposure and difficulty in obtaining continuous and representative samples, the channel sample data was also excluded.

A total of 8 drill holes comprised of 1,314m and 22 channel samples over 130m were included in the Deposit No. 2 database (see Table 17-2).

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A total of 11 drill holes comprised of 2,553m and 8 channel samples over a distance of 82 m were included in the Deposit No. 3 database (see Table 17-3). As the 2007 channel samples on Deposit 3 contained a complete deleterious dataset and were largely repeats of the 1960’s era channel sampling, only the 2007 channel samples were used for resource estimation.

Table 17-1: Exploration Database Used for Resource Estimation Mary River Deposit #1 Year Company Type No. of Collars 1964 WGM DDH 13 2004 BIM DDH 7 2005 BIM DDH 28 2006 BIM DDH 17 2007 BIM DDH 22 2008 BIM DDH 27 2009 BIM DDH 13 1963 WGM CHA 169 2006 BIM CHA 11

Table 17-2: Database Used for Resource Estimation Mary River Deposit #2

Year Company Type No. of Collars 2004 BIM DDH 1 2006 BIM DDH 7

1963 WGM CHA 22

Table 17-3: Database Used for Resource Estimation Mary River Deposit #3

Year Company Type No. of Collars 2006 BIM DDH 3 2007 BIM DDH 8

2007 BIM CHA 8

Terrapoint’s LiDAR topographic coverage at 30-cm resolution was used to estimate the mineral resources.

For the purposes of the current resource estimate, an overburden surface was based on data provided by Knight Piésold and was used, where appropriate, to clip the upper surface of the mineral resources for Deposit No 1. Due to a lack of evidence of significant overburden thickness and close proximity of the ore to surface, no overburden surface was constructed for Deposits Nos. 2 and 3.

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17.2 Solid Modeling

For Deposit No. 1, wireframes were interpreted for each approximately 75-m spaced drill-hole section. All wireframes were snapped to drill-hole assay intercepts. An assay boundary of ~55% Fe was used to define hanging- and footwall solid boundaries; however, where the massive iron formation graded along strike into a banded iron formation, a lower cut-off was used to allow grade interpolation parameters to define gradational extent of lower grade mineralization along strike between drill sections. A majority of the cross strike contacts were defined by very sharp declines in high grade iron content. A plan map outlining the location of the various zones described below is included in Section 7, Figure 7.4.

In order to manage the impact of highly skewed deleterious grades during grade interpolation, two high deleterious zones comprised of grades >1% S and >1% Mn were solid modeled separately. These two zones were comprised of for example, up to several percent sulphur or manganese along the footwall. These zones are limited in extent and occur over a limited distance along the south limb footwall (rock_code 1400) and the north limb footwall (rock_code 1300). These zones were characterized by fairly sharp increases in S and/or Mn.

High phosphorus was not solid modeled separately as its distribution was far more random from hanging wall to footwall. Three, small, poorly defined areas of elevated phosphorus occur within the Deposit 1 and have a limited extent. Higher grade phosphorus (0.1-0.3%) is concentrated in the extreme north end of the main rock code 100 solid, representing approximately 300 m of strike extent and also within portions of the two solids located in the inner hanging wall portion of the fold nose (rock codes 300 and 400). An additional area of high phosphorus is located at the extreme southeast end of the Deposit. Although included in the current resource estimate consideration should be given in future to excluding these composites and treating this zone as waste in order to further improve product quality.

High K2O (1-6%) occurs predominantly along the waste rock hanging wall and footwall contacts with the ore zone and should be avoided during mining. Otherwise, only rare

isolated 2-4m wide intervals of high K2O were noted within the ore zone and would be easily diluted during normal mining conditions.

Three internal waste zones were also solid modeled. These were at least 8m wide, and were either chlorite schist or low grade banded iron formation. A total of nine mineralized solids were modeled for Deposit No. 1. Within the main ore zone, four stacked stratigraphic ore zones were modeled. The lower zone or footwall zone represents the highest grade, most continuous and thickest zone. The overlying middle and two upper zones are limited in thickness, lateral extent, and concentrated

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largely within the fold nose. Finally, a narrow solid striking northeast along the hanging wall of the deposit is defined by drilling at its southern extent; however, airborne geophysics defines a continuous anomaly over several hundred metres to the northeast. The solid model was extended accordingly. Equally, airborne magnetics was used to help define the southeast extent of the South Limb mineralization.

These solids were used to constrain grade interpolation. Volumes for the solids are included in Table 17-4. This table should be considered in conjunction with the zone composite statistics Table 17-6 to Table 17-15 and has been included to allow the reader to assess product quality and quantity by zone.

Table 17-4: Deposit #1 Zone Solids Used to Constrain Grade Interpolation

Zone Solid No. Volume Dbase & Blk Mdl (m3) Zone No. Lower Zone 1 93,791,250 100 Middle Zone 5 15,265,125 200 Upper 1 Zone 7 7,398,000 300 Upper 2 Zone 9 1,393,875 400 Lower Zone Depth Ext 1 19,669,500 500 NE Zone 4 3,557,250 600 N Ext Zone 2 7,533,000 700 North Delet Zone 24 2,082,375 1300 South Delet Zone 8 1,606,500 1400 North Int Waste 12 124,875 1000 North Int Waste 11 631,125 1100 South Int Waste 14 2,082,375 1200

A lower zone and upper zone solid were defined for Deposit No. 2. Deposit No. 2 is narrower and bifurcates into two separate zones at its eastern extent. A single main zone was interpreted for Deposit 3.

17.3 Compositing

Samples which fell within a mineralized or waste solid were flagged in the database. In cases where more recent BIM drilling replaced earlier WGM samples, the replaced WGM samples were entirely excluded from compositing.

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The average sample lengths in the raw assay database ranged from a minimum sample length of 0.15 m to a maximum of 16.76 m, with an average of 2.29 m. A majority of the samples lengths greater than 2.29 m originated with the WGM sampling and drilling programs which have largely been replaced by more recent BIM drilling. Only four WGM holes remain in the 100 zone composite file. Three of these drill holes S-5, S-4 and S-16 are located in the near surface portion of the fold nose. The fourth hole S-22, is located beneath the south end of the bulk sample pit. These WGM

intervals are atypical as they only have Fe, SiO2 and S assay results. Samples were composited to 5-m composites for statistics and grade interpolation. Remaining composite intervals with a minimum length of at least 3.25 m were also included within these composite files.

An analysis of the impact on excluding the remaining non-5m composite intervals was completed in Table 17-5. The differences were not found to be material.

Table 17-5: Balance of Composite Length Analysis

Min. Comp. Min. % comp Fe Al2O3 SiO2 MgO Mn P S Avg. Length to Length to Length Include Include (m) (%) (%) (%) (%) (%) (%) (%) (%) (m) 2.5 50 66.48 1.31 2.22 0.87 0.186 0.042 0.20 4.95 3 60 66.47 1.31 2.22 0.87 0.187 0.042 0.20 4.96 3.25 65 66.48 1.31 2.22 0.87 0.185 0.042 0.20 4.97 3.5 70 66.48 1.31 2.212 0.87 0.185 0.042 0.20 4.98

17.3.1 Statistics

Statistics of 5-m composites were generated for each of the zone solids and for each of the assay elements (see Table 17-6 to Table 17-15).

Some grade differences were noted within the various zones. The lower zone of Deposit No. 1 showed the highest mean Fe grade at 66.4% Fe, while the middle zone average grade was 64% Fe and the upper zones 64% and 63.5% Fe. The difference in grade is a result of the middle zone containing a higher proportion of intervals which graded into enriched banded iron formation along strike and therefore higher corresponding SiO2 grades. Higher P mean grades are characteristic of the northern 700 Zone at 0.14%P. This zone also exhibits higher S mean grades at 0.45% S. Mn grades were also anomalously elevated in Zone 600 at 0.23%.

The two deleterious zone statistics highlight elevated S and Mn deleterious grades along the footwall of the Lower Zone. Zone 1300 along the North 100 footwall has a mean Fe grade of 60%, a mean S grade of 1.2%, while the mean Mn is 1.3%. The

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1400 Zone located on the South 100 footwall has a mean Fe grade of 61%, a mean S grade of 1.0% and mean Mn grade of 1.0%.

Deposit No. 2 differs from Deposit No. 1 in that the magnetite content is significantly lower, (see Table 17-6). This however, may be a result of having a proportionally larger near-surface dataset in Deposit No. 2 than in Deposit No. 1 and, therefore, a greater proportion of near surface martitization. Insufficient data precluded the generation of reliable statistics for the lower zone of Deposit No. 2 and the northern zone of Deposit No. 1.

Table 17-6: Comp_100 Statistics

Variable A 2 O 3 % CaO% Fe% FeO% H 2 O K 2 O LOI% Lump% Mag MgO Mn N 2 O P S Si 2 % l % % % % % a % % % O Number of samples 1754 1754 175 1754 1754 1754 1282 1754 1753 175 175 1754 1754 1754 1754 Minimum value 0.00 0.00 12.1924 0.00 0.02 0.0015 0.01 0.20 0. 0.004 0.004 0.00 0.00 0.00 0.0778 Maximum value 20.7965 59.63 71.658 32.53335 8.535 0.8634 11.01 0.9904 104.84331 16.6615 3.684 0.06521 5.291 3.87965 52.1 Mea 1.306354 0.163751 66.416 13.42711 0.523268 0.020354 1.5024468 0.8541974 43.30846 0.869422 0.1864372 0.0079 0.0419614 0.207697 2.294859 Mediann 0.5787 0.04 868.499 9.6047 0.2719 0.0074 0.91165 0.87 30.9976 9 0.2059 0.1076 0.00635 0.0086 0.0 3 0.69 Geometric Mean 0.704788 0.0491881 66.0374 7.882477 0.274909 0.008833 0.873435 0.8469978 25.56625 0.27558 0.103573 0.0060045 0.009679 0.0449533 0.940342 Variance 4.197307 0.252102 34.96591 116.4859 0.499831 0.003845 2.851518 0.009633 1209.95 2.640378 0.085114 4.68E- 0.033186 0.176265 16.91969 Standard Deviation 2.048733 0.502097 5.913207 10.79286 0.706987 0.062009 1.688644 0.098146 34.78433 1.6249249 0.29174 0.00683905 0.182171 0.419839 4.113355 Coefficient of variation 1.568283 3.066222 0.089033 0.803811 1.351121 3.046519 1.12393 0.114898 0.803176 1.868955 1.564833 0.860265 4.341446 2.021402 1.792427 2 9 Skewness 4.066668 9.568637 -3.86316 0.319783 3.553389 7.856257 2.259608 - 0.319105 4.312707 5.57704 3.206283 17.70575 3.885611 4.317523 Kurtosis 24.82496 132.9932 24.0054 1.44251 23.38189 77.39828 8.752685 9.5407912.03843 1.441872 29.37748 846.8163 19.00388 438.6487 23.38623 30.5693 5 Natural Log Mean - -3.01211 4.19022 2.064642 -1.29132 -4.7293 -0.13532 - 3.241273 -1.28885 -2.26748 -5.11525 -4.63784 - -0.06151 Log Variance 1.0563450.34986 1.802786 0.014271 1.514408 1.31267 0.907693 1.221048 0.0196920.16606 1.465525 2.395746 1.19685 0.591751 2.17339 3.183613.10214 1.57458 5 4 8 1 2 10.0 Percentile 0.2342 0.01 60.4782 1.47 0.0 0.0036 0.2 0.7353 4.7 0.04 0.0264 0.0022 0.0015 0.00 0.2335 20.0 Percentile 5 0.3175 0.01 64.8240 2.563 0.095 0.0046 0.3593012 0.80 8.266 0.076 0.0499 0.0034 0.00 0.005 0. 30.0 Percentile 0.3921 0.026 66.85385 4.013 0.14226 0.0056 0.50 0.83412 12.9453 0.10134 0.0685 0.0044 0.00423 0.01257 0.373 40.0 Percentile 0.47 0.034 5 67.872 5.921 0.1994 0.0064 30.6 0.8606 19.094 0.13 0.0865 0.0054 0.00 0.0197 0.48918 50.0 Percentile (median) 0.57875 0.042 68.499 9.60475 0.2719 0.0074 0.911659 0.87 30.9976 0.20592 0.1076 0.00635 0.00866 0.03 0.69 60.0 Percentile 0.7294 0.051 68.887 16.3403 0.36 0.0086 1.18735 0.8957018 52.7612 0.39 0.1302 0.0076 0.0116 0.05 1.082 70.0 Percentile 1.033401 0.074 69.186 23.3 0.502 0.0103 1.5384 0.91115 75.266 0.73088 0.1674 0.00 0.01 0.13025 1.715 80.0 Percentile 1.6909 0.1178014 69.5104 26.8924 0.796 0.01 2.2731 0.92 86.6642 1.24 0.23365 0.01089 0.02566 0.3118 3.05853 90.0 Percentile 3.10 0.38815 69.8773 28.6 1.304 0.024 3.5568 0.94638 92.394 2.34358 0.38375 0.01445 0.0884 0.6312 6.25 5 7 9 7 8 Trimean 0.7012 0.04795 68.0882 11.99735 0.3254 0.007875 1.01645 0.873651 38.68851 0.35595 0.11745 0.00662 0.010175 0.0675 1.02 Biweight 0.638953 0.041148 68.30865 12.84317 0.3093195 0.007328 0.973228 0.875667 41.42834 0.326662 0.10976 0.0064685 0.00882 0.055115 0.92863 6 5 2 2 Correlation Coefficient Table A 2 O 3 % CaO Fe% FeO H 2 O K 2 O LOI% Lump% Mag MgO Mn N 2 O P S SiO 2 % l % % % % % % % a % % % A 2 O 3 % 1 0.0405 -0.9007 0.1207 0.1171 0.4566 0.62 0.012 0.1206 0.8676 0.0688 0.3055 0.0845 0.1618 0.7474 CaOl 0.0405 1 -0.2089 0.2787 -0.0711 0.0198 0.31197 -0.06672 0.2789 0.2169 0.2262 -0.0041 0.3091 0.1918 0.0657 Fe% -0.9007 -0.2089 1 -0.1698 -0.0966 -0.4145 -0.7111 -0.0221 -0.1698 -0.8934 -0.2111 -0.2923 -0.1413 -0.2188 -0.8919 FeO% 0.1207 0.2787 -0.1698 1 -0.3328 0.1018 0.4277 0.094 1 0.3223 0.3388 -0.1849 0.1766 0.4361 0.2721 % 7 H 2 O 0.1171 -0.0711 -0.0966 -0.3328 1 -0.0187 0.0776 -0.2788 -0.3323 -0.0428 -0.0251 0.23 -0.0393 -0.0939 0.0098 % 9 K 2 O% 0.4566 0.0198 -0.4145 0.1018 -0.0187 1 0.2537 -0.0035 0.1018 0.4032 0.0613 0.46 0.0363 0.1123 0.3736 LOI% 0.62 0.3119 -0.7111 0.4277 0.0776 0.2537 1 0.051 0.4279 0.6361 0.3 30.1 0.0784 0.3764 0.5651 Lump% 0.0127 2 -0.0667 -0.0221 0.0947 -0.2788 -0.0035 0.0517 7 1 0.0934 0.0261 0.00573 -0.17292 -0.0826 0.0226 0.0641 Mag% 0.1206 0.2789 -0.1698 1 -0.3323 0.1018 0.4279 0.093 1 0.3222 0.3389 -0.1841 0.1767 0.4359 0.2719 MgO% 0.8676 0.2169 -0.8934 0.3223 -0.0428 0.4032 0.6361 0.02614 0.3222 1 0.1721 0.2139 0.1469 0.1985 0.759 Mn% 0.0688 0.2262 -0.2111 0.3388 -0.0251 0.0613 0.33 0.0057 0.3389 0.1721 1 0.0215 0.2013 0.2606 0.1239 Na 2 O% 0.3055 -0.0041 -0.2923 -0.1849 0.239 0.463 0.12 -0.1729 -0.1841 0.2139 0.0215 1 -0.0136 -0.0054 0.2347 P% 0.0845 0.3091 -0.1413 0.1766 -0.0393 0.0363 0.0784 -0.0826 0.1767 0.1469 0.2013 -0.0136 1 0.07 0.1328 S% 0.1618 0.1918 -0.2188 0.4361 -0.0939 0.1123 0.3764 0.0226 0.4359 0.1985 0.2606 -0.0054 0.07 1 0.1814 SiO 2 % 0.7474 0.0657 -0.8919 0.2721 0.0098 0.3736 0.5651 0.0641 0.2719 0.759 0.1239 0.2347 0.1328 0.1814 1

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Table 17-7: Comp_200 Statistics

Variable Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Number of samples 425 425 425 425 425 425 277 425 425 425 425 425 425 425 425 Minimum value 0.054 0.005 19.368 0.468 0.0356 0.002 0.01 0.162 1.508 0.025 0.004 0.001 0.001 0.005 0.142 Maximum value 15.72 5.486 70.556 29.5606 4.396 0.1859 11.2 0.9836 95.2639 16.46 0.99 0.0412 4.024 1.214 27.76 Mean 1.262395 0.295402 64.00598 14.15945 0.774921 0.011728 1.48774 0.801757 45.63966 1.427405 0.11777 0.007461 0.080697 0.146286 5.316978 Median 0.762 0.09 65.228 14.784 0.514 0.008 0.718 0.8195 47.642 0.832 0.0836 0.0064 0.0292 0.022 4.1774 Geometric Mean 0.812503 0.108252 63.71778 9.1731 0.455252 0.008657 0.810714 0.79149 29.57023 0.630031 0.082286 0.006206 0.02436 0.033896 3.275058 Variance 2.587593 0.334436 29.88059 98.13828 0.630031 0.000224 3.44549 0.012975 1019.397 3.057222 0.013624 2.77E-05 0.050416 0.052319 21.49978 Standard Deviation 1.6086 0.578304 5.466314 9.906477 0.793745 0.014976 1.856203 0.113909 31.928 1.748491 0.116724 0.005265 0.224536 0.228733 4.636786 Coefficient of variation 1.274244 1.957685 0.085403 0.699637 1.024291 1.276923 1.247666 0.142074 0.699567 1.224944 0.991118 0.705569 2.782438 1.563602 0.872072

Skewness 4.076936 4.867555 -2.2792 0.046006 1.745062 6.097548 2.632617 -1.25476 0.045843 2.890566 2.997129 2.940046 13.35419 2.157637 1.390358 Kurtosis 26.64574 32.84525 14.01614 1.428927 6.245142 55.05027 11.07374 6.87354 1.428837 18.14798 15.77608 15.427343 226.2982 7.959908 5.688759

Natural Log Mean -0.20764 -2.22329 4.154464 2.216275 -0.7869 -4.74938 -0.20984 -0.23384 3.386768 -0.46199 -2.49755 -5.082256 -3.7148 -3.38445 1.186336 Log Variance 0.791307 1.958215 0.010338 1.217196 1.223919 0.461034 1.332466 0.030471 1.2169 2.141081 0.754094 0.370073 2.850253 3.445769 1.267202

10.0 Percentile 0.316 0.02 56.632 1.7702 0.095 0.0042 0.198 0.6718 5.7062 0.072 0.0296 0.003 0.0017 0.005 0.586 20.0 Percentile 0.4045 0.03 60.6179 2.930001 0.1691 0.0052 0.33915 0.7171 9.445 0.1264 0.04215 0.004 0.0058 0.005 1.1812 30.0 Percentile 0.494 0.042 62.52 5.194 0.248 0.006 0.468001 0.7408 16.752 0.272 0.0588 0.0048 0.0098 0.005 2.0546 40.0 Percentile 0.6091 0.064 64.061 8.3069 0.359 0.0069 0.59 0.77735 26.7779 0.4835 0.06915 0.00555 0.0164 0.008551 2.9605 50.0 Percentile (median) 0.762 0.09 65.228 14.784 0.514 0.008 0.718 0.8195 47.642 0.832 0.0836 0.0064 0.0292 0.022 4.1774 60.0 Percentile 0.917 0.16 66.4652 18.929 0.67165 0.0092 1.09 0.8454 61.0007 1.2076 0.0973 0.0072 0.046 0.068 5.429 70.0 Percentile 1.124 0.2633 67.682 22.794 0.8344 0.0112 1.53655 0.8768 73.4594 1.7408 0.1236 0.0082 0.07 0.1504 7.052 80.0 Percentile 1.558 0.4329 68.3836 25.216 1.2482 0.01325 2.181 0.9026 81.2644 2.4551 0.15745 0.0096 0.1181 0.2794 8.56915 90.0 Percentile 2.874 0.645 69.058 27.262 1.972 0.0196 3.882 0.9356 87.8586 3.602 0.2356 0.0127 0.178 0.4851 11.6008

Trimean 0.819175 0.137025 64.9717 14.37475 0.56525 0.008475 0.91165 0.814575 46.32588 0.977162 0.089125 0.00655 0.040688 0.064 4.474575 Biweight 0.756672 0.135776 64.93495 14.20149 0.555909 0.008161 0.884237 0.812556 45.77437 1.027275 0.086179 0.006377 0.039972 0.063555 4.603067

Correlation Coefficient Table

Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Al2O3% 1 0.0836 -0.7487 0.0811 0.2343 0.335 0.6413 -0.0082 0.0811 0.7286 0.0211 0.1842 0.0642 0.1566 0.4105 CaO% 0.0836 1 -0.2896 0.2692 -0.0683 0.041 0.2823 0.0788 0.2691 0.332 0.301 0.0329 0.2832 0.2199 0.0761 Fe% -0.7487 -0.2896 1 -0.1711 -0.1911 -0.3083 -0.6237 -0.0099 -0.1712 -0.8692 0.0166 -0.1903 -0.1674 -0.1908 -0.8506 FeO% 0.0811 0.2692 -0.1711 1 -0.4232 0.1027 0.3415 0.2793 1 0.2946 0.2497 -0.2459 0.1086 0.5792 0.269

H2O% 0.2343 -0.0683 -0.1911 -0.4232 1 0.0632 0.0631 -0.3421 -0.4231 0.1237 -0.17 0.2466 -0.0058 -0.1632 0.0676

K2O% 0.335 0.041 -0.3083 0.1027 0.0632 1 0.2365 0.1836 0.1027 0.207 -0.0611 0.239 -0.0021 0.1203 0.2585 LOI% 0.6413 0.2823 -0.6237 0.3415 0.0631 0.2365 1 0.0949 0.3415 0.6239 0.1857 0.0891 0.0273 0.2542 0.3666 Lump% -0.0082 0.0788 -0.0099 0.2793 -0.3421 0.1836 0.0949 1 0.2793 -0.0124 -0.1854 -0.1167 0.056 0.2031 0.0641 Mag% 0.0811 0.2691 -0.1712 1 -0.4231 0.1027 0.3415 0.2793 1 0.2946 0.2497 -0.2459 0.1087 0.5792 0.2691 MgO% 0.7286 0.332 -0.8692 0.2946 0.1237 0.207 0.6239 -0.0124 0.2946 1 0.0489 0.1895 0.1683 0.2279 0.6066 Mn% 0.0211 0.301 0.0166 0.2497 -0.17 -0.0611 0.1857 -0.1854 0.2497 0.0489 1 -0.1802 0.0725 0.1492 -0.1344

Na2O% 0.1842 0.0329 -0.1903 -0.2459 0.2466 0.239 0.0891 -0.1167 -0.2459 0.1895 -0.1802 1 0.0366 -0.0623 0.0616 P% 0.0642 0.2832 -0.1674 0.1086 -0.0058 -0.0021 0.0273 0.056 0.1087 0.1683 0.0725 0.0366 1 0.0232 0.1397 S% 0.1566 0.2199 -0.1908 0.5792 -0.1632 0.1203 0.2542 0.2031 0.5792 0.2279 0.1492 -0.0623 0.0232 1 0.1679 SiO2% 0.4105 0.0761 -0.8506 0.269 0.0676 0.2585 0.3666 0.0641 0.2691 0.6066 -0.1344 0.0616 0.1397 0.1679 1

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Table 17-8: Comp_300 Statistics

Variable Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Number of samples 153 153 153 153 153 153 123 153 153 153 153 153 153 153 153 Minimum value 0.14 0.01 16.108 0.8375 0.05 0.002 0.04 0.452 2.7014 0.01 0.004 0.001 0.0022 0.005 0.354 Maximum value 14.38 1.8714 70.888 26.7 5.1128 0.1692 10.068 0.9767 86.042 19.74 0.454 0.0878 1.024 0.912 36.168 Mean 1.390427 0.227373 64.05142 8.661868 1.037273 0.008812 1.396788 0.780922 27.91896 1.57654 0.065067 0.009821 0.074968 0.051683 4.534331 Median 0.606 0.1328 66.562 6.65 0.564 0.006 0.632 0.795 21.4447 0.721 0.046 0.0054 0.0404 0.0065 2.4563 Geometric Mean 0.815603 0.123467 63.33435 6.326969 0.572223 0.006288 0.696288 0.773691 20.3934 0.664615 0.044961 0.00651 0.035251 0.014118 2.628011 Variance 4.654056 0.061439 61.91152 43.43159 1.376219 0.000236 4.049081 0.010386 451.1428 6.770963 0.0037 0.000176 0.015529 0.012054 31.93656 Standard Deviation 2.157326 0.247868 7.868387 6.590265 1.173124 0.015353 2.012233 0.101911 21.24012 2.602107 0.06083 0.01328 0.124615 0.10979 5.651244 Coefficient of variation 1.551557 1.090139 0.122845 0.760836 1.130969 1.742208 1.440614 0.130501 0.760778 1.650518 0.934886 1.352219 1.662238 2.124305 1.246324

Skewness 3.675988 2.714669 -3.01998 1.115485 1.699747 8.260746 2.659494 -0.66288 1.115144 3.905837 2.766981 3.432712 5.311276 4.198688 2.68596 Kurtosis 17.88095 15.54563 14.26602 3.459446 5.05083 81.40749 9.826284 3.239321 3.458369 22.19358 14.58903 15.3295 36.19291 27.50708 11.86728

Natural Log Mean -0.20383 -2.09178 4.148428 1.844821 -0.55823 -5.0691 -0.36199 -0.25658 3.015211 -0.40855 -3.10196 -5.03437 -3.34525 -4.26033 0.966227 Log Variance 0.788092 1.504358 0.028426 0.715106 1.275088 0.438413 1.359451 0.019545 0.715107 1.947783 0.845126 0.624177 1.656223 2.04205 1.033569

10.0 Percentile 0.3436 0.0185 56.009 1.6854 0.115 0.00295 0.1688 0.6479 5.4317 0.07875 0.0139 0.0036 0.00555 0.005 0.7051 20.0 Percentile 0.418 0.043 61.622 3.09765 0.226 0.004 0.28 0.6964 9.98505 0.255 0.0239 0.0042 0.0104 0.005 1.08835 30.0 Percentile 0.45565 0.0643 64.7308 4.457 0.28805 0.0049 0.395 0.7434 14.369 0.388 0.0327 0.0046 0.015 0.005 1.496 40.0 Percentile 0.542 0.0947 65.80025 5.8334 0.3993 0.0054 0.5245 0.7762 18.7998 0.539 0.03865 0.005 0.02515 0.005 1.8153 50.0 Percentile (median) 0.606 0.1328 66.562 6.65 0.564 0.006 0.632 0.795 21.4447 0.721 0.046 0.0054 0.0404 0.0065 2.4563 60.0 Percentile 0.781 0.247 67.4566 7.846 0.7431 0.00675 0.80575 0.8226 25.29275 1.009 0.0589 0.0066 0.0721 0.01 3.283 70.0 Percentile 1.1277 0.2986 67.966 10.2653 1.042 0.0078 1.1008 0.8384 33.0815 1.266 0.0729 0.0077 0.0878 0.014501 4.2756 80.0 Percentile 1.4506 0.3791 68.68645 12.7644 1.61345 0.00875 1.7793 0.86 41.1506 1.71445 0.0944 0.00925 0.1074 0.0722 6.0155 90.0 Percentile 3.0416 0.496 69.383 18.935 3.039751 0.0124 3.65835 0.90825 61.024 4.028 0.1296 0.0192 0.1491 0.172 12.64375

Trimean 0.741663 0.160975 66.14073 7.299175 0.677926 0.006138 0.750813 0.7885 23.53074 0.8296 0.05075 0.005975 0.047613 0.0125 2.7611 Biweight 0.705739 0.174901 66.46363 7.283536 0.628564 0.005895 0.683893 0.791164 23.47798 0.750597 0.050129 0.005657 0.051968 0.008451 2.465994

Correlation Coefficient Table

Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Al2O3% 1 0.2196 -0.9031 0.1246 0.1423 0.1249 0.9047 0.1216 0.1246 0.9287 0.2156 0.0544 0.0391 0.086 0.6694 CaO% 0.2196 1 -0.3092 0.048 0.2206 0.0257 0.3405 0.0092 0.048 0.3679 0.2268 0.2659 0.4651 0.0751 0.1853 Fe% -0.9031 -0.3092 1 -0.1437 -0.1221 -0.1373 -0.8522 -0.1879 -0.1438 -0.9216 -0.1629 -0.058 -0.149 -0.2362 -0.9046 FeO% 0.1246 0.048 -0.1437 1 -0.2805 -0.0527 0.3756 0.0364 1 0.1819 -0.1299 -0.1705 0.1388 0.4648 0.2388

H2O% 0.1423 0.2206 -0.1221 -0.2805 1 0.0008 0.2083 -0.4847 -0.2806 0.1172 0.0222 0.4697 0.0257 -0.1359 0.0309

K2O% 0.1249 0.0257 -0.1373 -0.0527 0.0008 1 0.0965 0.2354 -0.0527 0.0699 -0.0322 0.5001 -0.0227 -0.0058 0.1491 LOI% 0.9047 0.3405 -0.8522 0.3756 0.2083 0.0965 1 0.0828 0.3757 0.8759 0.2443 0.0307 0.091 0.2504 0.6241 Lump% 0.1216 0.0092 -0.1879 0.0364 -0.4847 0.2354 0.0828 1 0.0365 0.1568 0.1144 -0.1683 0.0644 0.1363 0.1936 Mag% 0.1246 0.048 -0.1438 1 -0.2806 -0.0527 0.3757 0.0365 1 0.1819 -0.1297 -0.1705 0.1389 0.4649 0.2388 MgO% 0.9287 0.3679 -0.9216 0.1819 0.1172 0.0699 0.8759 0.1568 0.1819 1 0.2083 0.0321 0.1476 0.1912 0.7038 Mn% 0.2156 0.2268 -0.1629 -0.1299 0.0222 -0.0322 0.2443 0.1144 -0.1297 0.2083 1 -0.005 0.3227 0.0863 0.0118

Na2O% 0.0544 0.2659 -0.058 -0.1705 0.4697 0.5001 0.0307 -0.1683 -0.1705 0.0321 -0.005 1 0.0475 -0.1229 0.0085 P% 0.0391 0.4651 -0.149 0.1388 0.0257 -0.0227 0.091 0.0644 0.1389 0.1476 0.3227 0.0475 1 0.1777 0.1486 S% 0.086 0.0751 -0.2362 0.4648 -0.1359 -0.0058 0.2504 0.1363 0.4649 0.1912 0.0863 -0.1229 0.1777 1 0.3158 SiO2% 0.6694 0.1853 -0.9046 0.2388 0.0309 0.1491 0.6241 0.1936 0.2388 0.7038 0.0118 0.0085 0.1486 0.3158 1

Project No.: 165926 Page 17-10 January 2011

Baffinland Iron Mines Corporation Mary River, Baffin Island Mary River Iron Ore Trucking NI 43-101 Technical Report

Table 17-9: Comp_400 Statistics

Variable Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Number of samples 27 27 27 27 27 27 26 27 27 27 27 27 27 27 27 Minimum value 0.224 0.026 48.5144 1.014 0.08 0.0036 0 0.724 3.268 0.034 0.0056 0.0036 0.009 0.005 0.97 Maximum value 2.204 0.988 68.984 9.2456 2.6484 0.023 3.396 0.868 29.7956 5.6984 0.2062 0.011 0.2078 0.0652 24.0096 Mean 0.830589 0.32723 63.50416 3.857963 0.705259 0.00847 1.156473 0.78053 12.43238 1.910159 0.046941 0.00633 0.077526 0.011167 5.041678 Median 0.668 0.282 64.172 3.248 0.408 0.0082 0.914 0.782 10.468 1.86 0.0356 0.0062 0.0654 0.005 4.034 Geometric Mean 0.677524 0.252398 63.31894 3.373639 0.455127 0.007338 Not Calculated0.779907 10.87166 1.192784 0.032455 0.006139 0.062089 0.00728 3.912263 Variance 0.311608 0.049086 21.70821 3.773718 0.448121 2.39E-05 0.617255 0.000979 39.18758 2.221292 0.001789 2.71E-06 0.002209 0.000225 19.51118 Standard Deviation 0.558218 0.221554 4.659206 1.942606 0.669418 0.004888 0.785656 0.031281 6.259999 1.4904 0.042299 0.001646 0.046999 0.015008 4.417147 Coefficient of variation 0.672075 0.677061 0.073369 0.503532 0.94918 0.577016 0.679355 0.040077 0.503524 0.780249 0.901106 0.259984 0.606239 1.344015 0.876126

Skewness 1.200145 1.198082 -1.41604 0.787445 1.452275 1.385991 1.334207 0.331888 0.787477 1.037348 2.005914 1.175562 0.791104 2.591739 2.943378 Kurtosis 3.374503 4.126845 4.96233 3.284316 4.403017 4.476041 4.13239 3.607315 3.284826 3.616169 7.869446 4.386192 3.176734 8.51976 12.87601

Natural Log Mean -0.38931 -1.37675 4.148184 1.215992 -0.78718 -4.91471 Not Calculated-0.24858 2.386159 0.17629 -3.42789 -5.09314 -2.77919 -4.92263 1.364116 Log Variance 0.400765 0.647574 0.006096 0.288951 0.927601 0.275654 Not Calculated0.001592 0.288947 1.471127 0.800707 0.059254 0.542487 0.560379 0.478481

10.0 Percentile 0.29 0.139 56.72 1.81 0.1395 0.0036 0.47435 0.736 5.832 0.32 0.00945 0.0046 0.0339 0.005 1.468 20.0 Percentile 0.385 0.16 60.0654 2.168 0.167001 0.004 0.571 0.752 6.986 0.398 0.014601 0.0052 0.0378 0.005 2.048501 30.0 Percentile 0.44 0.183 63.233 2.567 0.278 0.0052 0.741651 0.765 8.272 0.907 0.01985 0.0053 0.0436 0.005 3.058 40.0 Percentile 0.59 0.215 63.866 2.995 0.308 0.0062 0.853 0.775 9.652 1.6139 0.0282 0.0059 0.0495 0.005 3.821001 50.0 Percentile (median) 0.668 0.282 64.172 3.248 0.408 0.0082 0.914 0.782 10.468 1.86 0.0356 0.0062 0.0654 0.005 4.034 60.0 Percentile 0.739 0.3155 64.649 4.318 0.6199 0.00835 1.005 0.791 13.914 2.048 0.0385 0.0064 0.085501 0.005 4.278 70.0 Percentile 0.8502 0.379 66.302 4.74945 0.8188 0.00925 1.241 0.797 15.30495 2.221 0.0582 0.006701 0.09825 0.00605 4.872 80.0 Percentile 1.31955 0.504 67.4635 5.5715 1.348 0.0114 1.51745 0.8016 17.955 2.5826 0.0803 0.0072 0.1269 0.0081 7.0902 90.0 Percentile 1.7628 0.646 68.426 6.22155 1.499 0.01515 2.518 0.812 20.04815 4.25645 0.09235 0.0084 0.13835 0.0294 8.06075

Trimean 0.69655 0.28675 64.56925 3.477 0.49625 0.007763 0.971 0.7808 11.20525 1.67625 0.039 0.006101 0.070575 0.005288 4.1205 Biweight 0.670576 0.277945 64.65625 3.607774 0.494691 0.0075 0.909796 0.780108 11.62615 1.663466 0.038314 0.006005 0.071496 0.005211 3.962352

Correlation Coefficient Table

Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Al2O3% 1 0.5251 -0.6107 0.3277 0.3982 0.4541 0.8026 0.1963 0.3276 0.8005 0.2547 0.141 0.606 0.1592 0.2877 CaO% 0.5251 1 -0.6662 0.0906 0.2667 0.1349 0.7447 0.2554 0.0906 0.6051 0.6204 0.1623 0.6349 0.2068 0.4728 Fe% -0.6107 -0.6662 1 -0.3487 -0.2958 -0.3765 -0.6221 -0.4822 -0.3488 -0.8372 -0.1643 -0.4064 -0.7903 -0.5664 -0.9103 FeO% 0.3277 0.0906 -0.3487 1 0.1585 0.4309 -0.0231 -0.3133 1 0.0945 -0.4021 0.428 0.5129 0.4693 0.43

H2O% 0.3982 0.2667 -0.2958 0.1585 1 0.0487 0.6114 0.3112 0.1586 0.3647 0.2762 0.3461 0.0096 -0.0531 0.1384

K2O% 0.4541 0.1349 -0.3765 0.4309 0.0487 1 0.2279 0.0937 0.4309 0.3856 -0.1131 0.3632 0.3562 0.2603 0.3093 LOI% 0.8026 0.7447 -0.6221 -0.0231 0.6114 0.2279 1 0.4255 -0.023 0.8519 0.5886 0.2845 0.415 0.0827 0.261 Lump% 0.1963 0.2554 -0.4822 -0.3133 0.3112 0.0937 0.4255 1 -0.3133 0.4499 0.3748 0.2267 0.1677 0.1182 0.4308 Mag% 0.3276 0.0906 -0.3488 1 0.1586 0.4309 -0.023 -0.3133 1 0.0945 -0.4021 0.428 0.5129 0.4693 0.43 MgO% 0.8005 0.6051 -0.8372 0.0945 0.3647 0.3856 0.8519 0.4499 0.0945 1 0.3135 0.1944 0.5582 0.2754 0.5643 Mn% 0.2547 0.6204 -0.1643 -0.4021 0.2762 -0.1131 0.5886 0.3748 -0.4021 0.3135 1 0.0926 0.0137 -0.0228 -0.0504

Na2O% 0.141 0.1623 -0.4064 0.428 0.3461 0.3632 0.2845 0.2267 0.428 0.1944 0.0926 1 0.2886 0.5377 0.478 P% 0.606 0.6349 -0.7903 0.5129 0.0096 0.3562 0.415 0.1677 0.5129 0.5582 0.0137 0.2886 1 0.5526 0.7275 S% 0.1592 0.2068 -0.5664 0.4693 -0.0531 0.2603 0.0827 0.1182 0.4693 0.2754 -0.0228 0.5377 0.5526 1 0.6559 SiO2% 0.2877 0.4728 -0.9103 0.43 0.1384 0.3093 0.261 0.4308 0.43 0.5643 -0.0504 0.478 0.7275 0.6559 1

Project No.: 165926 Page 17-11 January 2011

Baffinland Iron Mines Corporation Mary River, Baffin Island Mary River Iron Ore Trucking NI 43-101 Technical Report

Table 17-10: Comp_600 Statistics

Variable Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Number of samples 28 28 28 28 28 28 22 28 28 28 28 28 28 28 28 Minimum value 0.084 0.01 41.264 3.7988 0.05 0.0026 0.29 0.758 12.2414 0.082 0.0149 0.001 0.001 0.005 0.254 Maximum value 1.546 4.976 68.6 29.08 1.66 0.12 11.5 0.982 93.714 8.162 0.7282 0.0106 0.1046 0.85 32.228 Mean 0.487057 1.181104 62.3316 20.92414 0.362707 0.015975 2.821518 0.853118 67.43166 2.185536 0.238779 0.005232 0.019686 0.240186 5.801354 Median 0.3152 0.90585 65.49385 23.59 0.221 0.0058 1.83 0.8379 76.024 1.542 0.2618 0.0044 0.0059 0.122 2.484 Geometric Mean 0.323756 0.363345 61.77515 18.51908 0.226815 0.008299 1.821667 0.85076 59.68046 1.258343 0.169496 0.004475 0.006517 0.081642 3.208717 Variance 0.195212 1.695092 58.99841 56.41645 0.14007 0.000634 6.831111 0.004124 585.9431 4.420529 0.025684 6.81E-06 0.000676 0.068608 56.59312 Standard Deviation 0.441828 1.301957 7.681042 7.511088 0.374259 0.025185 2.613639 0.064217 24.20626 2.102505 0.160263 0.002609 0.026003 0.261931 7.52284 Coefficient of variation 0.907138 1.102322 0.123229 0.358968 1.031848 1.576521 0.926324 0.075273 0.358975 0.962009 0.671176 0.498581 1.320912 1.090535 1.296739

Skewness 1.122089 1.196761 -1.65792 -1.17817 1.827252 3.009307 1.657735 0.578931 -1.17818 1.368081 0.843001 0.258757 1.618491 0.902631 2.177614 Kurtosis 2.959412 3.850088 4.46434 3.315738 6.09494 11.72163 5.981387 2.278148 3.315722 4.019114 4.129609 1.993523 5.07472 2.484478 6.98474

Natural Log Mean -1.12777 -1.0124 4.123501 2.918802 -1.48362 -4.79164 0.599752 -0.16163 4.089005 0.229796 -1.77492 -5.409325 -5.03329 -2.50541 1.165871 Log Variance 0.83319 3.799716 0.019509 0.345387 0.969563 1.028594 0.976549 0.005474 0.345417 1.42271 0.968513 0.370209 2.646254 3.262678 1.101295

10.0 Percentile 0.104 0.03 47.659 5.9863 0.05 0.0031 0.4325 0.779 19.2915 0.1888 0.0397 0.0022 0.001 0.005 1.15065 20.0 Percentile 0.133 0.0393 59.493 17.24 0.092 0.0034 0.6591 0.793 55.558 0.505 0.0854 0.003 0.001 0.01 1.5686 30.0 Percentile 0.159 0.054 62.07075 18.79665 0.129 0.0038 0.8753 0.808 60.57465 0.910001 0.1348 0.0034 0.001 0.031 1.806 40.0 Percentile 0.21505 0.219501 64.893 21.036 0.17 0.0044 1.71745 0.834 67.7916 1.2898 0.1973 0.0042 0.00285 0.0515 2.255 50.0 Percentile (median) 0.3152 0.90585 65.49385 23.59 0.221 0.0058 1.83 0.8379 76.024 1.542 0.2618 0.0044 0.0059 0.122 2.484 60.0 Percentile 0.341 1.152 65.816 25.03 0.3111 0.0096 2.998 0.8486 80.663 1.707001 0.26485 0.00605 0.013 0.192 3.202 70.0 Percentile 0.6654 1.989001 67.197 26.02 0.369 0.01495 3.457 0.885 83.852 2.99925 0.2842 0.0074 0.027 0.401 4.914 80.0 Percentile 0.809001 2.089 67.5292 26.37485 0.5181 0.01695 4.98 0.9121 84.9965 3.3016 0.348 0.008 0.03535 0.4814 6.5513 90.0 Percentile 1.2281 2.912 68.025 28.34 0.9368 0.0352 6.01435 0.963 91.332 5.989 0.4177 0.0089 0.0589 0.653 16.896

Trimean 0.38295 0.97825 64.85061 22.71416 0.238525 0.007812 2.004 0.842725 73.20016 1.771963 0.239176 0.005 0.0102 0.174375 3.12475 Biweight 0.393344 1.006973 65.40495 22.94955 0.221981 0.007549 2.181843 0.847171 73.95942 1.676082 0.230818 0.005079 0.011175 0.198511 2.674852

Correlation Coefficient Table

Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Al2O3% 1 -0.4062 0.2256 -0.3864 0.5606 0.3226 -0.4195 0.0051 -0.3864 -0.2571 -0.4466 0.0978 0.7637 0.0808 -0.1708 CaO% -0.4062 1 -0.4974 0.3105 -0.1725 -0.0392 0.5012 0.1502 0.3105 0.6182 0.5322 0.5766 -0.1682 0.0967 0.2343 Fe% 0.2256 -0.4974 1 0.0599 -0.2728 -0.076 -0.1534 -0.3523 0.0599 -0.9638 -0.2447 -0.7686 0.078 -0.4543 -0.9456 FeO% -0.3864 0.3105 0.0599 1 -0.0275 0.1263 0.2883 -0.2432 1 0.1062 0.379 0.0251 -0.3267 0.3729 -0.0687

H2O% 0.5606 -0.1725 -0.2728 -0.0275 1 0.4164 -0.3906 -0.0254 -0.0275 0.2192 -0.0741 0.4553 0.6285 0.3998 0.3552

K2O% 0.3226 -0.0392 -0.076 0.1263 0.4164 1 -0.2336 0.3141 0.1264 0.0784 -0.2947 0.2718 0.5496 0.5595 0.1152 LOI% -0.4195 0.5012 -0.1534 0.2883 -0.3906 -0.2336 1 0.4408 0.2883 0.1926 0.3266 0.0788 -0.3506 0.0932 -0.0078 Lump% 0.0051 0.1502 -0.3523 -0.2432 -0.0254 0.3141 0.4408 1 -0.2432 0.3045 -0.0582 0.2433 0.1682 0.3498 0.2284 Mag% -0.3864 0.3105 0.0599 1 -0.0275 0.1264 0.2883 -0.2432 1 0.1062 0.379 0.0251 -0.3267 0.3729 -0.0687 MgO% -0.2571 0.6182 -0.9638 0.1062 0.2192 0.0784 0.1926 0.3045 0.1062 1 0.3077 0.829 -0.0998 0.4751 0.8826 Mn% -0.4466 0.5322 -0.2447 0.379 -0.0741 -0.2947 0.3266 -0.0582 0.379 0.3077 1 0.1713 -0.5227 -0.1722 0.0816

Na2O% 0.0978 0.5766 -0.7686 0.0251 0.4553 0.2718 0.0788 0.2433 0.0251 0.829 0.1713 1 0.2785 0.4361 0.6676 P% 0.7637 -0.1682 0.078 -0.3267 0.6285 0.5496 -0.3506 0.1682 -0.3267 -0.0998 -0.5227 0.2785 1 0.172 -0.0704 S% 0.0808 0.0967 -0.4543 0.3729 0.3998 0.5595 0.0932 0.3498 0.3729 0.4751 -0.1722 0.4361 0.172 1 0.5158 SiO2% -0.1708 0.2343 -0.9456 -0.0687 0.3552 0.1152 -0.0078 0.2284 -0.0687 0.8826 0.0816 0.6676 -0.0704 0.5158 1

Project No.: 165926 Page 17-12 January 2011

Baffinland Iron Mines Corporation Mary River, Baffin Island Mary River Iron Ore Trucking NI 43-101 Technical Report

Table 17-11: Comp_700 Statistics

Variable Al2O3% CaO% Fe% FeO% H2O% K2O% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Number of samples 15 15 15 15 15 15 15 15 15 15 15 15 15 15 Minimum value 0.2378 0.1905 65.3129 27.4572 0.01 0.002 0.7556 88.4858 0.2308 0.0276 0.001 0.0555 0.005 0.6696 Maximum value 1.7462 0.8437 70.6029 30.0892 0.2426 0.103 0.8546 96.9676 1.9311 0.2308 0.0133 0.2658 2.0431 3.4222 Mean 0.698147 0.47236 69.20113 29.192467 0.090073 0.023587 0.79786 94.07855 0.68692 0.108933 0.003573 0.14582 0.44966 1.456553 Median 0.6452 0.4889 69.5074 29.3121 0.0934 0.0106 0.8001 94.4664 0.5223 0.0959 0.0018 0.1543 0.2152 1.2099 Geometric Mean 0.614595 0.430531 69.18755 29.183506 0.065171 0.013007 0.797375 94.04966 0.568627 0.089731 0.002406 0.132111 0.21446 1.322452 Variance 0.146056 0.037449 1.841282 0.514804 0.004106 0.000715 0.000776 5.348546 0.214629 0.003879 1.21E-05 0.003763 0.256839 0.48984 Standard Deviation 0.382173 0.193517 1.356938 0.717499 0.06408 0.026732 0.027864 2.312692 0.463281 0.062283 0.003482 0.061346 0.506793 0.699885 Coefficient of variation 0.547411 0.409682 0.019609 0.024578 0.711422 1.133332 0.034924 0.024583 0.674432 0.571756 0.974345 0.420694 1.127057 0.480508

Skewness 1.42049 0.233918 -1.563738 -0.937811 0.799473 1.7526 0.200429 -0.93795 1.422074 0.451751 1.53662 0.234706 1.972886 1.455864 Kurtosis 4.458071 1.943443 4.967774 3.126392 2.976839 5.503051 2.196308 3.125681 4.112394 2.110917 4.471238 1.971774 6.70763 4.679356

Natural Log Mean -0.48679 -0.84274 4.236821 3.373604 -2.73074 -4.34226 -0.22643 4.543823 -0.56453 -2.41094 -6.02969 -2.02411 -1.53963 0.279488 Log Variance 0.244527 0.197022 0.000397 0.000619 0.794159 1.237385 0.001214 0.000619 0.355679 0.435572 0.732392 0.211085 2.190129 0.181902

10.0 Percentile 0.393 0.2244 67.3615 28.0672 0.0156 0.0033 0.7633 90.4503 0.2521 0.031 0.001 0.0692 0.0226 0.7628 20.0 Percentile 0.412 0.27605 68.427 28.662 0.02815 0.00445 0.764901 92.3679 0.3389 0.04285 0.001 0.0826 0.1007 0.90895 30.0 Percentile 0.442 0.3312 69.1586 28.9464 0.0394 0.0069 0.7792 93.2874 0.3874 0.0742 0.001 0.1035 0.1289 1.1474 40.0 Percentile 0.48305 0.3604 69.32385 29.1692 0.06395 0.0086 0.7907 94.0019 0.49035 0.08 0.0016 0.10885 0.17645 1.1684 50.0 Percentile (median) 0.6452 0.4889 69.5074 29.3121 0.0934 0.0106 0.8001 94.4664 0.5223 0.0959 0.0018 0.1543 0.2152 1.2099 60.0 Percentile 0.70065 0.5644 69.68285 29.5491 0.0993 0.0189 0.8082 95.2279 0.5889 0.12365 0.00275 0.1757 0.4302 1.3888 70.0 Percentile 0.7244 0.6092 70.1742 29.6882 0.1041 0.0272 0.8116 95.6794 0.7204 0.134 0.0031 0.1873 0.5977 1.5018 80.0 Percentile 0.89215 0.6204 70.3208 29.78005 0.1383 0.04005 0.8228 95.9737 1.04645 0.1706 0.0068 0.1927 0.6917 1.9005 90.0 Percentile 1.3079 0.755 70.3905 30.006 0.1866 0.0568 0.8303 96.6985 1.3305 0.2063 0.0081 0.232 0.9343 2.4013

Trimean 0.623225 0.476175 69.56954 29.314438 0.08385 0.014375 0.7966 94.47308 0.5447 0.099288 0.002363 0.148663 0.2985 1.271525 Biweight 0.59624 0.471434 69.58398 29.326678 0.084561 0.014778 0.797806 94.51156 0.512894 0.102258 0.002467 0.146154 0.30716 1.208041

Correlation Coefficient Table

Al2O3% CaO% Fe% FeO% H2O% K2O% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Al2O3% 1 0.6126 -0.9399 -0.5212 0.3021 -0.1026 0.0251 -0.5213 0.9304 0.0117 -0.026 0.5947 -0.1999 0.9622 CaO% 0.6126 1 -0.7271 -0.2974 0.5295 -0.129 0.0388 -0.2974 0.6521 0.3744 0.1923 0.9994 -0.2399 0.7403 Fe% -0.9399 -0.7271 1 0.4772 -0.3895 0.0964 0.0976 0.4774 -0.9458 -0.165 0.0202 -0.7109 0.0452 -0.9656 FeO% -0.5212 -0.2974 0.4772 1 -0.1862 0.3622 0.2469 1 -0.4233 -0.3894 0.2241 -0.2733 0.2057 -0.4378

H2O% 0.3021 0.5295 -0.3895 -0.1862 1 -0.3145 -0.4004 -0.1864 0.5148 0.4818 -0.1122 0.5206 -0.2753 0.4106

K2O% -0.1026 -0.129 0.0964 0.3622 -0.3145 1 0.05 0.3619 -0.1123 -0.4444 0.6641 -0.1198 0.004 -0.1012 LOI% 0 0 0 0 0 0 0 0 0 0 0 0 0 0 Lump% 0.0251 0.0388 0.0976 0.2469 -0.4004 0.05 1 0.2471 -0.1327 -0.4507 0.2622 0.0578 -0.0148 0.0314 Mag% -0.5213 -0.2974 0.4774 1 -0.1864 0.3619 0.2471 1 -0.4235 -0.3895 0.2238 -0.2733 0.2055 -0.4379 MgO% 0.9304 0.6521 -0.9458 -0.4233 0.5148 -0.1123 -0.1327 -0.4235 1 0.129 -0.138 0.6336 -0.1866 0.9447 Mn% 0.0117 0.3744 -0.165 -0.3894 0.4818 -0.4444 -0.4507 -0.3895 0.129 1 -0.2957 0.362 -0.2147 0.1268

Na2O% -0.026 0.1923 0.0202 0.2241 -0.1122 0.6641 0.2622 0.2238 -0.138 -0.2957 1 0.2086 -0.0483 0.0293 P% 0.5947 0.9994 -0.7109 -0.2733 0.5206 -0.1198 0.0578 -0.2733 0.6336 0.362 0.2086 1 -0.2286 0.7248 S% -0.1999 -0.2399 0.0452 0.2057 -0.2753 0.004 -0.0148 0.2055 -0.1866 -0.2147 -0.0483 -0.2286 1 -0.2525 SiO2% 0.9622 0.7403 -0.9656 -0.4378 0.4106 -0.1012 0.0314 -0.4379 0.9447 0.1268 0.0293 0.7248 -0.2525 1

Project No.: 165926 Page 17-13 January 2011

Baffinland Iron Mines Corporation Mary River, Baffin Island Mary River Iron Ore Trucking NI 43-101 Technical Report

Table 17-12: Comp_1300 Statistics

Variable Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Number of samples 70 70 70 70 70 70 60 70 70 70 70 70 70 70 70 Minimum value 0.126 0.044 29.1074 20.818 0.03 0.0028 0.2 0.7672 67.0896 0.69 0.3166 0.001 0.001 0.046 0.38 Maximum value 15.7822 1.442 69.358 31.64 0.9094 1.6295 9.74 0.97 101.966 9.6326 2.739 0.0324 0.2164 5.878 30.3956 Mean 2.155044 0.393177 60.41358 27.99913 0.153913 0.06329 3.402927 0.916183 90.23692 2.380514 1.297326 0.006847 0.035986 1.221379 5.019233 Median 0.737 0.297 62.75295 28.19 0.0609 0.0102 3.01945 0.926 90.848 1.8425 1.32775 0.0056 0.0091 1.012 1.786 Geometric Mean 0.981971 0.273958 59.64987 27.90912 0.097157 0.013342 2.672705 0.915023 89.94702 1.910165 1.170694 0.004873 0.011746 0.917911 2.52782 Variance 9.701856 0.101704 72.79882 4.817478 0.036392 0.060969 4.334857 0.002044 50.00201 3.218929 0.278365 4.29E-05 0.002745 0.851076 44.12524 Standard Deviation 3.11478 0.31891 8.532222 2.194875 0.190768 0.246919 2.082032 0.04521 7.07121 1.794137 0.527603 0.006548 0.052396 0.922538 6.642683 Coefficient of variation 1.445344 0.811111 0.14123 0.078391 1.239454 3.901392 0.611836 0.049346 0.078363 0.753676 0.406685 0.956336 1.456022 0.755325 1.323446

Skewness 2.314415 1.151886 -1.86921 -0.69067 2.491227 5.471521 0.696256 -1.07769 -0.69228 1.822709 0.090772 2.527247 1.717299 2.075891 2.008932 Kurtosis 8.349806 3.84409 6.305036 3.336446 8.786493 32.23542 3.078793 3.614122 3.341223 6.355484 2.550305 9.548034 4.895796 10.31357 6.402757

Natural Log Mean -0.01819 -1.29478 4.088492 3.328953 -2.33143 -4.31684 0.983091 -0.08881 4.499221 0.647189 0.157597 -5.32395 -4.44426 -0.08566 0.927357 Log Variance 1.45681 0.813191 0.028969 0.006604 0.739756 1.437826 0.615861 0.002585 0.0066 0.4051 0.236948 0.703525 2.487011 0.68923 1.270995

10.0 Percentile 0.264 0.074 48.482 24.83 0.05 0.0045 0.895 0.846351 80.02 0.897 0.53025 0.001 0.0014 0.364 0.736 20.0 Percentile 0.349 0.118 57.524 26.12225 0.05 0.006101 1.4625 0.881 84.2106 1.055 0.80565 0.00245 0.0027 0.492 1.051501 30.0 Percentile 0.408 0.172 60.65665 27.12 0.05 0.0072 1.94 0.906 87.401 1.203 0.9873 0.004 0.0041 0.602 1.217301 40.0 Percentile 0.477 0.222 61.419 27.99 0.05 0.0085 2.56475 0.917 90.201 1.3934 1.2399 0.0048 0.0068 0.83185 1.368001 50.0 Percentile (median) 0.737 0.297 62.75295 28.19 0.0609 0.0102 3.01945 0.926 90.848 1.8425 1.32775 0.0056 0.0091 1.012 1.786 60.0 Percentile 1.085 0.376 64.082 28.86 0.106701 0.01235 3.808 0.93785 93.006 2.019 1.4438 0.0062 0.0143 1.213 2.841 70.0 Percentile 1.843 0.51165 65.144 29.29 0.1426 0.01685 4.528 0.94835 94.392 2.369001 1.5896 0.0066 0.0287 1.452 4.411 80.0 Percentile 2.873 0.6804 66.659 30.01 0.184 0.0211 4.77 0.956 96.711 3.556 1.7589 0.00805 0.06295 1.9352 7.71965 90.0 Percentile 7.0545 0.8186 67.694 30.58 0.399 0.0475 6.41705 0.962 98.5495 4.8069 1.9678 0.01335 0.1266 2.346 15.3458

Trimean 0.969175 0.325 62.65778 28.11643 0.0846 0.0112 3.093475 0.92335 90.61085 1.93125 1.300075 0.0056 0.0154 1.0355 2.53475 Biweight 0.842026 0.336049 63.201 28.15988 0.083387 0.010356 3.17645 0.92235 90.75447 1.848391 1.300141 0.005077 0.01179 1.061524 2.191944

Correlation Coefficient Table

Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Al2O3% 1 0.1983 -0.946 -0.5433 0.0768 0.7107 0.2185 -0.418 -0.5432 0.871 -0.2663 0.6504 0.5178 0.2213 0.9198 CaO% 0.1983 1 -0.2463 -0.0451 -0.0071 0.0842 -0.0293 -0.2724 -0.0449 0.2476 -0.1402 0.1506 0.3936 0.0391 0.192 Fe% -0.946 -0.2463 1 0.451 -0.016 -0.6417 -0.3436 0.3451 0.4509 -0.9325 0.123 -0.6085 -0.4494 -0.2752 -0.9328 FeO% -0.5433 -0.0451 0.451 1 -0.4702 -0.4371 0.0677 0.1614 1 -0.4283 0.5241 -0.2831 -0.2475 -0.1511 -0.5555

H2O% 0.0768 -0.0071 -0.016 -0.4702 1 0.0365 -0.2461 -0.1903 -0.4698 -0.0259 -0.239 0.0139 0.0938 -0.0442 0.045

K2O% 0.7107 0.0842 -0.6417 -0.4371 0.0365 1 0.0726 -0.218 -0.4373 0.4573 -0.203 0.7037 0.2002 0.2393 0.624 LOI% 0.2185 -0.0293 -0.3436 0.0677 -0.2461 0.0726 1 0.2526 0.0676 0.4199 0.2056 0.0355 -0.2036 0.0264 0.1249 Lump% -0.418 -0.2724 0.3451 0.1614 -0.1903 -0.218 0.2526 1 0.1602 -0.2128 0.2313 -0.2874 -0.6173 -0.0986 -0.4477 Mag% -0.5432 -0.0449 0.4509 1 -0.4698 -0.4373 0.0676 0.1602 1 -0.4284 0.5237 -0.2831 -0.2464 -0.1513 -0.5552 MgO% 0.871 0.2476 -0.9325 -0.4283 -0.0259 0.4573 0.4199 -0.2128 -0.4284 1 -0.0759 0.4708 0.3937 0.1839 0.815 Mn% -0.2663 -0.1402 0.123 0.5241 -0.239 -0.203 0.2056 0.2313 0.5237 -0.0759 1 -0.1742 -0.3309 0.0981 -0.3347

Na2O% 0.6504 0.1506 -0.6085 -0.2831 0.0139 0.7037 0.0355 -0.2874 -0.2831 0.4708 -0.1742 1 0.3014 0.1661 0.6047 P% 0.5178 0.3936 -0.4494 -0.2475 0.0938 0.2002 -0.2036 -0.6173 -0.2464 0.3937 -0.3309 0.3014 1 -0.0214 0.5736 S% 0.2213 0.0391 -0.2752 -0.1511 -0.0442 0.2393 0.0264 -0.0986 -0.1513 0.1839 0.0981 0.1661 -0.0214 1 0.2447 SiO2% 0.9198 0.192 -0.9328 -0.5555 0.045 0.624 0.1249 -0.4477 -0.5552 0.815 -0.3347 0.6047 0.5736 0.2447 1

Project No.: 165926 Page 17-14 January 2011

Baffinland Iron Mines Corporation Mary River, Baffin Island Mary River Iron Ore Trucking NI 43-101 Technical Report

Table 17-13: Comp_1400 Statistics

Variable Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Number of samples 73 73 73 73 73 73 69 73 73 73 73 73 73 73 73 Minimum value 0.16 0.021 42.802 15.4125 0.032 0.0026 0.08 0.4912 49.6678 0.132 0.0935 0.001 0.001 0.086 0.217 Maximum value 8.172 2.096 69.894 32.86 1.821 0.2152 10.56 0.9754 105.964 7.218 2.8148 0.0194 0.1912 5.774 22.3096 Mean 1.886518 0.567967 60.76743 27.7848 0.30739 0.019536 3.577754 0.881118 89.56471 2.354479 0.920166 0.006789 0.026551 1.374988 3.923064 Median 1.512 0.46 61.97 28.5 0.1668 0.0088 3.13 0.914 91.906 2.116 0.8008 0.006 0.01 0.976 2.558 Geometric Mean 1.393781 0.339227 60.51249 27.58625 0.187796 0.010648 2.498607 0.874803 88.92392 1.783229 0.734872 0.005687 0.011221 1.014366 2.509612 Variance 2.312604 0.243411 29.15051 8.959327 0.115421 0.001349 6.124815 0.009363 93.20805 2.313838 0.34803 1.43E-05 0.001603 1.140045 17.64935 Standard Deviation 1.520725 0.493367 5.399121 2.993213 0.339737 0.036728 2.474836 0.096763 9.654432 1.521131 0.589941 0.003779 0.040039 1.067729 4.201113 Coefficient of variation 0.806102 0.868654 0.088849 0.107728 1.10523 1.880072 0.691729 0.109818 0.107793 0.646058 0.641124 0.556614 1.508006 0.776537 1.070875

Skewness 1.753353 1.053204 -0.8158 -2.23795 2.198211 4.286036 0.850721 -1.84596 -2.23559 0.815116 1.039532 0.946321 2.434214 1.543413 2.584985 Kurtosis 6.708493 3.784921 3.55636 8.884415 8.507168 21.94007 3.302434 6.605323 8.878776 3.37657 3.86399 3.985314 8.526306 5.886673 10.56703

Natural Log Mean 0.33202 -1.08108 4.10285 3.317317 -1.6724 -4.54237 0.915733 -0.13376 4.487781 0.578426 -0.30806 -5.169656 -4.49 0.014264 0.920128 Log Variance 0.650343 1.350529 0.008698 0.016048 0.971696 0.793455 1.100809 0.015802 0.016067 0.730853 0.510985 0.417507 1.73216 0.691056 0.949868

10.0 Percentile 0.4855 0.056 53.218 25.288 0.05 0.0048 0.742 0.7474 81.5198 0.525 0.2464 0.0023 0.0021 0.313 0.713 20.0 Percentile 0.719 0.136 56.721 26.754 0.0735 0.0056 1.34165 0.8181 86.22 1.052 0.429901 0.00405 0.0038 0.538 1.058 30.0 Percentile 0.848 0.185 58.3575 27.2 0.1 0.006 2.16465 0.88 87.658 1.59705 0.54395 0.0048 0.0054 0.69 1.392 40.0 Percentile 1.103 0.25 60.374 28.01 0.128 0.0071 2.6856 0.904 90.296 1.7924 0.6277 0.0052 0.00675 0.8941 2.077 50.0 Percentile (median) 1.512 0.46 61.97 28.5 0.1668 0.0088 3.13 0.914 91.906 2.116 0.8008 0.006 0.01 0.976 2.558 60.0 Percentile 1.783 0.59 63.076 28.86 0.221 0.0101 3.727 0.925 93.007 2.43775 0.984601 0.007 0.0144 1.4097 3.811 70.0 Percentile 2.364 0.836 63.811 29 0.339 0.0124 4.36585 0.9316 93.519 2.783 1.0842 0.007951 0.0198 1.68 4.367 80.0 Percentile 2.884 0.94 65.433 29.62 0.5235 0.0176 5.60665 0.9533 95.488 3.454 1.4516 0.009 0.03425 1.994 4.997 90.0 Percentile 3.35 1.231001 67.284 30.52 0.7146 0.03785 7.104 0.966 98.42 4.801 1.643 0.0123 0.0854 3.087 8.5695

Trimean 1.617 0.4965 61.56188 28.3475 0.20665 0.009425 3.165501 0.90265 91.3995 2.149 0.832487 0.006275 0.013475 1.10275 2.75325 Biweight 1.590319 0.508085 61.39197 28.45947 0.208642 0.008681 3.277268 0.906806 91.74465 2.161909 0.822052 0.006156 0.01184 1.133821 2.788778

Correlation Coefficient Table

Al2O3% CaO% Fe% FeO% H2O% K2O% LOI% Lump% Mag% MgO% Mn% Na2O% P% S% SiO2%

Al2O3% 1 -0.2307 -0.7681 -0.1366 0.1266 0.5552 0.0994 0.0192 -0.1368 0.685 0.202 0.4328 0.1207 -0.1551 0.6897 CaO% -0.2307 1 -0.0476 0.2965 -0.3065 -0.1533 0.1423 0.087 0.2971 0.0729 0.1457 -0.106 0.1233 0.0223 -0.2133 Fe% -0.7681 -0.0476 1 -0.0405 0.0951 -0.4299 -0.5035 -0.1934 -0.0407 -0.8916 -0.6045 -0.3496 -0.1826 0.0748 -0.7246 FeO% -0.1366 0.2965 -0.0405 1 -0.3939 0.0353 0.091 -0.0579 1 0.1174 0.2934 -0.0089 0.0683 0.0333 -0.1007

H2O% 0.1266 -0.3065 0.0951 -0.3939 1 0.0227 -0.1011 -0.3553 -0.394 -0.2339 -0.2859 0.0537 -0.1217 0.1464 0.0244

K2O% 0.5552 -0.1533 -0.4299 0.0353 0.0227 1 -0.024 0.0032 0.0347 0.4032 0.1359 0.5065 0.0467 -0.1593 0.3989 LOI% 0.0994 0.1423 -0.5035 0.091 -0.1011 -0.024 1 0.1859 0.0923 0.5438 0.7085 -0.0493 -0.1111 0.2788 -0.0978 Lump% 0.0192 0.087 -0.1934 -0.0579 -0.3553 0.0032 0.1859 1 -0.0582 0.2116 0.1505 0.1874 0.0624 -0.0214 0.1594 Mag% -0.1368 0.2971 -0.0407 1 -0.394 0.0347 0.0923 -0.0582 1 0.1176 0.2941 -0.0104 0.0688 0.0338 -0.1013 MgO% 0.685 0.0729 -0.8916 0.1174 -0.2339 0.4032 0.5438 0.2116 0.1176 1 0.7003 0.2754 0.2045 -0.203 0.4854 Mn% 0.202 0.1457 -0.6045 0.2934 -0.2859 0.1359 0.7085 0.1505 0.2941 0.7003 1 0.0142 0.1313 -0.1219 0.0722

Na2O% 0.4328 -0.106 -0.3496 -0.0089 0.0537 0.5065 -0.0493 0.1874 -0.0104 0.2754 0.0142 1 0.0247 -0.0138 0.4356 P% 0.1207 0.1233 -0.1826 0.0683 -0.1217 0.0467 -0.1111 0.0624 0.0688 0.2045 0.1313 0.0247 1 -0.272 0.2586 S% -0.1551 0.0223 0.0748 0.0333 0.1464 -0.1593 0.2788 -0.0214 0.0338 -0.203 -0.1219 -0.0138 -0.272 1 -0.2532 SiO2% 0.6897 -0.2133 -0.7246 -0.1007 0.0244 0.3989 -0.0978 0.1594 -0.1013 0.4854 0.0722 0.4356 0.2586 -0.2532 1

Project No.: 165926 Page 17-15 January 2011

Baffinland Iron Mines Corporation Mary River, Baffin Island Mary River Iron Ore Trucking NI 43-101 Technical Report

Table 17-14: Statistics Report Deposit #2 Upper 200 Zone

Variable Al2O3 % Ca % Fe % FeO % H20 % K20 % LOI Lump % Mag % MgO % Mn % Na20 % P % S % SiO2 %

Number of samples 120 101 120 101 101 101 96 101 101 101 120 101 120 120 120 Minimum value 0.06 0.008 27.867 0.371 0.02 0.002 0.006 31 1.196 0.03 0.004 0.001 0.001 0.002 0.2 Maximum value 11.804 0.38 69.9 18.12 3.54 1.042 5.873 97.6 58.392 10.692 0.17 0.223 0.106 0.956 33.428

10.0 Percentile 0.132 0.013 54.902 0.591 0.03 0.002 0.074 71.1 1.904 0.03 0.004 0.002 0.004 0.005 0.431 20.0 Percentile 0.173 0.02 59.556 0.679 0.041 0.004 0.109 78.6 2.189 0.03 0.004 0.003 0.006 0.005 0.658 30.0 Percentile 0.2 0.024 62.491 0.82 0.048 0.005 0.139 80.6 2.643 0.036 0.008 0.004 0.008 0.005 1.108 40.0 Percentile 0.228 0.03 64.809 0.961 0.066 0.006 0.187 82.7 3.097 0.048 0.01 0.005 0.009 0.005 1.86 50.0 Percentile 0.291 0.04 66.789 1.092 0.092 0.011 0.235 84.2 3.518 0.086 0.013 0.006 0.011 0.005 3.643 60.0 Percentile 0.395 0.044 68.204 1.226 0.143 0.014 0.285 87.3 3.951 0.151 0.015 0.007 0.013 0.005 5.574 70.0 Percentile 0.545 0.056 68.752 1.382 0.251 0.026 0.458 89.8 4.452 0.291 0.02 0.009 0.017 0.006 8.85 80.0 Percentile 0.922 0.063 69.172 1.775 0.39 0.058 0.657 93.3 5.72 0.829 0.028 0.015 0.023 0.008 12.294 90.0 Percentile 2.625 0.109 69.492 3.689 0.863 0.255 1.458 94.7 11.889 2.211 0.056 0.026 0.038 0.016 15.217

Mean 1.022 0.053 64.006 2.014 0.285 0.079 0.557 83.8 6.491 0.639 0.024 0.013 0.016 0.019 6.437 Variance 3.838 0.003 49.995 9.869 0.225 0.033 0.783 0.01 102.499 1.968 0.001 0.001 0 0.008 54.142 Standard Deviation 1.959 0.058 7.071 3.142 0.475 0.18 0.885 0.102 10.124 1.403 0.032 0.027 0.016 0.089 7.358 Coefficient of variation 1.918 1.082 0.11 1.56 1.668 2.298 1.589 0.122 1.56 2.194 1.334 2.006 0.991 4.778 1.143

Skew ness 3.403 3.328 -2.076 3.579 3.914 3.254 3.44 -1.848 3.579 4.341 2.692 5.614 2.528 9.949 1.68 Kurtosis 15.191 16.147 8.578 15.339 23.662 13.896 17.264 9.106 15.338 27.73 10.132 40.953 11.163 104.726 5.908

Trimean 0.38 0.04 65.974 1.136 0.136 0.017 0.291 84.9 3.66 0.171 0.014 0.006 0.012 0.006 4.716 Biw eight 0.323 0.038 65.848 1.058 0.125 0.012 0.253 85.1 3.41 0.123 0.013 0.005 0.011 0.005 4.87

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Table 17-15: Statistics Report Deposit #3 Main Zone

Variable Al2O3 % CaO % Fe % FeO % H2O % K2O % Lump % Mag % MgO % Mn % Na2O % P % S % SiO2 %

Number of samples 128 128 128 128 128 128 128 128 128 128 128 128 128 128 Minimum value 0.09 0.005 22.338 0.478 0.025 0.002 62.2 1.54 0.1 0.006 0.002 0.005 0.005 0.152 Maximum value 18.158 3.184 69.818 28.84 0.484 3.184 97.6 92.923 5.459 0.313 0.485 0.199 1.724 41.772

10.0 Percentile 0.17 0.04 60.344 1.31 0.025 0.006 71 4.221 0.351 0.016 0.004 0.01 0.005 0.511 20.0 Percentile 0.241 0.085 63.916 1.592 0.03 0.007 76.3 5.13 0.48 0.02 0.005 0.014 0.005 0.663 30.0 Percentile 0.323 0.109 65.065 1.836 0.034 0.008 78.3 5.916 0.63 0.025 0.006 0.016 0.005 0.766 40.0 Percentile 0.485 0.191 65.812 2.033 0.043 0.01 80.4 6.55 0.852 0.03 0.008 0.024 0.005 0.992 50.0 Percentile (median) 0.628 0.319 66.693 2.307 0.054 0.013 84 7.434 1.142 0.044 0.009 0.028 0.005 1.321 60.0 Percentile 0.841 0.422 67.346 2.709 0.066 0.019 85.5 8.729 1.398 0.054 0.011 0.033 0.005 1.817 70.0 Percentile 1.07 0.559 67.945 3.287 0.101 0.037 87.6 10.592 1.558 0.07 0.014 0.043 0.007 2.319 80.0 Percentile 1.433 0.935 68.551 4.754 0.117 0.061 89.9 15.316 1.742 0.083 0.025 0.061 0.01 2.852 90.0 Percentile 3.358 1.43 68.987 9.828 0.151 0.16 92.6 31.666 2.275 0.11 0.098 0.099 0.018 7.301

Mean 1.567 0.536 64.807 4.403 0.082 0.113 82.5 14.188 1.284 0.059 0.036 0.041 0.029 3.084 Variance 9.624 0.395 50.619 29.773 0.006 0.149 0.007 309.084 0.932 0.003 0.006 0.001 0.028 32.534 Standard Deviation 3.102 0.628 7.115 5.457 0.077 0.386 0.082 17.581 0.965 0.054 0.076 0.038 0.168 5.704 Coefficient of variation 1.98 1.173 0.11 1.239 0.936 3.404 0.099 1.239 0.752 0.92 2.114 0.94 5.744 1.85

Skew ness 4.005 1.853 -3.975 2.638 2.562 5.839 -0.433 2.638 1.852 2.269 3.606 2.04 8.941 4.422 Kurtosis 19.54 6.317 20.617 9.545 11.018 40.3 2.589 9.545 7.648 9.13 16.995 7.365 85.716 25.085

Trimean 0.695 0.356 66.497 2.637 0.061 0.021 83.5 8.498 1.13 0.046 0.011 0.03 0.006 1.464 Biw eight 0.65 0.345 66.637 2.36 0.062 0.018 83.2 7.604 1.113 0.045 0.009 0.029 0.006 1.345

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17.4 Resource Estimation Methodology

Block models were constructed to cover the entire lateral and vertical extent of the Mary River Deposit Nos. 1, 2 and 3 potential pit limits. Block size for Deposit No 1 block model was 15x15x15m. Block size for Deposit Nos. 2 and 3 block model was 25x25x10m high.

The model block grade attributes were comprised of Fe%, SiO2%, Al2O3%, FeO%, Mag%, MgO%, S%, P%, Mn%, Na2O%, K2O%, and Lump%. Additional block attributes included resource classification, apparent density, distance to nearest sample, rock code and average distance to nearest sample. Resource classifications in the block model are coded as “1” for measured resources, “2” for indicated resources and “3” as inferred resources.

For the ore zones an iron grade based regression formula where apparent density = (0.0288* Fe%) + 2.65 was estimated for each block. Apparent Density is a calculated value accounting for porosity and was derived from SGA testwork. This formula was based on a Fe vs. Apparent Density regression derived from 387 metallurgical samples.

The apparent density methodology adopted by SGA was based on testing samples comprised of 60 g of 10-12.5mm crushed material via water immersion.

As a result of external audit questions related to the impact of sample type and size in the SGA methodology, a separate density testwork program was initiated in 2010 to confirm the SGA derived ore density regression and expand the waste density database.

In 2010, a further 492 half core ore intervals were submitted to SGS laboratories for a waxed coated water immersion density method. The regression curve formula of density versus iron grade was essentially identical to the SGA derived formula and thereby confirmed the appropriateness of the density regression formula used in the resource estimate.

Waste rock was assigned an assumed density of 2.65 g/cc. Overburden was assigned a density of 2.24 g/cc.

The waste density dataset was also expanded in 2010 to comprise of 72 samples which indicate a mean waste density of 2.82 g/cc. Although the results were received too late for the incorporation into this study, it is recommended that the 2.82 g/cc value be used for waste densities in future studies.

Knight Piésold (KP) provided density data for three overburden samples in the vicinity of the proposed Deposit No. 1 open-pit location. The average density was

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estimated at 2.24 g/cc. KP cautions that given the characterization of these samples, the results should be considered approximate.

17.4.1 Variography

The composite dataset was extracted to reflect separate areas of unique dip and strike. Variograms were then run to assess grade variance along each ellipse axis and to assess interpolation ranges (see Error! Reference source not found.).

Variograms were run for the Fe, SiO2, Al2O3, P, Mn and MgO.

In Deposit Nos. 2 and 3, drill hole spacing and limited drilling precluded the generation of reasonably informed variograms.

Table 17-16: Major Axis Variogram Results Deposit No. 1 – North Limb

Variography Deposit No. 1 North Limb Orientation Convention ZXY LRL Spherical

Variable C0 S1 S2 Range No. 1 Range No. 2 Bearing Dip Fe Mj 5 25.2 7.85 140 300 216 -36 Fe Sm 5 18.1 13.6 100 300 184 -49 Fe Mn 5 21.8 11.7 20 100 293 16

SiO2 Mj 0 7.5 10 100 200 200 -53

SiO2 SM 0 7.4 9.8 100 200 44 -34

SiO2 Mn 0 8.3 4.9 25 100 126 12

Al2O3 Mj 0 3 2.75 30 200 200 -52

Al2O3 SM 0 4.69 100 44 -36

Al2O3 Mn 0 5.57 40 126 6 P Mj 0.0008 0.0022 150 200 -53 P SM 0.0008 0.0023 100 44 -35 P Mn 0.0008 0.0019 70 126 12 Mn Mj 0.002 0.025 275 200 -53 Mn SM 0.002 0.026 200 44 -34 Mn Mn 0.002 0.032 100 126 12 MgO Mj 0.5 1.12 1.3 100 300 216 -36 MgO SM 0.5 0.66 1.7 130 160 52 -52 MgO Mn 0.5 2.3 70 131 8

17.4.2 Estimation Methodology

The Deposit No. 1 north limb interpolation was constrained to the area north of Y=7914100N and within the rock code 100 solid. The south limb grade interpolation was constrained to the area south of Y= 7914100N and within the rock code 100 solid. The north zone interpolation was further separated by the 460m elevation each with a separate search ellipse dip. The northern extent of the north limb was

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constrained by cross-cutting amphibolite intrusives exposed in outcrop. These intrusives also limited the interpolation of the northern and southern extent of the North Limb Extension solid (rock code 700). Interpolation of each rock code was restricted to using only composites contained within each corresponding domain.

For Deposit No 1, due to the relatively consistent drill hole spacing and lack of clustering an elliptical search was adopted. Surpac convention for ellipse rotation of ZXY LRL was used. A nested spherical model was used for interpolation. The number of descretisation points in the x, y and z were 3, 3 and 3. All boundaries were treated as hard during the interpolation. The maximum number of samples per drill hole for zones 100 through 600 was 3. The remaining solids did not have a limit of samples per drill hole.

Search ellipse parameters used for grade interpolation in Deposit No. 1 are included in Table 17-17. The maximum search distances and minimum and maximum composites used for interpolation were derived from directional variograms, zone statistic characteristics, deposit geometry and observations of mineralization characteristics in drill core.

Table 17-17: Deposit No. 1 Search Ellipse Parameters

Min/Max No. Mj/Sm Mj/Mn Zone Az Dip Range Comps Ratio Ratio

Lower South Limb_RC_100 308 -64 400 3/8 1 3 Lower North Limb_RC_100A 222 70 400 3/8 1 3 Lower North Limb_RC_100B 222 82 400 3/8 1 3 Middle _South_RC_200 0 0 300 3/8 1 1 Upper 1_South _RC_300 156 70 300 3/7 1 3 Upper 2 _South_RC_400 156 70 200 3/7 1 1 North RC_600 56 -80 400 3/7 1 4 Lower North Limb Ext 24 -90 400 3/8 1 4 RC_700 N1 Int Waste_1000 213 82 100 3/5 1 1 N2 Int Waste_1100 195 73 200 3/5 1 1 S Int Waste_1200 24 -90 200 3/5 1 1 North Delet_ RC_1300 195 71 250 3/5 1 1 South Delet_RC_1400 308 -64 250 3/5 1 1

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Search ellipse parameters used for grade interpolation in Deposit Nos. 2 and 3 are included in Table 17-18. Due to the limited population of pairs of points at various distances apart available for variography, the maximum search distances used for Deposit Nos. 2 and 3 were derived from variograms generated from Deposit No. 1 data. For Deposit No. 3, interpolation runs were divided into two areas of unique azimuth and dip of the iron formation. The 567900 E line was used to divide the two populations. Table 17-18 indicates that the dip of Deposit No. 3 at its western extent is sub vertical while the eastern extent develops a significantly flatter dip.

Table 17-18: Deposit No. 2 & 3 Search Ellipse Parameters

Zone Az Dip Range Mj/Sm Ratio Mj/Mn Ratio Dep 2 Lower 100 Zone 75 -81 450 1 4 Dep 2 Upper 200Zone 75 -81 450 1 4 Dep 3 West End 82 -90 450 1 3 Dep 3 East End 61 -70 450 1 3

The search methodology for the Deposit Nos. 2 and 3 block model used a minimum of three composite samples and a maximum of fifteen composite samples.

17.5 Mineral Resource Classification

Mineral Resources were classified by G Wahl (P. Geo) who is a QP as defined by National Instrument 43-101 (NI 43-101). Mineral resources were classified in accordance with NI 43-101 requirements and CIM definitions. Mineral Resources for both block models were classified according to the following criteria.

Measured Mineral Resources: Block informed by a minimum of three samples, within an interpreted mineralized solid, and within 45 m of the nearest informing sample.

Indicated Mineral Resources: Block informed by a minimum of three samples, within an interpreted mineralized solid and greater than 45 m and less than 120 m of the nearest informing sample.

Inferred Mineral Resources: Block informed by a minimum of three samples, within an interpreted mineralized solid and greater than 120 m and less than 450 m of the nearest informing sample.

Although the definition of measured is restrictive and results in “bull’s eyes” appearance of measured blocks, the distance chosen is more reflective of an elevated degree of uncertainty around the distribution of deleterious grades. Although confidence in Fe grades are high at the 75 m drill hole spacing,

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observations in drill core and outcrop suggest the ranges of deleterious grades may be smaller than current drill hole spacing. Thus the approach of assigning measured was retained in order to reflect this relative uncertainty.

17.6 Validation of Block Model

The local block grade estimates were validated by comparing block grades to drill- hole grades on a section-by-section and elevation-by-elevation basis.

The model was further validated by comparing the ordinary kriged grades versus the inverse distance squared estimate. Results indicate that grades between the two estimation methodologies on a global basis at 50% Fe cut off for all zones correlate well (see Table 17-19).

Table 17-19: Comparison of Inverse Distance Squared and Ordinary Kriging Interpolated Grades for Deposit No. 1

Grade Attribute Inv Dist Squared Ordinary Kriging Fe % 65.41 65.47 SiO2 % 2.89 2.90 Al2O3 % 1.17 1.15 FeO % 17.38 17.36 Mag % 56.0 56.0 MgO % 1.13 1.13 Mn % 0.25 0.25 Na2O % 0.007 0.007 K2O % 0.019 0.018 P % 0.055 0.054 S % 0.303 0.305 Lump % 84 84

Bivariate statistics comparing estimates interpolated with ordinary kriging versus inverse distance squared for all zones within Deposit No. 1 indicate reasonable correlation for each attribute between estimation methods.

Total assays in the interpolation were also assessed to see how closely they totalled 100%. The measured blocks totalled 100.15% while the indicated blocks totalled 100.11% and the inferred blocks totalled 100.53%. All blocks totalled 100.27%. Total blocks classified as either measured or indicated above 102.5% amounted to 20 Mt. Total interpolated assay control in the model indicated quite reasonable results.

Swath plots were generated along the strike extent of the lower 100 Zone reflects blocks and composites from 7914200N to 7914600N. See following Figure 17-1 to

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Figure 17-7. The results indicate good correlation of blocks and composites along the strike extent.

Figure 17-1: Swath Plot North 100 Zone Fe%

Figure 17-2: Swath Plot North 100 FEO%

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Figure 17-3: Swath Plot North 100 Zone Al2O3%

Figure 17-4: Swath Plot North 100 MgO%

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Figure 17-5: Swath Plot North 100 SiO2%

Figure 17-6: Swath Plot North 100 Mn%

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Figure 17-7: Swath Plot North 100 P%

Descriptions of open pit shell assumptions as well as mineral resource tables are contained in the following Pit Optimization Section.

17.6.1 Assessment of Reasonable Prospects of Economic Extraction

Open Pit Optimization

Whittle software was used to generate ultimate pit shells used to define the Mineral Resources for Deposit Nos. 1, 2, and 3. These pit shells, based on Measured, Indicated and Inferred resource categories, were used to support an assessment of reasonable prospects criteria.

The updated block model of May-June 2010 for Deposit No. 1 was used for the pit optimization of this deposit. The January 2008 block model was used for the optimization process of Deposit Nos. 2 and 3.

Optimisation Parameters

Parameters used to constrain the Mineral Resources to assess reasonable prospects of economic extraction were based on the rail option evaluated in the Aker Feasibility Study (Aker, 2008). Although the trucking study supersedes that

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estimate, the derivation of the rail cost parameters was considered more appropriate to support the open pit optimization as it is assumed the bulk of the Mineral Resources will likely be transported via rail rather than trucks to the coast.

The physical and economic parameters used for the pit optimization process for Deposit No. 1, and Deposit Nos. 2 and 3 are presented in Table 17-20.

Compared to the previous study the following modified parameters were used for the pit optimization for Deposit No. 1:

Slope Angle: Pit slope angles were adjusted by one to two degrees in order to allow for the ramping system, and a potential flatter slope requirement for the high wall. Mining Cost: An average mine operating cost was used instead of a base cost and incremental cost per bench. The selected average cost was higher than the estimated average mine operating cost in the 2008 Aker study. The selected average cost, allowed for longer haul distances of the waste rock to the larger waste dumps than the 2008 study, due to more tonnage associated with potentially larger pit shells, and due to potential larger mining limits in the horizontal and vertical dimensions, resulting in longer haul distances to the crusher. Metal Price: The long-term iron ore product price was provided by BIM. The following modified parameters were used for Deposit Nos. 2 and 3:

Mining Cost: A higher average mining cost was used compared to the Blue Sky pit optimization in 2008. This cost was also adjusted to allow for longer haul distances for the waste and ore rock types. Crushing/Conveying: These costs were adjusted accordingly for the 20Mtpa production rate and were provided by BIM Metal Price: The long-term iron ore product price was provided by BIM.

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Table 17-20: Optimisation Parameters

Slope Angles Overall Overburden Deposit 1 MI MII West Wall 38º 37º 27º North, South East 41º 40º 27º Sectors 4 and 8 37º 37º 27º Deposits 2&3 40º -

.Cost Item Deposit 1 Deposits 2&3 Average Mining Costs CAD 2.0/t CAD 2.2/t Mining Recovery / Mining Dilution 97.5% / - 1.5% 97.5% / 2.5%

CAD/tore Icebreakers and Tugs OPEX 0.50 Railway OPEX 1.45 General and Administration 4.19 3.77*1 Crushing, Screening and Ship loading 4.89 4.40*1

Conveying to Train Loading Station 2 - 0.83*

Rehabilitation Cost CAD 0.50/tore - Product Price (Average Lump/Fine) US ¢115.25/dmtu Process/Transport Recovery 100% Discount Factor 7%

*1 –Projected based on the assumed higher production rate compared to Deposit No. 1 *2 –Approximate adjustment for the Base Year

17.6.2 Mineral Resource Cut Off Grades

The optimization and final/ultimate pit shell selection for Deposit No. 1 was carried out by using a 59% Fe cut-off grade (CoG) for orebody 100 and 60% Fe for the remaining orebodies. This was to have consistency in generating the nested pit shell with the previous studies. However, in order to produce a direct shipping product with an average in-situ grade of +66.0% Fe, a mining CoG of 59% Fe was selected to estimate the mineral inventory inside the selected pit shells for Deposit No. 1.

The diluted iron grade product in Deposit Nos. 2 and 3 was selected at +64% Fe and the CoG was 60.5% Fe. This was achieved by applying 2.5% dilution at zero percent Fe in the optimization. As a result, the average insitu grade of the total mineral inside the selected ultimate pit shell for Deposits 2 and 3 is 65.8% Fe.

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17.6.3 Mineral Resource Statement

Mineral Resources for Deposit Nos. 1, 2, and 3 are shown in Error! Reference source not found. and Table 17-22 detail the mineral resource categories and deleterious grades. The Mineral Resource tables are inclusive of Mineral Reserves. Mineral resources that are not mineral reserves do not have demonstrated economic viability. The effective date of the estimate is 30 December, 2010. Mineral Resources were estimated in accordance with NI 43-101 and CIM resource classification definitions.

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Table 17-21: Deposit 1, 2 & 3 Mineral Resources

Deposit No Resource Category Million Tonnes Al2O3 % Fe % FeO % K2O % Mag % MgO % Mn % Na2O % P % S % SiO2 % 1 Measured 207 1.2 66.3 13.8 0.02 44.39 0.93 0.16 0.01 0.05 0.19 2.71 1 Indicated 211 1 66.3 15.7 0.02 50.54 0.98 0.21 0.01 0.05 0.26 2.54 Measured & 1 Indicated 418 1.1 66.3 14.8 0.02 47.49 0.96 0.19 0.01 0.05 0.23 2.62 1 Inferred 213 0.9 66.9 22.9 0.02 73.86 1.03 0.4 0.01 0.05 0.43 1.87 2 & 3 Indicated 26 1 65 2 0.07 6.3 0.60 0.0 0.01 0.02 0.02 5.2 2 & 3 Inferred 336 1.1 65.9 3.7 0.05 11.8 1.10 0.10 0.03 0.03 0.01 2.4 Note 1 : Deposit No 1 Mineral Resources are Above 59% Fe Cut off and within Pit Shell Note 2 : Deposit 2 & 3 Mineral Resources Above 60.5% Fe Cut off and within Pit Shell

Table 17-22: Combined Deposit 1, 2 & 3 Mineral Resources

Deposit No Resource Category Million Tonnes Al2O3 % Fe % FeO % K2O % Mag % MgO % Mn % Na2O % P % S % SiO2 % 1, 2 & 3 Measured 207 1.2 66.3 13.8 0.02 44.39 0.93 0.16 0.01 0.05 0.19 2.71 1, 2 & 3 Indicated 237 1.0 66.2 14.2 0.03 45.7 0.9 0.2 0.01 0.05 0.23 2.83 1, 2 & Measured & 3 Indicated 444 1.1 66.2 14.1 0.02 45.1 0.9 0.2 0.01 0.05 0.22 2.8 1, 2 & 3 Inferred 549 1.0 66.3 11.1 0.04 35.9 1.1 0.2 0.02 0.04 0.17 2.2 Note 1 : Deposit No 1 Mineral Resources are Above 59% Fe Cut off and within Pit Shell Note 2 : Deposit 2 & 3 Mineral Resources Above 60.5% Fe Cut off and within Pit Shell

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17.7 Mining and Trucking Based Mineral Reserves

The Mine Planning and Cost Estimation Study was carried out using the Mary River Iron Ore Deposit No. 1 orebody block model updated in 2010. In order to report the Mineral Resources beyond the designed practical pit, optimization was carried out using the inferred category as well resulting in an ultimate pit shell.

Whittle and Gems software was used for the pit optimization and pit planning purposes. Measured and indicated Mineral Resources were used for the pit optimization, practical pit design, and production scheduling.

Pit optimization parameters were selected based on the Conceptual Trucking Study (May-June 2010). The pit slope angles are based on the practical pit design in the Aker Feasibility Study (2008). The long term product price and the US$/CND exchange rate were provided by BIM (Table 17-23).

The optimization results were used in order to select an optimum pit outline for the trucking study containing a minimum of 60Mt of mineral in the measured and indicated categories. In order to design the open pit, and for practicality reasons a larger shell, containing 120Mt of mineral was selected and a 60Mt pit was designed within the larger shell accordingly.

This strategy allows for future expansion of the pit without waste backlog. Also, by designing this pit, the narrow and the scattered mining benches were eliminated. The engineered interim and final pits were used in reporting the rock inventory and for preparing the production schedule for the 20 years project life.

A mining CoG of 59% Fe and 60% Fe was used in the optimization process for the Upper Orebody, and the Middle and Lower orebodies respectively, to simulate a direct shipping product at an average >65 % Fe.

The reserving is carried out using an insitu Cut-off Grade (CoG) of 57% Fe for the Lower Orebody and a 50% CoG for the Middle and Upper orebodies. This allows for the continuity of the orebodies and is considered as part of the dilution during the mining operation.

For scheduling purpose additional mining dilution and mining recovery factors were assigned to each orebody according to the orebody geometry (Table 17-24). The lower and middle orebodies are solid and consist of some areas of internal waste. A one to two percent mining dilution at zero grade was applied to the in situ iron grade. The Upper orebody is smaller than the other orebodies with more internal waste, sharp transition between ore and waste rock, and irregular orebody geometry. Higher mining dilution and loss of 5% (at zero iron grades) were applied to calculate the diluted

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mineral grade in this area. Dilution was not considered for the in situ grade of the deleterious elements.

During the 20 year project life, the scheduled average iron grade in the Run-of Mine (RoM) is >66.0% Fe.

Geotechnical parameters were used to design pit slopes; average slopes are shown in Table 17-25.

The surveyed topography surface (LiDAR Surface) was used for Mineral Reserve estimation. This surface was adjusted to reflect the excavated rock during the bulk sample operation.

Table 17-23: Optimization Parameters

.Cost Item Value Average Mining Cost CAD 5.85/t Mining Recovery / Mining Dilution 97.5% / - 1.5%

CAD/tore Ship Loading 6.50 Road Haulage OPEX 7.20 General and Administration 0.93 Maintenance 0.64 Crushing/Screening 3.01

Rehabilitation Cost CAD 0.2/tore Product Price (75% Lump/ 25% Fine) US ¢115.25/dmtu Lump US ¢ 120/dmtu Fine US ¢ 101/dmtu Process/Transport Recovery 99.95% Exchange Rate 93% Discount Factor 7%

Table 17-24: Mining Recovery and Mining Dilution

Orebody Recovery Dilution Lower 99% 1% -% Fe Middle 98% 2% -% Fe Upper 95% 5% -% Fe

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Table 17-25: Summary of the Pit Slope Angle used in the Optimization

Slope Angle Overall Overburden Deposit 1 MI MII West Wall 38º 37º 27º North, South East 41º 40º 27º Sectors 4 and 8 37º 37º 27º

17.7.1 Mineral Reserve Statement

The Mineral Reserves for the Mary River Project Deposit No 1are tabulated in Table 17-26. Mineral Reserves have an effective date of 30 December 2010.

Table 17-26: Mineral Reserves within the Designed Pit for Trucking Study

Category Orebody Mt Fe% P% S% Mn% SiO2% Al2O3% FeO% Proven Lower 29.8 66.5 0.017 0.066 0.12 1.59 1.27 8.84 Middle 3.8 62.8 0.072 0.094 0.12 4.21 2.11 12.71 Upper 0.1 58.8 0.040 0.010 0.06 4.45 2.54 3.77 Total 33.7 66.1 0.023 0.069 0.12 1.89 1.37 9.26 Probable Lower 25.5 66.5 0.025 0.063 0.14 1.56 1.12 7.96 Middle 1.5 63.8 0.068 0.018 0.09 3.39 1.34 4.90 Upper 0.1 57.6 0.034 0.127 0.18 8.20 0.90 9.84 Total 27.1 66.3 0.027 0.061 0.14 1.69 1.13 7.80

Total Proven and Probable 60.7 66.2 0.025 0.065 0.13 1.79 1.26 8.60

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C ONTENTS

18.0 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES ...... 18-1 18.1 Proposed Mining Operation ...... 18-1 18.1.1 Pit Optimization ...... 18-1 18.1.2 Pit Design ...... 18-2 18.1.3 Proposed Crushing and Screening Plant ...... 18-4 18.1.4 Milne Inlet – Planned Transport, Stockpiling, Reclaiming and Ship Loading ... 18-4 18.1.5 Proposed Production Schedule ...... 18-5 18.2 Waste Rock Management ...... 18-9 18.2.1 Shift Roster and Required personnel ...... 18-9 18.3 Equipment Selection ...... 18-10 18.4 Infrastructure ...... 18-12 18.5 Environment ...... 18-12 18.6 Capital Cost Estimate ...... 18-12 18.7 Operating Cost Estimate ...... 18-13 18.8 Markets ...... 18-14 18.9 Taxation ...... 18-18 18.10 Financial Analysis ...... 18-19 18.11 Basis of Financial Analysis ...... 18-20 18.11.1 Results of Base Case Financial Analysis ...... 18-21 18.11.2 Base Case Sensitivity Analysis ...... 18-23 18.11.3 Alternate Case Financial Analysis ...... 18-24 18.12 Risk and Opportunity Analysis ...... 18-26

T ABLES

Table 18-1: Summary of the Pit Optimization Result ...... 18-1 Table 18-2: Summary of the RoM Production Schedule ...... 18-7 Table 18-3: Pit Rock Movement Schedule during the Life of Mine ...... 18-8 Table 18-4: Mining and Ancillary Fleet Requirement ...... 18-11 Table 18-5: Cost Summary ...... 18-13 Table 18-6: Operating Cost Summary ...... 18-14 Table 18-7: Iron Ore Consumption by Country ...... 18-16 Table 18-8: Capital Cost Estimate Summary ...... 18-19 Table 18-9: Operating Cost Estimate Summary ...... 18-20 Table 18-10: Financial Summary...... 18-20 Table 18-11: Revenue ...... 18-21 Table 18-12: Base Case Financial Analysis Summary ...... 18-21 Table 18-13: Financial Model ...... 18-22 Table 18-14: Summary of the Pre-Tax Annual Cash Flow ...... 18-23 Table 18-15: Owner Operated ($ 000) ...... 18-24 Table 18-16: Alternate Case, Summary of Contractor Operator Capital Costs ...... 18-25 Table 18-17: Alternate Case, Summary of Contractor Operator Operating Costs ...... 18-25 Table 18-18: Alternate Case Financial Analysis Summary ...... 18-26

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F IGURES

Figure 18-1: General Layout of the Designed Final Pit, Waste Dump. Haul Road and Crusher Location ...... 18-3 Figure 18-2: Scheduled Fe, P, and S Grade per Period ...... 18-8 Figure 18-3: World Production of Iron Ore and Crude Steel ...... 18-14 Figure 18-4: Iron Ore Price Forecast ...... 18-18

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18.0 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES

18.1 Proposed Mining Operation

18.1.1 Pit Optimization

Whittle software was used for open pit optimization to define the economic mining boundaries of the measured and indicated categories of Deposit No. 1 orebody using the parameters in Table 17-23.

The final pit shell for this project was selected based on a different approach than the usual pit optimization process due to utilization of just 60Mt of the potentially economical larger tonnages presented in the optimization summary. A 60Mt pit within the boundary of the selected larger shell (Pit Shell 8), containing 120Mt of above CoG mineralized rock, was designed. The designed pit allows for practicality and flexibility in the mining operation for in-pit blending, and providing sufficient exposed ore at any mining period. This also allows for the possibility of the pit expansion to 18Mtpa at any period during the 20 years project life.

A summary of the optimization result is presented in Table 18-1.

Table 18-1: Summary of the Pit Optimization Result

Cashflow Pit Revenue Undisc. Ave. Disc. RoM Waste Total Strip Shell Factor $M $M Mt Fe% Mt Mt Ratio 1 0.310 2,954 1,694 51 66.7 5 56 0.1 2 0.315 3,342 1,801 58 66.6 6 64 0.1 3 0.320 4,211 1,988 74 66.5 11 84 0.2 8 0.345 6,942 2,265 124 66.1 35 159 0.3 10 0.400 11,387 2,324 212 65.6 116 328 0.6 15 0.600 17,594 2,205 358 65.4 451 809 1.3 20 0.800 18,335 2,169 387 65.4 605 992 1.6 22 0.880 18,402 2,157 392 65.4 640 1,032 1.6 25 1.000 18,416 2,152 395 65.4 662 1,057 1.7

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18.1.2 Pit Design

The three interim and the final pits were designed using the following design parameters:

Bench height: 15 m Bench face angle: 65˚ Minimum berm width: 8 m Inter-ramp slope angle: 40˚ to 45˚ Overburden slope angle: 27˚ Ramp width: 33 m In-pit ramp gradient: 10% Minimum working width: 50m

The selected bench height of 15 m is the final height of each bench. The mining operation could be carried out in 7.5 m stacks.

Specific sumps and water collection areas were not included in the pit design. However, these were considered in the planning by allowing wide benches in each pushback for in-pit water management.

The designed Final Pit for the trucking Study and the general layout of the waste dump, main haul road to the crusher and the crusher location is presented in Figure 18-1.

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Figure 18-1: General Layout of the Designed Final Pit, Waste Dump. Haul Road and Crusher Location

N 600 574 PAG Non-PAG 660 OB

590 8

620 605 503 8

590

575

545

515

348

213

Crush er 283

The in-pit ramping system is designed to provide access for ore and waste rock haulage to the crusher, waste dumps, and the stockpile.

The 33m wide ramps are designed at a 10% gradient. The ramp width is wider than the required two-way road width for the selected 55t trucks. However, wider design allows for the pit expansion to the higher production rate of 18Mtpa and utilization of the larger trucks (i.e. 225t capacity) with no delays due to ramp issues.

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The outside access road on the East side of the pit will be used as the main access to the crusher. The road length from the mine start-up bench to the crusher is approximately 6.3km with a total width of 24m and a downhill gradient of a maximum of 10% in the hill side and 5% from the crusher to the toe of the hill. The outside road width is the standard width for the selected 55t truck. It is assumed the outside road could be widened to a final width of 33m when pit production rate is increased.

18.1.3 Proposed Crushing and Screening Plant

The crushing and screening plant is designed to operate 300 d/a, with an operating availability of 75%.

The primary objective of the plant is to maximize the production of lump product (-25/+7 mm), while at the same time, keeping fines product (-7 mm) at a minimum. The mine plant will be equipped with a 47‖ x 63‖ jaw crusher, one HP 800 cone crusher, one 1.8 m x 6 m double deck primary screen and one 3.66 m x 8.5 m double deck secondary screen. After crushing and screening, the product will be conveyed to separate lump and fine storage piles through a system of belt conveyors and radial stackers, where it will be reclaimed and loaded on to haul trucks.

The plant design throughput is planned at 600 t/h with the product streams consisting of 75% lump product and 25% fine product.

18.1.4 Milne Inlet – Planned Transport, Stockpiling, Reclaiming and Ship Loading

The crushed ore will be transported by haul trucks 100 km from Mary River to Milne Inlet, taking four to five hours for a round trip, depending on the time of the year. The haul trucks are planned to be twin trailer, side dump trucks with a payload of 134 t. The trucks and trailers will be custom designed for the arctic conditions and 21 haul trucks are required to be in operation to deliver the required 3 Mt/a in the 300 d/a allotted. Once the material is unloaded on the ground it will be transferred using front- end loaders, mobile feeders and mobile radial stackers to stockpiles. The stockpiles will be located on a plateau approximately 1 km southeast of the ore loading dock. Due to the short and variable shipping season (70 – 100 d/a) the total annual production (3.0 Mt) will be stockpiled at the port.

Front-end loaders will be used to reclaim the iron ore from the stockpiles. Two 3,000 t/h reclaim belts will run in-between and parallel to the two rows of longitudinal stock piles. Four movable reclaim hopper feeders, two on each reclaim belt, will be mounted

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on rails and run above and in line with the reclaim belts. The reclaim hopper feeders each will have a capacity of 1,500 t/h. The reclaimed product from the stockpiles will be conveyed by two 3,000 t/h belt conveyors to the floating ore loading dock located at the south western end of Milne Inlet.

Two 3,000 t/h truss-supported belt conveyors will be mounted on the causeway and the barge of the ore loading dock, perpendicular to the shore. The conveyors each will receive iron ore from either of the dual conveyors discharging at the land end of this barge. The conveyors will elevate the iron ore to a height sufficient to feed two elevated hoppers with support towers mounted on each side of the ship loader support barge. The elevated hoppers will feed the two conveyors going to the ship loaders. Two 3,000 t/h belt conveyors will each receive iron ore from their respective hoppers and deliver it to the two ship loaders. The ship loaders will rotate 180° and shuttle in and out giving the full loading coverage, without having to reposition the ship.

18.1.5 Proposed Production Schedule

The mine production schedule was prepared for a 3Mtpa of RoM for direct shipping product with an average grade of >66.0% iron. Negligible loss was assumed for crushing/screening, stockpiling/reclaiming, and truck transport.

The following mining periods were selected for the schedule:

Pre-production: six months prior to the mine production Quarterly: Year 1 Bi-annually: Year 2 and Year 3 Annually: Year 4 to Year 10 Five-year periods: Year 11 to Year 20

The pre-production period is projected to start in the first half of 2013 reaching the full 3Mtpa production rate in the second half of 2013.

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used for the Lower Orebody and a 50% CoG for the Middle and Upper orebodies due to limited tonnage (219 kt) of the mineralized rock in the 50%-57% Fe with minimal effect on the overall RoM grade.

The objective of the mine scheduling was to maximize the iron content in the RoM project and to avoid sharp fluctuation of the iron and deleterious elements grades during the defined 20 years. Product specification for the deleterious elements was not considered as a constraint in the schedule.

An average lump – fine ratio of 75% - 25% was suggested by BIM for the final product.

Each pushback was designed to provide RoM for a four to six year period if mined individually. However, in order to expose sufficient ore for in-pit blending, to eliminate waste backlog, and to maintain access between push backs, mining cuts were scheduled to be excavated concurrently.

Table 18-2 presents a summary of the RoM production schedule and the iron, phosphorous, and sulphur grades per period are presented in Figure 18-2. The scheduled pit rock movement is presented in Table 18-3.

It could be observed that the iron content gradually decreases from 66.5% Fe to 66.0% Fe in the first 15 years. Beyond Year 15, the average iron grade increases to a consistent 66.6% Fe. The average sulphur grade in the first 5 years increases to 0.97% and then decreases beyond Year 6, except that in Year 9 it reaches to 0.90%.

The average phosphorous grade is 0.020% during the first eight years and then it increases to about 0.022% in years nine and ten. The average phosphorous grade in years 11 to 15, and 16 to 20 is 0.025% and 0.034% respectively.

The stripping ratio in the preproduction period is 1.2 and then it is decreased to about 0.4 in the first five years. In Years 6 to 8 the stripping ratio is 0.2 and then it is increased to 0.3 until Year 15. In the final five year period it is decreased to 0.2. The total average stripping ratio is 0.3.

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Table 18-2: Summary of the RoM Production Schedule

Year Period kt Fe% P% S% Mn% SiO2% Al2O3% FeO% 2013 – Q1&2 Pre-Prod 600 66.1 0.015 0.014 0.13 1.85 1.44 8.32 2013-Q3 Y1-1 900 66.4 0.017 0.019 0.14 1.83 1.31 9.54 2013-Q4 Y1-2 900 66.5 0.018 0.029 0.13 1.88 1.30 10.04 2014-Q1 Y1-3 600 66.5 0.018 0.031 0.13 1.90 1.29 10.15 2014-Q2 Y1-4 600 66.5 0.018 0.031 0.13 1.90 1.29 10.15 Subtotal Y1 3,000 66.5 0.018 0.027 0.13 1.87 1.30 9.93 2014-2 Y2-1 1,800 66.4 0.018 0.056 0.12 2.03 1.35 12.19 2015-1 Y2-2 1,200 66.3 0.018 0.065 0.12 2.08 1.36 12.86 Subtotal Y2 3,000 66.3 0.018 0.060 0.12 2.05 1.35 12.46 2015-2 Y3-1 1,800 66.1 0.021 0.092 0.13 2.28 1.49 13.91 2016-1 Y3-2 1,200 66.2 0.017 0.060 0.11 1.52 1.22 6.58 Subtotal Y3 3,000 66.1 0.019 0.079 0.12 1.98 1.38 10.98 2016-17 Y4 3,000 66.1 0.020 0.090 0.13 1.88 1.35 10.03 2017-18 Y5 3,000 66.0 0.022 0.097 0.13 1.86 1.37 9.72 2018-19 Y6 3,000 66.0 0.020 0.079 0.12 1.92 1.42 10.01 2019-20 Y7 3,000 66.0 0.019 0.078 0.12 1.94 1.42 10.15 2020-21 Y8 3,000 66.0 0.019 0.075 0.11 1.94 1.42 10.17 2021-22 Y9 3,000 66.0 0.023 0.090 0.12 1.93 1.39 10.03 2022-23 Y10 3,000 66.0 0.022 0.081 0.12 1.90 1.36 9.62 2023-28 Y11-15 15,000 66.0 0.025 0.077 0.13 1.91 1.28 9.20 +2028 Y16-21 15,082 66.6 0.034 0.036 0.14 1.41 1.01 4.60 Total 60,682 66.2 0.025 0.065 0.13 1.79 1.26 8.60

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Figure 18-2: Scheduled Fe, P, and S Grade per Period

Fe (%) P & S (%) 66.7 0.11

66.6 0.10

66.5 0.09

66.4 0.08

66.3 0.07

66.2 0.06

66.1 0.05

66.0 0.04

65.9 0.03

65.8 0.02

65.7 0.01 PP 1 2 3 4 5 6 7 8 9 10 11-15 16-20 Production Period

Fe% P% S%

Table 18-3: Pit Rock Movement Schedule during the Life of Mine

Non-PAG Waste PAG & LG Stockpile Strip Year Period Ore Total Overburden Waste Total LG PAG Total Ratio 2013-Q1&2 PP 600 700 700 1.2 1,300 2013-14 1 3,000 1,021 1,021 24 255 279 0.4 4,300 2014-15 2 3,000 887 887 92 222 313 0.4 4,200 2015-16 3 3,000 977 977 79 244 323 0.4 4,300 2016-17 4 3,000 826 826 67 207 274 0.4 4,100 2017-18 5 3,000 812 812 86 203 288 0.4 4,100 2018-19 6 3,000 1 432 434 58 108 166 0.2 3,600 2019-20 7 3,000 1 436 437 54 109 163 0.2 3,600 2020-21 8 3,000 1 445 447 42 111 153 0.2 3,600 2021-22 9 3,000 35 578 614 142 145 287 0.3 3,900 2022-23 10 3,000 37 655 692 44 164 208 0.3 3,900 2023-28 11-15 15,000 187 2,949 3,136 26 737 763 0.3 18,900 +2028 16-20 15,082 279 2,296 2,575 - 574 574 0.2 18,231 Total 60,682 543 12,875 13,418 713 3,219 3,931 0.3 78,031 *Rock Tonnage is reported in kt

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18.2 Waste Rock Management

Waste and mineralized rock were categorized into potentially acid-generating (PAG) rock and clean waste. The PAG dump requires proper drainage in order to collect the potential run-off. Detailed studies were not carried out to quantify the PAG rock within the mining limit therefore a preliminary estimate of 20% of the total waste was considered as PAG. Mineralized rock below the CoG was included in this category and a separate pile was not assumed for the low-grade mineralized rock. The remaining waste rock, including the overburden was considered as non-PAG waste.

The PAG, overburden, and clean waste rock are dumped in the same area but in separate piles. Based on Knight Piésold’s June 28, 2010 recommendations, the bottom part of the waste dump consist of a layer of clean waste with a minimum thickness of 10-20 m. Parameters used for waste dump design were:

Maximum overall effective slopes: 2H:1V Individual bench slopes: 1.5H:1V Maximum bench height: 30 m

Additional studies are recommended to determine the amount of the PAG rock, and the waste dump set-up for the different materials allocated for each pile. The suggested method to develop the waste piles may be improved upon to allow for practicality of such dumps during the mine operation.

18.2.1 Shift Roster and Required personnel

A total of 161 personnel are required in the pre-production period and 177 personnel in Year 1 and beyond.

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18.3 Equipment Selection

Truck and shovel operation was selected for the load and haul operation at Mary River. The following is a list of the main mining and ancillary equipment:

Production drills equivalent to Atlas Copco DM45 Backhoe excavators with a bucket size of 4.75m3 Front-end-Loader with a bucket size of 6.1m3 Dump truck with 55t nominal capacity Trackdozer equivalent to CAT D9 Wheeldozer equivalent to CAT 834 Grader equivalent to CAT 14 The fleet estimation for drilling, loading, and hauling requirement was carried out based on the life-of-mine production schedule and the estimated haul distances to waste dump and crusher. Ancillary equipment was selected for road maintenance, snow clearing, equipment maintenance, secondary mining activities, and other related earthworks. The proposed fleet requirements are summarized in Table 18-4.

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Table 18-4: Mining and Ancillary Fleet Requirement

Equipment Period PP 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 +16 Production Required 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 Drill Budget 2 Ancillary Required 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Drill Budget 1 Required 2 2 2 2 2 2 2 2 2 2 2 2 3 2 2 2 2 Excavator Budget 2 1 1 Required 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Loader Budget 1 Required 6 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 Truck Budget 6 2 6 2 Required 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 Trackdozer Budget 2 2 1 Required 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Wheeldozer Budget 1 Required 1 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 Grader Budget 1 1 Required 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Watercart Budget 1 Required 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Utility Backhoe Budget 1 Service Required 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Truck Budget 1 1 Fuel/Lube Required 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 Truck Budget 2 2 Low-bed Required 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Truck Budget 1 Tire Required 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 handler Budget 1 Required 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 Pickup Truck Budget 7 7

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18.4 Infrastructure

Project infrastructure is discussed in Section 5.5.

18.5 Environment

Environmental considerations for the Project are discussed in Section 4.8.

18.6 Capital Cost Estimate

The capital cost estimate complies with the Association for the Advancement of Cost Engineering (AACE) Class 3 estimate standards. A Class 3 estimate has an accuracy range of -15% to +15% before contingency, and is typical for mining feasibility studies. Costs were divided into direct costs, indirect costs, closure costs, and contingency.

Capital cost estimates for the Mary River Iron Ore Trucking option includes mining, crushing, screening, stockpiling, tote road, port and surface facilities for a 3 Mt/a throughput rate. The total estimate is CAD$740.3 million (Table 18.5). Capital costs are inclusive of the costs up to and including plant commissioning and start up. There is no allowance for escalation, interest or financing during construction. There is also no foreign exchange component in the capital costing. Costs are expressed in fourth quarter 2010 dollars.

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Table 18-5: Cost Summary Mary River Iron Ore - Trucking Option Project Cost Summary

Closure Costs 27,000 Sustaining Capital 146,057

18.7 Operating Cost Estimate

Operating costs were derived for the last quarter of 2010. Estimates were prepared by area and component, and consider the mining plan and processing schedule. The cost estimate has no allowance for escalation.

Operating costs over the LOM are summarized in Table 18.6.

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Table 18-6: Operating Cost Summary CAD/DMT Mining 8.21 Processing 2.80 Road haul 8.85 Stock Piling 1.45 Ship Loading 3.50 Camp 2.06 Air Transportation 1.27 G&A 0.79 Total 28.93 Ocean Shipping* Excluded

18.8 Markets

Analysis of the iron ore industry must include virtually the industry’s only customer - the world crude steel industry. World growth of iron ore and steel generally expands and contracts in line with world economic growth. World markets collapsed in starting in mid-2008 and continuing through 2009. A collapse of confidence in the credit markets were acerbated by the failure of a large investment bank and the bailout of many giant financial institutions.

Figure 18-3: World Production of Iron Ore and Crude Steel

World Production of Iron Ore and Crude Steel 1994-2013 (Mt)

1800 1600 1400 1200 1000

Mt 800 600 400 200 0 1994 1996 1998 2000 2002 2004 2006 2008 2010e 2012e

Steel Production Year Iron Ore Production Seaborne Iron Demand China Iron Ore Imports

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Despite the economic collapse, steel production rebounded in late 2009 and into 2010 with demand steadily increasing not only in China, but in the developed countries (USA, EU and Japan) as well. The focus of Baffinland Iron Mines Corporation’s marketing efforts has been Europe and the Middle East. The European market has been a general stagnant or low growth market with iron ore consumption for the original fifteen European Union (EU-15) countries consuming between 145 and 150 million tonnes of iron ore annually. Although, this level of consumption increased during the period, EU-15 is generally a flat market with growth potential limited to the EU-12 countries and ―other Europe‖, which is generally Turkey.

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Table 18-7: Iron Ore Consumption by Country Million Tonnes 2005 2006 2007 2008 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020

France 19.5 19.9 20.1 18.2 10.0 15.4 15.5 15.8 16.4 17.0 17.7 18.4 19.1 19.7 20.7 21.2 Germany 42.4 45.3 46.6 45.9 29.2 35.7 36.8 39.2 41.8 43.7 45.3 46.7 48.2 49.2 50.4 51.9 Italy 17.6 17.8 17.0 16.3 8.1 11.5 12.3 13.5 14.2 15.0 15.8 16.4 17.0 17.5 17.9 18.4 Netherlands 12.3 11.3 12.1 10.8 5.6 10.1 10.5 10.9 11.4 11.9 12.2 12.5 12.7 12.9 13.2 13.5 United Kingdom 16.1 16.4 17.4 15.3 9.2 10.3 12.5 12.9 13.4 13.9 14.4 14.8 15.2 15.5 15.9 16.2 EU-15 142.6 146.4 148.5 144.3 80.5 111.2 116.6 122.7 129.5 135.4 140.7 145.1 149.8 153.6 157.8 161.8

Czech Republic 6.8 7.6 5.3 6.8 1.8 5.6 5.9 6.3 6.9 7.4 7.7 8.0 8.3 8.7 8.9 9.3 Poland 6.8 8.6 8.7 8.7 3.5 6.6 7.5 7.9 8.6 9.3 9.9 10.5 11.0 11.5 11.9 12.2 EU-12 29.3 32.3 29.7 27.5 12.0 23.4 25.5 27.1 29.1 30.8 32.4 33.9 35.4 36.7 37.8 38.9

Turkey 8.4 9.4 11.0 11.5 10.8 12.7 13.3 13.9 14.5 15.2 15.6 16.1 16.6 17.0 17.5 18.0 Europe Other 12.9 14.3 13.9 15.3 13.7 16.9 18.0 18.9 19.7 20.6 21.3 21.9 22.5 23.2 23.9 24.6

Europe 184.8 193.0 192.1 187.1 106.2 151.5 160.1 168.6 178.3 186.8 194.4 201.0 207.6 213.5 219.6 225.3

China 695.7 914.2 1,065.6 1,268.0 1,508.4 1,664.6 1,834.7 1,990.0 2,130.6 2,291.3 2,406.3 2,516.9 2,622.8 2,747.1 2,880.7 2,984.3

Other 661.4 705.5 734.4 762.9 623.0 671.9 724.4 766.0 809.9 846.5 887.3 929.2 971.1 1,016.2 1,061.6 1,105.2

World 1,541.9 1,812.7 1,992.1 2,218.0 2,237.6 2,488.0 2,719.2 2,924.6 3,118.8 3,324.6 3,488.0 3,647.1 3,801.5 3,976.8 4,161.9 4,314.8 y-o-y change 13.8% 17.6% 9.9% 11.3% 0.9% 11.2% 9.3% 7.6% 6.6% 6.6% 4.9% 4.6% 4.2% 4.6% 4.7% 3.7%

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Importantly China’s iron ore consumption figures are skewed as it includes imports of generally higher quality iron ore at a grade of 62 to 65% iron while domestic production figures are given as raw tonnage of iron ore grading between a reported grade between 20 and 30% iron. Official production figures are released by the Chinese government but average grades of production and other information are considered state secrets.

Europe is a mature market that has most steel companies producing high margin products. The limited ability to expand operations, due to emission expansion, has forced steel companies to focus on high productivity. This effort to maintain high productivity has seen costs escalate. Events in late 2008 have seen renewed efforts to reduce costs through capacity utilisation of higher quality iron ores. This has meant a focus on higher quality lump and fine iron ores as pellets are the most costly blast furnace feed.

Several steel companies have increased consumption of the South African Sishen lump and fine iron ores produced by Kumba Iron Ore. Discussions with blast furnace technical staff at several of Europe’s blast furnaces indicate that this attempt has challenges due to the high alkali content of the Sishen ores. High alkali content of the ore is anathema to the ceramic lining of the blast furnace and efforts to increase use of these ores beyond certain levels is not possible.

The consumption of a trial cargo by two of the European’s premier steel companies in 2009 confirmed the high quality of the Mary River lump and fine iron ores. The lump cargos were consumed in both high and moderate productivity environments and exhibited exceptional qualities in the blast furnace. Both lump cargos also exhibited excellent handling characteristics.

BIM’s natural market for its iron ores is Europe due to its proximity. BIM has letters of intent for sales of nine million tonne of iron ore per annum with five steel companies and expressions of interest for additional tonnage from other Groups.

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Figure 18-4: Iron Ore Price Forecast

Iron Ore Price Forecast - Atlantic Basin

300.0

250.0

200.0

per dmtu) per 150.0 ¢

100.0 Price (US Price

50.0

0.0 1998 1999 2000 2001 2002 2003 2004 2005 2006 2007 2008 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 Year

Lump Iron Ore Price Forecast Fine Iron Ore Forecast

Source: AME, CRU

18.9 Taxation

Taxation in Nunavut is subject to Federal and Nunavut Income taxes. In addition, mining income is also subject to the Nunavut Mining Royalty a graduated tax from 5 to 14% based upon net income. The Royalty calculation is the same as for income taxes except for interest on debt, which is not deductible for calculation of the Nunavut Mining Royalties payable.

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18.10 Financial Analysis

Table 18-8: Capital Cost Estimate Summary Mary River Iron Ore - Trucking Option Project Cost Summary

Closure Costs 27,000 Sustaining Capital 146,057

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Table 18-9: Operating Cost Estimate Summary

CAD/DMT Mining 8.21 Processing 2.80 Road haul 8.85 Stock Piling 1.45 Ship Loading 3.50 Camp 2.06 Air Transportation 1.27 G&A 0.79 Total 28.93 Ocean Shipping* Excluded

18.11 Basis of Financial Analysis

The tabulation below summarises the financial data developed for the project. Net Present Value calculations are annual and year 21 is assigned for mine closure.

Table 18-10: Financial Summary

(Cdn. $ x million) Production Statistics Proven and Probable Reserves 60.7 million tonnes Life of Mine Waste 17.3 million tonnes Stripping Ratio 0.30 Moisture Content Lump 1% Fines 3% Crusher recovery 100% Grade 66.2% Fe Production Schedule – Ore Shipped 58.7 million tonnes Annual Production Shipments 3.0 million tonnes Production Life 20 years Lump Shipments 75% Fines Shipments 25%

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Table 18-11: Revenue

Revenue (FOB Milne Inlet) Cdn$ million Cdn$/tonne of ore Iron Ore Sales Lump 5,745 128.03 Fines 1,541 100.66

Total Revenue 7,286

Product Pricing (Average) US$ /tonne US¢/dmtu

Lump Iron Ore Price 119.66 182.9 Fines Iron Ore Price 94.08 146.7

Foreign Exchange 1.07 Cdn$ 1.00 US$

18.11.1 Results of Base Case Financial Analysis

A summary of the cashflow analysis is included in Table 18-12. The cashflow, on an annualized basis, is shown in Table 18-12. A summary of the financial model is included as Table 18-13.

Table 18-12: Base Case Financial Analysis Summary

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Table 18-13: Financial Model

OWNER OPERATED OPTION Production Metrics (000 tonnes - dry, unless otherwise indicated)

Statistics Total 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031 2032 2033 Capital Expenditure Mining 59,469 - - 24,977 34,492 ------Crushing,Screening Plant, and Shiploading 77,120 - - 16,966 58,611 1,542 ------Tote Road Updgrade/Rail (DFS) 85,268 - - 42,634 42,634 ------Infrastructure at Mary River (Mine) 122,241 - - 53,786 56,231 12,224 ------Infrastructure at Milne Inlet (Port) 96,377 - - 64,573 26,986 4,819 ------Port Facilities 58,216 - - 34,930 23,286 ------Indirect costs ------Constructions Indirect costs 142,797 - 7,140 71,399 58,547 5,712 ------Owner Costs 8,970 - 2,691 3,140 3,140 ------Gravel Royalties 12,540 - - 8,778 3,762 ------Contingency 77,280 - - 38,640 38,640 ------Closure ------Sustaining Capital 147,236 - - - - 3,024 3,144 3,024 6,312 5,972 10,352 7,244 10,200 5,872 8,112 14,272 12,080 11,792 6,292 3,152 10,760 9,792 9,792 3,024 3,024 ------Total Capital expenditures (Excluding Exploration) CAD$000 897,514 - 19,831 359,822 346,328 27,321 3,144 3,024 6,312 5,972 10,352 7,244 10,200 5,872 8,112 14,272 12,080 11,792 6,292 3,152 10,760 9,792 9,792 3,024 3,024

Production Fe Grade (Average Annual) 66.5% 66.5% 66.2% 66.2% 66.0% 66.0% 66.0% 66.0% 66.0% 66.0% 66.0% 66.0% 66.0% 66.0% 66.0% 66.6% 66.6% 66.6% 66.6% 66.6% 66.6% Total Rock Mined 000 wmtu 100,137 - - - 3,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 5,000 2,137 Strip Ratio 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 2.40 Total Ore Mined 000 wmtu 60,082 - - - 1,800 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 3,000 1,282 Handling and Processing Losses 000 wmtu (599) - - - (18) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (30) (13) Total Ore Shipped 000 wmtu 58,691 - - - 990 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 2,970 1,269 Lump Ratio ------Total Lump Ore - moisture content of: 000 dmtu 44,018 - - - 743 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 952 Total Lump Ore - dry metrc tonne units (000 dmtu) 000 dmtu 43,578 - - - 735 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 2,205 942 Fines Ratio ------Total Fines Ore - moisture content of: 000 dmtu 44,018 - - - 743 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 2,228 952 Total Fines Ore - dry metric tonne units (000 dmtu) 000 dmtu 14,233 - - - 240 720 720 720 720 720 720 720 720 720 720 720 720 720 720 720 720 720 720 720 308

TOTAL REVENUE All metals CAD$000 7,286,248 - - - 172,398 439,712 389,249 366,156 356,676 354,526 356,204 359,980 359,980 359,980 359,980 359,980 359,980 359,980 359,980 363,252 363,252 363,252 363,252 363,252 155,230

Operating costs Mining CAD$000 ------Processing CAD$000 ------Road Haul CAD$000 ------Stock Piling CAD$000 ------Ship Loading CAD$000 ------Camp CAD$000 ------Air Transportation CAD$000 ------General & Administration CAD$000 ------Ocean Shipping (Excluded) CAD$000 ------Total CAD$000 1,739,937 - - - 38,754 85,470 87,660 87,360 86,160 87,480 86,760 87,270 87,060 85,680 85,290 85,290 85,290 85,290 85,290 90,990 90,990 90,990 90,990 90,990 38,883

Land Lease Costs CAD$000 38,400 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 1,600 Exploration CAD$000 ------Bonding deposits CAD$000 27,020 - 20 27,000 ------

Taxes Cash Taxes CAD$000 1,071,087 ------31,943 62,006 61,782 62,544 63,371 63,369 63,884 64,110 63,432 63,018 63,058 63,603 63,864 63,621 63,136 63,046 63,649 27,652 ------Working Capital CAD$000 - - - 19,377 6,264 657 (90) (360) 396 (216) 153 (63) (414) (117) - - - - 1,710 - - - - (27,297)

Royalty Nunavut Mining Royalty CAD$000 442,109 ------19,476 26,186 27,054 27,681 27,721 28,044 28,184 27,777 27,528 27,552 27,880 28,038 27,891 27,599 27,545 27,909 8,044 IIBA CAD$000 ------

Cumulative Cash Flow

Production Years 2,010 2,011 2,012 2,013 2,014 2,015 2,016 2,017 2,018 2,019 2,020 2,021 2,022 2,023 2,024 2,025 2,026 2,027 2,028 2,029 2,030 2,031 2,032 2,033 ------Net Project Cash Flow (Before Taxes) CAD$000 4,141,268 (1,600) (21,451) (388,422) (233,661) 319,056 296,188 274,262 243,488 232,892 230,654 236,032 233,462 239,198 236,910 231,041 233,482 233,746 238,917 237,762 232,011 233,271 233,325 239,730 130,976 Cumulative Net Cash Flows CAD$000 (1,600) (23,051) (411,473) (645,134) (326,078) (29,889) 244,372 487,860 720,752 951,406 1,187,438 1,420,900 1,660,098 1,897,008 2,128,049 2,361,530 2,595,276 2,834,193 3,071,956 3,303,966 3,537,237 3,770,563 4,010,292 4,141,268 Payback year Payback period Years 3 - - - 1 1 1 0 ------

------Net Project Cash Flow (After Taxes) CAD$000 3,070,181 (1,600) (21,451) (388,422) (233,661) 319,056 296,188 242,319 181,482 171,110 168,110 172,661 170,093 175,314 172,800 167,609 170,463 170,688 175,315 173,898 168,390 170,135 170,279 176,080 103,323 Cumulative Net Cash Flows CAD$000 (1,600) (23,051) (411,473) (645,134) (326,078) (29,889) 212,430 393,911 565,022 733,132 905,792 1,075,886 1,251,200 1,424,000 1,591,609 1,762,072 1,932,760 2,108,075 2,281,973 2,450,363 2,620,498 2,790,777 2,966,858 3,070,181 Payback year Payback period Years 2.6 - - - 1 1 1 0 ------

Summary of Cash Flow After Tax Pre-tax Cumulative net cash flow Undiscounted (from 2010 to 2025) CAD$000 3,070,181 4,141,268

Net present value (Excluding 2010) Discounted at 4% CAD$000 1,748,613 2,366,663 Discounted at 6% CAD$000 1,337,661 1,817,102 Discounted at 8% CAD$000 1,029,815 1,406,533 Discounted at 10% CAD$000 795,951 1,095,512 Discounted at 12% CAD$000 615,950 856,814 Discounted at 20% CAD$000 206,774 317,473 Internal rate of return 30.6% 34.4% Payback period Years 2.6 2.6

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Table 18-14: Summary of the Pre-Tax Annual Cash Flow Production Net Project Cash Years Year Flow (Pre-Tax) CAD$000

2010 -3 (1,600) 2011 -2 (21,451) 2012 -1 (388,422) 2013 1 (233,661) 2014 2 319,056 2015 3 296,188 2016 4 274,262 2017 5 243,488 2018 6 232,892 2019 7 230,654 2020 8 236,032 2021 9 233,462 2022 10 239,198 2023 11 236,910 2024 12 231,041 2025 13 233,482 2026 14 233,746 2027 15 238,917 2028 16 237,762 2029 17 232,011 2030 18 233,271 2031 19 233,325 2032 20 239,730 2033 21 130,976 TOTAL 4,141,268

18.11.2 Base Case Sensitivity Analysis

The project is subject to four main factors that have significant effect on financial returns of the project. These are the price of iron ore specifically lump iron ore, the currency exchange rate, and the Capital Cost (CAPEX) and Operating Cost (OPEX) matrix.

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Table 18-15: Owner Operated ($ 000)

Change in Factor SENSITIVITY OF NPV @ 8% (After-Cash Tax) Factor -30% -20% -10% 0% 10% 20% 30% Capital expenditure 1 1,175,956 1,127,963 1,079,529 1,029,815 980,582 931,272 880,643 Operating expenditure 2 1,171,493 1,124,144 1,077,156 1,029,815 982,499 935,556 888,368 Metal price 3 434,605 637,140 835,451 1,029,815 1,223,516 1,415,686 1,606,577

Factor Exchange Rate 4 434,605 637,140 835,451 1,029,815 1,223,516 1,415,686 1,606,577 Grade 5 434,605 637,140 835,451 1,029,815 1,223,516 1,415,686 1,606,577

Change in Factor SENSITIVITY OF NPV @ 10%(After-Cash Tax) Factor -30% -20% -10% 0% 10% 20% 30% Capital expenditure 1 937,095 890,840 843,990 795,951 748,329 700,619 651,567 Operating expenditure 2 915,021 875,231 835,746 795,951 756,177 716,733 677,073 Metal price 3 295,519 466,298 633,036 795,951 958,218 1,119,024 1,278,358

Factor Exchange Rate 4 295,519 466,298 633,036 795,951 958,218 1,119,024 1,278,358 Grade 5 295,519 466,298 633,036 795,951 958,218 1,119,024 1,278,358

Change in Factor SENSITIVITY OF NPV @ 12% (After-Cash Tax) Factor -30% -20% -10% 0% 10% 20% 30% Capital expenditure 1 752,450 707,793 662,424 615,950 569,839 523,633 476,085 Operating expenditure 2 717,315 683,446 649,835 615,950 582,085 548,509 514,744 Metal price 3 189,629 335,529 477,583 615,950 753,687 890,034 1,024,767

Factor Exchange Rate 4 189,629 335,529 477,583 615,950 753,687 890,034 1,024,767 Grade 5 189,629 335,529 477,583 615,950 753,687 890,034 1,024,767

18.11.3 Alternate Case Financial Analysis

An alternate financial analysis was performed to assess the economics of a mining operation that was performed under contract. Financial analysis of the alternate case showed the pre-tax project NPV at 8% to be CAD$795 million, the after-tax NPV at 8% to be CAD$577 and the internal rate of return (IRR) to be 22.3%. The payback period is 3.6 years.

The capital and operating cost assumptions for this case are included as Table 18-16 and Table 18-17, and a summary of the resulting financial analysis as Table 18-18.

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Table 18-16: Alternate Case, Summary of Contractor Operator Capital Costs Project Area Description Cost ($ millions) Direct Costs Mining 19.7 Crushing, Screening Plant 38.2 Stockpiling & Shiploading - Infrastructure at Mary River (Mine) 119.7 Infrastructure at Milne Inlet (Port) 96.4 Ore Dock Facilities 20.5 Tote Road 102.0 Total Direct Costs 396.5 Indirect Costs Project Indirect Costs 128.1 Owner’s Costs 7.3 Gravel Royalties 12.5 Total Indirect Costs 147.9 Subtotal Direct and Indirect Costs 544.4 Contingency 60.3 Total Project Costs 604.7 Additional Life of Mine Costs Closure Costs 27.0 Sustaining Capital n/a

Table 18-17: Alternate Case, Summary of Contractor Operator Operating Costs

$/t ore Mining 17.29 Processing — Road haul 17.51 Stockpiling 17.00 Shiploading — Non-Productive Days/Contingency 5.18 Camp 2.20 Air Transportation 1.04 Tug Operations 1.68 G&A 0.81 Total 62.71

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Table 18-18: Alternate Case Financial Analysis Summary

18.12 Risk and Opportunity Analysis

As part of the feasibility study work, a risk assessment workshop was carried out to identify and categorize the key potential risks and opportunities associated with developing Deposit No. 1, based upon available information at the feasibility study stage. Wherever it was possible and practical, mitigation measures were included in the feasibility study, both in the design and operation of the facilities. However, some measures require further investigations and data collection; these will need to be addressed in the basic engineering phase. Some of the issues that are not fully defined also represent an opportunity to optimise designs and reduce costs once the data are available.

Key risk/opportunity areas were recognised as follows:

Delays arising from the Project approvals process or permitting Engineering for road alignment, port facilities and foundation designs for structures is based on limited geotechnical data Construction of the Project will be a significant logistical challenge Availability of skilled labour is a risk, but it also presents an opportunity to maximise Inuit participation The feasibility study does not include the cost of the ships performing the ore transport and assumes that they will be chartered from interested parties Provision for cost escalation will need to be considered.

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C ONTENTS

19.0 OTHER RELEVANT DATA AND INFORMATION ...... 19-1

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19.0 OTHER RELEVANT DATA AND INFORMATION

To the best of the Authors knowledge there is no additional information required to make this technical report understandable and not misleading.

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C ONTENTS

20.0 INTERPRETATIONS AND CONCLUSIONS ...... 20-1

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20.0 INTERPRETATIONS AND CONCLUSIONS

In the opinion of the QPs, the following interpretations and conclusions are appropriate to the Project:

Legal opinion provided to AMEC indicates that the mining tenure held by BIM in the Project area is valid, and sufficient to support declaration of Mineral Resources and Mineral Reserves. Legal opinion provided to AMEC indicates that BIM holds sufficient surface rights in the Project area to support the planned mining operations Water for the Project can be sourced from Camp Lake (Mary River site) and Phillips Creek and Thirty Two Kilometre Lake (Milne Inlet). Water is required only for domestic use and dust control operations in the mine and ore haulage operations. Water is not required for the crushing and screening operation. Easements or rights of way will be required to support the trucking operation proposed; it is a reasonable expectation that the rights of way will be obtainable to support planned mining operations Permitting for the Project requires that an environmental impact assessment (EIA) be prepared and approved prior to any mining activity being permitted. It is currently expected that The draft EIS was submitted in January, 2011 BIM will need to apply for additional permits as appropriate under Inuit, Provincial, and Federal laws to allow mining operations. Current field and exploration activities are permitted The existing and planned infrastructure, availability of staff, the existing power, water, and communications facilities, the methods whereby goods are transported to the mine, and any planned modifications or supporting studies are well- established, or the requirements to establish such, are well understood by BIM, and can support the declaration of Mineral Resources and Mineral Reserves The geologic understanding of the deposit settings, lithologies, and structural and alteration controls on mineralization is sufficient to support estimation of Mineral Resources and Mineral Reserves The mineralization style and setting is well understood and can support declaration of Mineral Resources and Mineral Reserves. Work completed on the Project includes geochemical sampling, trial mining for bulk sampling, mineral resource estimation, core drilling including geotechnical, hydrological, confirmation and condemnation drill holes, baseline environmental studies, metallurgical testwork, and engineering and design studies. Completed

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exploration and development programs were appropriate to the mineralization style. Sampling methods are acceptable, meet industry-standard practice, and are acceptable for Mineral Resource and Mineral Reserve estimation purposes The quality of the analytical data used in Mineral Resource and Mineral Reserve estimation is reliable and sample preparation, analysis, and security are generally performed in accordance with exploration best practices and industry standards. Metallurgical testwork completed on the Project has been appropriate to establish process routes that are applicable to the mineralization types and was performed on samples that were representative of the mineralization. Metallurgical testwork has shown that the mineralization is amenable to processing as lump ore, with three bulk tests performed at European blast furnaces. Process design envisages a crushing and screening plant, followed by trucking to a port at Milne Inlet, then shipping to European markets. Product will average 66% Fe. The lump ore ratio of the direct shipping material will be high and is expected to be in a range over 75% (plus 6.3 mm) Mineral Resources and Mineral Reserves, which were estimated using core drill and surface channel sample data, have been performed to industry best practices, and conform to the requirements of CIM Definition Standards (2005). Reviews of the environmental, permitting, legal, title, taxation, socio-economic, marketing and political factors and constraints for the Project support the declaration of Mineral Reserves using the set of assumptions outlined The proposed open pit mining method is appropriate to the style of mineralization Mining will proceed using a conventional truck-and-shovel fleet Production forecasts are achievable with the proposed equipment and plant The predicted mine life of 20 years is achievable based on the projected annual production rate and the Mineral Reserves estimated There is some additional upside if the rail scenario considered in the superceded 2008 feasibility study can be integrated into this trucking option feasibility study.

“Indicative” purchase terms contained within the sales quotations received are typical and consistent with standard industry practice and are similar to contracts for the supply of lump iron ore elsewhere in the world; these were used to establish marketing costs

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for the supply of lump iron ore elsewhere in the world; these were used to establish marketing costs Taxation considerations were included in the financial analysis commensurate with the Canadian setting of the Project Total life-of-mine average operating costs are projected at CAN$28.93/dmt Total life-of-mine capital costs are estimated at CAD$740.3 million. Provision was made, in addition, for CAN$27 M in projected closure costs, and CAN$146 M of sustaining capital. The base case financial data is assumes an Owner-mining scenario. Financial analysis of the base case showed the pre-tax project NPV at 8% to be CAD$1,407 million, the after-tax NPV at 8% to be CAD$1,030 million and the internal rate of return (IRR) after tax to be 30.6%. The payback period is 2.6 years. Sensitivity analysis shows that the Mary River Project is most sensitive to Iron ore price. An alternative case, which envisages contractor mining, was evaluated. Financial analysis of the alternate case showed the pre-tax project NPV at 8% to be CAD$795 million, the after-tax NPV at 8% to be CAD$577 and the internal rate of return (IRR) to be 22.3%. The payback period is 3.6 years for this scenario. A number of risks and opportunities were identified. Most will require further investigations and data collection; these will need to be addressed in the basic engineering phase. Some of the issues that are not fully defined also represent an opportunity to optimise designs and reduce costs once the data are available Significant exploration potential remains within the Project, with four additional deposits identified in 2010, all of which remain as encouraging drill targets.

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C ONTENTS

21.0 RECOMMENDATIONS ...... 21-1 21.1 Environmental ...... 21-1 21.2 Geology ...... 21-1 21.3 Geotechnical ...... 21-2 21.4 Mining ...... 21-2 21.5 Process, Materials Handling, Transport and Infrastructure ...... 21-3 21.6 Tote Road ...... 21-4 21.7 Dock and Port Management ...... 21-5 21.8 Construction ...... 21-5 21.9 Budget ...... 21-6

T ABLES

Table 21-1: Drilling Budget for 2011 ...... 21-6 Table 21-2: Overall Budget (CAD Million) ...... 21-6

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21.0 RECOMMENDATIONS

The Feasibility Study identifies a number of areas where additional data or studies are required to address the requirements of the next phases of work. There are no contingent proposed phases of work

Study recommendations are summarised by category below, with more detail being provided in the body of the Feasibility Study. Many of the recommendations are a result of the risk and opportunity evaluation described in Section 19.14.

21.1 Environmental

Carry out additional studies as follows.

Continue archaeological investigations. Continue with environmental baseline studies. Provide construction guidelines and sequence construction activities to reduce potential impact on wildlife. Continue compliance monitoring of site activities.

21.2 Geology

Waste rock: The waste rock types have to be tested to determine the potential for acid generation. This would affect the capacity, location and mitigation requirements associated with the waste dumps. The overburden surface should be defined accurately within the perimeter of the Final pit. This allows for more accurate slope design and a better estimate of the required mining fleet. BIM should continue their infill and regional exploration program.

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21.3 Geotechnical

Pit Slope Angle: Detailed geotechnical test work and study are required to define the pit slopes for different rock types, i.e. overburden, mineralized rock, and waste rock. Some approximations are included in the provided slope angles due to the limited geotechnical data. This becomes more important when the 3Mtpa mining operation is expanded to the 18Mtpa production rate with a final pit wall height of few hundred meters. Similar detailed studies for the waste dump area and the dump slope angle will result in a better understanding and design of the dumps. Water condition within the pit and waste dumps must be defined in order to finalize the slope stability analysis for both open pit and waste dumps. This also assists in designing the dewatering systems to divert water from the mining operation. Drill and blast-ability tests must be conducted to better estimate the drill and blast pattern. This will also assist in the estimation of the drilling equipment performances and explosives consumption. Geotechnical investigations for the port and dock facilities, tote road, site roads, stockpile locations, crushing and screening facilities and site buildings.

21.4 Mining

A detailed study is required to define the quantity and quality of the inpit water. Certain areas should be allocated for dumping of rock/overburden/snow that is contaminated by fuel/oil spills from the mining and ancillary equipment. These areas should be lined to prevent any migration to the local environment. A better estimation of the Potentially Acid Generating (PAG) material within the extent of the 20 year pit is required in order to better estimate the size of the waste dumps. Proposed waste dump setup for the PAG and overburden material appears to be impractical for a smooth mining operation. In the early years and for the suggested area for the waste dump, sufficient clean waste rock may not be available to provide the suggested 20m base for the PAG material, due to the sloped topography. The waste dump configuration must be revisited in the Basic Engineering Phase of the project. Due to the extreme cold weather in the winter, metal fatigue on body and hydraulic systems of the mining and ancillary equipment may result in an increase in the

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maintenance cost. Other mining operations in extreme cold weather conditions must be assessed to optimize the mine planning and equipment utilization at Mary River. Optimize the design of the access road between the pit and the crusher with a view towards reducing the capital cost. Confirm used tire disposal in the mine waste dump is viable and can be permitted. If not provision must be made for backhaul and disposal in the south.

21.5 Process, Materials Handling, Transport and Infrastructure

In the course of the study and during the risks and opportunities evaluation, the following opportunities were identified. These were not pursued due to study timing constraints but should be considered in the future due to the potentially very large positive influence on project financial result:

Optimize production rate to approximately 4 Mtpa, with the same port and stockpile facilities, larger secondary crusher and additional mobile equipment by accepting greater demurrage costs (which are small compared to additional revenue). Use simulation modeling of the shiploading process and all influencing factors to optimize the production rate. Reduce operating unit cost (due to fixed operating costs distributed over more tonnes). Optimize construction forces and hence camp size Review, prioritize and resolve all items of risk and opportunities identified in Section 19.14. Consider alternative power generators with lower capital cost but higher operating cost. Consider a lower cost accommodation design that incorporates an ablution area rather than shared wash facilities between individual rooms. Consider fabric covered accommodation and other buildings (similar to current Weatherhaven camp). Reconsider a single shiploader and reclaim conveyor rather than a dual system together with consideration of the increased production potential associated with acceptance of higher demurrage costs.

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Consider covered conveyors rather than enclosed conveyors. Optimize the crushing/screening layout and consider lower cost building alternatives. Optimize the mining and haulage workshop size and consider integration into a single building. Consider the use of a larger capacity road train that incorporates a third trailer with a powered dolly. Reconsider the 100 year flood criteria used for design of the bridges and culverts for the Tote Road upgrade, particularly if a short, temporary trucking operation is considered as a precursor to a long term rail operation. Review, prioritize and resolve all items on the Risk/Opportunities registers included in Section TBA. Consider a temporary trucking operation with a modest mine life of say 5 years as a precursor to a larger rail operation. This may enable lower cost design features and alternative operating philosophies. Conduct a competitive tendering process and/or negotiate a lower cost of contractor services.

21.6 Tote Road

Complete of geotechnical work along Tote Road to further refine and optimize the road structure design. Complete of geotechnical work for quarries for aggregate sourcing. Completion of material properties testing for aggregates and concrete. Complete of paleontological surveys of quarry sites (if required). Commence preliminary design for the bridges. Begin dialogue with approving agencies so as to better define the design constraints. The Tote road design has several areas where there may be potentially high snow accumulations. This design requires a careful review at the next design stage including consideration of road design modifications and/or construction of snow fence structures.

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21.7 Dock and Port Management

Carry out further optimisation of the dock design; there is potential for a reduction in costs. It is suggested that all ship manoeuvres in the harbour area are simulated through a single short open water period. The effectiveness of ballast discharge should be reviewed.

21.8 Construction

Develop a more detailed schedule for construction considering various constraints, including possible environmental restrictions. Develop an early partnership with a qualified contractor with a proven track record for working in remote northern environments. Communicate closely with a qualified contractor to anticipate changes in the future labour market. Communicate closely with a qualified contractor when developing the IIBA to ensure all issues are addressed with respect to training, Inuit labour targets, etc. Where possible, construct camp accommodations off-site and mobilise to site (in part) in 2012. Where possible, construct fuel storage tanks off-site and mobilise to site in 2011/20012, to ensure quicker set up and operation. Develop an early partnership with a qualified contractor to ensure appropriate lead times for equipment and material procurements are optimised. Further Investigate potential borrow sources to ensure anticipated construction materials are of sufficient quality in the quantities required. Further investigations for the infrastructure borrow sources are required.

Engage a qualified contractor in the detailed design phase of the Project to address constructability issues. Also, set up a contractual arrangement that deals seamlessly. Prepare a detailed execution plan, particularly with respect to early activities and commitments during Q1 and Q2 2011 that is required for the 2011 sealift, 2011 field work and for procurement of long lead items.

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21.9 Budget

A budget for additional works has been prepared by BIM and is shown in Table 21-1 and Table 21-2.

Table 21-1: Drilling Budget for 2011

Year Year 1 Drilling Type metres Geotechnical Deposit 1 2000 Deposit No 1 Infill 2000 Regional Exploration Deposit 6 & 8 2000 Total 6000

Budget (Cdn $ Million) 6

Table 21-2: Overall Budget (CAD Million)

Year Year 1 Drilling 6 Operations Support 10 Technical Services 12 Human Resources Health and Safety 2 Engineering &Construction 7 Corporate 4 Budget Total 41

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C ONTENTS

22.0 REFERENCES ...... 22-1 22.1 Geology and Mineral Resources ...... 22-1 22.2 Metallurgy ...... 22-3 22.3 Geotechnical References ...... 22-3

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22.0 References

22.1 Geology and Mineral Resources

Aker Kvaerner. 2008 Expansion Study, Mary River Iron Ore Project Northern Baffin Island, Nunavut. (Unpublished study).

Aker Kvaerner. 2008 Technical Report of the Definitive Feasibility Study, Mary River Iron Ore Project Northern Baffin Island, Nunavut.

Aker Kvaerner. 2006 Technical Report, Mary River Iron Ore Project Northern Baffin Island, Nunavut.

Baffinland Iron Mines Corporation. 2005 Annual Report 2004 32 p., 2006 Annual Report 2005, 44 p.

Baffinland Iron Mines Limited. 1964 Engineering Report No. 2 for Baffinland Iron Mines Limited on the Mary River Iron Deposits. Watts, Griffis and McOuat Ltd., Report, v. 1 (text) – 179 p., v. 2, (maps).

Baffinland Iron Mines Limited. 1965 Engineering Report No. 3 for Baffinland Iron Mines Limited on the Mary River Iron Deposits. Watts, Griffis and McOuat Ltd., Report, v. 1 (text) – 221 p., v. 2 (maps).

Blais, R.A. (1964). Geological report on the Baffinland iron deposits, Mary River, Baffin Island, N.W.T. Report for Baffinland Iron Mines Ltd., 50 p.

Harmsworth, R.A., Kneeshaw, M., Morris, R.C., Robinson, C.J., and Shrivastava, P.K. (1990). BIF-derived iron ores of the Hamersley Province. In: Geology of the Mineral Deposits of Australia and Papua New Guinea, (ed.) F.E. Hughes; Australian Institute of Mining and Metallurgy, p. 617-642.

Iannelli, T.R. 2005. Geology of the Mary River Iron Deposits, Unpublished Paper.

Jackson, G.D. (2006). Field data, NTS 37 G/5 and G/6, northern Baffin Island, Nunavut, Canada. Geological Survey of Canada, Open File 5317, 2 sheets.

Jackson, G.D. (2000). Geology of the Clyde – Cockburn Land map area, north-central Baffin Island, Nunavut. Geological Survey of Canada, memoir 440, 303 p.

Jackson, G.D. (1965). Geology of NTS map sheets 37G/5 and 37G/6, 1:50,000 scale. Geological Survey of Canada (unpublished geology maps).

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Jackson, G.D. and Berman, R.G. (2000). Precambrian metamorphic and tectonic evolution of northern Baffin Island, Nunavut, Canada. Canadian Mineralogist, v. 38, p. 399-421.

Jackson, G.D. and Morgan, W. (1978). Geology, Conn Lake, District of Franklin. Geological Survey of Canada, Map 1458A, scale 1:250,000.

Jackson, G.D., Morgan, W.C., and Davidson, A. (1978a). Geology, Icebound Lake, District of Franklin. Geological Survey of Canada, Map 1451A, coloured, with expanded legend, scale 1:250,000.

Jackson, G.D., Morgan, W.C., and Davidson, A (1978b). Geology, Buchan Gulf-Scott Inlet, District of Franklin. Geological Survey of Canada, Map 1449A, coloured, with expanded legend, scale 1:250,000.

Jackson, G.D., Morgan, W.C., and Davidson, A (1978c). Geology, Steensby Inlet, District of Franklin. Geological Survey of Canada, Map 1450A, coloured, with expanded legend, scale 1:250,000.

Jackson, G.D. (1966). Geology and Mineral Possibilities of the Mary River Region, Northern Baffin Island. Canadian Mining Journal, 87, No. 6, 57-61.

Johnson, R.C. (2007). Petrographic Description of Selected Drill Core and Channel Samples from the North Limb, Baffinland Iron Mines Corporation Deposit No. 1 Mary River Unpublished report.

MacLeod, M. (2009) Manganese and Phosphorous Enrichment in Baffinland Iron Mines Corp. Deposit No. 1, Mary River Iron Formation, Northern Baffin Island, Nunavut. BSC Thesis

Taylor, D., Dalstra, H.J., Harding, A.E., Broadbent, G.C., and Barley, M.E. (2001). Genesis of high-grade hematite orebodies of the Hamersley Province, Western Australia. Economic Geology, v. 96, p. 837-874.

von Guttenberg, R. and Farquharson, G. (2003). A review of the Mary River Iron Ore Project, northern Baffin Island, Nunavut. Strathcona Mineral Services Ltd., 175 p.

Yukovenko, A. New-Sense Geophysics. Logistics Report for the High Resolution Helicopter Magnetic Airborne Geophysical Survey. Nov. 2008.

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22.2 Metallurgy

Studien-Gesellschaft fur Eisenerz-Aufbereitung – Status-Report on Chemical, Physical and Metallurgical Characterisation of Mary River Iron Ore, Baffin Island, December 2007.

Studiengesellschaft fur Eisenerzaufbereitung SGA – Summary of Results – Mary River 1 Samples (Batch 1-4), 27.11.2007.

Studiengesellschaft fur Eisenerzaufbereitung SGA – Drop Tests.

Metso Minerals – Test Report on MP Cone Pilot Crushing Tests, November 28, 2006.

Metso Minerals – Test Summary Report, November 29, 2006.

ProMet Engineers – Report C5305-PR-002-P2 – Brief Technical Review of the Mary River Deposit Based on 2006 Reports, September 2007.

22.3 Geotechnical References

Knight Piésold Ltd., Memo Ref. No. NB07-00641, July 27, 2007.

Knight Piésold Ltd., “ Inlet Port Site – 2007 Investigation Summary Report”, Ref. No. NB102-0031/8-1, November 2007.

Knight Piésold Ltd., “Mine Site Infrastructure, Pit Overburden and Waste Dumps 2006 Site Investigation Report”, Ref. No. NB102-0031/2, February 28, 2007.

Knight Piésold Ltd., “Mine Site Infrastructure 2007 Site Investigations and Foundation Recommendations Summary”, Ref. No. NB102-0031/8-1, November, 2007.

Knight Piésold Ltd., “Rail Infrastructure 2006 Site Investigation Summary Report”, Ref. No. NB102-0031/3-3, March 26, 2007.

Knight Piésold Ltd., “Rail Infrastructure 2007 Site Investigation Summary Report”, Ref. No. NB102-0031/8-3, November 2007.

Knight Piésold Ltd., Memo Ref. No. NB07-00744, August 28, 2007.

Knight Piésold Ltd., Memo Ref. No. NB07-00380, May 25, 2007.

Knight Piésold Ltd., Memo Ref. No. NB07-00392, May 30, 2007.

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Knight Piésold Ltd., Memo Ref. No. NB07-00475, June 8, 2007.

Knight Piésold Ltd., Memo Ref. No. NB07-00873, October 12, 2007.

Knight Piésold Ltd., Memo Ref. No. NB07-00548, June 15, 2007.

Knight Piésold Ltd., Memo Ref. No. NB07-00495, May 31, 2007.

Knight Piésold Ltd., Memo Ref. No. NB07-00669, August 2, 2007.

Knight Piésold Ltd., Memo Ref. No. NB07-00784, September 20, 2007.

16. Knight Piésold Ltd., Memo Ref No. NB07-00805, September 20, 2007.

AMEC, Memo: TC101510-2010_08_06-Rev0 - Construction Aggregate Resource Assessment Trip Report.

AMEC, Memo: TC101510-2010_09_14-Rev0 - Trucking Feasibility Mary River Foundation Recommendations.

AMEC, Memo: TC101510-2010_10_13-290-MEM-014rev0 -Milne Inlet Foundation Considerations.

AMEC, Memo: TC101510-2010_10_19-290-MEM-015rev0 -Tote Road Design Considerations.

AMEC, Memo: TC101510-2010_10_21-290-MEM-016rev1 - Gravel Tote Road Design.

AMEC, Memo: TC101510-2010_12_09-290_MEM-017rev1-ToteRoadGeochem.

AMEC, Memo: TC101510-2010_11_25-290-DRAFT_MEM-018rev0-Review of Waste Dump Design.

AMEC, Memo: TC101510-2010_11_30-290-DRAFT_MEM-020rev0 - Open Pit Geo- mechanics Review.

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C ONTENTS

23.0 DATE AND SIGNATURES ...... 23-1

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Baffinland Iron Mines Corporation Mary River Baffin Island Mary River Iron Ore Trucking NI 43-101 Technical Report

23.0 DATE AND SIGNATURES

The undersigned prepared this technical report titled “Mary River, Baffin Island Mary River Iron Ore Trucking NI 43-101 Technical Report” effective date January 13th, 2011.

“Signed and sealed”

G. Wahl, P. Geo 15 Feb, 2011

“Signed and sealed”

R. Gharapetian, P. Eng. 15 Feb, 2011

“Signed and sealed”

James Jackson, P. Eng. 15 Feb. 2011

“Signed and sealed”

Vikram Khera, P. Eng. 15 Feb, 2011

“Signed and sealed”

Gregory G. Wortman, P. Eng. 15 Feb, 2011

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