43-101_bloom_lake_nov_08[1]

630 boul. René-Lévesque Ouest, Suite 2500, Montréal (Québec) Canada H3B 1S6 Tel.: (514) 866-2111 • Fax: (514) 866-2116 • [email protected] • www.bba.ca

TECHNICAL REPORT

BLOOM LAKE PROJECT LABRADOR TROUGH, QUEBEC

Technical Report 43-101 on the

Feasibility Study for the Bloom Lake Project 8-million tonnes per year of Iron Concentrate

Prepared by:

Consolidated Thompson Iron Mines Ltd.

and BBA inc.

André Allaire, Eng., M. Eng., Ph.D. Enzo Palumbo, M. Eng. Patrice Live, Eng. René Scherrer, Eng.

November 12, 2008

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TABLE OF CONTENTS

1. SUMMARY.....................................................................................................................1-1 1.1 Introduction ....................................................................................................................1-1 1.2 Location..........................................................................................................................1-1 1.3 Ownership ......................................................................................................................1-1 1.4 Project Status.................................................................................................................1-1 1.5 Geology and Mineral Resources....................................................................................1-1 1.6 Mineral Reserves ...........................................................................................................1-3 1.7 Metallurgical Testwork ...................................................................................................1-5 1.8 Processing .....................................................................................................................1-5 1.9 Infrastructure and Support Systems...............................................................................1-5 1.10 Railway & Port Facilities.................................................................................................1-6 1.11 Schedule ........................................................................................................................1-6 1.12 Environmental and Permitting ........................................................................................1-6 1.13 Financial Analysis ..........................................................................................................1-7 1.14 Recommendations .......................................................................................................1-10

2. INTRODUCTION............................................................................................................2-1 2.1 Introduction ....................................................................................................................2-1 2.2 Purpose of the Technical Report....................................................................................2-1 2.3 Basis of the Technical Report ........................................................................................2-1 2.4 Qualified Persons...........................................................................................................2-2 2.5 Site Visit .........................................................................................................................2-2

3. RELIANCE ON OTHER EXPERTS ...............................................................................3-1

4. PROPERTY DESCRIPTION AND LOCATION .............................................................4-1

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

5.1 Access............................................................................................................................5-1 5.2 Climate ...........................................................................................................................5-1 5.3 Local Resources and Infrastructure ...............................................................................5-1 5.4 Physiography .................................................................................................................5-1

6. HISTORY .......................................................................................................................6-1

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7. GEOLOGICAL SETTING ..............................................................................................7-1 7.1 Regional Geology...........................................................................................................7-1 7.2 Local Geology ................................................................................................................7-5

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

9. MINERALIZATION ........................................................................................................9-1 9.1 Lithology and Mineralization...........................................................................................9-1 9.2 Structural Geology .........................................................................................................9-4

10. EXPLORATION ...........................................................................................................10-1 10.1 1998 Program ..............................................................................................................10-1 10.1.1 General .......................................................................................................................10-1 10.1.2 Line Cutting and Surveying.........................................................................................10-1 10.1.3 Mapping and Bulk Sampling .......................................................................................10-2 10.2 QCM Airborne Geophysical Survey .............................................................................10-6 10.3 2005 Program ..............................................................................................................10-6 10.3.1 General .......................................................................................................................10-6 10.3.2 Mini-bulk Sampling Program.......................................................................................10-6

11. DRILLING ....................................................................................................................11-1 11.1 General ........................................................................................................................11-1 11.1.1 Phase I/II.....................................................................................................................11-6 11.1.2 Phase III......................................................................................................................11-6 11.2 Hole Collar and Down-Hole Attitude Surveys ..............................................................11-6 11.3 Core Handling and Logging Protocols .........................................................................11-6 11.3.1 Descriptive Logging ....................................................................................................11-6 11.3.2 Geotechnical Logging .................................................................................................11-7 11.3.3 Video Recording .........................................................................................................11-7 11.3.4 Magnetic Susceptibility ...............................................................................................11-7 11.3.5 Core Storage...............................................................................................................11-7

12. SAMPLING METHODS AND APPROACH.................................................................12-1

13. SAMPLE PREPARATION, ANALYSES AND SECURITY..........................................13-1 13.1 Sample Preparation .....................................................................................................13-1 13.2 Assaying.......................................................................................................................13-1 13.3 QA/QC..........................................................................................................................13-3

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13.4 Security ........................................................................................................................13-3

14. DATA VERIFICATION.................................................................................................14-1

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

16. MINERAL PROCESSING AND METALLURGICAL TESTING...................................16-1 16.1 Historical Testwork.......................................................................................................16-1 16.2 Recent Testwork ..........................................................................................................16-2 16.2.1 Shaking Table and Davis Tube Magnetic Separation Tests .......................................16-2 16.2.2 Grindability Tests and AG Mill Simulations .................................................................16-7 16.2.3 Weight Recovery Versus Total Fe Relationships........................................................16-9 16.2.4 Weight and Iron Recoveries and the Financial Model ..............................................16-11 16.3 Confirmatory Testwork ...............................................................................................16-12 16.3.1 Metallurgical Test Work ............................................................................................16-12 16.3.2 Overall Weight and Fe Recovery ..............................................................................16-21 16.3.3 Confirmatory test work conclusions ..........................................................................16-21

17. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES..............................17-1 17.1 General ........................................................................................................................17-1 17.2 Geological Interpretation ..............................................................................................17-1 17.3 Compositing .................................................................................................................17-2 17.4 Block Model Parameters ..............................................................................................17-4 17.5 Grade Interpolation and Statistics ................................................................................17-4 17.6 Density Assignment ...................................................................................................17-11 17.7 Mineral Resource Definitions .....................................................................................17-16 17.8 Mineral Resource Statement......................................................................................17-17 17.9 Mineral Reserve Estimate ..........................................................................................17-17

18. ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES....................................................18-1

18.1 Mineral Reserves .........................................................................................................18-1 18.1.1 Weight Recovery Model..............................................................................................18-1 18.1.2 Pit Optimization...........................................................................................................18-1 18.1.3 Mining Planning ..........................................................................................................18-9 18.2 Production Equipment Selection ................................................................................18-20 18.3 Service Equipment .....................................................................................................18-23 18.4 Process Design ..........................................................................................................18-24

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18.4.1 Design Ore Type.......................................................................................................18-24 18.4.2 Design Criteria ..........................................................................................................18-25 18.5 Metallurgical Processing Plant ...................................................................................18-25 18.5.1 Crushing, Ore Stockpiling and Grinding....................................................................18-37 18.5.2 Plant Water and Services .........................................................................................18-39 18.6 Tailings Disposal ........................................................................................................18-41 18.6.1 Site Description.........................................................................................................18-41 18.6.2 Fundamental Design Criteria and Assumptions........................................................18-41 18.6.3 Operation ..................................................................................................................18-41 18.7 Infrastructure and Support Systems...........................................................................18-42 18.7.1 Service Building ........................................................................................................18-42 18.7.2 Access and Site Roads.............................................................................................18-44 18.7.3 Power Supply and Distribution..................................................................................18-44 18.7.4 Control System .........................................................................................................18-45 18.7.5 Fresh Water Supply ..................................................................................................18-46 18.7.6 Reclaim Water Supply ..............................................................................................18-46 18.7.7 Fire Protection...........................................................................................................18-46 18.7.8 Fuel Storage .............................................................................................................18-47 18.7.9 Communications System (Local and External) .........................................................18-47 18.7.10 Effluent Water Treatment..........................................................................................18-47 18.7.11 Sanitary Treatment and Waste Disposal ..................................................................18-48 18.7.12 Camp Accommodation..............................................................................................18-48 18.8 Railway Transportation and Port ................................................................................18-48 18.9 Environment ...............................................................................................................18-53 18.9.1 Applicable regulations...............................................................................................18-53 18.9.2 Native Land Status....................................................................................................18-53 18.9.3 Capital Cost Estimate ...............................................................................................18-53 18.10 Operating Cost Estimate ............................................................................................18-55 18.10.1 Manpower .................................................................................................................18-55 18.10.2 Mine ..........................................................................................................................18-56 18.10.3 Process Plant............................................................................................................18-56 18.10.4 Administration ...........................................................................................................18-56 18.10.5 Rail Transportation & Port.........................................................................................18-57 18.11 Financial Analysis ......................................................................................................18-57

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18.11.1 Introduction ...............................................................................................................18-57 18.11.2 Basis of Evaluation ...................................................................................................18-57 18.11.3 Results of Financial Evaluation.................................................................................18-59 18.11.4 Discussion.................................................................................................................18-62

19. PROJECT IMPLEMENTATION...................................................................................19-1

20. OTHER RELEVANT DATA AND INFORMATION ......................................................20-1

21. INTERPRETATION AND CONCLUSIONS .................................................................21-1

22. RECOMMENDATIONS................................................................................................22-1

23. REFERENCES/SOURCES OF INFORMATION..........................................................23-1

24. DATE ...........................................................................................................................24-1

25. CERTIFICATES ...........................................................................................................25-1

Appendix A List of CLM Mining Claims

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LIST OF TABLES

Table 1-1: Total in situ Mineral Resources at a Cutoff Grade of 15% Total Fe..........................1-2 Table 1-2: Mineral Reserves by Category in Ultimate Pit Design ..............................................1-3 Table 7-1 : Regional Stratigraphic Column ................................................................................7-2 Table 9-1 : Summary – Rock Type Codes for Descriptive Logs ................................................9-1 Table 9-2 : Summary of Means and Medians ............................................................................9-2 Table 10-1 : Analytical Results for Grab Samples before Blasting ..........................................10-4 Table 10-2: Analytical Results for Grab Samples after Blasting .............................................10-4 Table 10-3 : Description of Mini-Bulk Samples ........................................................................10-5 Table 10-4 : 2005 Mini-bulk Sample Locations and Descriptions ............................................10-7 Table 11-1 : List of 1998 Drillholes, Locations and Depths......................................................11-2 Table 14-1 : Statistical Summary- Soluble vs Total Iron for 41 samples..................................14-1 Table 14-2 : Comparison between Iron Assays in hole 72-02 and its Twin 98DN-070............14-2 Table 16-1 : Summary of 1976 Test Results ...........................................................................16-1 Table 16-2 : Composition of Lock-Cycle Composite Samples.................................................16-2 Table 16-3 : Lock-Cycle Test Weight and Iron Recoveries......................................................16-4 Table 16-4 : Jeffrey Magnetic Cobbing Results for Composite A ............................................16-5 Table 16-5 : Davis Tube Results for Composite A at Four Grinds...........................................16-6 Table 16-6 : Chemical Analysis of Bloom Lake Concentrate...................................................16-7 Table 16-7 : JKSimMet Simulation results for Autogenous and Ball Mill Circuits ....................16-8 Table 16-8: Mineralogical Compositions................................................................................16-11 Table 16-9 : Summary of Gravity Separation Results............................................................16-14 Table 16-10: Assumptions to Estimate Overall Recoveries for an Average Bloom Lake Ore16-21 Table 17-1 : Total InSitu Mineral Resources at a cutoff grade of 15% Total Fe ....................17-17 Table 18-1 : Pit Optimization Parameters for LG 3D ...............................................................18-2 Table 18-2 : Cut-off Grade Calculation ....................................................................................18-3 Table 18-3 : Mineral Reserves by Ore Category in Ultimate Pit Design ..................................18-8 Table 18-4 : 20-Year Annual Mine Production Schedule.......................................................18-11 Table 18-5 : Haul and Lift Distances (m) ...............................................................................18-22 Table 18-6 : Major Mining Equipment Fleet ...........................................................................18-23 Table 18-7 : Investment Cost Estimate (US Million Dollars) ..................................................18-54 Table 18-8 : Manpower Requirements...................................................................................18-55

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Table 18-9 : Investment Schedule .........................................................................................18-58 Table 18-10 : IRR and Cumulative Cash Balance .................................................................18-59 Table 18-11 : IRR and NPV @ 5% on Total Investment........................................................18-60

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LIST OF FIGURES

Figure 1-1 : Bloom Lake Iron Ore Project – Sensitivity Analysis (before tax) – IRR%...............1-9 Figure 1-3 : Bloom Lake Iron Ore Project – Sensitivity Analysis (before tax) ..........................1-10 Figure 4-1 : Bloom Lake Property - Location Map .....................................................................4-2 Figure 4-2 : Bloom Lake Property - Location Map .....................................................................4-3 Figure 7-1 : Bloom Lake Property – Location Map ....................................................................7-3 Figure 7-2 : Regional Geology (note that the name of the company was changed from

Consolidated Thompson-Lundmark Gold Mines Ltd. to Consolidated Thompson

Iron Mines Ltd.) ....................................................................................................7-4 Figure 7-3 : Geology Plan Map of the Property .........................................................................7-6 Figure 7-4 : Schematic X-Section through the Deposit ..............................................................7-7 Figure 7-5 : Representative Section: D2 - Looking Northeast ...................................................7-8 Figure 7-6 : Representative Section: K – Looking East .............................................................7-9 Figure 11-1 : Drillhole Location Plan........................................................................................11-5 Figure 13-1 : Routine Sample Analysis Flowsheet ..................................................................13-2 Figure 16-1 : Lock -Cycle Flowsheet .......................................................................................16-3 Figure 16-2 : Fe Grade vs Recovery........................................................................................16-4 Figure 16-3 : Fe Recovery vs %SiO2 ......................................................................................16-5 Figure 16-4 : Weight (wt) Recovery vs Head Grade ................................................................16-9 Figure 16-5 : Weight (wt) and Fe Recoveries vs Head Grade ...............................................16-10 Figure 16-6: Weight and iron recoveries for gravity separation .............................................16-15 Figure 16-7: Weight recovery versus head grade for the four different areas of the Bloom Lake

pit ......................................................................................................................16-16 Figure 16-8 : MgO versus MgO in head sample ....................................................................16-17 Figure 16-9: Deportment of magnetite to tails as a function of magnetite in the ore .............16-18 Figure 16-10 : Magnetite in tails as a function of magnetite in the ore ..................................16-19 Figure 16-11: Concentrate silica content versus magnetite in gravity rougher tails...............16-20 Figure 16-12 : Magnetite Recovery in Cobbing stage as a Function of Magnetite in Rougher

Tails.................................................................................................................16-20 Figure 17-1 : Lithological Block Model .....................................................................................17-3 Figure 17-2 : Total Fe% Histogram..........................................................................................17-5 Figure 17-3 : Total Fe% Cumulative Frequency Plot ...............................................................17-6 Figure 17-4 : Total Fe% Probability Plot ..................................................................................17-7

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Figure 17-5 : Magnetite % Histogram ......................................................................................17-8 Figure 17-6 : CaO% Histogram................................................................................................17-9 Figure 17-7 : MgO% Histogram .............................................................................................17-10 Figure 17-8 : Fe% Grade Block Model...................................................................................17-12 Figure 17-9 : Magnetite Grade Block Model ..........................................................................17-13 Figure 17-10 : Ca0% Grade Block Model ..............................................................................17-14 Figure 17-11 : Mg0% Grade Block Model..............................................................................17-15 Figure 18-1 : Detailed Open-Pit Mine Design ..........................................................................18-4 Figure 18-2 : Pit X-Section @ 614600 E (Looking West).........................................................18-5 Figure 18-3 : Pit X-Section @ 615600 E (Looking West).........................................................18-6 Figure 18-4 : Pit X-Section @ 5855600 N. (Looking North).....................................................18-7 Figure 18-5 : End of Year 1....................................................................................................18-12 Figure 18-6 : End of Year 2....................................................................................................18-13 Figure 18-7 : End of Year 3....................................................................................................18-14 Figure 18-8 : End of Year 4....................................................................................................18-15 Figure 18-9 : End of Year 5....................................................................................................18-16 Figure 18-10 : End of Year 10................................................................................................18-17 Figure 18-11 : End of Year 15................................................................................................18-18 Figure 18-12 : End of Year 20................................................................................................18-19 Figure 18-13 : Process Block Diagram ..................................................................................18-26 Figure 18-14 : Crushing and Ore Stockpiling.........................................................................18-27 Figure 18-15 : Grinding ..........................................................................................................18-28 Figure 18-16 : Spiral Plant .....................................................................................................18-29 Figure 18-17 : Concentrate & Tailings Dewatering, Loadout .................................................18-30 Figure 18-18 : Magnetic Plant................................................................................................18-31 Figure 18-19 : Water Supply ..................................................................................................18-32 Figure 18-20 : Reagent & Services (Flocculant) ....................................................................18-33 Figure 18-21-: Reagent & Services (Air) ................................................................................18-34 Figure 18-22: Reagent & Services (Steam generation) .........................................................18-35 Figure 18-23: Reagent & Services (Fuel oil) ..........................................................................18-36 Figure 18-24 : Ship loading proposed site plan (Phase 1).....................................................18-50 Figure 18-25 : Ship loading proposed site plant (Phase 2)....................................................18-51 Figure 18-26 : Proposed ship loader arrangement ................................................................18-52

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Figure 18-27: IRR Bloom Lake Iron Ore Project - Sensitivity Analysis (before tax) - IRR% ..18-61 Figure 18-28: NPV Bloom Lake Iron Ore Project – Sensitivity Analysis (before tax) .............18-62 Figure 19-1: Schedule for Bloom Lake Project ........................................................................19-1

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LIST OF ABBREVIATIONS

AG autogenous grinding MDDEP Ministère du Développement Durable de

l’Environnement et des Ressources

AFWR adjusted formula weight recovery mm millimetre

APSI Administration portuaire de Sept-Îles

(Sept-Îles Port Authority)

MRN Ministère des Ressources Naturelles

BBA Breton, Banville and Associates NPV net present value

bcm bank cubic metre opex operating expenditure

°C degrees centigrade PAC poly aluminium chlorhydrate

capex capital expenditure PLC programmable logic controller

CCIC Cleveland-Cliffs Iron Company PMF probable maximum flood

CIM Canadian Institute of Mining,

Metallurgy and Petroleum

ppm parts per million

CLM Consolidated Thompson Iron Ore

Mines Limited

psi pounds per square inch

d50 median particle diameter psig pounds per square inch gauge pressure

EIA Environmental Impact Assessment QA/QC quality assurance/quality control

EPCM engineering, procurement and

construction management

QCM Quebec Cartier Mining Company

F80 80% by weight of the particles in the

feed are smaller than this size

QNS&L Quebec North Shore & Labrador Railway

FeR iron recovery Queco Quebec Cobalt and Exploration Ltd.

FWR formula weight recovery RMA Roussy, Michaud and Associates

g gram ROM run of mine

h hour RQD rock quality designation

ha hectare (10,000 m2) s second

H&S Heath and Sherwood SCADA supervisory control and data acquisition

HDPE high density polyethylene SG specific gravity

HEM horizontal loop electromagnetic SMC Sursho Mining Corporation

HMI human machine interface SPI MinnovEX SAG Power Index (in minutes)

hp horsepower S.R. stripping ratio

IF iron formation st short ton

IRR internal rate of return t tonne (metric)

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IVD2 inverse distance squared t/h tonnes per hour

J&L Jones and Laughlin t/m3 tonnes per cubic metre

kg kilogram t/t tonnes per tonne

kg/t kilogram per tonne t/y tonnes per year

km kilometre TFe total iron

kPa kilopascal UPS uninterrupted power supply

kt kilotonne V volt

kV kilovolt VA volt.ampere

kW kilowatt VLF very low frequency

kWh kilowatt-hour VFD variable frequency drive

kWh/t kilowatt-hours per tonne WGM Watts, Griffis & McOuat Limited

L litre WLIMS wet, low-intensity magnetic separator

LG 3D Lerchs-Grossman 3D algorithm WRA whole rock analysis

LOI loss on ignition WtR weight recovery

M million XRD x-ray diffraction

m metre XRF x-ray fluorescence

m3 cubic metre y year

mag magnetite #/h pounds per hour

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1. SUMMARY

1.1 Introduction

Following the completion by Breton, Banville and Associates (BBA) in October 2005 of a Conceptual Study for the development of a 5,000,000 t/y iron ore concentrate mine and concentrator in northern Quebec, BBA was selected by Consolidated Thompson Iron Mines Ltd. (CLM) in November of 2005 to undertake a feasibility study. The feasibility study results were presented in the technical report of May 2006.

In 2007, BBA carried out a feasibility study to expand the Bloom Lake project to 7 million tonnes per year of concentrate and the results were presented in the technical report of May 2007. This report presents the feasibility study results of the Bloom Lake project at a production rate of 8.0 million tonnes per year of concentrate

The present study is based on the resources and reserves published in the May 2006 report. The scope of the present study includes an update of the mining plan, the mine and concentrator infrastructure, the capital and operating costs at 8.0 Mt/y as well as a review of the financial analysis.

1.2 Location

The Bloom Lake property forms part of the south western corner of the Labrador Trough iron range and is located in close proximity to a number of producing mines. The site is approximately 940 km northeast of Montreal and is serviced by road, rail and air.

1.3 Ownership

CLM holds 240 mining claims covering a total area of approximately 10,500 ha in Normanville Township of the Province of Quebec.

1.4 Project Status

The Environmental Impact Study (EIS) was presented to the Ministère du Développement Durable de l’Environnement et des Ressources (MDDEP) at Québec in December 2006. The certificate of authorization was issued by the MDDEP in April 2008. Detailed engineering was advanced in March 2008. Construction is underway with plant start-up scheduled for September 2009. Construction of the magnetic separation plant is scheduled for 2011 with start-up the following year.

Construction of the rail line is subject to authorization from the Newfoundland and Labrador government. An environmental review report was submitted to the Newfoundland and Labrador Government on August 29, 2008 and an approval was obtained on October 27, 2008.

1.5 Geology and Mineral Resources

The Bloom Lake deposit and nearby iron mines such as Mont-Wright, are located within the Grenville Province of the Canadian Shield just south of the Grenville Front. The rocks of the Mont-Wright area form part of the highly folded and metamorphosed south western branch of the Labrador Trough or Labrador-Québec fold belt.


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Iron formations ("IF") and associated rocks within the Grenville are highly metamorphosed and complexly folded and occur as numerous isolated segments. Iron mines and undeveloped deposits in the area include the Mont-Wright, Humphrey and Scully mines, the Bloom Lake and Fire Lake deposits, Lac Jeannine (mined out) and Smallwood (mined out).

IF consists of tabular to folded and anastomosing bands of specularite, magnetite and intercalated actinolite in a dominantly quartz matrix. The sequence also contains narrow bands of amphibolite and quartzite or chert containing very little iron oxide. In the west end of the property the IF consists mostly of specularite-rich sequences with only minor intercalated actinolite. In the east and north parts of the property, actinolite-rich sections are more uniformly interwoven at a fine scale with magnetite and or specularite.

In 1998, Watts, Griffis and McOuat Limited (“WGM”) designed and managed an exploration program carried out on the Bloom Lake property. The program consisted of geological mapping, 18,705 m of diamond drilling, preparation of a "mineral resource" estimate, bulk sampling and metallurgical testing. The "mineral resource" estimate was based on the results of the WGM-managed program and exploration data gathered by previous operators.

Lakefield Research Limited ("Lakefield") conducted all sample analyses and the metallurgical testing to authenticate historic drilling assay results, establish the relationship between Soluble and Total iron (old drill results were in the form of soluble iron while 1998 assays were total iron), and characterize sample mineralogy. All split cores were analyzed for total iron, magnetic iron, sulphur and specific gravities (“SG”) were determined.

In May 2005, WGM reviewed its 1998 "mineral resource" estimate and concluded that the estimate complied with the requirements of the Canadian Securities Administrators’ National Instrument 43-101 (“NI 43-101”) of February 2001 and the standards adopted by the Council of the Canadian Institute of Mining, Metallurgy and Petroleum in August 2000 (“the CIM Standards”) and reclassified it according to the CIM standards.

Based on WGM’s knowledge of the geology and mineralogy of the deposit, the limited metallurgical testwork carried out and the economics of the iron deposits of the area, a 15% Fe cut-off grade was used for the Mineral Resource estimate. Using this 15% Fe cut-off and an inverse distance squared ("IVD2") grade interpolation technique, the total Measured and Indicated block model Mineral Resource for the Bloom Lake deposit on the CLM property was estimated at 638 million tonnes grading 29.76% total Fe and 10.54% magnetite. The block model Mineral Resource is summarized in Table 1-1.

Table 1-1: Total in situ Mineral Resources at a Cutoff Grade of 15% Total Fe Resource Volume Tonnage Average Grades Category bcm* x 1,000 (kt) Total Fe% Magnetite% CaO% MgO%

Measured 141,350 488,465 29.91 10.54 2.32 2.18

Indicated 43,372 149,232 29.29 10.55 2.37 2.15 Total Meas. + Ind. 184,722 637,697 29.76 10.54 2.33 2.17

Inferred 10,322 35,697 30.97 8.47 0.84 0.82 * Bank cubic meter

The Inferred Resource is in addition to the Measured and Indicated Resource.

A site visit was conducted by André Allaire of BBA and Richard Quesnel of CLM in December 2005 for general acquaintance with the project setting and verification of certain drill hole


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locations as well as mini-bulk sample pit locations. André Allaire did not take any independent samples, as the mineralogy and iron content of the deposit are well established. The Bloom Lake deposit has been extensively sampled by past workers during the course of diamond drilling programs and most recently during the mini-bulk sampling program carried out in 2005.

On May 16, 2006, René Scherrer of CLM conducted a site visit of the property for the locations of the mini-bulk samples taken in 2005 under the supervision of WGM. During the visit, the proposed locations of the surface infrastructure including waste dumps, tailings, concentrator and the connecting rail line profile and the port and ship loading facilities in Sept-Îles were also visited.

1.6 Mineral Reserves The planned mining of the Bloom Lake iron ore project is based on the normal practice of open pit operations.

Pit optimization and pit design were carried out to develop the Mineral Reserves from the Mineral Resources. The original resource model was prepared by WGM in its Technical Report filed on SEDAR in May 2005. BBA has accepted the original Mineral Resource estimate and has developed a formula-based weight recovery model in preparation for open-pit optimization. The MedSystem Lerchs-Grossman 3D (“LG 3D”) pit optimization algorithm was used to develop the pit shell at the end of its economic life. In accordance with the regulations governing the preparation of an NI 43-101 report, the pit optimization has used only blocks classified in the Measured and Indicated categories to generate revenue. Other mineralized blocks containing Inferred Resources bear no economic value, regardless of grade and weight recovery and are treated as waste rock, unless proven otherwise by additional geological work.

The pit optimization exercise was based on economic parameters derived from an internal cost model prepared by BBA and a price of US$ 38.94/ tonne of concentrate at 66% Fe in the concentrate and a 0.855 exchange rate. The final pit geometry was adjusted to include haulage ramp, practical wall positions and appropriate pit slope stability as well as benching arrangement.

The Mineral Reserves have been classified in accordance with the requirements of NI 43-101 (as amended in December 2005) and the CIM Standards (as amended in December 2005). Table 1-2 provides a summary of the mine-life mineral reserves for the Bloom Lake Iron Ore Project using a cut-off of 15% Fe.

Table 1-2: Mineral Reserves by Category in Ultimate Pit Design (using a cut-off grade of 15% Fe)

Tonnes WtR(*) TFe MagFe CaO MgO MagClassification

(million) (%) (%) (%) (%) (%) (%)

Proven 463.4 38.3 30.1 7.6 2.2 2.1 10.5

Probable 116.2 37.7 29.7 7.7 2.3 2.1 10.7

Total Ore 579.6 38.2 30.0 7.6 2.3 2.1 10.5

Total Waste

(Waste+Inferred)563.8

Stripping Ratio (t/t) 0.97

(*) For hematite ONLY in block model


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Figures may vary slightly due to rounding effects and the number of decimal points used.

The Mineral Reserves contained in the detailed pit design amount to 463.4 million tonnes in the Proven category and 116.2 million tonnes in the Probable category for a total combined of 579.6 million tonnes at a grade of 30% for total Fe and 38.2% hematite weight recovery, based on a cut-off of 15%. The stripping estimate, including Inferred Resource material, is 563.8 million tonnes for an overall mine-life stripping ratio of 0.97 tonnes of waste per tonne of ore.

Based on the results of the metallurgical testing, the average calculated weight recovery is 38% in hematite and 41% with the magnetite recovery plant, equivalent to a concentration ratio of 2.63 tonnes of ROM per tonne of concentrate in hematite and 2.43 with the magnetite recovery plant.

The Bloom Lake iron ore deposit will be mined using conventional open pit mining methods based on a truck/shovel operation and will be developed to support a nominal capacity of 8 million tonnes of concentrate per annum. The mining equipment will be leased, operated and maintained by the personnel of Bloom Lake. Following drilling and blasting, run-of-mine (ROM) ore will be delivered to a primary crusher located approximately 1,000 m to the north-east of pit edge at about 680 m elevation.

Selective mining and blending of material will be carried out in order to:

• Mine only hematite ore in the first year of operation with magnetite ore in the subsequent years following the commissioning of the magnetite plant;

• Maintain the MgO and CaO levels under 4.5% in the ore;

• Insure that the concentrate contains below 6.0% of silica plus alumina;

• Ensure that concentrate produced contains less than 4.5% SiO2;

• Optimize the use of high weight recovery ore to maximize the concentration ratio for project net present value optimization;

• Feed the mill with ore of constant hardness for a stable process plant operation.

For blending purposes, it is likely that three to four production faces will be developed in ore at any one time on 2 to 3 benches. A 20-year mine production has been prepared on an annual basis in the first 5 years and by 5-year increments in the subsequent years. The average total combined ore and waste movement is fairly constant at approximately 30-31-million tonnes per annum during the first 5 years of operation.

Any inflow water will be collected in sumps and pumped out of the pit using submersible pumps. In addition, two (2) main waste disposal dumps will be constructed during the life of the mining operations. The waste dump capacity has been estimated using a swell factor of 30%.

No allowance has been made for backfilling in mined-out areas, except for deposition of some tailings on the west side of the pit at the end of operation.

The selection of the mine equipment was made using the following criteria:

• An average annual material movement of 30-31-million tonnes in the first 5 years of operation, both run-of-mine and waste combined;

• A truck size of 240 st (218 tonnes);


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• The main loading equipment is the hydraulic shovel with a lifting capacity of 64 tonnes per bucket (ONK 340);

• Availability of a minimum of one (1) front end loader having a lifting capacity of 45 tonnes per bucket (Letourneau Model L-1850 or similar) as support loading equipment. The flexibility of the loader with its fast response time in replacing a shovel or for a particular blending requirement, justifies its use;

• Drills are electric with a capacity to drill 15” diameter blast hole with high pull down pressure (P&H 120-A or similar).

1.7 Metallurgical Testwork

Although field work in the Bloom Lake area began in 1952, the initial metallurgical testwork was undertaken only in 1973-76. At this time Lakefield Research (now SGS Lakefield) produced gravity concentrates on a Wilfley Table using drill core samples crushed to -410 microns in size (35 mesh). In 1998 further metallurgical testing was carried out by Lakefield for Quebec Cartier Mining Company who had an option on the claims at the time. In mid-2005, eleven mini-bulk samples were taken from iron formation outcroppings on the property and sent to SGS Lakefield for testing. These samples were used to confirm and supplement testwork carried out in the 1970s and 1998, and to establish design criteria for the feasibility study.

Based on the results of the metallurgical testing, average weight recoveries of 38% for hematite and 41% with the magnetite recovery plant have been used in the financial evaluation.

The confirmatory test work run in parallel to the engineering activities confirmed the results of the previous metallurgical testing. The 38% weight recovery for hematite concentrate and 41% with the magnetite concentrate are maintained for financial evaluation. The recent test work showed improved overall recoveries.

1.8 Processing

Metallurgical processing is similar to that found at neighbouring mining operations. Ore from the mine will be crushed, stockpiled then fed to a single wet autogenous grinding mill operated in closed circuit with vibrating screens. The liberated hematite and magnetite minerals will be recovered using three stages of spirals. Tailings from the spiral plant will be passed over wet drum magnetic separators to recover unliberated magnetite particles. This material will be reground in a ball mill followed by several more magnetic separator stages to produce a high-grade magnetite concentrate. Concentrates from the spiral and magnetic plants will be filtered and the filter cake loaded into railcars for transportation to the port. The tailings from the magnetic plant will be pumped to a disposal area located several kilometres from the plant.

1.9 Infrastructure and Support Systems

The main service building will be attached to the concentrating plant and will incorporate maintenance shops, warehousing, the steam plant, a compressed air plant, emergency vehicle garage, offices, change rooms, laboratory, a communication room and the security office. Mine vehicles will be serviced in an adjacent garage.

Bloom Lake will provide fresh water for domestic and boiler use. Drinking water will be supplied to the mine site in bottles.


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Reclaim water will be pumped from a barge located on either the settling pond or the polishing pond in the tailings retention area back to the process water reservoir at the concentrator to reduce the demand for fresh make-up water. Reclaim water will be used for equipment cooling as well as for general processing use.

Storage tanks for #2 light fuel oil and gasoline, delivered by road tankers, will be provided adjacent to the concentrator. The 200 000 litres capacity of the #2 light fuel oil tanks represents a minimum on-site storage capacity of 4-day’s consumption in the boiler house at peak load. The boiler house will provide steam for heating and concentrate drying. Number 2 light oil will also be used for heating the crusher building and as a back-up supply for the boiler house.

The major infrastructure required to develop the project is already present in the area. A 12 km power line will connect the mine site with the Hydro-Québec power grid. Road access will be provided by the construction of a 6 km section of unpaved highway to meet Highway 389. Thirty-one kilometres long of railway will also be built to transport concentrate to the existing common carrier railway line for delivery to the port.

Approximately 300 peoples will be employed at the plant site. Accommodation for construction crews and plant operating personnel is provided in a camp constructed in the town of Fermont, 13 km from the mine site.

1.10 Railway & Port Facilities

Concentrate will be loaded into railcars and transported over a new section of track to join the existing common carrier railway. A marshalling yard and a rolling stock repair shop will be constructed near the junction. The frequency will be approximately one full train per day in each direction.

Loaded cars will be dumped at the port of Sept-Îles and the concentrate loaded directly onto the boat or stockpiled and loaded onto boats later for delivery to customers.

1.11 Schedule

The projected length of the design/construction period is 41 months. Basic engineering, environmental studies and metallurgical testwork began in April 2006 with plant start-up scheduled for September 2009. The global environmental certificate was issued in April 2008. Construction of the magnetic separation plant is scheduled for 2011 with start-up the following year.

An Environmental Review Report was submitted to the Newfoundland and Labrador Government on August 29, 2008 and was obtained on October 27, 2008. Construction of the railway can now proceed.

1.12 Environmental and Permitting

The project is subject to an Environmental Impact Assessment (EIA) in the Province of Quebec where the plant will be located and another in Newfoundland & Labrador through which the railway line will pass. No separate federal EIA will be required.

An ongoing program will be implemented to monitor the effects of the Project on water quality, groundwater, effluents, fish population, benthic invertebrate communities, geotechnical matters and sediment quality during both the construction and operational periods.


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A rehabilitation plan will be submitted to the Quebec Ministry of Natural Resources and Wildlife prior to the start of production. Some monetary allocation will be provided in the first eight years of operations for on going rehabilitation.

1.13 Financial Analysis

The financial evaluation for the Bloom Lake project is carried out by the preparation of a discounted cash flow model to which the capital and operating cost estimates as well as the production schedule developed in the mining section are input data. The Internal Rate of Return (IRR) on total investment and the Net Present Value (“NPV”) resulting from the net cash flows generated by the project have been calculated. The payback period is also indicated as a financial measure. The main inputs to the financial analysis are summarized as follows:

• Exchange rate: US $1 = CAN$1.10

• Project timing: production phase of 20 years out of a total mine life reserve of more than 27 years. The construction phase is two (2) years excluding the construction of the magnetite recovery plant in year 3.

• Financing plan: 100% equity

• Income tax: pre-tax and after tax basis (a requirement for a mining operation in the Province of Quebec)

• Capital cost and disbursement schedule is as follows:

2008 - Initial Capex $96.2 M

2009 – Initial Capex $445.9M

2010 – Mine Garage $13.7M

2011 (Magnetite Recovery Plant) $27.8 M

2013 (Leased Mining Equipment

Purchase)

$30.8 M

Working Capital $17.1 M

On-going Capital (Year 2010 –

2030)

$127.7M

• Operating cost: owner operated leased mining equipment, processing, contracted rail transportation and port handling, environment and G/A have been estimated to average US$24.18 per tonne of concentrate for the first five (5) years of operation and US$24.76 over a period of 20 years.

• Concentrate sale price (3-year rolling average for 2006 to 2008): US$1.069 / Fe% or US$71.09/tonne at 66.5% Fe content.

• Working capital: 32 days of operating expenses


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On the basis of the assumptions given above, the IRR and the cumulative cash balance at various discount rates have been computed and the results are summarized in the table below:

100% Equity Funding Before Tax After Tax

IRR % 58.9 48.6 NPV @ 0% ($M) 6 813 4 493 NPV @ 5% ($M) 3 847 2 520 NPV @ 8% ($M) 2 838 1 847 NPV @ 10% ($M) 2 348 1 521 NPV @ 15% ($M) 1 520 966 NPV @ 25% ($M) 708 418

The results of the financial analysis indicate that the Bloom Lake project has an economic potential before-tax IRR of 58.9% and a NPV of US$6 813 M and US$ 3 847 M, at a discount rate of 0% and 5% respectively. On an after-tax basis, the IRR is 48.6% and the NPV amounts to US$4 493 M and US$2 520 M, at a discount rate of 0% and 5% respectively.

The results of the sensitivity analysis of the IRR for iron ore price, capital cost and operating cost and pre-tax NPV @ 5% discount rate are presented in graphical form in Figures 1 and 2, respectively. The analysis shows that the project is mainly sensitive to revenue, that being iron ore price, ore grade and weight recoveries.


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Figure 1-1 : Bloom Lake Iron Ore Project – Sensitivity Analysis (before tax) – IRR%

Bloom Lake- 8 m tpy Concentrate FS Sensivity Analys is (pre-tax) Internal Rate of Return (IRR)

20

30

40

50

60

70

80

90

-20% -15% -10% -5% 0% 5% 10% 15% 20%

Change (%)

IRR

(%)

OPEX CAPEX SALES


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Bloom Lake- 8 m tpy Concentrate PFS Sensitivity Analys is (pre-tax)

Net Present Value (NPV) @ 5% Discount rate

1000

1500

2000

2500

3000

3500

4000

4500

5000

5500

-20% -15% -10% -5% 0% 5% 10% 15% 20%

Change (%)

NPV

(M$)

OPEX CAPEX SALES

Figure 1-2 : Bloom Lake Iron Ore Project – Sensitivity Analysis (before tax) NPV @ 5% Discount Rate

1.14 Recommendations

• Start the environmental restoration studies to replace the lake and river sections affected by the tailings ponds in the summer of 2007 (completed).

• Finalize the agreements with the mining fleet equipment suppliers (completed).

• Finalize the leasing agreements with the railway and port operators (completed).

• Perform exploration drilling to expand the reserves on the west side of the pit (completed).

• Perform definition drilling in the proposed initial mining areas (underway).


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2. INTRODUCTION

2.1 Introduction

Following the completion by Breton, Banville and Associates (BBA) in October 2005 of a Conceptual Study for the development of a 5,000,000 t/y iron ore concentrate mine and concentrator in northern Quebec, BBA was selected by Consolidated Thompson Iron Mines Ltd. (CLM) in November of 2005 to undertake a feasibility study. The feasibility study results were presented in the technical report of May 2006.

In 2007, BBA carried out a feasibility study to expand the Bloom Lake project to 7 million tonnes per year of concentrate and the results were presented in the technical report of May 2007. This report presents the feasibility study results of the Bloom Lake project at a production rate of 8.0 million tonnes per year of concentrate

2.2 Purpose of the Technical Report

The objective of the study was the evaluation of the technical feasibility and economic viability for a period of twenty years at 8 Mtpy. The study included the review and validation of all pertinent existing data, including ongoing studies such as laboratory testing, estimation of the Mineral Reserves, mine planning, infrastructure and service facilities, environmental evaluation, investment and operating costs along with the financial analysis.

2.3 Basis of the Technical Report

The Feasibility Study was carried out using information contained in, but not limited to, the following reports and documents:

• Bloom Lake Conceptual Study, November 2005.

• A report titled “A Technical Review and Mineral Resource Estimate for the Bloom Lake Iron Deposit, Labrador Trough, Québec” prepared by WGM in compliance with the standards of NI 43-101.

• Report prepared by Journeaux, Bédard & Associates inc on the feasibility and cost evaluation to contain and manage the tailings and process water for 20 years of exploitation of the Bloom Lake property.

• A review prepared by Geostat International Inc of the geological block model.

• An environmental evaluation prepared by Roche Ltd which outlined mitigation measures, a restoration plan and costs, and identified the various certificates of authorization required.

• An environmental impact study prepared by Genivar Inc. was submitted to the MDDEP in December 2006.

• Lakefield/SGS Lakefield testwork results from 1975-1976 and 2005-2007.

• Process flowsheets of similar existing operations.

• Commercially available database and cost models.


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• Budgetary prices from equipment suppliers.

• Feasibility Study for 8 Mtpy Iron Concentrate – Bloom Lake Project, Prepared by BBA inc. for Consolidated Thompson, September 2008.

2.4 Qualified Persons

The qualified persons responsible for the preparation of this report are: André Allaire, Eng, M. Eng, PhD; Patrice Live, Eng; Enzo Palumbo, M. Eng; René Scherrer, Eng.

2.5 Site Visit

A site visit was conducted by André Allaire of BBA and Richard Quesnel of CLM in December 2005 for general acquaintance with the project setting and verification of certain drill hole locations as well as mini-bulk sample pit locations. André Allaire did not take any independent samples, as the mineralogy and iron content of the deposit are well established. The Bloom Lake deposit has been extensively sampled by past workers during the course of diamond drilling programs and most recently during the mini-bulk sampling program carried out in 2005.

On May 16, 2006, René Scherrer of CLM conducted a site visit of the property for the locations of the mini-bulk samples taken in 2005 under the supervision of WGM. During the visit, the proposed locations of the surface infrastructure including waste dumps, tailings, concentrator and the connecting rail line profile and the port and ship loading facilities in Sept-Îles were also visited.

This report is based on information from the Bloom Lake Feasibility Study for 8-million tonnes per year iron concentrate and was prepared by a team of Qualified Persons from BBA Inc. and Consolidated Thompson Iron Mines Limited following the guidelines of the Canadian Securities Administrators’ National Instrument 43-101 and Form 43-101F1 and in conformity with generally accepted CIM “Exploration Best Practices” and ‘Estimation of Mineral Resources and Mineral Reserves Best Practices” Guidelines.

In the later half of 2006 and the beginning of 2007, several visits to the site, under the direction of René Scherrer, were conducted for the purpose of carrying out a drilling program and a collection of bulk samples for confirmatory testwork and to obtain geotechnical data (for plant location and soil samples).

In the summer of 2008, a drilling campaign was underway on the west side of the property, under the direction of René Scherrer. Results will be presented at a later date by Consolidated Thompson Iron Mines in a separate report.


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3. RELIANCE ON OTHER EXPERTS

The authors have compiled this report using information contained in the Bloom Lake Iron Ore Project Feasibility Study dated September 2008, consultants’ reports, and other documents supporting the Feasibility Study.

The authors of the present report have not carried out a thorough review of each consultant’s work. The information provided to BBA was supplied by reputable consultants and BBA has no reason to doubt the validity of the information.


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4. PROPERTY DESCRIPTION AND LOCATION

The Bloom Lake property is located in the province of Quebec in Normanville Township, Kaniapiskau County. The centre of the property is at approximately latitude 52° 50’ 30” North, longitude 67° 17’ West. It is part of the south western corner of the Labrador Trough iron range. The property is located about 940 km northeast of Montreal, 8 km north of Quebec Cartier Mining Company’s (“QCM”) Mont-Wright concentrator and 13 km northwest of the Town of Fermont. The municipalities of Wabush and Labrador City are located 30 km to the east of the property. The general location of the mine is shown in Figure 4-1.

CLM holds 240 contiguous map designated mining claims in the Bloom Lake area covering a total surface area of approximately 10,500 ha as shown in Figure 4-2. Sixteen of the claims date from 1951 and 1952 and the remainder were acquired by map designation during 2005, at which time the original claims were also converted to map designated claims. There are sufficient assessment credits to hold a large portion of the claims for several years.

The claims convey no surface rights, which are the property of the Crown.

The list of mining claims, along with their expiration dates, is given in Appendix A.


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Figure 4-1 : Bloom Lake Property - Location Map


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Figure 4-2 : Bloom Lake Property - Location Map


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5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Access

The Bloom Lake deposit lies 4 km north of Highway 389 between Mont-Wright and the Town of Fermont, which connects with Highway 138 at the Town of Baie-Comeau on the north shore of the St. Lawrence River, a distance of 565 km between Fermont and Baie-Comeau. Highway 138 connects Baie-Comeau to Montréal. The Wabush Airport, 30 km from the proposed mine site, provides frequent flights to Montreal and the Island of Newfoundland. Currently, access is provided by the construction of a 6 km section of unpaved highway to meet Highway 389.

5.2 Climate

The area has a sub-Arctic climate with temperatures ranging from -40°C in winter to 25°C in summer. The average annual temperature is -3.6°C and the average total precipitation is 880 mm. The prevailing winds are from the west and have an average speed of 14 km per hour, based on 30 years of records at the Wabush Airport.

5.3 Local Resources and Infrastructure

The town of Fermont has a population of approximately 4,000 and is the residential town for QCM employees who work at the Mont-Wright operations. The town was originally built by QCM, which still owns some of the units in town. Fermont has schools, a 72-room hotel, municipal and recreational facilities, and a business and shopping complex.

There is only a very small labour pool in the area.

There is grid electrical power available nearby, a plentiful supply of water and ample surface area available for constructing mining and processing infrastructure, for placing waste piles and tailings impoundment areas, although CLM holds no surface rights as yet.

5.4 Physiography

Topography is gently rolling with the occasional more precipitous area. The general elevation is 600 metres above sea level and relief is about 100 m. The property is poorly drained and covered by typical northern boreal forest consisting of often stunted spruce, alders and willows. Hill tops are generally barren and the property has extensive swampy areas.


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6. HISTORY

This section of the report deals with work carried out prior to the feasibility study of April 2006, with the exclusion of metallurgical testwork. Metallurgical test programs are reviewed in section “Mineral Processing and Metallurgical Testing”

In 1951, James and Michael Walsh staked claims for Mr. Bill Crawford of Sursho Mining Corporation (“SMC”) following the discovery of a cobalt showing at Bloom Lake. In February 1952, Quebec Cobalt and Exploration Limited (Queco) was incorporated to acquire the claims held by SMC. Quebec Cobalt and Exploration Limited initially optioned 1 million shares from SMC and gradually accumulated more shares until full control was obtained in February 1956. In 1952, Thompson Lundmark Gold Mines, Arnold Hoffman, and Thayer Lindsay advanced money to Queco for its field program

Today, CLM has 100% ownership of the claims and property formerly held by Queco.

In 1952, a crew of six prospectors initiated a program to prospect an area that included the Bloom Lake property. The purpose of the program was to sample the cobalt showing and to prospect in the Mt-Wright-Carheil Lake area. The results of the cobalt showing work were disappointing but several zones of magnetite-hematite iron formation between Bloom Lake and Pignac Lake were identified and sampled. Further work was done in 1953, but little is known about this work.

In 1954, further work was conducted to investigate and explore the iron occurrences. The 1954 program consisted of line cutting, geological mapping and magnetometer surveys.

In 1955, Jones and Laughlin (“J&L”) started the first of several field programs after obtaining a lease from Queco. Cleveland-Cliffs Iron Company (“CCIC”) joined J&L on an equal participation. The lease was terminated in 1968.

In 1971, Queco sponsored a program consisting of line cutting, geological mapping, gravity and magnetometer surveys, and diamond drilling that continued into 1972. The program was managed by H.E. Neal & Associated Ltd.

In 1973, Republic Steel Corporation optioned the property and H.E. Neal & Associates Ltd prepared a “Preliminary Evaluation” of the property consisting of currently held property and claims further to the west. This work was carried out until 1976. The evaluation included “mineral reserve” estimates, a metallurgical test program and preliminary mine design. The mine design included pit outline, dump area, access roads, and railway spur. Dames and Moore prepared the mine design and “reserve” estimates. Lakefield Research carried out the metallurgical testwork.


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7. GEOLOGICAL SETTING

7.1 Regional Geology

The rocks of the Mont-Wright area form part of the highly folded and metamorphosed south-western branch of the Labrador Trough or Labrador-Québec fold belt. The Labrador Trough extends along the margins of the eastern boundary of the Superior-Ungava craton for more than 1,200 km and is up to 75 km wide. The Bloom Lake deposit and the Mont-Wright mine are located within the Grenville Province of the Canadian Shield just south of the Grenville Front. The Grenville Front, the northern limit of the Grenville Province, truncates the Labrador Trough; separating the Churchill Province greenschist metamorphic grade Trough rocks from the highly metamorphosed and folded Grenville Province Trough rocks.

The Labrador Trough consists of a succession of folded Proterozoic sedimentary and volcanic rocks and mafic intrusions deposited in interconnected sub-basins. These Aphebian successions, characteristic of the Labrador Trough, are known as the Kaniapiskau Supergroup and include the economically important iron formations. The Kaniapiskau Supergroup is divided into the Knob Lake Group and the Doublett Group. The Kaniapiskau Supergroup unconformably overlies Archean gneisses and is stratigraphically overlain by the Helikian Shabogamo Group. Table 7-1 summarizes the regional stratigraphic column.

IFs and associated rocks within the Grenville are highly metamorphosed and complexly folded and occur as numerous isolated segments. Iron mines and undeveloped deposits within the Grenville include in addition to QCM's Mont-Wright mine and the Bloom Lake deposit, Lac Jeannine (mined out), Fire Lake, Smallwood (mined out), and the Humphrey and Scully mines (Figure 7-1).

Figure 7-2 shows the geology of the Bloom Lake deposit and surrounding area modified after Cunningham, 1954; Gross, 1968; Alexander, 1975 and QCM, 1998 program results.


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Table 7-1 : Regional Stratigraphic Column

Stratigraphic Column

PROTEROZOIC Helikian

Shabogamo Group

Gabbro, Diabase

-------------------------------------Intrusive Contact--------------------------------------------------------------

PROTEROZOIC Aphebian

Kaniapiskau S

Churchill Province Grenville Province

(Low-Grade Metamorphism) Knob Lake Group

(High-Grade Metamorphism)

Menihek Formation Nault Formation Black shale, siltstone

Graphitic, chloritic and micaceous schist

Sokoman Formation Cherty iron formation

Wabush Formation Quartz magnetite-specularite- carbonate iron formation

Wishart Formation Quartzite, siltstone

Carol Formation Quartzite, quartz-muscovite-garnet schist

Denault Formation Dolomite, calcareous

siltstone

Duley Formation Meta-Dolomite and calcite marble

Attikamagen Formation Gray shale, siltstone

Katsao Formation Quartz-biotite-feldspar schist and gneiss

------------------------------------ --------Unconformity--------- -----------------------------------

ARCHEAN Ashuanipi Complex

Granitic and granodioritic gneiss, mafic intrusives

Note: The Duley, Carol and Wabush Formations are included in the Gagnon Group.


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Figure 7-1 : Bloom Lake Property – Location Map


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Figure 7-2 : Regional Geology (note that the name of the company was changed from Consolidated Thompson-Lundmark Gold Mines Ltd.

to Consolidated Thompson Iron Mines Ltd.)

Graphics by WGM


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7.2 Local Geology

The geology and geological interpretation for the Bloom Lake property and the Bloom Lake deposit are based on data from a number of sources. These sources include the diamond drilling and mapping done on the property as part of the 1998 program, as well as the drilling conducted in 1956, 1957, 1967, 1971 and 1972. The geological interpretation relies heavily on the mapping programs conducted in 1952 and the ground magnetic surveys carried out in 1967 and 1971/72 as compiled in 1973.

Figure 7-3 is a geology plan map of the property. Figure 7-4 is a schematic cross section through the stratigraphy illustrating patterns in the lithological sequence and consequently lithogeochemical patterns and signatures. Figures 7-5 and 7-6 are typical drill cross sections through the deposit. One section (D2) cuts through the deposit northeast of Pignac Lake and the other section (K) cuts through the main zone of IF west of Pignac Lake.


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Figure 7-3 : Geology Plan Map of the Property


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Figure 7-4 : Schematic X-Section through the Deposit

Graphics by WGM


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Figure 7-5 : Representative Section: D2 - Looking Northeast

Graphics by WGM


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Figure 7-6 : Representative Section: K – Looking East

Graphics by WGM


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8. DEPOSIT TYPES

The Bloom Lake deposit is composed of iron formations of the Lake Superior-type. Lake Superior–type iron formation consists of banded sedimentary rocks composed principally of bands of iron oxides, magnetite and hematite within quartz (chert)-rich rock, with variable amounts of silicate, carbonate and sulphide lithofacies. Such iron formations have been the principal sources of iron throughout the world (Gross, 1996).

Lithofacies that are not highly metamorphosed or altered by weathering are referred to as taconite. Strongly metamorphosed taconites are known as meta-taconite or itabirite. The iron deposits in the Grenville part of the Labrador Trough in the vicinity of Fermont and Wabush are meta-taconite. The Bloom Lake and Mont-Wright deposits are examples.

A number of models have been considered for the origin of iron formation and associated lithofacies.

According to Gross (1996) the two principal genetic but controversial models are:

1. Volcanogenic and hydrothermal effusive or exhalative; and

2. Hydrogenous-sedimentary with derivation of the iron, silica and other constituents by deep weathering of a landmass.

Gross reports that iron-oxidizing micro-organisms might have played a role. Oolites are generally common in iron formation.


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9. MINERALIZATION

9.1 Lithology and Mineralization

Lithological coding and rock type nomenclature used for the Bloom Lake deposit is based on the coding used at QCM’s Mont-Wright mine. WGM used this coding, with slight modification, for logging the 1998 drillholes and applied it to the previous drillholes based on the interpretation of the older drill logs. Table 9-1 is a summary of the rock type codes and rock type names used. More detail and description follows for the main rock types.

Several rock type codes are hybrids of codes for principal rock types and are not described separately.

Iron Formations (IF) on the property are of the Lake Superior type (Gross, 1996).

Table 9-1 : Summary – Rock Type Codes for Descriptive Logs

Code Rock type AMP Amphibolite OIF Oxide Facies Iron Formation SIF Silicate Iron Formation MS Mica Schist (60-100% mica) MSIF Mica Schist with more than 15% iron oxide QR Quartzite (0-40% mica) QRIF Quartzite with less or equal to 15% iron oxide QRMS Quartz Rock Mica Schist (between 40 and 60 % mica) GN Gneiss

Amphibolite ("AMP") is dominantly a competent, dark green to black medium to coarse grained rock consisting mainly of hornblende, biotite and feldspar. A mass of amphibolite covering a large portion of the claims northeast of Pignac Lake is exposed at surface. Many of the drillholes drilled on the NE baseline encounter a considerable thickness of amphibolite, the "amphibolite cap" up to 200 metres thick, before penetrating into IF structurally below the amphibolite. The amphibolite-IF contacts are sharp although the whole rock analysis (“WRA”) results indicate that IF immediately adjacent to the contact is often contaminated by aluminium, suggesting either metasomatic alteration of the IF or microscopic threads or inclusions of amphibolite in the IF. A narrow argillized zone of amphibolite often occurs immediately above the IF contact. Often this argillization can be seen to be recent with water presently running along this contact. In addition to the large massive body of amphibolite described above, narrower intercepts of amphibolite also occur throughout the IF sequence. These amphibolite units tend to occur towards the bottom of the IF sequence (see schematic cross section Figure 7-4). In Table 9-2, the means and medians for total iron, magnetic iron, sulphur, SG and the other whole rock oxides are listed for each of the major rock types. One hundred and seven samples of amphibolite have a mean SG of 3.20, and a median SG of 3.18.


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Table 9-2 : Summary of Means and Medians

Oxide Iron Formation ("OIF") comprises sequences rich in specular hematite (specularite) and magnetite. Such OIF sequences are commonly 25% to 40% iron oxide (remainder mostly quartz) and can have thicknesses (ignoring minor intercalated bands of amphibolite and quartz rock) of up to 200 m. It is these sequences that are of economic importance. Intervals of core up to a couple of meters long can comprise near-massive magnetite but commonly, like specularite, magnetite most often occurs in anastomosing to discontinuous stringers and bands less than 10 cm thick in a quartz rock or actinolite-quartz rock matrix. Bands tend to be folded and deformed but also can be regular and tabular. The actinolite-IF contains about 15% magnetite on average, higher than the other rock types except magnetite-OIF.

Specularite is silvery-grey and non-magnetic. It tends to be coarser grained than magnetite and like magnetite often occurs in bands. Magnetite-rich OIF ("MOIF") commonly, but not always, occurs beneath (structurally) the main mass of amphibolite, i.e., Sections C2, D, D2 or immediately under a thin interval of grunerite-SIF (Figure 7-5). This sequence of magnetite-rich OIF is often about 30 m to 50 m thick. Thereafter, down sequence, the amount of magnetite decreases as the amount of specularite increases until specularite is predominant with little or no magnetite present. The amphibolite cap is missing from the IF sequence west of Pignac Lake and the magnetite-rich OIF sequence typically under the amphibolite cap is also missing. Magnetite also often occurs in intervals bordering discrete units of lean OIF or quartz rock, and usually is prominent in actinolite-SIF adjacent to the narrow intercalated amphibolite units or bands.

Specularite-rich OIF ("SOIF") dominates the OIF sequence between the amphibolite cap related magnetite OIF sequence and the sequence characterized by actinolite-SIF, magnetite (±


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amphibolite, pyrite, garnets, and chalcopyrite). In the IF branches north, east and northeast of Pignac Lake, this actinolite-SIF sequence (as described above) is also characterized by the occurrence of narrow intervals of amphibolite. In the IF branch to the west of Pignac Lake, the narrow amphibolite bands are missing.

This specularite-rich OIF sequence is commonly up to 100 m to 150 m thick but also thins as the entire sequence narrows on the edges of the deposit as shown in holes 98DN-005 on Section D3 and 98DN-073 on Section G4.

As shown on Table 9-2, the mean SG of OIF is about 3.47 regardless of whether it is magnetite or specularite-rich.

Silicate Iron-Formation ("SIF") - Two main types have been recognized on the property. One of these is dominated by actinolite ("AIF" in Table 9-2), while in the other grunerite is most prevalent. Both types are coded as SIF in the logs but the two are generally distinguished on the cross sections and level plans. The two types can be transitional into one another and likely there is also some tremolite-rich SIF present.

During core logging, intervals were defined as SIF partly on the basis of colour change. The IF becomes distinctly greener with increased amounts of actinolite. Actinolite-SIF often also shows an increase in calcite content.

The WRA (whole rock analysis) results show that intervals with actinolite-SIF have increased levels of MgO as compared to more oxide-rich IF. Work undertaken by Lakefield has shown that the actinolite content is roughly proportional to MgO levels with actinolite contents about equal to MgO values multiplied by a constant of 4.5 to 4.7 WGM used an MgO level of 5%, which roughly translates to 23% to 25% contained actinolite by weight, as a visual guide for defining actinolite-SIF and distinguishing it from OIF. It is important to note that all rock with >5% MgO, >15% total Fe is not SIF since amphibolite also contains more than 5% MgO. SIF can contain significant amounts of oxide iron.

In the IF units west of Pignac Lake the demarcation between OIF and actinolite-SIF is often clearer than in the IF northeast of Pignac Lake. In the IFs to the west of Pignac Lake there are distinct anastomosing to continuous bands of actinolite-SIF that often occur towards the bottom of the IF sequence adjacent to the basal QRIF or QR unit (Figure 7-4). The OIF in this area is often much "cleaner" with very low levels of MgO and associated CaO as compared with the IF coded as OIF northeast of Pignac Lake. Northeast of Pignac Lake the OIF contains generally higher, and more erratically distributed levels of MgO and associated CaO making the demarcation of distinct bands of SIF less clear and consequently the correlation and continuity of distinct bands of SIF less certain.

Actinolite-SIF also tends to occur on the borders of the thinner amphibolite units towards the lower part of the stratigraphic sequence. The IF in these areas is also often enriched in magnetite as compared with specularite. These SIF intervals often containing multiple thin units of amphibolite, sometimes sheared and segmented, can also be enriched in sulphide. Such sulphides consist mostly of pyrite sometimes in coarse, up to 1 cm crystals but also minor chalcopyrite. Actinolite-SIF also preferentially occurs on IF, QRIF or QR contacts. In such places it again is more likely associated with increased prevalence of magnetite at the expense of specularite.

Quartz rock ("QR") and its related variant Quartz Rock Iron formation ("QRIF"), are used to describe various units that occur throughout the IF sequence and a unit that occurs at the structural base of the IF immediately above the basal mica schist. During logging, QR was used to define a rock type consisting mostly of quartz, 95%+, with minimal to nil specularite and/or


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magnetite content. This material may have been derived from chert, quartzite or quartz pebble conglomerate and the various textural varieties are not distinctly coded or distinguished.

QRIF intervals were defined on the basis of a quartz dominant rock containing less than 15% total iron but containing some iron in the form of specularite and/or magnetite or silicate. QRIF is therefore a rock often transitional between OIF and QR, or SIF and QR. The QRIF can contain significant actinolite-SIF with MgO contents above 5% but is not coded as SIF because total iron is less than 15%. The QR and QRIF units above the basal schist can often contain significant muscovite, garnet and feldspar and are transitional into the basal schist or gneiss unit.

Quartz Rock Mica Schist ("QRMS") is used mainly for the schist sequence at the base of the IF sequence beneath the QR unit. QRMS has occasionally, however, been used for coding biotite-rich units within the IF sequence, which are likely genetically related to AMP. These QRMS rocks are described under AMP.

Gneiss ("GN") - The basal QRMS rocks appear to be transitional into GN. Hole 98DN-006 on Section K and hole 98DN-004 on Section J terminate in gneiss. These GN rocks contain less mica but more feldspar and quartz than QRMS.

The basal QRMS sequence consists mostly of muscovite and biotite schist with characteristic porphyroblasts of garnet and feldspar.

9.2 Structural Geology

In gross terms, the Bloom Lake deposit comprises two doubly plunging synforms on a main east-west axis separated by a gently north to northwest plunging antiform. One of these synforms is centred on Triangle Lake, while the centre for the other is located just north of Bloom Lake. The Bloom Lake property is centred primarily on the eastern synform but covers a portion of the northern limb of the western synform.

These synforms are the result of a minimum of two episodes of folding and are of regional scale.

In addition to these regional scale folds, which have created the deposit scale synforms shaping Bloom Lake deposit, there are several other folds of diverse orientation on the property. It is not clear if all folding directions represent distinct folding episodes or progressive change in fold orientation with time.

As a result of various folding events or transposition of stratigraphy by folding, local reversals and repeats in the stratigraphic sequence in the IF are common. The lensitic, crenulated and stringer nature of the IF as observed in drill core with anastomosing discontinuous interwoven bands of magnetite and/or specularite and actinolite is largely a result of this complex folding.

There appears, however, to be no wholesale duplication, repeating or truncation of the IF sequence resulting from isoclinal folding within the drill tested area. One exception may be along the northwesterly edge of the northeast branch of the deposit (northwestern end of Section D - lower parts of drillholes 98DN-007 and 98DN-070) where interference between two fold sets might be contributing to complex patterns.

Faulting appears to have played a much less significant role than folding on the property. Cunningham's crews mapped a northeast trending fault that displaces IF west of the west edge of Pignac Lake. Gross (1968) shows a series of northeast and northwest trending faults, as well as a series of stratigraphy-parallel faults. There has probably been some faulting but in correlating IF from hole to hole and section to section, no major displacements due to faulting have been


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defined. Areas of poor ground encountered by drilling west of Pignac Lake may be due to faulting or simply slumping and weathering. As stated previously, occasionally a weathered zone was encountered between the amphibolite cap and IF suggesting movement of water along this contact.

In accordance with this outline, the gross masses of IF are interpreted to be generally continuous from one drill section to the next. However, the continuity and correlation of thin, individual units of OIF, SIF, amphibolite or quartz rock can be poor and difficult due to isoclinal folding, refolding, transposition and segmentation.


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10. EXPLORATION

10.1 1998 Program

10.1.1 General

In 1998, a major exploration program was carried out by WGM for QCM, which then held the Bloom Lake property under option from CLM. QCM held the option on the property until 2001, but no work was carried out between 1998 and 2005.

The 1998 program included line cutting, surveying, road building, camp construction, diamond drilling, geological mapping, mini-bulk sampling, bench-scale preliminary metallurgical testwork, preparation of a “mineral resource” estimate, camp demobilization and site clean-up. The WGM-managed work was carried out on 16 of the 19 original claims of the Bloom Lake deposit group only. Portions of the program, such as the “mineral resource” estimate and metallurgical testwork (carried out by Lakefield) are discussed in detail in separate sections of this report.

The drilling was carried out in three phases and QCM had the option of terminating the program at the end of each phase. Phase I drilling commenced on May 14. Phase I continued into Phase II without pause. The Phase I/II drilling was completed on July 14 with the completion of 36 holes. After a scheduled break, Phase III drilling commenced on August 12 and terminated on September 25. Seventy-five drillholes totalling 18,705 m were completed during the entire program.

All sample analyses was conducted by Lakefield. A total of 2,529 samples of split core was analysed over the duration of the program. Laboratory work included routine sample analysis, heavy liquid testing (Phase I/II), mineralogical studies and liberation testing. In addition, further testing was carried out to authenticate old drilling results, establish weight recoveries and the applicability of Mont-Wright procedures for determining weight recoveries for Bloom Lake property samples.

10.1.2 Line Cutting and Surveying 10.1.2.1 General

To provide control for locating drillhole collars and drilling directions, a grid was re-established over a portion of the property. Old drillhole collars located in the field and new collars were surveyed once they were drilled. Roussy, Michaud and Associés ("RMA") of Sept-Îles, Québec, was contracted to cut the grid, survey the baselines and survey drillhole collars.

Minor line cutting was done during the program by the WGM geological and geotechnical crew to extend lines and spot drillholes not located on lines cut by RMA.

10.1.2.2 Grid Surveying and Cutting

Three baselines (main east-west, E-W BL; northeast, NE BL; and southeast, SE BL) were laid out to enable drilling on cross section lines perpendicular to the strike of three main branches of the iron deposit. The NE BL and SE BL were subsidiary lines to the main east-west baseline and were turned off from the main east-west baseline. The new grid was to a substantial degree coincident with the original (1952) property grid.


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A total of 12.4 km of grid lines was cut in 1998, consisting of 4.0 km of baseline and 8.4 km of cross section lines.

The baselines were picketed at 50 m intervals using 2 by 2 wooden pickets. Intermediate 25 m stations and the cross lines were picketed using 1 by 2s. The 50 m and 100 m baseline pickets were labelled with aluminium tape.

The baselines were surveyed providing X, Y and Z co-ordinates for stations at 50 m intervals. A Total Station instrument was used to control line position and station location. Cross lines were turned off with the Total Station but beyond a 25 m distance from their respective baseline were not surveyed-in but sighted-in, by visually aligning pickets. Section lines were labelled relative to origin point on the baseline. The drill cross section line labels are rounded off to the closest meter.

Surveying was generally restricted to hole collars, baselines and the section-baseline intersections. Elevations for constructing the topographic profile on the cross sections were derived from an aerial photographic survey and topographic model produced by Western Photogrammetry Ltd. in 1973. This same topographic model was used for the Mineral Resource estimate. Close agreement between the surveyed hole collar elevations and the topographic model derived from the 1973 survey indicates that the topographic model is valid for the Mineral Resource estimate. Hole collar elevations between the 1973 survey and the 1998 survey are generally within ±2 m.

10.1.3 Mapping and Bulk Sampling

10.1.3.1 General

Two areas on the property, known as the north hill and the west hill were selected for re-mapping. These areas are situated on the hilltops forming two topographic highs which occur respectively towards the north and west ends of the property. The north hill area is centred at UTM (NAD 83) co-ordinates 616100mE/5855900mN while the west hill area is centred at 614900mE/5855300mN.

Geological mapping was necessary to obtain a more accurate picture of surface geology to complement detailed diamond drilling results so that a cohesive geological interpretation for the property could be completed and as a preliminary step in selecting bulk sample sites.

The north and west hills were selected because the two main zones of IF on the property were exposed on these hills. The IF from the north hill area is distinctly different from the IF exposed on the west hill. The NE branch of the deposit tends to be more uniformly actinolite-rich than the IF from the west branch of the deposit, which is more actinolite-poor and specularite-rich.

There is also good outcrop exposure on both hills and thus these two areas represented good sites for bulk sampling IF outcrop. Re-mapping of these areas was also important because of the uncertainty of outcrop locations. Outcrop and drillhole locations as shown on the 1975 geological compilation map of the property did not agree with the initial 1998 land survey results.

The detailed mapping is incorporated into the geological plan for the property (Figure 7.3)


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In 2005, eleven sample sites were selected for bulk sampling on the basis of range of rock types, magnetite and actinolite content. Nine sites were sampled in the north hill area and two samples were taken from the west hill area.

10.1.3.2 Mapping

Control for mapping was established by turning off intermediate grid cross lines, at 25 m intervals, from the NE or E-W surveyed baseline between the existing surveyed drill cross section lines. Station markers were established at 25 m intervals along these grid lines. Outcrop locations were determined by chaining from the nearest grid line.

Outcrops were assigned a unique outcrop number for reference purposes and for locating samples. The prefix OC1 designated outcrops from the west hill area while the prefix OC2 was used for north hill outcrops.

Due to magnetic deviation caused by magnetite IF, strike measurements shown on the map were measured from reference lines, established by stretching a chain over the subject outcrop between known points on the grid.

10.1.3.3 Sampling

Three sets of samples were collected from the north and west hill areas. These samples consisted of two sets of grab samples and one set of mini-bulk samples. One set of grab samples was taken prior to blasting for the mini-bulk samples and the second set of grab samples was taken subsequent to blasting. The grab samples consisted of 5 kg to 10 kg of material from each site. These samples were sent to the Mont-Wright mine laboratory for assay. The analytical results are listed in Tables 10-1 and 10-2.

A total of nine mini-bulk samples was taken from the north hill area and one sample was taken from the west hill area. Each of these samples consisted of 500 kg to 750 kg of material. A gasoline powered plugger was used to drill three to seven holes about 0.5 m deep at each sample site. The holes were subsequently loaded and blasted. Castonguay Frères Ltée of Rock Forest, Québec was contacted to do the blasting. The blasting was all completed in one day.

The blasted rock was hand loaded into steel drums, which were then labelled and sealed. The drums were carried to the highway and delivered to the QCM warehouse for shipment. They were sent to Lakefield but there was no testwork undertaken.

In Table 10-3, location and geological data for the mini-bulk sample sites are summarized.


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Table 10-1 : Analytical Results for Grab Samples before Blasting

Site Sample Outcrop Fe Total

wt%

SiO2

wt%

TiO2

wt%

Mn

wt%

P

wt%

K2O

wt% AL2O3

wt%

MgO

wt%

CaO

wt%

Na2O4

wt%

Fe3O4

wt%

S

wt%

Lot 1 199838963 OC2-150 44.29 40.38 0.01 0.040 0.007 0.016 0.37 0.084 0.078 0.038 5.7 0.005 Lot 2 199838959 OC2-143 29.55 54.18 0.01 0.079 0.033 0.017 0.41 2.307 1.188 0.068 2.5 <0.001 Lot 2 199838965 OC2-144 43.87 39.63 0.02 0.049 0.007 0.015 0.37 0.350 0.448 0.036 10.0 0.01 Lot 3 199838961 OC2-4 45.63 37.64 0.03 0.034 -0.004 0.016 0.39 0.736 0.275 0.048 0.9 0.016 Lot 4 199838960 OC2-10 40.47 40.99 0.02 0.061 0.027 0.018 0.41 1.927 1.388 0.072 12.5 0.006 Lot 5 199838966 OC2-63 23.23 65.29 0.01 0.050 0.006 0.016 0.36 0.75 1 0.684 0.061 1.0 <0.005 Lot 6 199838964 OC2-73 38.25 41.70 0.02 0.228 0.017 0.017 0.41 2.872 2.044 0.058 43.1 0.015 Lot 7 199838968 OC2-43 23.14 59.38 0.01 0.169 0.037 0.020 0.47 3.798 2.298 0.111 2.0 <0.001 Lot 8 199838962 OC2-173 41.94 41.68 0.01 0.055 0.015 0.019 0.40 0.788 0.5 16 0.061 7.5 0.004 Lot 9 199838967 OC2-1 30.28 56.47 0.01 0.055 -0.003 0.016 0.36 0.554 0.274 0.043 5.5 <0.003 Lot 10 199838958 OC1-2 39.95 44.41 0.01 0.019 -0.005 0.015 0.38 0.075 0.03 1 0.035 2.8 0.003 Average 36.42 47.43 0.01 0.08 0.01 0.02 0.39 1.29 0.84 0.06 8.50 0.01

Table 10-2: Analytical Results for Grab Samples after Blasting

Site Sample Outcrop Fe Total

wt%

SiO2

wt%

TiO2

wt%

Mn

wt%

P

wt%

K2O

wt%

Al2O3

wt%

MgO

wt%

CaO

wt%

Na2O

wt%

Fe3O4

wt%

S

wt% Lot 1 199839887 OC2-150 28.13 56.54 0.01 0.102 0.004 0.015 0.36 0.255 1.659 0.039 3.8 <0.001Lot 2 199839883 OC2-143 39.24 41.52 0.02 0.109 0.015 0.016 0.41 3.039 1.203 0.064 5.0 0.003Lot 3 199839892 OC2-4 32.56 53.13 0.01 0.045 0.013 0.016 0.39 0.167 0.426 0.037 7.5 <0.002Lot 4 199839888 OC2-10 36.30 44.44 0.01 0.092 0.02 1 0.017 0.40 0.772 2.589 0.050 15.2 0.005Lot 5 199839884 OC2-63 22.52 59.36 0.01 0.200 0.022 0.018 0.41 3.349 2.995 0.110 2.5 <0.001Lot 6 199839889 OC2-73 29.80 52.42 0.01 0.212 0.011 0.018 0.39 2.632 1.737 0.056 31.3 <0.002Lot 7 199839890 OC2-43 21.11 62.10 0.02 0.154 0.02 1 0.020 0.49 1.949 3.111 0.090 3.1 0 Lot 8 199839891 OC2-173 31.45 50.45 0.01 0.125 0.03 1 0.022 0.44 2.084 2.060 0.095 7.5 0 Lot 9 199839885 OC2-1 33.87 50.96 0.01 0.069 0.006 0.017 0.37 0.532 0.63 1 0.046 3.2 0 Lot 10 199839886 OC1-1 26.89 61.49 0.01 0.010 0.007 0.015 0.36 0.137 0.060 0.038 1.1 <0.006Lot 10 199839893 OC1-2 27.19 60.96 0.01 0.012 <0.003 0.014 0.36 0.092 0.055 0.042 1.2 <0.003 Average 29.91 53.94 0.01 0.10 0.01 0.02 0.40 1.36 1.50 0.06 7.40 0.00


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Table 10-3 : Description of Mini-Bulk Samples


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10.2 QCM Airborne Geophysical Survey

In the summer of 1998 and independent of the WGM-managed exploration program, QCM carried out a regional high-resolution helicopter-borne aeromagnetic/Horizontal Loop Electromagnetic ("HEM”)/Very Low Frequency ("VLF") survey over an area of approximately 4,100 km2. This survey was roughly centered on the Mont-Wright operation and the 19 original CLM claims (which are part of the actual 178 claims) were included within the northwest quarter of the survey area. The survey was flown by High-Sense Geophysics Limited of Toronto. Line direction and spacing were NW-SE and 200 m respectively, and mean terrain clearance was 30 to 45 m for the geophysical birds and 60 m for the helicopter.

The magnetic survey outlined the Bloom Lake iron formation, confirming its folded nature and relatively large aerial extent. A bulls-eye magnetic response was recorded over the three-claim (Roach Lake Extension) group. Neither the HEM nor the VLF survey recorded anomalies over either of the CLM claim groups.

10.3 2005 Program

10.3.1 General

In 2005, CLM retained WGM to carry out a technical review, including the preparation of a Mineral Resource estimate for the Bloom Lake iron deposit to assist CLM in making business decisions and future planning. The technical review was carried out and prepared in compliance with the standards of NI 43-101 in terms of structure and content and the Mineral Resource estimate was carried out in accordance with provisions of NI 43-101 guidelines and the CIM Standards.

WGM also assisted CLM in consolidating its land position as discussed in Section 4 above.

In addition, WGM planned and carried out a mini-bulk sampling program in order as discussed below.

10.3.2 Mini-bulk Sampling Program

Eleven mini-bulk samples were drilled, blasted and collected on the North and West Hill areas between July 26 and 28, 2005. The access trail from Highway 389 had been rehabilitated and the sample sites marked out earlier in the month. Sample sites were chosen to match as closely as possible those of the similar program carried out for QCM in 1998.

Nine of the sample locations were on the North Hill and two samples were located on the West Hill, as described in Table 10-4 and shown on Figure 7-3. All samples were taken in the same locations sampled in 1998 and designated with the same “lot number” as that used in 1998, with the exception of Lots 10 and 11. In 1998 the Lot 10 sample was taken from two locations. In 2005 the two locations where the 1998 Lot 10 was taken were sampled separately and identified as Lot 10 and Lot 11 respectively. Each sample consisted of approximately 500 kg of material. RSM Mining Services of Labrador City prepared the sites, gathered and bagged the samples and transported them to the highway.

Castonguay Frères Ltée of Labrador City conducted the drilling and blasting using approximately 6-7 blast holes at each sample location. Blast holes were drilled by a compressed air powered plugger. Twenty “rice bags” of blasted material were collected at each sample location. Each bag weighed approximately 25 kg for a total sample weight of 500 kg each. Serialized sample


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tags were enclosed in each “rice bag” and these were then sealed with plastic cable ties. Each sample, consisting 20 rice bags, was then loaded into a separate one cubic metre cargo bag. The cargo bags were hauled approximately 4 kilometres to the paved road on a skidder, which had been adapted to carry the samples. At the road the cargo bags were placed individually on eleven pallets, wrapped in plastic, loaded into a transport truck and shipped to SGS Lakefield in Lakefield, Ontario.

The entire program, including loading the samples onto the transport truck, was supervised and directed in the field by Peter Legein, P.Geo., a WGM Senior Associate Geologist.

The results of the metallurgical testwork, which continued into early 2006 are discussed in Section 16.

Table 10-4 : 2005 Mini-bulk Sample Locations and Descriptions

Lot# Area

Location

Grid Location UTM

Easting

UTM

Northing

Mineralogy

1 North Hill L8+23mNE 0+38mSE 616214 5856000 35% specularite; 2% Mg, 0% actinolite

2 North Hill L7+75mNE 0+75mSE 616219 5855941 15% specularite, 10% Mg, 15% actinolite



5 North Hill L6+12mNE 0+50mSE

616112 5855815

30% specularite, 0% Mg, actinolite, 7%

epidote


7 North Hill L7 + 12mNE 0+llmSE 616132 5855923 20-25% specularite, 2-5% Mg, 15% actinolite



10 West Hill 17+35mW 0+75mS 614873 5855229 30% specularite, 0% Mg, 0% actinolite

11 West Hill L17+30mW 0+25mS 614883 5855275 30% specularite, 0% Mg, 0% actinolite


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11. DRILLING

11.1 General

The 1998 drill program was divided into Phases I, II and III. QCM reserved the option to terminate work after each phase. The program was designed to drill on cross sections at 150 m intervals with drillholes generally spaced at 150 m along the lines.

Heath & Sherwood Drilling (1986) Inc. ("H&S") was the principal drill contractor. H&S supplied one H&S HS-35A computerized hydraulic diamond drill and one JKS-300 diamond drill with BQ and NQ rod strings, casings, hexagonal core barrels (stabilized) and all ancillary supplies. Both drills were unitized and skid mounted. Most drilling was done with BQ sized equipment, but where drilling problems were encountered or anticipated due to friable rock, NQ sized holes were collared.

A total of 18,705 m in 75 holes was completed in three phases of drilling. Table 11.1 provides a summary of pertinent drilling statistics. Figure 11.1 shows the drillhole locations by drilling phase and the cross section lines, A through N (except for sections E and F which fall outside the drilled area). Core was delivered to camp after each shift by the drillers.

Drilling took place on a two shift per day basis, 22 hours per day, 7 days per week. The remaining 2 hours per day were taken up with crew shift changes.

H&S was also responsible for establishing the camp located at the northwest corner of Pignac Lake (same location as the 1956, 1971 and 1972 camps), managing and servicing the camp and providing room and board for drillers, WGM geological and geotechnical crew and technical subcontractors, as required.

Core recovery was generally very good, 98% to 100% for most of the drilling, however, in certain areas due to extensive alteration, IF was extremely friable. Recovery in these intervals was low despite NQ size coring, extensive use of mud, use of 5 foot core barrels, slow watchful advance and face discharge bits. In the geotechnical logs core recoveries were tabulated by run.


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Table 11-1 : List of 1998 Drillholes, Locations and Depths

Hole Baseline Section Co-ord. 1 Co-ord. 2 Claim Core Size Start Com Collar Az Collar Inclin Length (m)

Phase I/II - May 14 to July 4, 1998 98DN-001 E-W G-1 L3+00m W 0+15m S 51229-4 BQ 14-May 17-May 210 -45 263.0098DN-002 E-W I L7+92.5m W 0+40m N 60065-5 BQ 15-May 18-May 180 -75 329.0098DN-003 E-W H L4+88m W 0+75m N 51229-4 BQ 17-May 21-May 0 -90 322.0098DN-004 E-W J L10+36m W 0+10m N 60065-5 BQ 18-May 21-May 180 -60 290.3098DN-005 NE D3 L3+35m NE 1+10m SE 5 1229-4 BQ 22-May 24-May 302 -45 262.0098DN-006 E-W K L12+50m W 0+25m S 60064-5 BQ 21-May 24-May 180 -75 279.0098DN-007 NE D L5+43m NE 2+50m SE 5 1229-3 BQ 24-May 28-May 302 -45 362.0098DN-008 E-W L L15+24m W 0+00m N 60064-5 BQ 24-May 25-May 0 -90 64.0098DN-08A E-W L L15+24m W 0+05m S 60064-5 NQ 25-May 27-May 0 -90 55.8498DN-009 E-W M L17+07m W 0+17m S 62390-2 NQ, BQ 27-May 30-May 360 -75 237.0098DN-010 NE C L8+48m NE 1+70m SE 51229-2 BQ 28-May 1-Jun 302 -45 326.0098DN-011 E-W N L20+12m W 1+85m S 62390-2 NQ,BQ 31-May 2-Jun 360 -90 221.0098DN-012 NE C-2 L6+95m NE 2+07m SE 51229-3 BQ 1-Jun 4-Jun 302 -45 326.0098DN-013 E-W M-2 L18+56m W 1+35m N 62390-2 NQ,BQ 2-Jun 3-Jun 0 -90 176.8098DN-014 E-W M L17+07m W 3+75m S 60066-1 BQ 4-Jun 6-Jun 360 -60 210.0098DN-015 NE B-2 L9+96m NE 1+79m NW 60065-3 BQ 5-Jun 9-Jun 122 -45 378.0098DN-016 E-W L-2 L16+00m W 1+00m S 60066-1 NQ 6-Jun 9-Jun 360 -50 225.5098DN-017 E-W L L15+24m W 2+80m S 60066-1 BQ 9-Jun 11-Jun 360 -70 191.0098DN-01 8 NE D L5+43m NE 5+65m SE 51229-4 BQ 9-Jun 25-Sep 302 -45 539.0098DN-019 E-W K-2 L13+86m W 3+60m S 60066-1 BQ 11-Jun 12-Jun 360 -45 164.0098DN-020 E-W K-2 L13+85m W 0+50m S 60064-5 NQ 12-Jun 14-Jun 360 -90 154.8098DN-021 NE D-2 L4+39m NE 4+00m SE 51229-4 BQ 14-Jun 18-Jun 302 -59 398.0098DN-022 E-W K-2 L13+85m W 0+50m S 60064-5 NQ, BQ 15-Jun 17-Jun 360 -45 183.0098DN-023 E-W J L10+36m W 0+10m N 60065-5 NQ, BQ 17-Jun 20-Jun 360 -90 254.0098DN-024 E-W G-1 3+00m W 0+10m S 51229-4 BQ 18-Jun 19-Jun 360 -90 351.0098DN-025 E-W I-2 L9+27m W 0+05m N 60065-5 BQ 20-Jun 22-Jun 360 -90 267.0098DN-026 E-W H-2 6+41m W 0+25m N 51229-4 BQ 21-Jun 23-Jun 360 -90 298.00


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(continued)

Hole Baseline Section Co-ord. 1 Co-ord. 2 Claim Core Size Start Com Collar Az Collar Inclin Length (m)Phase I/II - May 14 to July 4, 1998 98DN-027 NE D-5 L0+35m NE 0+30m NW 60065-5 BQ 22-Jun 25-Jun 360 -55 271.5098DN-028 E-W H-2 L6+40m W 0+25m N 51229-4 BQ 23-Jun 25-Jun 179 -45 230.0098DN-029 E-W H-2 L6+40m W 1+50m N 51229-4 BQ 22-Jun 26-Jun 360 -90 333.0098DN-030 NE D-4 L1+90 m NE 1+30m SE 51229-4 BQ 25-Jun 27-Jun 322 -70 348.0098DN-031 NE D-2 L4+39m NE 1+85m S 51229-4 BQ 27-Jun 30-Jun 302 -50 317.0098DN-032 NE D-3 L3+35m N 1+12m SE 51229-4 BQ 28-Jun 30-Jun 360 -90 356.0098DN-033 E-W J-2 L11+44m W 0+12m N 60065-5 BQ 30-Jun 2-Jul 360 -90 195.1098DN-034 SE G-3 L3+60m SE 0+20m NE 51225-4 NQ 30-Jun 2-Jul 250 -75 183.0098DN-035 E-W L-2 L15+96m W 2+40m S 60066-1 BQ 3-Jul 4-Jul 359 -80 181.00Subtotal 36 holes, 52 elapsed days, 9,542 m, 46 m average per drill per shift

Hole Baseline Section Co-ord. 1 Co-ord. 2 Claim Core Size Start Com Collar Az Collar Inclin Length (m)Phase III - August 12 to September 25, 1998

98DN-036 E-W H L4+88mW 0+25m N 51229-4 BQ 12-Aug 16-Aug 180 -75 254.0098DN-037 SE G-2 L2+45m SE 1+60m N 51225-4 BQ 12-Aug 15-Aug 250 -75 324.0098DN-038 NE D-3 L3+35m NE 6+00m SE 51229-4 BQ 15-Aug 17-Aug 360 -90 270.0098DN-039 E-W H L4+88m W 0+75m N 51229-4 BQ 16-Aug 19-Aug 360 -65 401.0098DN-040 E-W I L7+92m W 0+05m N 60065-5 BQ 17-Aug 18-Aug 360 -50 104.0098DN-041 E-W I L7+92m W 0+05m N 60065-5 BQ 18-Aug 20-Aug 180 -50 257.0098DN-042 NE D-3 L3+35m NE 3+70m SE 51229-4 BQ 20-Aug 23-Aug 0 -90 369.0098DN-043 NE D-3 L3+35m NE 0+10m SE 60065-5 BQ 20-Aug 21-Aug 301 -45 110.0098DN-044 NE D-2 L4+39m NE 0+80m SE 51229-3 BQ 21-Aug 22-Aug 302 -45 180.0098DN-045 NE C-2 L6+95m NE 0+70m SE 51229-3 BQ 22-Aug 25-Aug 302 -45 287.0098DN-046 NE D-2 L4+39m NE 3+00m SE 51229-4 BQ 23-Aug 26-Aug 302 -50 366.0098DN-047 NE B-2 L 9+96m NE 0+50m NW 60065-3 BQ 25-Aug 26-Aug 122 -45 246.0098DN-048 NE C-2 L 6+95m NE 2+55m SE 51229-3 BQ 26-Aug 29-Aug 360 -90 302.0098DN-049 NE D-2 L4+39m NE 4+55m SE 51229-4 BQ 27-Aug 30-Aug 302 -70 384.0098DN-050 NE C-2 L 6+95m NE 2+35m SE 51229-3 BQ 29-Aug 2-Sep 302 -65 402.0098DN-051 NE D-2, G-1 L4+39m NE 4+75m SE 51229-4 BQ 30-Aug 1-Sep 0 -90 303.0098DN-052 E-W K L12+51m W 0+25m S 60064-5 NQ 1-Sep 4-Sep 0 -78 188.28


November 2008

11-4

(continued)

Hole Baseline Section Co-ord. 1 Co-ord. 2 Claim Core Size Start Com Collar Az Collar Inclin Length (m)Phase III

98DN-053 NE C-2 L6+95m NE 2+90m SE 51229-3 BQ 2-Sep 4-Sep 122 -60 252.00 98DN-054 E-W M-2 L18+57m W 0+02m N 62390-1 BQ 4-Sep 5-Sep 360 -45 137.40 98DN-055 E-W L 15+00m W 0+25m S 60064-5 NQ, BQ 4-Sep 7-Sep 180 -75 185.00 98DN-056 E-W M-2 L18+57m W 0+02m N 62390-1 BQ 5-Sep 7-Sep 360 -90 212.00 98DN-057 E-W L 15+00m W 0+25m S 60064-5 NQ 7-Sep 8-Sep 360 -45 153.41 98DN-058 E-W M-2 L18+57m W 1+40m S 62390-2 BQ 7-Sep 11-Sep 360 -55 324.00 98DN-059 E-W J-2 L11+44m W 0+25m S 60065-5 NQ 8-Sep 13-Sep 0 -55 218.24 98DN-060 E-W M-2 L18+57 W 2+15m S 62390-2 BQ 11-Sep 12-Sep 360 -90 117.00 98DN-061 E-W M-2 L18+57m W 2+70m S 62390-2 BQ 12-Sep 13-Sep 360 -50 150.00 98DN-062 E-W M L17+07m W 2+20m S 62390-2 BQ 13-Sep 14-Sep 360 -60 170.00 98DN-063 E-W J-2 L11+44m W 0+25m S 60065-5 NQ 13-Sep 15-Sep 180 -55 227.38 98DN-064 E-W L L15+24m W 2+45m S 60066-1 BQ 14-Sep 15-Sep 360 -45 110.00 98DN-065 E-W I-2 L9+14m W 0+30m S 60065-5 BQ 15-Sep 16-Sep 180 -45 97.00 98DN-065A E-W I-2 L9+14m W 0+33m S 60065-5 NQ 16-Sep 19-Sep 180 -45 220.68 98DN-066 E-W K L12+51m W 4+25m S 60066-3 BQ 15-Sep 16-Sep 360 -45 96.00 98DN-067 SE G-4 L4+73m SE 0+44m NE 51225-4 BQ 16-Sep 17-Sep 360 -90 147.00 98DN-068 SE G-5 L5+75m SE 0+25m NE 51225-4 BQ 17-Sep 18-Sep 250 -75 105.00 98DN-069 SE G-3 L3+59m SE 0+77m N 51225-4 BQ 18-Sep 19-Sep 360 -90 194.00 98DN-070 NE D L5+61m NE 4+35m SE 51229-3 BQ 19-Sep 23-Sep 303 -46 491.00 98DN-071 SE G-2 L2+45m SE 2+30m NE 51225-4 BQ 19-Sep 22-Sep 360 -90 329.00 98DN-072 SE G-3 L3+60m SE 2+30m NE 51225-4 BQ 22-Sep 24-Sep 249 -75 264.00 98DN-073 SE G-4 L4+75m SE 2+05m NE 51225-4 BQ 24-Sep 25-Sep 249 -75 216.00 Subtotal 39 holes, 45 days, 9,163 m, 50.9 m per shift per drill

Total Phase I, II and III 75 holes, 97 elapsed days, 18,705 m, 48 m per shift per drill

Note: 1. Hole 98DN-01 8 was initially collared in Phase I. It was re-entered and deepened in Phase III.

2. Holes 98DN-055 and 57 are not actually located on Section L but are located about 22 m east of section L.


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Figure 11-1 : Drillhole Location Plan

Graphics by WGM


November 2008

11-6

11.1.1 Phase I/II

The program completed July 4, comprising holes 98DN-001 to 35, was designated the Phase I/II drill program. The WGM geological and geotechnical crew demobilized from the property July 12 after completing all Phase I/II core logging, core sampling and sample shipping.

11.1.2 Phase III

The Phase III program commenced August 12 with essentially the same drilling, geological and geotechnical crew as had been engaged for Phases I and II. Phase III drilling continued to September 26 when the last hole of the program, hole 98DN-073, was completed.

Core logging, sampling, drill demobilization and camp clean-up extended to October 2. By this date, all drilling company personnel, surveyors and WGM geological and geotechnical crew had left the area.

11.2 Hole Collar and Down-Hole Attitude Surveys

Drillhole collars were spotted prior to drilling by chaining in the collar locations from the nearest line picket. Drilling azimuths were established by lining up the drill by eye on the cut grid line. Drill inclinations were established using a Brunton compass on the drill head. Downhole inclinations were tested by the drillers using acid tests. Acid tests were taken every 50 m to 75 m down each hole as designated by the drill geologist. RMA was not involved in this inclination determination. Once the hole was finished the drill geologist was responsible for placing a marker in the collar of each hole and labelling it with an aluminium tag. Subsequently the X, Y and Z co-ordinates for these collar markers were surveyed by RMA. Where the holes contained casing, these casings were surveyed by RMA for azimuth. Holes without casing were assumed to have azimuths the same as originally designated at the time of drill set-up. Stabilized core barrels were used for all drilling and deflections in azimuth and inclination are believed to be minimal and acceptable considering the scale of the deposit and the drilling density with holes collared on nominal 150 m centers.

The grid line co-ordinates with respect to the appropriate baseline are shown on the drill logs. The logs also show the NAD 83 co-ordinates for each of the holes. All previous drill collars (1956, 1967, 1971, 1972) that were located were also surveyed. For collars not located, the collar co-ordinates were calculated using old grid co-ordinates as recorded on the old logs. WGM has confidence in these calculated collar locations because the drillholes located in the field were generally found close to their recorded position.

11.3 Core Handling and Logging Protocols

11.3.1 Descriptive Logging

Descriptive logs include a description of the various geological features noted during examination of core including lithological and structural description. Lithological coding was generally according to guidelines supplied by QCM and followed at its Mont-Wright operation. Lithological coding and general rock type descriptions for the various rock units encountered are compiled in Section 9.1.

Descriptive logs also record hole locations and all down hole survey data, all sample locations, iron and sulphur assays, SG and magnetite determinations for each sample.


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11-7

The Log II computer program of Log II Systems Inc., Toronto, Ontario was used for descriptive logging.

11.3.2 Geotechnical Logging

The geotechnical logging followed QCM guidelines. Logging was done on a drill run by run basis, generally at 3 m intervals. The geotechnical logging is based on a digital code system with the code representative of the drill run as a whole.

Geotechnical data recorded included core recovery as a percentage of recovered core per run; Rock Quality Designation ("RQD"), a measure of fracture density; type of foliation, i.e., thin bands, wide bands, undulating, anastomosing, discontinuous or folded; foliation angle with respect to core axis; friability; colour and granulometry index with values from 1 to 3 with 1 representing aphanitic grains, 2 representing medium grains and 3 representing grains greater than 2 mm in size. A summary of codes is in the footer to each of the geotechnical logs.

11.3.3 Video Recording

A video record of all core was made as part of the logging process. An 8 mm video camera was used and the core was photographed tray by tray. Audio highlights were added as appropriate. Hole number and meterage blocks are clearly legible in the record. The video was subsequently converted to VHS format

11.3.4 Magnetic Susceptibility

As part of the logging routine, magnetic susceptibility measurements were taken on core from all of the drillholes as an approximation of the magnetite content. A Mico Kappa KT-5 instrument distributed by Scintrex was used for the determinations. The unit has a sensitivity of 1 by 10-5 SI units. Readings were taken every metre (approximate position) on the core's un-split surface down each of the holes. Readings were taken by the geotechnical logger. The magnetic susceptibility measurements are listed with the geotechnical log. The readings listed are "apparent" susceptibility, uncorrected for core size.

11.3.5 Core Storage

After core logging and sampling were completed, core trays containing the reference half core and the un-split parts of the holes were stacked in core racks. No core was discarded. All core was sent to the Mont-Wright mine site for storage is racks after drilling was completed.

WGM visited the Mont-Wright mine and the core rack area in August 2005. All the 1998 core remains intact, in racks and is easily identifiable.


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12. SAMPLING METHODS AND APPROACH

The demarcation of sampling boundaries was made generally on a geological basis as selected by the drill geologist during logging. The sampling guidelines required minimum sample lengths of 3 m, with maximum lengths of 4.5 m for NQ core and 7 m for BQ core. This scheme was closely followed for the most part. In total, 2,529 samples were taken from the 75 drillholes. No samples were taken from holes 98DN-040 and 98DN-053 because all rock intersected was amphibolite. The average sample length was close to 6 m as most of the drilling was BQ and most samples ranged in weight from 5 kg to 12 kg.

All rock estimated to contain at least 10% iron in the form of oxide was sampled. In addition, one sample on either side of all IF was taken in wall rock to bracket all IF sequences.

Some samples include more than one rock type because of the 3 m minimum sample length stipulation. This situation most often occurs in sequences of IF alternating with narrow, less than 1 m intervals of amphibolite. In such cases, individual samples may include both amphibolite and IF.

Sample intervals were marked out by the drill geologist with a china marker during descriptive logging. Three-part sample tickets, from sample books with unique sequential ticket numbers, were used to number and label core and subsequently sampled intervals and bagged samples. Sample numbers and intervals were recorded for later computerization and collation in a sample database. A second record of sample number and interval was saved with the descriptive log.

One portion of the sample ticket remained in the sample book. One portion was placed under the last few centimetres of core in the sample interval in the core tray. The third portion was inserted by the core splitter/sampler into a 40 x 60 cm canvas sample bag with the split core sample. The sample number was also written on the sample bag.

Core samples were split using a hydraulic core splitter. The second half of the spit core sample was returned to the core tray. These reference split core segments were put back into the core trays in the original order and fitted back together as much as possible. Core trays were then stacked on timber core racks located on site.

Sample bags were stored in the splitting tent until removed for transport to the highway and then to Mont-Wright for loading onto transport trucks and shipment to Lakefield Research Limited ("Lakefield"), Lakefield, Ontario.

Sample shipments from camp to Mont-Wright were nominally twice per week. Sample bags were loaded onto H&S tracked vehicles (GT-1,000s) and transported to the highway; a 1 to 1.5 hour trip. At the highway the sample bags were loaded into 25 gallon steel drums. The sample bags were as much as possible loaded into the drums in numerical sequence. Each drum held from 10 to 15 sample bags depending on sample and drill core size, i.e. BQ or NQ. Each of the sample shipments, consisting usually of 4 to 8 drums, was assigned a sequential sample shipment number and each drum was assigned a consecutive (all-shipments-inclusive) drum number. Shipment numbers and drum numbers were recorded in the sample database. A list of samples comprising the sample shipment was placed into the first drum in each shipment.

Sample drums pre-loaded onto a pick-up truck were then sealed and transported to the Mont-Wright warehouse where they would await next-day transport to Lakefield. The entire sample shipment procedure from camp to the Mont-Wright mine was carried out by the sampler/splitter assisted by either a geotechnician or a geologist. Once delivered to the warehouse the samples were in the hands of Mont-Wright mine warehouse staff.


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13. SAMPLE PREPARATION, ANALYSES AND SECURITY

13.1 Sample Preparation

All split core samples were sent to Lakefield (now part of the SGS group). Each sample was weighed, then jaw crushed to 6 mesh and riffle split. A 2 kg split was then passed through a roll crusher to 100% minus 20 mesh. A 150 g split was then pulverized to minus 200 mesh. A 1.5 g sub-sample was used for WRA, a 5 g sub-sample for determining magnetic iron and a 30 g sub-sample was used for SG determinations. These various sub-samples were then subjected to analysis as described below.

13.2 Assaying

All split core samples were analyzed for a suite of whole rock elements including: SiO2, TiO2, Al2O3, Fe2O3, MnO, MgO, CaO, Na2O, K2O, P2O5 and loss on ignition ("LOI"). Analysis was done on lithium tetraborate fused pressed pellets by X-ray Fluorescence ("XRF") following sample crushing and pulverization. Additional analysis included determining magnetite by the Satmagan method, and total sulphur by infrared detection analysis using LECO instrumentation following sample combustion. SG for each sample was determined using an air comparison pycnometer. It should be noted that this method does not take into account existing porosity in a rock and some of the IF does contain vugs due to calcite removal. Although the degree of porosity has not been quantified, it is estimated on the basis of visual examination of drill core to be generally less than 2%.

Total iron was calculated from Fe2O3 by dividing total iron expressed as Fe2O3 by a factor of 1.4295. A flow sheet (Figure 13-1) provides details for the individual procedural steps.


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Figure 13-1 : Routine Sample Analysis Flowsheet

Graphics by WGM


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13.3 QA/QC

In 1998, Lakefield was accredited by the Standards Council of Canada under ISO Guide 25. The routine quality control program at Lakefield was modeled after guidelines provided by the International Standards Organization, the Ontario Ministry of Environment and Energy, Environment Canada and the Canadian Association of Environmental Analytical Laboratories and included the processing of method blanks, replicate samples and standard reference materials.

Quality control for the routine sample analysis included both Lakefield's own quality control procedures involving both internal and external checks. Approximately 8% of laboratory throughput was quality control material, 4% blanks, 2% duplicates and 2% reference material (standards). These procedures and the QA/QC results are described in the WGM Technical Report, filed on SEDAR in May 2005, under Internal Checks Section.

Limited checking that was undertaken independently of Lakefield is described in the WGM Technical Report, filed in SEDAR in May 2005, under External Checks Section.

The Internal Check work showed that precision was excellent for total iron, magnetite and SG. Duplicate and replicate assays were generally within 1 to 2% of original determination.

External Checks indicated that some inter-laboratory bias was present with respect to total iron and SG, however, only a few samples were used to assess repeatability of assays between laboratories.

13.4 Security

The chain of custody of samples has been described previously. There is no reason to believe that samples or analytical results were tampered with.


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14. DATA VERIFICATION

A limited amount of analytical work was undertaken in 1998 to investigate certain problems related to authenticating, combining and using the analytical data from the old drillholes for the Mineral Resource estimate.

For the 1998 program, as described previously, drillhole samples were analyzed for total iron by XRF and magnetite was determined by Satmagan. Total iron was not determined for either the QC or 71/72 series drillhole samples. The QC and 71/72 series drillhole samples were analyzed for soluble iron rather than total iron and this soluble iron was likely determined using the hydrochloric acid-stannous chloride procedure. Furthermore, magnetite in the QC holes (expressed as magnetic iron) was determined by Davis Tube, while for the 71/72 and 1998 drillholes, magnetite was determined by Satmagan. In addition, although QC samples were likely analyzed at the Jones & Laughlin laboratory in Negaunee, Michigan under a quality control situation, neither the analytical and quality control procedures, nor the laboratory actually responsible for the analysis are documented in reports on-hand. It is known that the 71/72 drillhole samples were analyzed at Lakefield.

Analytical work was undertaken to investigate the relationship between soluble and total iron analysis as part of the 1998 program. It was believed that this research was required before analytical data for the QC, 71/72 and the 1998 holes were combined for the purpose of preparing the Mineral Resource model.

To this end two investigations were undertaken:

1. Forty-one samples were selected from the 1998 holes and analyzed for both soluble iron and total iron so that the differences between soluble iron and total iron could be assessed for a group of samples containing varying amounts of iron silicates.

2. A 1972 drillhole was twinned. Hole 98DN-070 was drilled as close as possible to hole 72-02. The result was a string of samples for which new assays of total iron and magnetite were obtained. These assays could then be compared against 1972 assays for soluble and magnetite for a closely comparable rock sequence.

Soluble Iron versus Total Iron

A statistical summary for the analytical results is provided in Table 14-1. Appropriately the total iron results are slightly higher than results for soluble iron as shown by the mean value, and median values. Soluble iron and total iron generally increase in a linear fashion.

Table 14-1 : Statistical Summary- Soluble vs Total Iron for 41 samples

Total Fe Soluble Fe Total Fe Soluble Fe Mean (wt%) 34.21 32.70 Skewness 1.77 -0.15 Median (wt%) 33.36 33.20 Range (wt%) 26.80 38.80 Mode (wt%) 31.34 33.70 Minimum (wt%) 25.25 12.10 Standard Deviation(wt%) 5.06 6.08 Maximum (wt%) 52.05 50.90 Sample Variance (wt%) 25.63 36.95 Sum (wt%) 1,402. 1,340.50 Kurtosis 4.64 4.66 Count 41 41


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The sample with the largest difference between its total iron and soluble iron concentrations is sample 100984. This sample was found to contain 76.3% grunerite during X-ray Diffraction Analysis ("XRD") (Section 15.3.2). Although there is some correlation between the magnitude of the difference between total and soluble iron and MgO content, the degree of correlation is poor and not entirely explained by the actinolite content, as reflected by its MgO concentration.

Drillhole Twinning

In Table 14-2, a statistical summary for total, soluble and magnetite results is presented. Hole 72-02 was original assayed for soluble iron. The 1998 hole was analyzed for total iron. Magnetite in both cases was by Satmagan.

Table 14-2 : Comparison between Iron Assays in hole 72-02 and its Twin 98DN-070

Hole 98DN-070 Hole 72-02

Total Fe Magnetite Soluble Fe Magnetite

Mean 29.6 19.68 28.8 22.25

Standard Error 1.0 2.43 1.0 2.01

Median 29.7 11.72 30.4 16.24

Standard Deviation 7.2 36.62 10.5 0.69

Sample Variance 51.8 335.64

110.4 412.94

Kurtosis 0.4 -1.2 0.4 -1.1

Skewness -0.8 0.6 -0.5 0.5

Range 33.3 53.4 51.9 69.8

Minimum 8.4 0.98 1.9 0.41

Maximum 41.7 54.4 53.8 70.2

Sum 1,689.4 1121.8

2,935.0 2269.7

Count 57 57 102 102

Note: The table only includes assays for that part of hole 98DN-070 from 88.65 m

to 393.8 m

Best efforts were made to drill hole 98DN-070, as close as possible to hole 72-02 and to select the interval in hole 98DN-070 that represented the same lithological section as intersected in hole 72-02. The section from 88.65 m to 393.80 m in hole 98DN-070 was selected. WGM does not believe that the correlation between the two holes is of sufficient quality to allow for a direct comparison of the two holes on a sample by sample basis.


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The statistical analysis shows that the results for soluble iron are slightly lower than those for total iron, fitting with the ideal that soluble iron is a partial analysis. The average for magnetite in hole 72-02 is slightly higher than the average for magnetite in hole 98DN-070 but the medians show the reverse pattern. The distribution of magnetite results is log-normal and accordingly the median should be a better estimate of central tendency for magnetite results than the arithmetic average. The total iron assays have a normal distribution.

WGM concludes that the assays for total iron for the old holes appear to be unbiased quality data and that the use of unadjusted soluble iron assays for the QC and 71/72 series holes in the Mineral Resource estimate is appropriate and if anything probably slightly undervalues the grade of the deposit in terms of total iron content. The magnetite results, for the 71/72 holes, may be a little less reliable and a bit too high but there are not many of these holes and their volume influence for the deposit is relatively small.


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15. ADJACENT PROPERTIES

The Bloom Lake property is located in a very active mining area, which includes the following mining operations:

• Mont-Wright Mine, Quebec Cartier Mining (QCM) – 13.5 M tpy iron ore concentrate (commercial operations started in 1977)

• Carol Lake Mine, Iron Ore Company of Canada (IOC) – about 15.0 M tpy iron ore concentrate

• Wabush Mine, Cleveland Cliffs – about 6 M tpy iron ore concentrate

The three mining operations combined represent the largest iron ore production in North America.


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16. MINERAL PROCESSING AND METALLURGICAL TESTING

All the recent test work results are presented in section 16.3, confirmatory testwork.

16.1 Historical Testwork The initial metallurgical testwork was carried out by Lakefield Research in the period 1973-76 on 17 drill core composites of magnetite-specularite. Nine samples were taken from the “North-East Extension” and eight from the “West Extension”. Four other composite samples of magnetite-silicate waste from an area east of the “North-East Extension” were also tested. All samples were stage crushed to –35 mesh and tested on a Wilfley table to produce a gravity concentrate. A summary of the results is shown in Table 16-1.

* Composite sample numbers 3, 4, 12 & 13 only.

Table 16-1 : Summary of 1976 Test Results In 1998, further metallurgical testing was carried out on a total of 1,267 drill core samples for QCM, who had an option on the claims at this time. The majority of the samples did not provide acceptable concentrate grades at a grind of -20 mesh using a heavy liquid of 2.95 SG. The low concentrate grades, generally between 45% and 60% total Fe and with high MgO and CaO, were attributed to a lack of liberation and low heavy liquid rejection of the actinolite, which has an SG of 3.2. A small number of the samples yielded concentrate grades in the range of 60% to 68% total Fe, with a weight recovery ranging between 30% and 45%. These samples were primarily from the West Extension where the MgO, CaO and magnetite contents are low.

Composite No. Drill Hole Head Table Concentrate Table + Mag Conc.

% Sol

Fe

% Mag

Fe

Wt.

%

% Sol

Fe

% Dist.

Sol Fe

Wt

%

% Sol

Fe

% Dist.

Sol Fe

North-East Extension

1-5, 12-15 71-1, -2 30.26 13.02 38.1 66.3 83.4 52.8* 66.6* 94.6*

West Extension

6-10, 19-21 71-3, -6, -

8,-9,-10 31.4 0.74 38.1 67.9 82.6

Mag-silicate waste

11,16-18 72-2, -3,

-5 16.1 12.5 18.3 67.8 75.2 21.5 68.2 89.2


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Further small scale liberation tests were carried out on selected samples using either heavy liquid separation or gravity separation with a Mozley mineral separator. It was concluded that use of a Mozley shaking table is a better method for predicting weight recovery than a heavy liquid separation where the specific gravity of the liquid medium is close to that of the mineral actinolite, (Fe-CaO-MgO silicate), which occurs in localized concentrations of up to 10% in the orebody. The Mozely results showed that, at a grind of -35 mesh, a concentrate grade of 67.5-68.8% Fe and 1.6 – 2.1% SiO2 could be obtained with a 94% rejection of MgO and CaO.

16.2 Recent Testwork

16.2.1 Shaking Table and Davis Tube Magnetic Separation Tests

In mid-2005, a further eleven (11) mini bulk samples were taken from outcroppings on the property and sent to SGS Lakefield for testing.

Testwork consisted of Wilfley table tests followed by cleaning of the Wilfley concentrate on a Mozely table. The results of these tests confirmed that in most cases a low silica concentrate could be produced with a high rejection level of actinolite.

Middlings and tailings from the Wilfley table were combined for each test and ground to a nominal 75% -44 microns (-325 mesh) and passed through a Davis tube to recover fine magnetite as would be the case in a magnetic separation plant. Of the 11 samples tested, only two provided sufficient magnetic concentrate for assaying and it was recommended that larger scale tests be conducted in order to determine the potential magnetite recovery from the tailings.

In order to assess the impact of recycle streams on the concentrate quality and total iron recovery for the commercial plant flowsheet, a 6-step lock-cycle test program was undertaken. For these tests, two composite samples were prepared from the eleven (11) mini-bulk samples. Assays on these samples are given in Table 16-2.

Table 16-2 : Composition of Lock-Cycle Composite Samples

Figure 16-1 illustrates the lock-cycle procedure followed for each of the composites.

For each step, 15 kg of composite sample were treated on a Wilfley table to produce a rougher concentrate, which was then treated on a second Wilfley table to produce a final cleaner concentrate. The rougher middlings and the cleaner tails were then combined with new feed material and fed to the succeeding roughing stage.

At the end of the six (6) stages, tails from cycles 3 to 6 were combined for magnetic separation tests and concentrates from cycles 3 to 6 were combined as final gravity concentrate.

Composite Assay, %

Fe SiO2 MgO Fe2O3 Fe3O4

A (Direct) 34.6 47.1 1.98 34.2 14.7

B (Direct) 31.8 49.4 2.12 33.9 11.2

A (Calc) 31.8 49.9 1.96 30.9 14.1

B (Calc) 33.3 49.4 1.98 36.5 11.4


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The results of the lock-cycle gravity separation results for composites A and B are shown in Table 16-3.

Combined tails of cycle 3,4,5, and 6

Wilfley table(Cleaner 1)






Wilfley table(Rougher 1)






cleaner tails

cleaner tails

cleaner tails

cleaner tails

cleaner tails

cleaner tails

Middlings

Middlings

Middlings

Middlings

Middlings

Middlings 6

Feed 2 (15 kg)

Feed 1 (15 kg)

Feed 3 (15 kg)

Feed 4 (15 kg)

Feed 5 (15 kg)

Feed 6 (15 kg)

Tails 1

Tails 2

Tails 3

Tails 4

Tails 5

Tails 6

Cleaner tails 6

Cleaner Conc. 1

Cleaner Conc. 2

Cleaner Conc. 3

Cleaner Conc. 4

Cleaner Conc. 5

Cleaner Conc. 6

Jeffrey Separator

Grind to 250, 270, 325, 400 mesh

Davis Tube

20 g samples

Magnetic Concentrate

Tails

Figure 16-1 : Lock -Cycle Flowsheet


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Table 16-3 : Lock-Cycle Test Weight and Iron Recoveries

Weight Assay % % Distribution Product

Kg % Fe SiO2 MgO Ht* Mt** Fe SiO2 MgO Ht* Mt**

Comp. A

Concentrate 24.9 28.4 67.0 3.53 0.34 71.6 23.5 59.9 2.01 4.97 65.8 47.4

Tails 61.5 70.1 17.4 68.4 2.66 14.1 10.5 38.5 96.3 94.9 32.0 52.1

Head (Calc) 87.7 100 31.8 49.9 1.96 30.9 14.1 100 100 100 100 100

Comp. B

Concentrate 37.4 43.0 66.1 5.58 0.53 75.5 18.4 85.5 5.01 11.44 91.2 68.5

Tails 50.3 57.9 9.77 78.9 3.06 7.56 6.19 17.0 95.3 89.1 12.3 31.0

Head (Calc) 87.6 100.9 33.8 47.9 1.99 35.6 11.6 102.5 100.3 101 103.5 99.5

* Ht = hematite, Fe2O3; ** Mt = magnetite, Fe3O4

The weight and iron recoveries were respectively 28.4% and 59.9% for composite A, and 43.0% and 85.5% for composite B. It should be noted that the silica content of composite A concentrate is significantly lower than that of composite B, while the iron grade is slightly higher.

The effect of iron and silica grades on the overall iron recovery is illustrated by the predicted performance curves developed by Lakefield shown in Figures 16-2 and 16-3. These curves are based on the iron and silica grades and recoveries for the individual stages of the lock-cycle tests.

Figure 16-2 : Fe Grade vs Recovery


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Figure 16-3 : Fe Recovery vs %SiO2

In the case of composite A, the curves predict an iron recovery of about 60% for an iron grade of 67%, which is in line with the overall lock-cycle result. If the iron and silica grades are 66% and 5.5% respectively, the curves predict an iron recovery of about 72%. For composite B, the predicted iron recovery is about 78%, which is lower than the 85.5% determined from the lock-cycle test. One possible explanation for this is that for lower iron grades, lock-cycle tests generally give better recoveries than those predicted by batch tests because recycling of middlings has a bigger impact on iron recoveries.

For each composite sample, approximately 20 kg of the combined tails were passed through a Jeffrey drum separator to produce a magnetic cobbing concentrate and tails.

To determine the effect of grind fineness on iron recovery and concentrate grade, samples of the cobbing concentrate for Composite A were ground to pass 250, 270, 325, and 400 mesh and a 20-g sample of each was then passed through a Davis tube to produce a magnetite concentrate and tails.

Table 16-4 shows the Jeffrey magnetic cobbing separation results for composite A, while Table 16-5 shows the Davis tube magnetic separation results for the 4 different grind finenesses.

Table 16-4 : Jeffrey Magnetic Cobbing Results for Composite A

Weight Assay % % Distribution Product

Kg % Fe SiO2 MgO Ht* Mt** Fe SiO2 MgO Ht Mt

Comp. A

Concentrate 6.6 31.7 26.6 56.6 2.17 9.4 27.6 52.8 25.6 25.8 23.8 88.1

Tails 14.1 68.3 11.1 76.3 2.90 14.0 1.74 47.2 74.4 74.2 76.2 11.9

Head (Calc) 20.6 100 16.0 70.0 2.67 12.5 10.0 100 100 100 100 100

* Ht = hematite, Fe2O3; ** Mt = magnetite, Fe3O4


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Table 16-5 : Davis Tube Results for Composite A at Four Grinds

Grind Sample Weight Assays, % Distribution % Mesh micro

n g % Fe SiO2 MgO Fe SiO2 Mg

O

250 63 Conc

4.8

28.2

67.8

4.15

0.13

74.7

2.05

1.62

270 53 Conc

5.5

28.4

68.1

4.16

0.18

75.1

2.07

2.28

325 44 Conc

5.1

26.3

69.0

2.66

0.10

71.4

1.21

1.07

400 38 Conc

5.5

28.7

69.0

2.83

0.13

75.2

1.44

1.65

Head (direct) - - 26.6 56.6 2.17

The Davis tube results show that the grind fineness has no impact on iron recovery. However, a finer grind has a significant impact on the rejection of silica with a slight improvement in iron grade. The iron grade increases from 67.8% at 250 mesh to 69% at 325 mesh, while the silica decrease from 4.15% to 2.66%. These results suggest that all of the recoverable iron is liberated at the coarser grind and that it is not necessary to grind finer than 270 mesh in order to obtain a high-grade magnetite concentrate.

Based on these results, the overall weight and iron recoveries for Composite A are about 34.7% and 75%, respectively. The overall weight and iron recoveries for Composite A could have been significantly higher if the gravity concentrate had been of lower grade. Due to the higher grade of Composite A, significantly more hematite and magnetite were rejected to the tails than for Composite B. Consequently, much of the hematite was not recovered by magnetic separation.

In a further series of tests, cobber concentrates for Composites A and B were stage-ground to 100% passing 325 mesh in a laboratory ball mill, followed by three (3) cleaner stages of WLIMS (wet low-intensity magnetic separators). In the case of Composite A, the weight and iron recoveries for the cleaner stage were 27% and 72.2% respectively, which were similar to the Davis tube results. The overall weight and iron recoveries for Composite A were 34.4% and 74.5%. The overall weight and iron recoveries for Composite B were 45.3% and 90.2%.

Table 16-6 shows a complete analysis of the concentrate from the lock-cycle and magnetic separation tests which indicate that there are no major impurity problems associated with the Bloom Lake deposit.


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Table 16-6 : Chemical Analysis of Bloom Lake Concentrate

16.2.2 Grindability Tests and AG Mill Simulations

In December 2005, Lakefield Research carried out a grindability study using composite samples of mini-bulk lots 10 and 11 from the west zone of the ore deposit. The grindability study included JKTech drop-weight tests, Bond ball and rod mill grindability tests, Bond low energy impact tests, Bond abrasion tests and autogenous mill circuit design simulations using JKSimMet. Two SMC tests, which are abbreviated low cost drop-weight tests, were also carried out on two other lots to estimate the ore hardness of other zones of the deposit. In addition, comparative work index determinations were made on samples of gravity tails, cobber concentrate and cobber tails.

The JKSimMet results for the primary grinding circuit autogenous (AG) mill and the secondary magnetite circuit ball mill are given in Table 16-7

Typical Specification Chemical Analysis Formula

(%) Min. (%) Max. (%)

Iron Fe 66.0 65.0 68.0

Silica SiO2 4.5 5.5

Alumina Al203 0.15 0.3

Lime CaO 0.25 0.5

Magnesia MgO 0.3 0.5

Titania TiO2 0.02 0.2

Manganese Mn 0.02 0.15

Phosphorus P 0.015 0.03

Sulphur S 0.02 0.04

Soda Na2O 0.018 0.15

Potash K2O <0.01 0.03

LOI 0.02 1.0

Size Analysis (From 1976 Testwork)

Larger than 0.4 mm 2.0 6.0

Average particle size (d50) 0.26 mm 0.26 mm

Smaller than 0.15 mm 25.0 40.0

Smaller than 0.045 mm 5.0 25.0


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Table 16-7 : JKSimMet Simulation results for Autogenous and Ball Mill Circuits


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For the AG simulation, a 10.97 m (36’) diameter x 4.57 m (15’) long mill with an effective grinding length of 43.46 (13.25’) was assumed. At a throughput rate of 1,700 t/h ore, the simulation predicts an AG mill power requirement of 4.2 kWh/t for a screen size of 35 mesh. The total installed power required is 8,952 kW (12,000 HP). The maximum throughput for this mill size, assuming the same screen opening and a maximum mill charge of 30% by volume, is 2,100 t/h. The power requirement is 3.8 kWh/t and the installed power is 10,071 kW (13500 HP).

16.2.3 Weight Recovery Versus Total Fe Relationships

Figure 16-4 shows a plot of the tabling and magnetic separation results of the 17 drill core samples tested for Republic Steel Corporation in 1975-76. As can be seen from the plot, the weight recovery for a head grade of 30% Sol Fe ranges from about 38% for gravity separation to as high as 42% with magnetic separation. It should be noted that the drill core samples used contained both low and high magnetite. However, the level of MgO and actinolite is not known.

Figure 16-4 : Weight (wt) Recovery vs Head Grade (17 composite samples – 1976 Report)

Weight and total iron recoveries for recent gravity separation tests carried out on a Wilfley table are plotted in Figure 16-5, along with Davis tube magnetic separation results on two (2) of the eleven (11) mini-bulk lots. Once again, the weight recovery for a head grade of 30% Fe is about 38% for a gravity concentrate of 65-66.5% Fe and about 42% with magnetic separation. The iron recovery from gravity separation alone is about 83%.


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Figure 16-5 : Weight (wt) and Fe Recoveries vs Head Grade (1 gravity table)

Based on the results of the eleven (11) mini-bulk samples, the weight recovery versus head grade relationship for a gravity concentrate of 65-66.5% Fe was established to be:

% weight recovery, Y = 1.3788X – 3.1746

where: X = Head grade, %Fe

With the magnetic separation included, the overall weight recovery was estimated to be: Y = 1.3788X

Both relationships were used in the geological block model and little difference was found in the economic pit area. Therefore, it was decided that the relationship for gravity separation would be used in the geological block model and adjustments would be made in the financial analysis to take into account magnetite recovery. Furthermore, it was recommended that an additional adjustment be made to take into account the effect of MgO in the feed on weight and iron recoveries. It was assumed that in the worst case, any iron tied up with MgO in the form of actinolite would not be recovered. It was determined that this effect would be well within our experimental precision of ±10% and therefore would never be detected by our testwork. An explanation of the recommended adjustments to the financial analysis is given below.

Adjustment for magnetite recovery

Based on the magnetic separation results of two (2) of the eleven (11) mini-bulk samples, it was estimated that magnetite recovery would increase the overall weight recovery by about 3.2%. For a head grade of 30% Fe, this would increase the weight recovery from 38.2% to 41.4%. Iron


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recovery would increase from 83% to 90.25%, assuming a magnetite concentrate grade of 68% Fe (i.e. Fe recovery = ((1700*0.30*.83 + 1700*0.032*0.68)/(1700*0.3))*100%).

Adjustment for MgO in Feed

As mentioned above, any iron tied up with MgO in the form actinolite, potentially would not be recovered.

In the 43-101 report from WGM, there is an explanation on the nature of the actinolite as well as references to the mineralogical work done by Lakefield research. Actinolite grains from 16 drill core samples from the CMQC drilling campaign were examined by X-Ray diffraction and electron microprobe analysis. Lakefield confirmed the presence of actinolite in the iron formation. The actinolite grains contained on average 21.3% MgO and 12.7% CaO and the composition was quite uniform at plus or minus 10%.

Actinolite is a mixture of two perfectly miscible minerals called tremolite and ferro-actinolite. Tremolite does not contain iron but ferro-actinolite contains 12.8% iron in the form of FeO. The amount of iron that can potentially be lost depends on the relative proportions of the two minerals. Based on the MgO content determined from previous testwork at Lakefield, it was estimated that the Bloom Lake actinolite contains about 68% tremolite and 32% ferro-actinolite.

Table 16-8: Mineralogical Compositions Mineral % CaO (%) MgO (%) FeO (%) Fe(%) SiO2 (%) H2O (%) Tremolite 68 13.81 24.81 - - 59.17 2.22 Ferro-Actinolite

32 12.81 13.81 16.41 12.8 54.91 2.06

Bloom Lake Actinolite

100 13.5 21.3 5.3 4.12 57.8 2.1

Table 16-8 shows that the Bloom Lake actinolite is estimated to contain 5.3% FeO or 4.12% Fe. This means that for each 1% MgO in the ore, there is 0.194% Fe attached to it.

The Bloom Lake deposit is estimated to contain 2.2% MgO. Therefore, the iron recovery is expected to be reduced by 0.43% (2.2*0.194%). At a concentrate grade of 65%, this translates to a reduction in weight recovery of 0.66% (0.43/0.65). Such a small reduction of the weight recovery is not detectable in our laboratory or piloting testwork.

16.2.4 Weight and Iron Recoveries and the Financial Model

The financial model is adjusted to take into account a credit for magnetite recovery as well as a debit for the reduction in weight yield and iron recovery resulting from the loss to tailings of iron tied up with MgO as actinolite.

Relationships used in the model are:

% weight recovery = 1.3788 * %Fe feed – %MgO * 0.194 *100 / %Feconc

% iron recovery = 1.3788 * %Feconc – %MgO * 0.194

where 0.194 is the ratio of Fe to MgO in actinolite, as mentioned in the previous section


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16.3 Confirmatory Testwork

As recommended in the April 2006, 5 Mtpy feasibility study report, a drilling campaign was carried out by Consolidated Thompson Iron Mines Ltd between November 2006 and January 2007 to obtain drill core samples for metallurgical and grindability testwork. Thirty-two drill-core samples for metallurgical testing and 32 drill-core samples for grindability testing were obtained from 12-bore holes located in the west, central, northeast and southeast areas of the Bloom Lake pit. The samples were collected at different depths.

The grindability tests confirm that a 36 x 19.8 ft autogenous mill is appropriate for processing 2156 tph of ore. The specific power consumption of the autogenous mill was estimated to be 3.82 kWh/t at the shell. The samples representing the Bloom Lake ore were characterized as soft to very soft with the presence of some hard material. The SPI (The MinnovEX SAG Power Index) profile of the Bloom Lake ore is very similar to other iron ores where the majority of the ore is soft, but a small portion is hard enough to be used as grinding media in a fully autogenous mill. SPI is the time (in minutes) required to grind a sample from 80% passing 1/2” to 80% passing 10 mesh. SPI values are used to predict throughputs for SAG/AG mills and in the determination of power requirements for such mills via the MinnovEX CEET program.

16.3.1 Metallurgical Test Work

Metallurgical test work was carried out on 32 drill core samples from four areas of the Bloom Lake pit at different depths. The test work included rougher gravity separation by Wilfley shaking table, cleaner gravity separation by Mosely shaking table, magnetic separation by Davis tube on selected Wilfley rougher tails, and Jeffrey cobbing followed by Davis tube cleaning on other selected Wilfley rougher tail samples.

16.3.1.1 Gravity Separation Results

The gravity separation test results are summarized in Table 16-9. For most samples, the results of the rougher stage are reported, as the grades and recoveries are high and no further processing is necessary. In the cases where the silica contents of the rougher concentrates were about 5% or higher, the results of either the cleaner concentrate or cleaner concentrate plus middlings are reported, depending on which had better recoveries while achieving concentrate silica contents of less than 5%. The average composition of the drill cores tested is 29.8% Fe, 1.94% MgO, and 10.3% Fe3O4, which is similar to the average ore composition used in the feasibility study carried out in 20061. The average drill core concentrate composition was 67.7% Fe, 2.76% SiO2, 0.26% MgO, and 25% Fe3O4, while the average weight and Fe recoveries were 38.8 and 88.3%, respectively.

The concentrate grades and Fe recoveries achieved in these tests were significantly better than those achieved in the metallurgical tests carried on mini-bulk surface samples in 2005 and 2006. The average concentrate grade and Fe recovery for the mini-bulk samples were about 65% and 83%, respectively. The average weight recovery was slightly higher than the 38% achieved with the mini-bulk samples tested in 2005 and 2006. In both series of tests, the ore samples were ground to -35 mesh. The drill core results seem to suggest that the Fe was well liberated in most cases and may not have required such a fine grind as in the

1 BBA, 5 Mtpy Feasibility Study – Bloom Lake Iron Ore Project, Revision 0, project no. 5743002, May 2006


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16-13

earlier tests carried out in 2005 and 20062. The main differences between the two series of tests are that the mini-bulk samples were surface samples and were taken mainly from the northeast and west sections of the pit, whereas the drill cores samples were taken in all areas of the pit and at different depths.

16.3.1.2 Weight and Iron Recovery Relationships

Figure 16.6 shows the weight and iron recoveries as a function iron head grade. Based on the test results, the weight recovery versus head grade relationship is as follows:

WtR = 1.3015*X

Where WtR = Weight recovery, %

X = Head grade, % Fe

For an average Fe content of 30%, the weight recovery is 39%. The iron recovery is 88.3% This is better than the mini bulk sample test work which gave a weight recovery of 38.2% and an iron unit recovery of 83 % for a feed grade of 30%.

2 SGS Lakefield Research Limited, An Investigation into The Beneficiation of Iron Ore Samples from the Bloom Lake Property, prepared for Watts, Griffis and McOuat on behalf of Consolidated Thompson Lundmark Gold Mines Ltd., February 23, 2006.


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Table 16-9 : Summary of Gravity Separation Results

Head Assay Concentrate assay Recovery/Recjection Product Type Sample

ID Location From (m) %Fe % MgO % Fe3O4 %Fe %

SiO2 %

MgO % Fe3O4 Fe Rec MgO Rej Wt Rec

Rougher conc. 3053 central 80 29.7 0.74 6.71 68.1 2.36 0.17 12.8 91.7 89.4 40.7

Rougher conc. 3054 central 89 21.7 0.98 13.8 68.6 2.73 0.25 42.6 86.8 92.2 27


Cl. Conc. + Mid 2-1 3058 central 141 42.8 0.32 2 66.3 4.73 0.1 2.43 93.6 77.5 59.2

Rougher conc. 3059 central 150 43.6 1.87 3.07 67.6 1.81 0.26 4 96.1 89.6 63.7

Cl. Conc. + Mid 2-1 3061 central 67 31.3 0.42 10.2 68.3 2.08 0.1 21.1 82.5 90.8 36.7

Rougher conc. 3063 central 75 34.6 0.14 3.95 66 4.32 0.05 7.33 90.9 80.9 42.6


Cl. Conc. 3070 northeast 140 25.5 3.2 21.2 66.7 3.51 0.69 59.3 81.7 93 28.9

Rougher conc. 3071 northeast 203 29 3.66 4.99 68.5 1.42 0.25 11.5 87.3 97.3 37

Cl. Conc. + Mid 2-1 3076 southeast 228 23.2 5.1 0.63 64.3 3.87 1.07 1.64 86.4 91.4 34.3

Rougher conc. 3083 southeast 240 23.3 3.32 30.8 68.2 3.99 0.46 95.8 82.5 95.9 30

Rougher conc. 3084 southeast 246 33.1 4.87 21.8 69.2 1.71 0.42 48.8 91 96.5 40.9

Rougher conc. 3085 southeast 264 20.4 2.7 8.96 69.5 1.38 0.23 30.2 91.1 97.2 29.8

Cl. Conc. 3089 southeast 99 32.9 0.74 40.9 67.9 5.01 0.16 86.6 70.1 90.9 34.7

Rougher conc. 3090 southeast 105 35 1.95 25.7 69.2 1.54 0.18 52.3 92.7 94.5 50.9

Rougher conc. 3092 southeast 123 41 0.99 8.57 68.9 1.42 0.08 14.5 92.1 94.9 50

Rougher conc. 3094 southeast 243 30.7 4.49 1.22 65.7 3.39 1.07 2.1 89.7 89.6 44.7

Cl. Conc. + Mid 2-1 3099 west 98.7 33.4 0.69 67.8 2.92 0.05 0.84 81.5 40.9

Cl. Conc. + Mid 2-1 3100 northeast 80 20.8 1.73 24.8 66.5 4.99 0.5 83.7 81.7 26.7

Rougher conc. 3101 northeast 98 23.7 0.38 6.13 65.6 4.92 0.12 12.6 92 86.8 35

Rougher conc. 3106 west 50 29.9 0.05 0.15 68.1 2.36 0.02 0.7 92 39

Rougher conc. 3108 west 74 32.3 0.05 0.46 67.3 3.16 0.02 0.6 93.3 44.4

Rougher conc. 3109 west 86 25.5 0.16 0.35 67.6 2.56 0.03 0.7 90.8 94.8 34.3

Rougher conc. 3114 northeast 134 24.7 3.91 3.22 68.8 0.55 0.14 7.6 90.7 98.7 35.8

Rougher conc. 3115 northeast 143 25 3.06 2.26 68.1 0.83 0.15 5 88.3 98.6 28.2



Average 29.79 1.94 10.29 67.73 2.76 0.26 24.97 88.32 92.42 38.75


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Wt and Fe recoveries vs Head Grade (Best Recoveries with %SiO2 in Conc. less than 5%- Wilfley or Mosely Table)

y = 1.3015xR2 = 0.8386

y = 0.2498x + 80.875R2 = 0.083

0

10

20

30

40

50

60

70

80

90

100

110

0 5 10 15 20 25 30 35 40 45 50

Head Grade, % Fe

Wt a

nd F

e re

cove

ry %

Head grade vs Fe recovery Head grade vs wt recovery overall Fe recovery (incluing mag sep)overall wt recovery (incluing Mag sep) Linear (Head grade vs wt recovery) Linear (Head grade vs Fe recovery)

Average Head Grade = 29.8% Fe, 1.94% MgO, 10.3% Fe3O4

Average Conc. Comp.% Fe - 67.7% SiO2 - 2.8% MgO - 0.26% Fe3O4 - 25

Figure 16-6: Weight and iron recoveries for gravity separation


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Magnetic separation results are also shown in Figure 16.6 and will be discussed in section 16.3.1.2.2.

Figure 16-7 shows the weight recovery relationships by pit area. As can be seen, the weight recoveries are essentially the same in all areas of the pit.

Wt Recovery vs Head Grade( by Area)

y = 1.2734xR2 = 0.5398

y = 1.3028xR2 = 0.6811

y = 1.3146xR2 = 0.9038

y = 1.3073xR2 = 0.7498

0

10

20

30

40

50

60

70

0 5 10 15 20 25 30 35 40 45 50

Head Grade (%Fe)

Wt R

ecov

ery

(%)

head grade vs wt recovery (central) head grade vs wt recovery (southeast)head grade vs wt recovery (northeast) head grade vs wt recovery (west)Linear (head grade vs wt recovery (west)) Linear (head grade vs wt recovery (northeast))Linear (head grade vs wt recovery (southeast)) Linear (head grade vs wt recovery (central))Linear (head grade vs wt recovery (west))

Figure 16-7: Weight recovery versus head grade for the four different areas of the Bloom

Lake pit


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16.3.1.2.1 MgO Rejection

Figure 5 shows MgO rejection by gravity separation as a function of MgO in the head samples.

Mgo rejection vs Head MgO

50

60

70

80

90

100

0 1 2 3 4 5 6

% MgO (head)

MgO

reje

ctio

n (%

)

MgO head vs MgO rejection Log. (MgO head vs MgO rejection) Figure 16-8 : MgO versus MgO in head sample

For an average MgO content of 2% in the ore, the rejection of MgO to the gravity tailings is about 94%, which is in line with previous testwork. Thus rejection of over 90% of the actinolite during gravity separation is confirmed for all areas of the orebody.


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16.3.1.2.2 Magnetite to Gravity Tails

Figure 16-9 shows the relationship between magnetite in the head samples versus the deportment of magnetite to the gravity tails.

Magnetite in Ore vs Magnetite to Tails

y = 26.928x-0.2036

R2 = 0.2636

0

10

20

30

40

50

60

70

80

90

0 5 10 15 20 25 30 35 40 45

Magnetite in Ore, %

Mag

netit

e to

Tai

ls, %

dist

Magnetite in feed vs magnetite to tails Magnetite in feed vs magnetite to tails (composites) Power (Magnetite in feed vs magnetite to tails) Figure 16-9: Deportment of magnetite to tails as a function of magnetite in the ore

As can be seen from the graph, the deportment of magnetite to the gravity tails is about 15% for an ore magnetite content of 5% or greater.


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Figure 16-10 shows the relationship between magnetite in the ore and magnetite in gravity tails.

Magnetite in Gravity Tails vs Magnetite in Ore

y = 0.263xR2 = 0.7939

0

1

2

3

4

5

6

7

8

9

10

11

12

13

14

15

16

17

18

0 5 10 15 20 25 30 35 40 45

Magnetite in Tails, %

Mag

netit

e in

Gra

vity

Tai

ls, %

Magnetite in Feed vs Magnetite in Tails Linear (Magnetite in Feed vs Magnetite in Tails) Figure 16-10 : Magnetite in tails as a function of magnetite in the ore

For an average ore magnetite content of 10%, the magnetite concentration in gravity tails would be about 2.5%.

16.3.1.3 Magnetic Separation Results

In a typical magnetic separation circuit, a cobbing concentrate is produced before further grinding and magnetic recovery is done. Davis tube tests were carried out on the gravity tails and on the cobbing concentrate. The samples tested were 3061 and 3066, respectively. The results showed that the magnetite level in the gravity tails has an impact on the silica content of the concentrate produced and that a cobbing stage was necessary in order to achieve a silica content of less than 4% in the final concentrate.

Additional Davis tube tests carried out on the cobbing concentrates of samples 3083 and 3100, showed that in order to obtain concentrate silica levels of about 3% or less, the cobbing concentrate would have to be ground to 53 microns (270 mesh) or finer. This is in line with previous testwork.


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y = -0.3004x + 7.0354R2 = 0.3944

0

1

2

3

4

5

6

7

8

9

10

0 2 4 6 8 10 12 14 16 18 20

% Fe3O4 in Feed

% S

iO2

in C

onc.

New drill cores

Comp B

Comp A

Figure 16-11: Concentrate silica content versus magnetite in gravity rougher tails

16.3.1.3.1 Magnetite Recoveries

Figure 16-12 shows magnetite recovery to cobbing concentrate as a function of magnetite concentration in gravity tails. As was seen in Figure 16-10, the magnetite concentration in gravity tails is expected to be around 2.5% for Bloom Lake ore containing an average of 10% Fe3O4. At this level, magnetite recovery to cobbing concentrate is expected to be around 55%.

The results from Davis tube tests carried out on cobbing concentrate showed that the magnetite recovery during the cleaning stage was consistently around 97%.

Magnetite Recovery in Cobbing Stage vs Magnetite in Rougher Tails

0

10

20

30

40

50

60

70

80

90

0 1 2 3 4 5 6 7 8 9 10

% Fe3O4 in Rougher Tails

% F

e3O

4 R

ecov

ery

in C

obbi

ng S

tage

Magnetite in rougher tails vs magnetite recovery in cobbing stage Log. (Magnetite in rougher tails vs magnetite recovery in cobbing stage) Figure 16-12 : Magnetite Recovery in Cobbing stage as a Function of Magnetite in

Rougher Tails


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16.3.2 Overall Weight and Fe Recovery

Based on the recent metallurgical test work on drill core samples, there is sufficient information to estimate the overall weight and iron recoveries for an average Bloom Lake ore composition of 30% Fe, 1.94% MgO, and 10% Fe3O4. Based on test data, the assumptions used to estimate overall recoveries for an average Bloom Lake ore are shown in the following table.

Table 16-10: Assumptions to Estimate Overall Recoveries for an Average Bloom Lake Ore

Stage Fe2O3 Recovery, %

Fe3O4 Recovery, %

Fe Recovery,

%*

Gravity sep. – Wilfley Table

90 85 88.6

Cobbing – Jeffrey 10 55 24.5

LIMS – Davis Tube

25 97 77

* Calculated from Fe2O3 and Fe3O4 recoveries

The average overall weight and iron recoveries were estimated to be 40% and 90.8%, respectively. Magnetic separation would provide an average additional weight and iron recovery of 1% and 2.2%, respectively. Therefore, the magnetite concentrate would account for about 2.5% of the total concentrate produced. This is equivalent to the overall weight recovery of 41.4% and iron unit recovery of 90.3% measured on the minibulk samples. The lower weight recovery is due to a better average grade acheived on the drill core samples. The average silica level was 2.7 % for the drill core concentrates versus 4.5% for the mini bulk concentrates.

16.3.3 Confirmatory test work conclusions

The metallurgical tests confirmed the overall Fe recovery for the Bloom Lake project at 90.8% for gravity and magnetic separation combined. The average concentrate grade and Fe recovery for gravity separation was also significantly better than for the previous mini-bulk samples tested. The average concentrate grade and Fe recovery were 67.8% and 88.3%, respectively.

The relatively high grade and Fe recovery for gravity separation suggests that the ore is well liberated at 35 mesh throughout the ore body and that it may be possible to perform gravity separation at a coarser grind.

Due to the high Fe recoveries achieved in the gravity separation tests, only about 15% of the magnetite in the ore samples reported to the gravity tails. Assuming this is representative of the Bloom Lake pit, magnetic separation of gravity tails would provide additional weight and iron recoveries of 1% and 2.2%, respectively.

The weight recoveries in all four areas of the ore body tested is essentially the same as shown in Figure 16-7. Thus the 35 mesh grind size is appropriate for all zones and ore depth.


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MgO reduction of 90% and over was confirmed for all zones and depth of the ore body. This clearly shows that the actinolite is rejected with the silica during the gravity concentration process and does not pose any operating problems for the Bloom Lake concentrator.


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17. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

17.1 General

As part of the 1998 exploration program, WGM prepared a “mineral resource” estimate for the Bloom Lake iron deposit. WGM produced cross sections, level plans and block models for geology and selected elements and used this model to determine the tonnage and grade of the deposit at various cut-off grades. For the purposes of its May 2005 report WGM reviewed the 1998 estimate and determined that it was prepared in a manner that was compliant with NI 43-101, which came into effect in 2002. The remainder of this Section 17, up to Section 17.9 is taken with minor editing from the WGM Technical Report filed on SEDAR on May 2005.

All pertinent Log II data were converted into Gemcom Software International Inc. ("Gemcom") format for the estimation procedure. The imported assay data were checked with hard copies received from the lab and the database was validated within Gemcom to search for errors such as missing or overlapping intervals, correct hole lengths, azimuths and dips, duplicate samples, etc. Over the course of the project, meetings and discussions were held with QCM personnel regarding important elements of the new database and parameters and methodology to be used for the estimation.

WGM used the previous and new geological data, in conjunction with agreed upon cut-offs and parameters, to outline the lithology and mineralized zones on sections. The interpretations on the sections were then transferred to level plans where the final interpretations and refinements were completed. The block model grades (Total Fe%, Fe3O4%, CaO%, MgO%) were interpolated into these outlines using an inverse distance squared ("IVD2") method and weighting these by specific gravity ("SG"). WGM produced an interim tonnage and grade after the Phase I/II drilling program, and a final block model estimation was completed after Phase III drilling was finished.

The final density model was created by using the actual SGs supplied by Lakefield, or the calculated SGs, based on the Fe% grades, in the case of the historic sample intervals. Mean SGs were used for the other geological units.

17.2 Geological Interpretation

The Bloom Lake deposit is a complexly folded and metamorphosed sedimentary OIF (see Table 9-1) with abundant intercalated layers and bands of SIF, QRIF, QR, MS and AMP. The OIF above 15% Total Fe represents the potential economic mineralization, and the remaining lithologies are considered waste material and were subdivided where possible. While WGM believes that the Soluble Fe% content in the QC and 71/72 series holes is probably slightly underestimated related to Total Fe%, no correction has been made to these previous holes to adjust them to a Total Fe% value (Section 14 – Data Validation). For the digitizing of the geological boundaries and the grade interpolations, Soluble Fe% and Total Fe% were considered to be equivalent.

Digital topographic data were imported into Gemcom (dxf format) from information supplied by QCM and data collected by WGM from various sources. A triangulated surface was constructed from these data and used as the base map for subsequent work.

Cross sections at 100 to 150 m spacing were generated across the strike of the deposit (Figure 11.1). The sections included the intersection of the topographic surface and drillhole traces with geological codes (Section 9.1) and both Total Fe% and Mag Fe%. The geological interpretations were drawn on the plotted sections and subsequently digitized, attempting to connect any continuous units from section to section (Figure 7.5 and 7.6).


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A surface plan was created from the old and new mapping and was used in the projection of the units to surface. Since the sections were at various orientations, the folding complex in some areas and numerous discontinuous or isolated waste bands are present, it was very difficult to build a true 3-D model by correlating every individual band or unit across sections.

The sectional information was transferred to level plans that were generated at 15 m intervals. A more detailed geological interpretation was prepared on these level plans and the final level geology was used to build the lithological block model. This was done by projecting each level outline half-way to the adjacent levels to make solids (or slices) of each defined unit in order to update the block model matrix. Any narrow or discontinuous bands within the OIF that could not be considered mineable units (i.e., approximately 6 m or less horizontal width) were included as internal waste.

Figure 17-1 shows the detailed geological interpretation and resulting lithological block model on Level Plan 647.5 m, which has been chosen as a representative level.

WGM therefore established controls on the deposit by digitizing an envelope around the OIF at a cut-off grade of 15% Total Fe, based on the nominal 6 m crude assay intervals. These constraining envelopes would prevent grades from being smeared along the search directions and would approximate the actual geological and possible mining controls of the deposit.

17.3 Compositing

The compositing was done within all the lithological units, however, only the OIF units were of interest for this exercise. Fixed length (6 m) composites were produced starting from the upper boundary of each of the units. Each composite was flagged with the rock code within that particular outline. At the base of each unit, the last composite may not be the full 6 m fixed length. In order for the composite to be flagged as OIF for the grade interpolation, it had to be at least 40% of the fixed length (i.e., 2.4 m).

Total Fe%, Fe3O4%, CaO%, MgO% and SG were composited using this method over the assayed intervals and were used to build the block model. Only the composites defined as OIF were used for the grade interpolation within the previously defined Mineral Resource boundaries. Other elements of interest, if the project progresses further, could also be composited and used in the block model, as the Mineral Resource intervals are already calculated.


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Figure 17-1 : Lithological Block Model

Graphics by WGM


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Using this method of compositing, only the narrow, <6 m horizontal width or isolated bands of QR, SIF or amphibolite are incorporated into the model as internal dilution. The grades produced for this model represent an in situ Mineral Resource and do not take into account any external mining dilution or mining recoveries. Bench compositing would more closely reflect actual mining units and grade distribution, especially at the upper and lower contacts of the deposit.

17.4 Block Model Parameters

The block model was created using Gemcom's PC-Mine and Gem4 Win programs and was used to estimate tonnes and grade. The block model size at Bloom Lake is set at 10 x 10 x 14 meters in the (x,y,z) directions, respectively.

The level plan solids generated previously were used to build a 3-D lithological model, and codes identifying the rock types were assigned to each block that fell within the solids model. For OIF, 40% or more of the block had to be included in the solids to be identified as OIF, and for waste, 50% or more of the block had to be included. This model was used as the boundaries for the grade interpolations, however, grades were only interpolated into the OIF units.

Block sizes used were the same as those used at the Mont-Wright Mine since the deposits are similar in nature.

17.5 Grade Interpolation and Statistics

Statistics were calculated for the core samples for each rock type and are summarized in Table 9.2. A histogram, a cumulative frequency plot and a probability plot (Figures 17.2 to 17.4) for Total Fe%, were produced using the 6 m fixed length composites for the oxide iron formation. The histogram shows a robust normal distribution (what one would expect from an iron ore deposit), therefore an inverse distance squared ("IVD2") search appears to be appropriate.

Both the frequency and probability plots indicate that no cutting of high grade Total Fe% values is necessary. There are few high-grade outliers, which will have negligible effects on the overall grade distribution of the deposit after interpolation. These values are from composites and WGM believes that they represent real high-grade areas and should not be reduced. Histograms were also produced for Magnetite%, CaO% and MgO% for the 6 m composites (Figures 17.5 to 17.7). Variograms were also constructed and were used for the search ellipse parameters.

The 6 m composite grades calculated within the Mineral Resource zones (based on the previously stated compositing and identification procedures) were used for interpolation into the grade block model. Only values with a matching zone designation of OIF were used in the Mineral Resource block model creation. This avoids the transfer of assay values between unspecified or waste zones and reduces the smearing effect inherent in block modelling when no such restrictions are imposed.


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Figure 17-2 : Total Fe% Histogram

Graphics by WGM


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Figure 17-3 : Total Fe% Cumulative Frequency Plot

Graphics by WGM


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Figure 17-4 : Total Fe% Probability Plot

Graphics by WGM


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Figure 17-5 : Magnetite % Histogram

Graphics by WGM


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Figure 17-6 : CaO% Histogram

Graphics by WGM


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Figure 17-7 : MgO% Histogram

Graphics by WGM


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The parameters for the grade modelling in the zones were as follows:

Method:

Search Ellipse:

Measured

Inverse Distance Squared

No Rotation, Anisotropic Weighting

Range 1 (X axis - principal strike direction): 75 m

Range 2 (Y axis - width): 75 m Range 3 (Z axis - thickness): 75 m

Indicated Range 1 (X axis - principal strike direction): 150 m Range 2 (Y axis - width): 150 m Range 3 (Z axis - thickness): 150 m

Inferred Range 1 (X axis - principal strike direction): 400 m Range 2 (Y axis - width): 400 m Range 3 (Z axis - thickness): 400 m

Composite Length: 6 m Minimum Samples per Block: 2 Maximum Samples per Block: 12 Maximum Samples per hole: 8

Figures 17.8 to 17.11 show level plans of the interpolated grade blocks for the four elements at elevation 647.5 m, the same location as the plan showing the geological interpretations and block lithological model.

17.6 Density Assignment

A density model was created using PC-Mine and Gem4Win software. The 71/72 and QC series of drillholes did not have SGs for the assayed intervals. WGM calculated a SG for each interval for the old holes by using the formula derived from the relationship of SG to total iron content in the new drilling. While there were a few samples with an abnormally high SG, it was recognized that these samples had a negligible effect on the overall density model, however, WGM elected to recalculate (using the formula) any SG over 4.0 for any samples below 50% Fe in the new holes.


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Figure 17-8 : Fe% Grade Block Model

Graphics by WGM


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Figure 17-9 : Magnetite Grade Block Model

Graphics by WGM


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Figure 17-10 : Ca0% Grade Block Model

Graphics by WGM


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Figure 17-11 : Mg0% Grade Block Model

Graphics by WGM


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The calculated and Lakefield SGs were composited using the methodology described above and these 6 m intervals were used in the IVD2 interpolation for the density model for the OIF. WGM used average SGs for the remaining geological units as the database was not as complete for the waste rock.

The following SGs were used: OIF: Variable Density Model (IVD2) QR: 2.77 SIF: 3.37 AMP: 3.19 QRIF: 2.94 MS: 2.80

17.7 Mineral Resource Definitions

When the Mineral Resource estimate was prepared in 1998, NI 43-101 had not yet been implemented. There were, however, draft Mineral Resource and Reserve definitions being reviewed by regulators at that time and these draft definitions were used by WGM to classify the Bloom Lake Mineral Resource.

For the purposes of May 2005 report, WGM reviewed its 1998 estimation procedures and reclassified the Mineral Resource estimate using NI 43-101 guidelines and CIM Standards. The definitions in effect in May 2005 (they were amended slightly in December 2005) for the CIM Standards were as follows:

A Mineral Resource is a concentration or occurrence of natural, solid, inorganic or fossilized organic material in or on the Earth's crust in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge.

An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes.

An Indicated Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough for geological and grade continuity to be reasonably assumed.

A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough to confirm both geological and grade continuity.


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17.8 Mineral Resource Statement

The Bloom Lake Mineral Resource classifications are based on drillhole and surface sample density and confidence in mineralization continuity. Drillhole spacing is nominally at 150 m on and between sections. A 15% Fe cut-off was used and was based on that employed at operating mines in the Labrador Trough exploiting deposits similar to Bloom Lake.

Measured Resources are defined as OIF mineralization no greater than 75 m from drillhole intersections within the interpreted geological boundaries.

Indicated Resources are defined as OIF mineralization between 75 m and 150 m from drillhole intersections. This mineralization is the deeper OIF and represents larger drillhole spacing and less confidence in the geological interpretation.

Inferred Resources are defined as OIF mineralization between 150 m and 400 m from drillhole intersections. This mineralization only occurs on the fringes of the deposit and represents a very minor amount (about 5%) of the OIF.

Table 17-1 summarizes the results for the final block model Mineral Resource estimation. The Measured plus Indicated block model Resources at Bloom Lake using a 15% Fe cut-off and an IVD2 grade interpolation technique total 638 million tonnes grading 29.76% Total Fe and 10.54% Magnetite.

Table 17-1 : Total InSitu Mineral Resources at a cutoff grade of 15% Total Fe

Resource Volume Tonnage Average Grades Category bcm* x 1,000 (kt) Total Fe% Magnetite% CaO% MgO%

Measured 141,350 488,465 29.91 10.54 2.32 2.18

Indicated 43,372 149,232 29.29 10.55 2.37 2.15 Total Meas. Indicated

184,722 637,697 29.76 10.54 2.33 2.17

Inferred 10,322 35,697 30.97 8.47 0.84 0.82

* Bank cubic metre.

The Inferred Resource is in addition to the Measured and Indicated Resource.

17.9 Mineral Reserve Estimate

This is discussed in Section 18.


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18. ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES

18.1 Mineral Reserves

Pit optimization and pit design were carried out to convert the Mineral Resources into Mineral Reserves for the deposit.

18.1.1 Weight Recovery Model

The original resource model was prepared by Watts, Griffis and McOuat Limited (WGM) in its Technical Report filed on SEDAR in May 2005. Geostat System International (“Geostat”) reviewed the WGM original Mineral Resource modeling in January 2006 and concluded that the Mineral Resource compilation for the Measured and Indicated categories using similar classification parameters is within 1% of the estimate by WGM based on a cut-off grade of 15% Fe. BBA has accepted the WGM original Mineral Resource estimate and has created a formula-based weight recovery model in preparation for open-pit optimization. The weight recovery formula was developed by BBA and details are presented in Section “Mineral Processing and Metallurgical Testing” in this Technical Report (Section 16).

18.1.2 Pit Optimization

The Lerchs-Grossman 3D (LG 3D) pit optimization algorithm in MineSight/MedSystem software was used to develop the configuration of the open-pit at the end of its economic life. The LG 3D algorithm is a true pit optimizer based on graph theory, an operational research technique. It operates on a net value calculation for all the ore blocks as defined in the block model, i.e. revenue from concentrate sale less operating costs. The revenue generated by the sale of concentrate is proportional to the weight recovery of the mineral resource.

18.1.2.1 Pit Optimization Criteria and Parameters

In accordance with the regulations governing the preparation of an NI 43-101 report, the pit optimization has used only ore blocks classified in the measured and indicated categories to generate revenue. Other mineralized blocks containing inferred resource bear no economic value, regardless of grade and weight recovery and are treated as waste rock, unless proven otherwise by additional geological work.

The pit optimization exercise was based on economic parameters derived from an internal cost model prepared by BBA. A summary of all the LG 3D optimization parameters is presented in Table 18.1.


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Table 18-1 : Pit Optimization Parameters for LG 3D

Sale Revenue

Concentrate Price (66% Fe) 38.94 US$/t con

Operating Cost

Mining (ore and waste) 3.03 US$/t rom

Crushing and Processing 1.57 US$/t rom

Transport and Port 3.56 US$/t rom

G/A 0.46 US$/t rom

Others - US$/t rom

Incremental Cost per Bench -

Metallurgy

Weight Recovery As per Block Model %

Pit Parameters

Overall Pit Slope 50o

18.1.2.2 Cut-off Grade Calculation

The economic cut-off grade is used to determine the material inside the pit design as ore or waste. The economic cut-off grade is based on metallurgical weight recovery or concentrate yield to cover beneficiation and administration costs.

The cut-off grade for the Bloom Lake project was determined using the data presented in Table 18.2.


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Table 18-2 : Cut-off Grade Calculation

Sale Revenue

Concentrate Price 38.94 US$/t con

Operating Cost

Mining (ore and waste) 3.03 US$/t rom

Crushing and Processing 1.57 US$/t rom

Transport and Port 3.56 US$/t rom

G/A 0.46 US$/t rom

Others (royalty…) - US$/t rom

Total 8.62 US$/t rom

Weight Recovery Cut-off 22.1%

Total Fe Cut-Off (with Mag Rec) 16%

Used 15%

18.1.2.3 Detailed Pit Design

The detailed pit design work was carried out based on the optimum pit outline obtained from the LG 3D described above. Detailed mine design has taken into account the following parameters:

• Haul road gradient, width and location

All in-pit ramps will be 35 m wide to accommodate 240-ton class off-highway dump trucks. This ramp width is sufficient to support the two-way traffic system required to maintain an uninterrupted haulage cycle. Based on hauling experience from other mines operating under similar freezing conditions in winter, the maximum recommended ramp gradient will be 8% for all mine roads. Temporary ramps will be used in the early years of mine operations to shorten haulage distances to the primary crushers or to the waste dump. As the pit gets deeper over the years, a final ramp will be used to access the lower ore.

• Geotechnical pit slope

Due to limited available geotechnical strength data, an overall slope angle of 50o and a face angle of 70o have been used for the detailed pit design based on experience from similar iron ore operations in the Labrador – Mont Wright area. Double-benching is also adopted for the pit arrangement as per operating iron ore mines in the area. Further analysis on pit slope stability in the hanging-wall will be undertaken to determine whether this is the optimal configuration with respect to rock structures and properties.

The general layout of the detailed mine design is shown in Figure 18-1. As can be seen in this drawing, the Bloom Lake ultimate pit design is divided into 2 areas: the West pit and the Main Pit. Three typical cross-sections, showing the pit profiles and the ore blocks are presented in Figures 18-2, 18-3 and 18-4.


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Figure 18-1 : Detailed Open-Pit Mine Design


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Figure 18-2 : Pit X-Section @ 614600 E (Looking West)


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Figure 18-3 : Pit X-Section @ 615600 E (Looking West)


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Figure 18-4 : Pit X-Section @ 5855600 N. (Looking North)


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18.1.2.4 Mineral Statement

Based on the economic input parameters described in the sections above, the mineral reserves have been determined and classified in the Proven and Probable categories in accordance with the criteria of the Canadian National Instrument 43-101 (“CNI 43-101”) for Standards of Disclosure of Mineral Project of February 2001 and the classifications adopted by the CIM Council in August 2000 as follows:

A “Probable Mineral Reserve” is the economically mineable part of an Indicated, and in some circumstances a Measured Mineral Resource demonstrated by at least a Preliminary Feasibility Study. This study must include adequate information on mining, processing, metallurgical, economic, and other relevant factors that demonstrate, at the time of reporting, that economic extraction can be justified.

A “Proven Mineral Reserve” is the economically mineable part of a Measured Mineral Resource demonstrated by at least a Preliminary Feasibility Study. This study must include adequate information on mining, processing, metallurgical, economic, and other relevant factors that demonstrate, at the time of reporting, that economic extraction can be justified.

Table 18-3 provides a summary of the mine-life mineral reserves for the Bloom Lake Iron Ore Project using a cut-off of 15% Fe.

Table 18-3 : Mineral Reserves by Ore Category in Ultimate Pit Design (using a cut-off grade of 15% Fe)

Tonnes WTRec(*) TFe MagFe CaO MgO MagClassification

(million) (%) (%) (%) (%) (%) (%)

Proven 463.4 38.3 30.1 7.6 2.2 2.1 10.5

Probable 116.2 37.7 29.7 7.7 2.3 2.1 10.7

Total Ore 579.6 38.2 30.0 7.6 2.3 2.1 10.5

Total Waste

(Waste+Inferred)563.8

Stripping Ratio (t/t) 0.97

(*) For hematite ONLY in block model

Figures may vary slightly due to rounding effects and the number of decimal points used.

The mineral reserve contained in the detailed pit design amount to 463.4 million tonnes of proven category and 116.2 million tonnes of probable category for a total combined of 579.6 million tonnes at a grade of 30.0% Fe based on a cut-off of 15%. The stripping estimate (including inferred material) is 563.8 million tonnes for an overall mine-life stripping ratio of 0.97 tonne of waste per tonne of ore.


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18.1.3 Mining Planning

18.1.3.1 Mining Method

The Bloom Lake iron ore deposit will be mined using conventional open pit mining methods based on a truck/shovel operation. Following drilling and blasting, run-of-mine (ROM) ore will be delivered to primary crusher located approximately 1000 m to the north-east of pit edge at about 680 m elevation. The mining equipment will be leased, operated and maintained by the personnel of Bloom Lake.

18.1.3.2 Annual Production Requirement

For the current expansion study, the mine will be developed to support a nominal capacity of 8-million tonnes of concentrate per annum. Based on the results of the metallurgical testing and the geological block modeling work, the average calculated weight recovery is 38% in hematite and 41% with the magnetite recovery plant, equivalent to a concentration ratio of 2.63 tonnes of ROM per tonne of concentrate in the hematite and 2.43 with the magnetite recovery.

18.1.3.3 Blending

According to the production schedule, the magnetite recovery plant will be built in year 3 following additional test work and will begin operation in 2011 with a production target in 2012. As described in the section on Geology, the Bloom Lake deposit is divided into 2 areas, each with the distinctive geological and quality features as follows:

• The West Pit contains mainly hematite material with a low MgO% and CaO% content (under 0.03% on average);

• The Main Pit is characterized by up to 10% magnetite in the ore together with higher MgO% and CaO% contained in the actinolite material.

Considering these conditions, selective mining and blending of material from both pit areas will be carried out in order to:

• Mine only hematite ore in the first year of operation with magnetite ore in the subsequent years following the commissioning of the magnetite plant;

• Maintain the MgO and CaO levels under 4.5% in the feed and 0.3% in the concentrate;

• Ensure that concentrate produced contains less than 4.5% SiO2;

• Optimize the use of high weight recovery ore to maximize the concentration ratio for project net present value optimization;

• Feed the mill with constant ore hardness for a stable process plant operation.

18.1.3.4 Mining Plan

As the top of the orebody is near to the surface, little pre-production stripping will be required in preparation for mining. However, sufficient waste rock will be mined during the pre-production period to build the tailings dam, ROM pad and other general construction work such as site access roads.


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Generally, the most easily accessible areas of the deposit will be scheduled for development in the initial years as this will maximize cash flow. However, some less easily accessible areas may be mined in the initial phase at a slightly higher unit cost to meet blending requirements. Pit staging using pushback technique will be used as more stripping work needs to be done with the deepening of the pit in the northeast side. As a general rule, waste stripping in each mine phase or pushback should be kept relatively constant with respect to mine fleet requirements due to increasing hauling distance and pit depth.

For blending purposes, it is likely that three to four production faces will be developed in ore at any one time on 2 to 3 benches. A 20-year mine production has been prepared on an annual basis in the first 5 years and by 5-year increments in the subsequent years. The average total combined ore and waste movement is fairly constant, at approximately 30-31-million tonnes per annum during the first 5 years of operation.

Table 18-4 summarizes the 20-year mine production plan for the development of the Bloom Lake Project.

For reference purposes, pit layouts at the end of years 1, 2, 3, 4, 5, 10, 15 and 20 extracted from the April 2006 Feasibility Study are shown in Figures 18-5 to 18-12.


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Table 18-4 : 20-Year Annual Mine Production Schedule

8 Mtpy Iron Concentrate

Inferred Waste Total S.R.Year (kt) Tfe FWR MagFe CaO MgO Mag (kt) Tfe FWR MagFe CaO MgO Mag (kt) Tfe WR MagFe CaO MgO Mag AFWR (2) (kt) (kt) (kt) (kt) (t/t)

(%) (%) (%) (%) (%) (%) (%) (%) (%) (%) (%) (%) (%) (%) (%) (%) (%) (%)Pre-Prod 9 000 -

3 months 14 116 32.42 39.65 2.24 0.17 0.32 3.10 5 184 32.34 39.55 5.42 1.25 1.36 7.48 4 825 32.40 39.62 3.09 0.46 0.60 4.28 41.50 2 002 61.56 2 158 2 220 0.461 (1) 14 116 32.42 39.65 2.24 0.17 0.32 3.10 5 184 32.34 39.55 5.42 1.25 1.36 7.48 19 300 32.40 39.62 3.09 0.46 0.60 4.28 41.50 8 010 246.19 8 632 8 878 0.46

2 12 789 31.93 40.84 6.80 0.68 1.09 9.40 6 511 30.86 39.38 10.57 1.90 2.22 14.62 19 300 31.57 40.34 8.07 1.09 1.47 11.16 41.50 8 010 413.23 10 588 11 001 0.57

3 15 536 31.27 39.60 6.63 1.45 1.74 9.17 3 764 30.30 38.51 9.17 2.26 2.48 12.67 19 300 31.08 39.39 7.13 1.61 1.88 9.85 42.48 8 199 34.22 10 774 10 808 0.56

4 17 181 30.56 38.78 8.98 1.91 1.99 12.41 2 119 29.47 37.19 8.09 2.18 2.26 11.18 19 300 30.44 38.60 8.88 1.94 2.02 12.27 41.58 8 025 5.18 11 575 11 580 0.60

5 17 315 30.56 38.82 9.34 1.89 1.96 12.91 1 985 29.38 37.06 8.49 2.06 2.14 11.73 19 300 30.44 38.64 9.25 1.91 1.98 12.79 41.58 8 026 5.86 11 574 11 580 0.60

6-10 86 096 30.50 38.73 8.37 1.67 1.81 11.58 10 799 29.11 36.39 6.92 2.19 2.31 9.55 96 895 30.34 38.47 8.21 1.73 1.87 11.35 41.47 40 183 29.75 85 277 85 307 0.886 17 315 30.56 38.82 9.34 1.89 1.96 12.91 1 985 29.38 37.06 8.49 2.06 2.14 11.73 19 300 30.44 38.64 9.25 1.91 1.98 12.79 41.58 8 026 5.86 14 662 14 668 0.767 17 210 30.65 38.98 8.71 1.65 1.81 12.05 2 090 29.14 36.70 7.21 2.18 2.34 9.95 19 300 30.48 38.73 8.55 1.71 1.87 11.82 41.67 8 042 4.25 16 208 16 212 0.848 17 142 30.70 39.08 8.31 1.50 1.72 11.49 2 158 29.00 36.49 6.45 2.25 2.45 8.91 19 300 30.51 38.79 8.10 1.58 1.80 11.20 41.72 8 051 3.22 16 981 16 984 0.889 17 142 30.70 39.08 8.31 1.50 1.72 11.49 2 158 29.00 36.49 6.45 2.25 2.45 8.91 19 300 30.51 38.79 8.10 1.58 1.80 11.20 41.72 8 051 3.22 18 139 18 142 0.9410 17 287 29.88 37.70 7.20 1.80 1.85 9.95 2 408 29.00 35.20 5.98 2.23 2.16 8.27 19 695 29.77 37.40 7.05 1.86 1.88 9.74 40.68 8 012 13.20 19 288 19 301 0.98

11-15 87 561 28.77 35.41 6.73 2.36 2.11 8.84 14 269 28.44 33.32 6.35 2.38 1.99 8.78 101 830 28.91 35.35 6.70 2.37 2.10 8.88 39.45 40 173 128.40 82 571 82 700 0.8111 17 664 27.63 34.42 5.48 2.17 1.98 7.56 2 786 29.01 33.79 5.47 2.21 1.84 7.56 20 450 28.79 35.54 5.64 2.22 2.00 7.80 39.30 8 037 26.50 17 356 17 383 0.8512 17 664 28.75 35.82 5.67 2.22 2.02 7.84 2 786 29.01 33.79 5.47 2.21 1.84 7.56 20 450 28.79 35.54 5.64 2.22 2.00 7.80 39.30 8 037 26.50 16 742 16 769 0.8213 17 664 28.75 35.82 5.67 2.22 2.02 7.84 2 786 29.01 33.79 5.47 2.21 1.84 7.56 20 450 28.79 35.54 5.64 2.22 2.00 7.80 39.30 8 037 26.50 16 947 16 974 0.8314 17 293 29.34 35.50 8.34 2.59 2.27 10.42 2 952 27.62 32.64 7.62 2.63 2.20 10.54 20 245 29.09 35.08 8.24 2.59 2.26 10.44 39.67 8 031 25.32 15 968 15 994 0.7915 17 276 29.36 35.48 8.46 2.61 2.28 10.53 2 959 27.56 32.59 7.71 2.64 2.21 10.66 20 235 29.10 35.06 8.35 2.61 2.27 10.55 39.68 8 030 23.56 15 557 15 581 0.77

16-20 86 364 29.36 35.48 8.46 2.61 2.28 10.53 14 791 27.56 32.59 7.71 2.64 2.21 10.66 101 155 29.10 34.49 8.22 2.57 2.23 10.38 39.69 40 150 117.81 109 067 109 185 1.0816 17 276 29.36 35.48 8.46 2.61 2.28 10.53 2 959 27.56 32.59 7.71 2.64 2.21 10.66 20 235 29.10 35.06 8.35 2.61 2.27 10.55 39.68 8 030 23.56 21 830 21 854 1.0817 17 259 29.36 35.48 8.46 2.61 2.28 10.53 2 956 27.56 32.59 7.71 2.64 2.21 10.66 20 215 29.10 32.21 7.68 2.40 2.08 9.68 39.72 8 029 23.56 22 213 22 237 1.1018 17 276 29.36 35.48 8.46 2.61 2.28 10.53 2 959 27.56 32.59 7.71 2.64 2.21 10.66 20 235 29.10 35.06 8.35 2.61 2.27 10.55 39.68 8 030 23.56 22 437 22 461 1.1119 17 276 29.36 35.48 8.46 2.61 2.28 10.53 2 959 27.56 32.59 7.71 2.64 2.21 10.66 20 235 29.10 35.06 8.35 2.61 2.27 10.55 39.68 8 030 23.56 21 426 21 449 1.0620 17 276 29.36 35.48 8.46 2.61 2.28 10.53 2 959 27.56 32.59 7.71 2.64 2.21 10.66 20 235 29.10 35.06 8.35 2.61 2.27 10.55 39.68 8 030 23.56 21 161 21 185 1.05

Total 336 958 29.99 37.29 7.59 1.96 1.91 10.09 59 422 28.89 35.16 7.33 2.29 2.15 10.13 396 380 29.88 36.91 7.60 2.02 1.95 10.17 40.56 160 774 980.64 330 058 331 038 0.84

(1) Magnetite Recovery Plant Commissioned by the middle of Year 3(2) AFWR: Adjusted Formula Weight Recovery for MgO% content (see details in FS Report)

Proven Reserve Total ROM Stripping

Concentrate Probable Reserve


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Figure 18-5 : End of Year 1


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18.1.3.5 Waste Dumps

During the construction period, a quantity of waste equivalent to 1.35 million bank cubic meters (bcm) of backfill material will be required for site construction, tailings dam and access roads. Two waste dumps will be built in the early years of pre-production stripping to prevent water from surrounding two lakes from flowing into the pit. These dumps will be located at:

• Confusion Lake (during mine access road construction);

• West Overburden Dump (5.4 million bcm capacity);

The main waste disposal dump will be constructed North of Bloom Lake pit and West of Mazare Lake. The total waste dump capacity has been estimated at 204.3 million bcm from 563.8 million tonnes of waste rock using an average swell factor of 25% with compaction. The waste dump can be elevated several meters on the north-western side for additional capacity if needed.The waste dump capacity has been estimated using a swell factor of 30%.

The design parameters are as follows:

- Face angle: 35°

- Overall slope: 20o

- Bench height: 20 m

- Berm width: 20 m

No allowance has been made for backfilling in mined-out areas.

18.2 Production Equipment Selection

The selection of the mine equipment for the Bloom Lake mine is made using the following criteria:

• An average annual material movement of 30-31-million tonnes in the first 5 years of operation, both run-of-mine and waste combined;

• A truck size of 240 t (218 tonnes);

• The main loading equipment is the hydraulic shovel with a lifting capacity of 64 tonnes per bucket (ONK 340).

• The equipment fleet must have a minimum of one (1) front end loader having a lifting capacity of 45 tonnes per bucket (Letourneau Model L-1850 or similar), as a supplementary production loading equipment. The flexibility of the loader with its fast response time in replacing a shovel or for a particular blending requirement, justifies its use;

• The drills are electric with a capacity to drill 15” diameter blast hole with high pull down pressure (P&H 120-A or similar);


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The major mine equipment will be leased and operated and maintained by the owner’s personnel.

Average hauling distances and lifts for ROM and waste to primary crusher to dumps respectively are presented in Table 18-5.

Table 18-6 provides a list of the expected major mine equipment.


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Table 18-5 : Haul and Lift Distances (m)

Year 1 2 3 4 5 Y 6-10 Y 11-15 Y 16-20 ROM Bench 760-705 746-705 732-704 732-704 718-705 704-705 676-705 662-705 Grade Level 305 460 395 395 489 541 388 570 Uphill 8% 350 525 Downhill 8% 700 525 350 350 175 Downhill 1% 1801 1801 1801 1801 1801 1801 1801 1801 Flat 300 300 300 300 300 300 300 300 Primary Crusher 200 200 200 200 200 200 200 200

Waste Pignac Lake

South dump

South dump

South dump

North dump

North dump

North dump

North dump

Grade

Bench 760-760 746-760 732-760 732-760 718-704-

690 704-690 676-704-

690 662-704-

690 Level 305 300 200 200 314 388 570 Uphill 8% 175 350 350 350 525 Downhill 8% 350 175 175 175 Downhill 1% Flat 391 1609 859 859 878 752 1284 2001 Waste Dump 200 200 200 200 200 200 200 200


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Table 18-6 : Major Mining Equipment Fleet

Period Years 1 to 5 Years 6 to 15 Years 16 to 20

Haul trucks 8-10 12-13 16

Shovels 2 2 2

Blast Hole drills 2 2 2

Wheeled loaders

2 2 2

* Float + auxiliary equipment

18.3 Service Equipment

The following service equipment was selected to carry out the routine road maintenance and other miscellaneous work within and around the mining areas: a) Track-dozers: one (1) CAT D-10 or equivalent with ripper and one (1) Cat D-9 or equivalent are

used to maintain the waste dumps as well as to perform general work within the mine. Track-dozers are used on ore stockpiles, waste dumps and in the pit when required;

b) Graders : two (2) Cat 16G or equivalent are used to maintain the roads of the entire mine site;

c) One (1) truck with interchangeable boxes depending on the season: a box to hold crushed stone for icy road conditions in winter, and one box to carry water for dust abatement in the summer;

d) One (1) truck for fuel and lube will be used to service equipment directly in the mine, preferably during the lunch breaks;

e) Secondary drilling of oversized boulders is performed by a contractor;

f) Front-end loaders: two (2) Cat 980 or equivalent are used for road maintenance, blast hole stemming and snow removal. The units will be complete with quick release couplings to allow easy changes from one attachment to another. Attachments include a bucket designed for blast hole stemming and a snow removal bucket;

g) Other service equipment: boom truck, pickup truck, welding truck, fire truck, ambulance, tire handler and lineman’s truck to insure the proper operation of the mine.

h) A radio communication system to link all the vehicles together and to insure that the equipment is effectively managed and truck dispatching controlled.


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18.4 Process Design

18.4.1 Design Ore Type

Drill core and bulk samples indicate that there is little variation in the total iron content of the deposit but that there are wide localized variations in the percentages of magnetite and of actinolite. Based on the mineral reserves, an average mill feed grade of 30% total iron and 10% magnetite content has been used in this study.


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18.4.2 Design Criteria

The basic process design criteria directed by the client, taken from testwork or assumed at the outset of the Feasibility Study were the following:

• Production will be 8-million tons per year of concentrate from a single line but the layout should allow for future expansion.

• The design mill feed rate is 2372 dry tonnes per hour.

• The plant will operate 365 days a year with 93% equipment utilization.

• Concentrate grade should be 65% Fe or greater with a maximum of 4.5% SiO2.

• Weight recovery for year 1 operation (gravity separation only) will be 38%. Weight recovery in year 3 onward with gravity and magnetic concentrate production will average 41.4%.

• The iron minerals are liberated at 410 microns (35 mesh).

• The power consumption in primary grinding is 4.2 kWh/t of concentrator feed. Power consumption for magnetite regrinding is 23.8 kWh/t of feed to the regrind circuit (Lakefield testwork, 2005).

• Half of the magnetite is recovered in the spiral concentrate and half in the magnetite concentrate.

• Concentrate will be dried to less than 2% moisture in the winter months to avoid freezing in the railcars.

18.5 Metallurgical Processing Plant

The process flowsheet is shown in block form in Figure 18-13 below. Detailed process flowsheets are shown in Figures 18-14 to 18-23.


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Figure 18-13 : Process Block Diagram

CRUSHING AND STOCKPILING

GRINDING AND

CLASSIFICATION

GRAVITY CIRCUIT

HEMATITE

CONCENTRATE

FILTRATION

MAGNETIC SEPARATION

CIRCUIT

MAGNETITE

CONCENTRATE FILTRATION

TRAIN LOAD-OUT

STATION

TAILINGS DEWATERING

AND THICKENING

TAILINGS DISPOSAL

& RECLAIM WATER

PUMPHOUSE

PROCESS WATER

RESERVOIR

TAILS

CONCENTRATE CONCENTRATE

RUN -OF- MINE ORE

CRUSHED ORE

GROUND ORE

TAILS

DECANT WATER

THICKENER OVERFLOWWATER

FRESH WATER MAKE-UP

WATER TO PROCESS

DENSIFIED TAILINGS


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Figure 18-14 : Crushing and Ore Stockpiling


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Figure 18-15 : Grinding


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Figure 18-16 : Spiral Plant


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Figure 18-17 : Concentrate & Tailings Dewatering, Loadout


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Figure 18-18 : Magnetic Plant


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Figure 18-19 : Water Supply


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Figure 18-20 : Reagent & Services (Flocculant)


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Figure 18-21-: Reagent & Services (Air)


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Figure 18-22: Reagent & Services (Steam generation)


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Figure 18-23: Reagent & Services (Fuel oil)


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18.5.1 Crushing, Ore Stockpiling and Grinding

Ore from the mine will be delivered in 240 t trucks to the gyratory crusher. The crusher will be provided with two dump points though only one will be used initially. A hydraulic rock breaker will be installed adjacent to the crusher to break lumps too large to enter the crusher. Crusher, discharge feeder and belt conveyor parts will be removable for maintenance using a 75 t capacity overhead crane or a monorail. The crusher building will be enclosed and provided with a venturi-type wet scrubber and an air make-up unit. Scrubber effluent, floor washdown water and drainage will be collected in a sump and pumped to the reclaim tunnel sump.

Ore crushed to -200 mm (8”) will be fed by an 1830 mm wide by 6700 mm long apron feeder with a design capacity of 3900 tones per hour and driven by a 75 HP variable speed hydraulic drive onto the 1524 mm (60”) wide fixed-speed crushed ore belt conveyor. The conveyor with walkways on both sides will be enclosed in an unheated gallery and will discharge onto the crushed ore stockpile. The 26 000-tonne live capacity of the stockpile will be sufficient for an 11-hour operation..

Ore will be withdrawn from the stockpile by two 1830 mm wide by 7000 mm long apron feeders (one operating and one standby) located inside a heated reclaim tunnel. The two feeders, each driven by a 125 HP variable speed drive, will feed crushed ore onto a 1524 mm wide mill feed conveyor. Each feeder will be able to feed the autogenous (AG) mill at a rate of up to 2600 tones per hour with an average rate of 2372 tones per hour. A belt magnet will be installed over the mill feed conveyor to remove scrap steel and protect the belt.

A venturi scrubber and air make-up unit will be installed close to the feeders. Scrubber effluent, wash down water and drainage water will be collected in a sump and pumped to the autogenous mill feed chute. Monorail hoists will be installed around the feeders for maintenance of the equipment.

The 10.97 m diameter x 6.10 m long mill will be driven by dual 5590 kW (7500 HP) motors. The mill feed tonnage will be controlled electronically by varying the feeder speed with a signal from the belt scale. A belt magnet will be installed over the mill feed conveyor to remove scrap steel and protect the belt.

The reclaim tunnel sump pump will pump scrubber effluents, wash down water and drainage water to the scalping screen undersize pump box.

Both the crushed ore conveyor and the mill feed conveyor will be provided with walkways along each side. Conveyors and walkways will be enclosed for protection from wind and snow to facilitate maintenance.

Process water will be added to the mill to maintain a 72 percent solids density at the mill discharge.

Product discharged from the mill will pass over two scalping screens to remove large material. The scalping screen undersize material will be collected in pump boxes and pumped together with the cleaner spiral tailings stream to four vibrating classification screens making a nominal 420 microns (35 mesh) separation. Water sprays will be provided over the screens decks to aid the separation of fines if necessary. Oversize from the scalping and classification screens will be collected on conveyor belts and returned to the feed end of the mill. Material passing through the classification screens will be diluted with the recleaner spiral tailings and pumped to the spirals circuit.


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An overhead crane will be used for maintenance in the grinding bay with two smaller cranes provided for maintenance of the screens and recirculating conveyors. A liner handling machine will be provided to facilitate liner changes.

Spillage and wash-down water in the grinding and screening area will be directed towards sump pumps and pumped to the scalping screen undersize pump box.

18.5.1.1 Spiral Plant, Concentrate and Tailings Dewatering, Load Out

Slurry from the grinding circuit will be distributed to a total of 32 banks of rougher spirals arranged in four lines. Each bank will have two rows of eight back to back double start spirals Concentrate from each bank of rougher spirals will be combined and distributed by a 24-way rubber-lined cleaner spiral distributor to feed a bank of twelve double start back to back cleaner spirals. Concentrate from each cleaner spiral will feed directly to a recleaner spiral below for final upgrading. In total, there will be 1024 rougher spirals, 768 cleaner spirals and a similar number of recleaner spirals. Tailings from the recleaner spirals will be recirculated to the rougher spiral feed through gravity flow to the classification screen undersize pump boxes. Tailings from the cleaner spirals will return, also by gravity, to the scalping screen undersize pumpboxes to dilute the feed to the classification screens.

A monorail hoist will be provided over each line of spirals for maintenance.

The concentrate from the recleaner spirals will be dewatered on four horizontal pan filters, one filter per spiral line. Filters will be fitted with steam hoods to dry the concentrate during the winter months to facilitate transportation. Vacuum will be provided by four vacuum pumps connected onto a common header. Space will be left for the future installation of a fifth vacuum pump. Concentrate will be discharged from the filters to a 1067 mm (42”) wide belt conveyor and transported to a 24,000 t capacity silo at the train load-out station. The silo will hold approximately twenty-four hours concentrate production. Railway cars, drawn by a locomotive, will be loaded through a retractable gate mechanism, whilst in motion at a rate of 6000 tph or or one ralicar per minute. The load-out requirement is for approximately 1 train per day (240, 100-tonne capacity cars per train).

If the silo is full, filter cake will be discharged to an 80,000-tonne capacity emergency concentrate stockpile in the load-out area. When the level in the silo is sufficiently low, concentrate will be reclaimed by the front-end loader and returned to the conveyor feeding the silo.

In Phase 1 (spiral concentrate production only), rougher spiral tailings will be pumped to tailings dewatering cyclones. The overflow from these cyclones will be directed to a 31 m diameter high-rate thickener where flocculant will be added. The clarified thickener overflow stream will flow by gravity to the process water reservoir and constitute the main source of process water for the plant’s operation. The underflows from the dewatering cyclones and the thickener will be combined and pumped to the tailings disposal area where the solids will be allowed to drain. The decant water will be collected and pumped back to the process water reservoir for reuse.

Air compressors will provide compressed air for crusher and mill use. Part of the compressed air supply will be dried for operation of control valves and other instrumentation.


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Process water will be supplied to the mill from an in-ground reservoir by vertical turbine pumps. The main source of process water (85%) will be the tailings thickener overflow. Tailings decant water and fresh water make-up will constitute the remainder. Fresh water make-up will be negligible under normal operating conditions.

Two, light fuel #2 fired boilers will provide steam for concentrate drying and indirectly for building heating through heat exchangers and a closed-circuit glycol system. The two boilers will be rated at 23 000 kg/h.

Heavy oil #6 fuel-fired boilers will provide hot water for building heating and steam required for concentrate drying. Two 735 KW auxiliary hot water boilers burning #2 light oil will provide steam heat for the crusher building. The #2 light oil will also be used as a back-up supply to the main boilers, for the mine trucks, emergency generators and operation of the emergency fire water and fresh water supply pumps.

Stand-by generators will provide emergency power at the concentrator and crusher to operate critical items of equipment in case of failure of the main power supply.

18.5.1.2 Magnetic Plant

In Phase 2 of the Project (production of both spiral and magnetic plant concentrates), tailings from the four lines of rougher spirals will be passed through low intensity magnetic drum separators to recover any magnetite not picked up in the spirals. The recovered magnetic material will be pumped to hydrocyclones. Coarse, mixed grains in the cyclone underflow will be grounded to approximately 80% passing 44 microns (325M) in a ball mill and fed to a set of rougher drum separators. Rougher concentrate will be fed back to the cyclones. Fine magnetite in the hydrocyclone overflow will feed a set of finisher magnetic separators to produce a final magnetite concentrate.

Magnetite concentrate will be collected in an agitated storage tank and pumped to a disc filter for dewatering. As with the pan filters, the disc filter will also be fitted with a steam hood for winter operation. A dedicated compressor will be installed to supply air for snap-blow at the disc filter. Concentrate will be discharged to a collecting conveyor and transferred together with the spiral plant concentrate to the load-out silo.

The various tailings streams from the magnetite separation plant will be combined and pumped to the tailings dewatering cyclones used in Phase 1. In Phase 2, the underflow and the overflow from these hydrocyclones will be handled in the same way as for Phase 1.

18.5.2 Plant Water and Services

18.5.2.1 Fresh Water

Fresh water from Bloom Lake will be stored in the fresh water tank in the Concentrator. From the tank two distribution pumps, one operating, one stand-by, will pump the water to the points of use. To reduce fresh water requirements, spent cooling water will be returned to the gland seal water tank.


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18.5.2.2 Gland Seal Water

Pump seal water will be taken from the gland seal water reservoir. Two low pressure pumps will supply seal water to most of the pumps in the plant. Two high pressure water pumps will supply seal water to the tailings pumps.

18.5.2.3 Process Water

Vertical turbine pumps will supply the Concentrator with process water from an in-ground reservoir. Thickener overflow will be the principal supply of water to the tank. Reclaim water will be a secondary source. It will also be possible to add fresh water to the reservoir but this will not be necessary under normal operating conditions.

18.5.2.4 Compressed Air

One air compressor with integral dryer will provide compressed air at the crusher to meet requirements for service air and for air to operate control valves and other instrumentation. Two air compressors with integral dryers will provide compressed air for mill service use and for operation of control valves and other instrumentation. Two low pressure blowers (one operating) will supply compressed air at 40 kPa pressure to the horizontal table filters to clean the cloths.

18.5.2.5 Emergency Power Supply

A 1000 kW standby diesel generator will be on stand-by to provide emergency power to operate critical items of equipment in case of failure of the main power supply.

18.5.2.6 Mobile Equipment

A bob cat and front end loader will be used for clean-up work and material handling around the mill.

18.5.2.7 Process Control

The concentrator will be fully automated with operation controlled from a single control room adjacent to the primary grinding mill and with minimum operator involvement locally.

For process control and metallurgical accounting purposes, automatic samplers will be installed in the final tailings stream and on the concentrate collecting conveyor. Because of the difficulty in sampling material up to 200 mm in size from the mill feed conveyor and the presence of recirculating loads and flows in the grinding circuit, the mill feed head grade will be back-calculated from mass flow measurements and samples on the three streams: spiral feed, cleaner spiral tails and recleaner spiral tails.

During the construction of the magnetic plant, three more samplers will be added to sample the feed to the magnetic circuit, the magnetic concentrate produced and the combined spiral concentrate and magnetite concentrate on the concentrate transfer conveyor.


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Weigh scales will be installed on the mill feed conveyor and on the concentrate load-out conveyor to balance material tonnages into and out of the mill.

18.6 Tailings Disposal

18.6.1 Site Description

Some 230-million metric tonnes of inert mill tailings will be produced over 20 years and will be stored in a 741 ha area situated on the north side of the pit and west of the plant site. The topography is characterized by a low, poorly drained, gently undulating surface of overburden deposits with some bedrock outcrops.

The initial pipeline length will be 2 km but will be extended progressively to 6 km.

Information is presently collected and analyzed for this site and geotechnical parameters will be used to optimize the design criteria.

The site will be developed so as to optimize the use of the surrounding topography to accommodate the required tailings and water storage volumes, to maximize the safety of the installations (for example by locating spillways in rock or natural ground), and to minimize operating costs.

18.6.2 Fundamental Design Criteria and Assumptions

The site will comply with standard practices for industrial water impoundment and will

• be able to contain and treat, without spilling to the environment, an extreme design flood event,

• have an emergency spillway that is capable of safely evacuating surplus water flows up to the Probable Maximum Flood (PMF).

Based on guidelines from the (Ministère des Ressources Naturelles), water dams will have a minimum safety factor of 1.3 under static loading and 1.1 under dynamic loading.

The average surplus volume of 0.4 Mm3 of water retained annually in the tailings disposal area will be pumped out of the system during a period of up to 5 months per year. If required, the water will be treated to remove colour and turbidity before disposal

The retention time required to ensure that the effluent will comply with the environmental regulations is approximately 30 days and was based on preliminary laboratory test work. This must be confirmed by laboratory testing once the mill process starts operation.

Additional geotechnical and survey work will be required during the detailed engineering phase of this project.

18.6.3 Operation

By appropriate utilization of the local topography minimum dyking will be required to retain the tailings solids for the first twelve years of operation. After this time dykes will be required to


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prevent encroachment of the disposal area on the plant site and to maintain a drainage channel between the mine waste pile and the deposited tailings.

Surface run-off water and water draining from the deposited solids will be collected into a settling pond of 2 km long by 27 m high dam. The water level behind this dam will be maintained by pumping water on a continuous basis back to the process water reservoir at the Concentrator or by intermittent pumping to a clarification pond with discharge to the environment. Water pumped to the clarification pond will be treated chemically, if required, to flocculate fine solids and to remove the red colour typical of iron minerals.

Periodically, the solids which collect on the bottom of the clarification pond will be pumped back within the tailings disposal area.

The water dams will be made of rock fill with a center till core protected by filter and drain layers. The tailings dams will be built in lifts using the coarse fraction of the tailings from a starter lift of rock waste with filter.

A number of ditches will be dug around the tailings disposal area to divert clean surface water away from the site.

18.7 Infrastructure and Support Systems

18.7.1 Service Building

The 2485 m2 (35 m x 71 m) service building attached to the concentrator building will provide the following services:

• Unloading and warehousing area for parts and supplies; a boiler plant to provide steam for heating and filtercake drying; electrical/instrument repair shop; a first-aid room; offices for administration, purchasing, human resources, technical services (engineering and geology), training and plant operating personnel; a laboratory equipped for metallurgical testwork, wet and dry assaying; lunchroom, men’s and women’s change rooms, sanitary and locker facilities; a communications room; compressor room to provide service air and instrument air to the concentrator; a blower room to supply low pressure air to the concentrate filters; a fresh water storage tank and water treatment facilities, and an electrical room.

The steam plant will occupy the full 16.4 m height of the eastern end of the building. The warehouse and electrical shop will occupy two floors (9.5 m) in the western half of the building. Services in the remainder of the building will be arranged on three-floor levels.

18.7.1.1 Shop & Warehouse

The warehouse floor area will cover an area of 630 m2 (21 m by 30 m) and will be 9.5 m high. Trucks to be unloaded will descend a ramp to bring the truck bed level with the loading dock and floor inside the warehouse. A fenced outdoor warehouse yard will also be used to stock material.

18.7.1.2 Utilities Area

The 820 m2 utilities area includes the boiler room, fresh water storage tank and water treatment, blower and compressor rooms and the emergency MCC room. The emergency generator will be located outside the Service Building


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Two 50 000 #/h capacity fire tube boilers will supply low pressure steam to the concentrate filters and to the hot water heat exchangers for building heating. At peak load, both boilers will be in operation. Light fuel oil #2 will be used as fuel.

18.7.1.3 Emergency Vehicle Station

The emergency vehicle station is sized for an ambulance and one fire truck. The first-aid station will also be located in this area. Hot and cold water will be provided as well as emergency power.

18.7.1.4 Offices, Change Rooms and Lunch Room

An office space of 1379 m2 for administration, human resources, accounting, purchasing, engineering, plant operating and maintenance personnel will be provided on the second floor of the service building. Washrooms and a first aid room will also be located on this floor. Offices along the outer walls will be provided with windows. There will also be a direct access from the offices to the concentrator operating floor. Change rooms, showers and toilets for men will be located on the ground floor and on the first floor for women. A lunch room will be provided on the first floor.

18.7.1.5 Laboratory

The laboratory located on the ground floor will have 266 m2 of floor space for the preparation and analysis of samples by wet methods and XRF.. The preparation area will be equipped for splitting, drying, crushing, grinding, screening and filtering of samples from both the mine and the concentrator. A dust collection system will be provided in the preparation area. Fume hoods will be installed in the wet assay room. A storage room and shelving will be provided for samples and supplies. There will also be a direct access from the laboratory to the concentrator.

18.7.1.6 Heating, Ventilation and Air Conditioning

Systems are designed for outdoor temperatures of -40°C in winter and 17°C in summer and inside temperatures of 19-21°C. Fresh air changes vary from 1 in the offices to 10 in change rooms.

The shops, warehouse and concentrator will be heated with steam from the boiler plant by a central system for each sector, which includes supply fan, return/exhaust fan, steam heating coil, filter and air/air energy recovery system.

The office, laboratory and lunch room will be air conditioned by a variable volume central unit with a 700 kW steam heating coil and 40-tonne roof-mounted cooling unit. Heating of cold perimeter areas will be supplemented by using electrical baseboard heaters.

A steam aerotherm heater will be installed near each garage door in the service building and concentrator to compensate for the heat loss through air infiltration in winter.


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18.7.1.7 Water Distribution and Drainage Network

Hot and cold water will be distributed to all sanitary facilities in the office sector, warehouse and shops.

Emergency showers and eye-washes will be installed in the laboratory, shops, and warehouse and at the flocculant preparation area in the concentrator.

Drainage systems from wash bays in the truck shop and warehouse will incorporate sand and sludge interceptors, oil interceptors and water/soap settling tanks with oil skimmers.

Drainage from shop repair bays will pass through an oil separator only.

Effluents from the wash bays and repair bays will be discharged into the drainage ditch surrounding the plant site. Water in this ditch will pass through two cleaning basins before it is discharged into Confusion Lake.

18.7.2 Access and Site Roads

The sole access road to the Bloom Lake deposit will be from Highway 389, between the Mont-Wright mine and the town of Fermont. This 6 km access road will be provided with a barrier gate to control passage. A telephone and camera will link the gate with the security station in the Service Building. Other roads will be constructed from the concentrator to the mine, the crusher, along the route of the tailings line and to the freshwater pump houses and Bloom Lake.

18.7.3 Power Supply and Distribution

The major infrastructure required to develop the project is already present in the area. Two 17 km power lines will be constructed by Hydro-Québec (HQ). The first 11.5 km will run from the HQ NORMAND station alongside the 389 road to the mine power metering station at the junction of the 389 road and the plant access road. The second section of 6.5 km will be constructed alongside the mine access road to complete the circuit to the Bloom 34.5 kV main station. The power demand anticipated for the project 8 million tonnes per year of concentrate capacity is 40 MW with a power factor of 0.95.

The two Hydro Power Lines will terminate on a 34.5 kV power structure inside the mine main station and will be fed through a SF-6 gas insulated type 34.5 kV indoor switchgear. The main switchgear will be used to connect various 34.5 kV power line. A total of 22.5 km of 34.5 kV aerial lines will be built to supply power to various plant and mine loads such as: Guard gate, Crusher station, Fresh water pumping station, Reclaim pumping station, Lac Confusion pumping station, Fire pump station, Loadout station, Explosive storage, Garage station and Mining station.

The main 34.5 kV switchgear will also connect two 34.5-4.16 kV power transformers rated 21/28 MVA equipped with on-load tap changer. The two transformers will feed the main 4.16 kV switchgear.

The mine site will be supplied at 7.2 kV via two portable substations 34.5 kV to 7.2 kV, 7.5 MVA located at each end of the pit. The mine load will be connected to a total of five switch houses throughout the site and fed from a total of 8 km of 7.2 kV power lines. The 7.2 kV power lines will be built according to the standards applicable for the climatic conditions prevailing in the northern


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region of the province of Québec. In order to meet the safety aspect of the mining code, the phase-to-ground fault level of the 7.2 kV system will be limited to 25 A. This 7.2 kV system will be connected to a distinct remote ground mat. This technology has the advantage of limiting the ground potential rise to a prescribed value whether the ground fault appears on the 34.5 kV or the 7.2 kV systems.

The main 4.16 kV switchgear fed from the two 21/28 MVA transformer will be located in the main electrical room of the concentrator building. The 4.16 kV switchgear buses will be compensated to maintain 0.95 PF with the help of two 8 Mvar capacitor banks tuned to the 4.85th harmonic. The 4.16 kV power system phase-to-ground fault level of each will be limited to 25 A for personnel safety, thus limiting damages to the equipment during a ground fault condition.

The main 4.16 kV switchgear will feed various 4.16 kV motor control centers, 4.16 kV variable frequency drives (VFD) and 4.16 kV to 600 V substations through the plant. The AG Mill twin 7,500 HP motor will also be connected to the main 4.16 kV switchgear. The AG Mill 7,500 HP motors are wound rotor motors equipped with liquid rheostat to limit power inrush on startup as required by Hydro-Quebec. The AG Mill motors will be connected to appreciatively 3 MVA capacitor bank to help raise the power factor to Hydro-Quebec prescribed level.

Motor loads of 300 HP and higher, using across-the-line starters, will be supplied at 4.16 kV. Smaller motors and motors with a Variable frequency drive (VFD), with loads up to 1000 HP, will be supplied from 600 V substations.

The 600 V substations are sized at 2.5/3.3 MVA to feed the various plant loads as follows: Crusher CD-36, Concentrator CD-31, CD-31A, CD-33, CD-40, CD-40A, Loadout CD-35 and Service building CD-37.

The various 600 V power supplies will operate with a resistance grounded neutral, limiting the phase-to-ground fault to 5 A. Such a configuration avoids service outages for a single phase-to-ground fault and is considered safe for the staff responsible for operations.

The 600/347 V lighting and building services will use small isolating transformers, thus limiting the 600 V fault level to a safe level of less than 10 kA, which allows for the use of distribution of equipment currently used in large industrial complexes.

Power supply for the workers residential camp will come directly from Hydro-Québec since the camp will be installed in the town of Fermont.

The emergency power supply for the concentrator and service building will be assured by a 1000 kW diesel generator set which will be installed adjacent to the Service Building and will supply critical loads such as lighting and the steam plant during a power outage. A 600 kW diesel generator set will supply emergency power to the crusher building.

Two small 6 kVA UPS will supply the control system units (PLC, DCS, HMI and the servers).

18.7.4 Control System

The control system for the Bloom Lake operation will consist of a highly integrated programmable logic control/human machine interface (PLC/HMI) system. Programmable controllers will be recent technology offering high performance in an easy-to-use environment. Ethernet communication will be used in addition to fieldbus networks. Input/output will be distributed through the installation.


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SCADA software (Supervisory Control and Data Acquisition) will be based on server/multi-user applications for the supervisory level monitoring and control applications.

Smaller graphic HMI terminals will be installed locally for smaller applications. PLC/HMI will also be installed for major equipment items (crusher, AG mill and ball mill). PLCs, HMI and graphic terminals will be linked through the Ethernet plant-based network.

A fibre optic network will be installed across the plant facilities to integrate the various control, telephone, security and fire alarm systems. Fibre optic cables will be run in cable trays or on 34.5 kV transmission lines.

During the first year of operation, a dedicated historian system will be installed to provide rapid access to operational data for production reports and other needs.

18.7.5 Fresh Water Supply

Fresh water will be required for make-up to the boilers and for domestic consumption. The average hourly consumption on an annual basis will be approximately 6 m3, the peak consumption will be approximately 40 m3/h.

Fresh water will be supplied to the fresh water tank at the Concentrator by gravity flow from Bloom Lake The 1.5 km long, 152 mm diameter HDPE fresh water pipeline will be buried to prevent the water from freezing in winter.

18.7.6 Reclaim Water Supply

Reclaim water will be pumped from the decanted water in the tailings retention area. There will be two vertical turbine type reclaim pumps with 400 HP drives, one operating and one standby. Each pump will have a capacity of 1385 m3/hr.

The pipeline will be constructed of 610 mm diameter HDPE pipe for the entire length (approximately 4.5 km) and will be buried. In the event of a power failure, water in the line will drain back to the decant water pond. No relief valve will be required on the line due to the uniform slope of the pipe.

The reclaim water line will terminate at the process water reservoir outside the Concentrator. A branch line will supply reclaim water to the industrial water tank.

18.7.7 Fire Protection

The fire protection system will include fire water pumps, a fire water distribution network, fire water hose stations and water sprinkler systems.

A water sprinkler system will be installed over covered conveyor belts and over the lubrication and hydraulic systems in the process areas.

Fire water pumps will be located in a pump house at Bloom Lake. There will be three pumps, two main pumps and one jockey pump to maintain the pressure in the fire water pipe network. One of


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the main pumps will be driven by an electric motor and the other will be a diesel engine driven with a control panel for automatic start.

Alarm signals will be automatically transmitted to the security station in the service building.

18.7.8 Fuel Storage

Number 2 light fuel oil and gasoline will be delivered to the site by road tanker.

The fuel storage system will include:

• Four 50,000 L capacity tanks for light fuel oil #2 for the concentrator heating and steam generation. The minimum on-site storage capacity is 4 days.

• One 10,000 L tank for light fuel oil #2 for the crusher heating. The minimum on-site storage capacity is 5 days.

• Two 50,000 L capacity tanks for diesel fuel for the emergency generator and the diesel fire water pump.

• One 5,500 L capacity tank for gasoline with a storage capacity for 14 days requirements. Fuel oil tanks will have integral holding sections to retain leakage and prevent contamination of the ground.

A gasoline fuel station for pick-up trucks and other vehicles will be located close to the storage tank area.

If the ore in the crushed ore stockpile is frozen and cannot be withdrawn by the apron feeders, propane burners will be used to thaw the material. Propane will be delivered by tanker to a 22,700 L storage tank. This will be provide for over 22 days of continuous gas supply to the burners.

18.7.9 Communications System (Local and External)

The communications system will comprise a vocal radio communication system, a digital radio communication system from trucks, shovels and drills, a telephone system with a minimum capacity and modular expansion, an FM broadcast system for music in all vehicles and one transmitter/receiver station including antenna tower and housing for radio communication equipment. The location of the tower will be selected to optimize communication transmissions.

18.7.10 Effluent Water Treatment

Ageing tests were carried out at SGS Lakefield on two shaking table tailings samples to determine whether water from the Bloom Lake tailings pond would have to be treated prior to being released to the environment. The tailings slurry samples contained 2.8% and 6.7% solids. Each slurry sample was divided into several sub-samples. Some sub-samples were treated with a flocculant and coagulant and others were not. The sub-samples were allowed to settle undisturbed over several days. Sub-samples were analysed on Day 0, 2, 14 and 30 for total suspended solids, color, PH, turbidity, conductivity, EMF and heavy metals. .


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After 2 days, the total suspended solids for both slurry samples were well below the Directive 019 limit of 25 mg/L. Heavy metals were also well below the Directive 019 limits. The PH was consistently around 8. Based on visual observation, all sub-samples were clear after 14 days or earlier (i.e. no red color). Given that the residence time of the slurry water at the polishing basin of the Bloom Lake tailings pond will be at least 30 days before being released to the environment, it is recommended that no treatment of the effluent water be carried out. Based on the Lakefield results, no water treatment would be required to comply with the Directive 019 limits of MDDEP and that no colorants would be required for “red water” elimination.

It is recommended that during the initial year of the Bloom Lake operation, no effluent water treatment facilities will be installed. The polishing basin will be built as originally proposed, but the monitoring pond will be deferred until after effluent sampling is carried out. If the suspended solids surpass the allowable limit, preliminary scoping studies indicate that the effluent can be treated with the coagulant Flomin 45VHM and anionic floculant Flomin 910 SH at dosages of 0.5 ppm and 0.3 ppm, respectively.

18.7.11 Sanitary Treatment and Waste Disposal The sewage disposal system is designed to accommodate the concentrator and service building and the truck shop. The sewage treatment system will include collecting and pumping stations, a septic tank and aerobic treatment stages. The treated water will meet mandated discharge parameters and will be released to the environment by ground infiltration.

An electric toilet will be provided at the crusher building.

Solid waste materials which cannot be recycled and domestic waste will be sent to the Fermont municipal waste site.

Used oil and lubricants and oil skimmed from ponds will, if possible, be burned in the boiler or disposed of off-site by a recognized waste disposal company.

18.7.12 Camp Accommodation Temporary lodgings to accommodate construction workers and a permanent facility to accommodate workers at the mine will be built within the town of Fermont. Potable water supply, sewage treatment, fire protection and electric utilities will be provided by the town.

Workers in the temporary camp will be housed in trailers with six rooms to a unit. In the permanent facility, each unit will accommodate eight people at a time in single rooms.

18.8 Railway Transportation and Port A new 31 km long single-track railway line will connect the mine site with the existing common carrier near Labrador City. A marshalling yard near the junction with the existing railway and a section of double track for passing will add another 6 km of track. All of the line will be in the Province of Newfoundland and Labrador.

The terrain through which the proposed railway line will pass is low lying with boreal forest, muskeg and a number of lakes. From an elevation of 700 m at the Quebec-Labrador border the line will descend to 540 m at the junction. Construction of the line will require the erection of several bridges and the installation of a level crossing where the railway line traverses Highway 389.

Apart from a short section of the line near Labrador City the railway is remote from any residential areas.


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Measures will be taken to mitigate disturbances to wildlife, wetlands and water courses by noise, dust generation or leakage of fuel oil or lubricants during the construction and operating periods.

Empty cars will pass beneath the concentrate load-out silo and will be loaded automatically at a rate of one car per minute. Loaded cars will be held on a loop line before beginning the outbound trip. With a mine production rate of 8-million tones per year of concentrates, train frequency will average 1.0 per day in each direction (240 cars per train, 100 tones of concentrate per car basis). Other traffic will be variable and consist mainly of track inspection and maintenance vehicles. Fuel oil and other operating supplies will be brought to the mine site by road.

For the purpose of the study a train operating cycle of 72 hours has been assumed. This cycle time includes the time of loading and unloading operations, inspection and maintenance, fuelling, crew changes and miscellaneous delays.

Use of a cable belt conveyor to transport the concentrate from Bloom Lake to load-out facilities and a short section of track to connect to the QNS&L railway at Labrador City was considered as an alternative to the 31 km long rail line but was rejected because of logistics.

Concentrate will be unloaded from railcars at the Port of Sept- Îles and either loaded directly onto a vessel or stored in the yard to be reclaimed and loaded onto a vessel at a later date. New equipment to be installed will include a car dumper to accommodate 86 iron ore car trains, a 5,500 t/h ship loader and a concentrate reclaim system. In addition space will be required for a ± 1,000,000 t capacity stockyard.

CLM is currently working with the Sept-Îles Port Authority, who is committed to assist and help develop emerging new companies in the area, mainly by giving access to port facilities and maintaining these facilities (dredging and dock maintenance).

Assets owned and work undertaken by the Port of Sept-Îles:

Docks Q30 and Q31 are owned by the Port of Sept-Îles and are used to accommodate other users. Wabush Mines presently own and operate two ± 3,500 t/h ship loaders on Dock Q30.

CLM will use the land owned by the Port of Sept-Iles for its facilities (car dumper, stock yard and ship loader).

CIMA+ and Lassing Dibben are presently working with CLM to develop a concept for car dumping, stockyard and ship loading facilities (see figures 18-24 and 18-25). The flow sheets for potential scenarios will be evaluated and reviewed with the Sept-Îles Port Authority and also with Logistec who are interested in building and operating the required facilities at Pointe Noire on an operating lease basis. The first phase is planned to be operational by September 30, 2009.


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Figure 18-24 : Ship loading proposed site plan (Phase 1)


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Figure 18-25 : Ship loading proposed site plant (Phase 2)


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Figure 18-26 : Proposed ship loader arrangement


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18.9 Environment

18.9.1 Applicable regulations

Since the production rate is more than 7000 tonnes per day, the project will be subject to an Environmental Impact Assessment (EIA) in the Province of Quebec. No separate federal EIA will be required for the disturbance to water bodies even though within the context of the present project, 300 ha of fish habitat will be affected.

Construction of the railway line which is subject to authorization from the Newfoundland & Labrador government will also require an Environmental Impact Assessment. Although Newfoundland & Labrador’s environmental regulations are not identical to the Quebec regulations, the procedure and time limits are similar.

The principal environmental issues expected to be addressed in the environmental studies are:

• Land usage • Air quality and ambient noise • Destruction of fish habitats

To meet provincial and federal requirements, an on-going program will be implemented during both the construction and operational periods to monitor the effects of the Project on geotechnics, water quality, effluents, groundwater, fish population, benthic invertebrate communities and sediment quality and to include mitigation or compensatory measures as applicable.

A rehabilitation plan will be submitted to the Ministry of Natural Resources and Wildlife prior to the start of production. Monetary provision will be provided in the first eight years of operations.

18.9.2 Native Land Status

Following discussions with ITUM (Innu Takuaikan Uashatmak Mani-Utenam representing the aboriginal people), an agreement was signed with CLM on May 30, 2008.

18.9.3 Capital Cost Estimate

The capital cost estimate presented in this report is obtained from the Bloom Lake Iron Ore 8.0 mt/y feasibility study and is based on written budget quotations from equipment suppliers and other consultants. The hourly labour rate for construction and installation manpower was obtained from a contractor familiar with the working conditions of the Bloom Lake area. This rate includes the following costs: salaries, benefits, mobilization and demobilization, operation of construction equipment, consumables, supervision, miscellaneous and profit. All quotations received in Canadian dollars were converted to U.S. dollars using an exchange rate of 1.10 $CND/ 1.00 $US.

The capital investment for the Bloom Lake project, including equipment and materials during the construction period (years 0-1) and working capital, is estimated to be 463.6 million USD. The total capital investment over a 20-year period is 557.2 million USD


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A summary of the investment costs is given in Table 18-7.

Table 18-7 : Investment Cost Estimate (US Million Dollars)

Direct Costs Year 0, 1, 2 Yrs 0 to 20

Construction Phase $309.5 $309.5 Allowance Train Loading Silo $22.7 $22.7

Magnetic Separation Plant $0 $27.7

Sub-Total Direct Cost $332.2 $359.9

Owner’s Cost $12.7 $12.7 EPCM $35.1 $35.1

Contingency $27.3 $27.3

Sub-Total (w/ Direct, owner’s, EPCM) $407.3 $435.0

Other Costs Schedule Acceleration $9.1 $9.1

Tailings Impoundment Dams & Ditches $14.1 $14.1 Environment (Mitigation) $4.5 $4.5

Preparation Mine Development $5.6 $5.6 Working Capital $16.4 $16.4

Other (Mine garage) $2.7 $2.7 Dams $3.9 $3.9

Deferred capital $0 $65.9

TOTAL PROJECT COST $463.6 $557.2

Mining equipment will be leased, and therefore, the leasing cost is included in the operating cost. Leased equipment will be bought back in year 4. An investment cost during operation is also considered to cover the replacement of ageing equipment as well as the addition of equipment required throughout the project life.

The rail line, rail equipment and port facilities will be owned and operated by a sub-contractor. The costs associated with the use of these facilities are included in the operating costs except for an initial investment of $95.2 M by CLM.

The owner’s costs consist of salaries and expenses of Consolidated Thompson Iron Mines Limited (CLM) personnel assigned to the project during the preparation, construction, training, and start-up phases, costs for core drilling program and metallurgical testing, environmental impact study, geotechnical study and value engineering studies. The total owner’s cost is estimated to be about $12.7 million.

Working capital was estimated to be equal to approximately 32 days of operation of the mine.


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18.10 Operating Cost Estimate

Present and historic costs from other operations of similar capacity, vendor quotations, commercial data bases and BBA’s expertise have all been used to estimate the total Bloom Lake operating costs.

The average operating costs for the first five years of operation(when in full year operation) are as follows:

• Average mining cost: $ 1.76/tonne mined

$ 6.58/tonne concentrate

• Average crushing and processing cost: $ 4.18/tonne concentrate

• Rail transport, port handling & ship loading $ 11.87/tonne concentrate

• General and Administration (mine site & corporate) $ 1.54/tonne concentrate

• Total $ 24.18/tonne concentrate

The operating cost for a twenty-year operation is $24.76/tonne concentrate. This average compares favourably to the three mining operations in the Labrador-Mont-Wright area, Bloom Lake being situated in the middle of the group.

18.10.1 Manpower

Approximate manpower requirements are summarized in Table 18-8. Table 18-8 : Manpower Requirements

Area Staff Hourly Total

Mine 17 130 147

Concentrator & Tailings Disposal 21 71 92

General and Administration 21 4 25

Sub-total CLM 264

Railway 20

Camp 8

Port 8

Total Project 300

Salaries and wages are typical of similar northern operations. A standard 35% is added for benefits.

Manpower costs for each area are included in the appropriate sections below.


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18.10.2 Mine

Operating cost estimates for the mine have been developed based on the use of CLM -operated leased mining equipment.

It is anticipated that all mobile equipment repairs will be performed at the shop, shovels and drills will be repaired in the mine and major components will be repaired by specialized shops offsite.

Hourly equipment operating costs were obtained from suppliers and adjusted to the conditions of the Bloom Lake area.

18.10.3 Process Plant

Consumption rates for crusher liners, autogenous and ball mill liners and grinding balls are based on published rates for similar operations or on experience. Consumption rates for screen cloths and filter cloths are based on vendor information. Unit costs for all these items are based on budget quotations.

Three percent of the total mechanical equipment cost is included annually to cover the cost of other maintenance supplies and parts.

Electrical costs are estimated following application of an utilization factor to the horsepower of operating equipment.

Fuel oil consumption rates for heating are estimated from building floor areas. Annual fuel oil consumption figures are estimated assuming a graduated monthly heating load. No fuel is consumed for the two months of July and August. Fuel oil used to generate steam for concentrate drying is calculated based on a 5-month drying period.

18.10.4 Administration

General and Administration costs cover:

• Service building maintenance and utilities

• Office supplies

• Maintenance costs of auxiliary equipment (pick-ups, forklift, bobcat, ambulance, fire truck, lineman’s truck and boom truck)

• Employee transportation by plane and bus

• Environment monitoring

• Insurance and fire protection

• Recruiting, training and R&D costs

• Consultant and legal fees

• Communication services

• Miscellaneous items (social activities, association dues etc).


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18.10.5 Rail Transportation & Port

The costs of construction and operation of the railway system to transfer concentrate to the port will be covered under a construction/operation agreement with a third party. A similar agreement will be negotiated for the construction and operation of the port facilities.

Port improvements will be covered by the Sept-Iles Port Authority (APSI) at no cost to the project. The Owners will enter into an initial 20-year lease agreement with APSI for the estimated 150,000 m2 storage area required with provision for renewal at five year intervals..

18.11 Financial Analysis

18.11.1 Introduction

The purpose of this section is to assess the economic viability of the proposed mining project as described in the previous sections, for the production of 8 million tonnes of iron concentrate per annum.

The financial evaluation for the Bloom Lake project is carried out by the preparation of a discounted cash flow model to which the capital and operating cost estimates as well as the production schedule developed in the mining section are input data. The Internal Rate of Return (IRR) on total investment and the Net Present Value (NPV) resulting from the net cash flows generated by the project have been calculated. The payback period is also indicated as a financial measure.

18.11.2 Basis of Evaluation

This section defines the fundamental assumptions that have been incorporated in the financial model.

18.11.2.1 Exchange Rate

When applicable, the exchange rate is assumed to be US$1 = CDN$1.10

18.11.2.2 Project Timing

The financial evaluation is carried out over a period of 20 years out of a total mine life reserve of more than 27 years. The construction phase is two (2) years excluding the construction of the magnetite recovery plant in year 3.

18.11.2.3 Owner’s Equity Contribution and Loan Capital

The financial plan assumes owner equity will be utilized to cover 100 % of the capital requirement.

18.11.2.4 Income Tax

The financial evaluation is carried out on a pre-tax and after-tax basis (a requirement for mining operation in the Province of Quebec).


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18.11.2.5 Depreciation

No depreciation rate has been considered since the financial appraisal is on a pre-tax basis.

18.11.2.6 Escalation and Inflation

The project financial analysis is carried out using the constant money basis.

18.11.2.7 Capital Investment Costs, Disbursement Schedule and Allowance

For the purpose of the cash flow analysis, all the pre-production capital expenditures have been assumed to occur in Years -1 and 1, with mining and processing commencing in the third quarter of Year 1. In addition to the initial capital of $446.9 M for the mine and the mill, other infrastructures investment amount to $108.9 M for maintenance shops, railway (31km), port and loadout facilities

The major annual disbursements are summarized in Table 18-9 as follows:

Table 18-9 : Investment Schedule

2008 - Initial Capex $96.2 M

2009 – Initial Capex $445.9M

2010 – Mine Garage $13.7M

2011 (Magnetite Recovery Plant) $27.8 M

2013 (Leased Mining Equipment Purchase) $30.8 M

Working Capital $17.1 M

On-going Capital (Year 2010 – 2030) $127.7M

A salvage value of $24.4 M in year 2030, representing approximately 5% of the equipment items is also accounted for the cash flow model.

18.11.2.8 Operating Costs

Operating costs including owner’s operated leased mining equipment, processing, contracted rail transportation and port handling, environment and G/A have been estimated to average $24.18 per tonne of concentrate for the first five (5) years of operation and $24.76 over a period of 20 years.


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18.11.2.9 Sales Revenue

According to international accepted standards, the appropriate price for the iron concentrate is the moving average of the last 3 years of iron price given as follows:

• 2006: $0.76 per iron unit;

• 2007: $0.834 per iron unit;

• 2008: $1.340 per iron unit.

The resulting FOB moving average for the last three years (2006-2008) is $1.069 per iron unit. Based on a concentrate grade of 66.5% Fe, the selling price is $71.09/tonne of concentrate.

18.11.2.10 Working Capital

Working capital required to meet expenses after start-up of operations before revenue becomes available is assumed to be equal to approximately 32 days of operating expenses in a full year operation. The working capital is shown separately on the cash flow schedule in Year 1 and will be recovered at the end of the life of the project. The working capital has been estimated at $17.1 million.

18.11.2.11 Other Assumptions

The other premises and assumptions applied in the preparation of the financial projections are outlined below:

• no provision for tax losses carried forward;

• no provision for dividend payment.

18.11.3 Results of Financial Evaluation

18.11.3.1 Determination of Internal Rate of Return & Cumulative NPV Cash Balance

On the basis of the assumptions described above in Section 18.11.2, the IRR and the cumulative cash balance at various discount rates have been computed and the results are listed in Table 18-10.

Table 18-10 : IRR and Cumulative Cash Balance

100% Equity Funding Before Tax After Tax

IRR % 58.9 48.6 NPV @ 0% ($M) 6 813 4 493 NPV @ 5% ($M) 3 847 2 520 NPV @ 8% ($M) 2 838 1 847 NPV @ 10% ($M) 2 348 1 521 NPV @ 15% ($M) 1 520 966 NPV @ 25% ($M) 708 418


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The calculation of the IRR is essentially an analysis based on the flow of the real resources, as opposed to financial resources providing the performance indicator related to the total investment. Therefore, the financial flows, such as financial sources, debt service, dividends are eliminated in this calculation.

18.11.3.2 Payback Period

Capital investment criteria must also consider payback. The payback is the time required for the cash income from the project to recoup the initial investment, i.e. the point in time when the cumulative undiscounted cash flows become zero. In general, the shorter the payback period, the more attractive is the investment project.

In the present calculation, the payback period has been established at approximately 2.6 years and 2.7 years after the end of the construction period for pre-tax and after-tax, respectively.

18.11.3.3 Sensitivity Analysis

The three major parameters affecting the net cash flow are sales prices, operating costs and the initial fixed investment. The sensitivity analysis is the process whereby the IRR is computed from the variations of these input data in the financial model to determine their impact on the project profitability. The data elements are changed independently of one another.

The result of the sensitivity analysis should help in identifying and focusing on the strategic parameters to improve the overall performance of the project.

The sensitivity analysis on IRR and NPV @ 5% discount rate is summarized in Table 18-12:

IRR (%)

-20% -15% -10% -5% 0% 5% 10% 15% 20%Item

OPEX 65.7 64.0 62.3 60.6 58.9 57.2 55.6 53.9 52.3CAPEX 71 67.6 64.4 61.5 58.9 56.5 54.2 52.2 50.2SALES 40.6 45.2 49.8 54.3 58.9 63.5 68 72.6 77.2

NPV ($M) @ 5% Discount Rate -20% -15% -10% -5% 0% 5% 10% 15% 20%Item

OPEX 4330 4209 4089 3968 3848 3728 3607 3487 3366CAPEX 3955 3928 3902 3875 3848 3821 3794 3768 3741SALES 2464 2810 3156 3502 3848 4194 4540 4886 5232

Table 18-11 : IRR and NPV @ 5% on Total Investment


November 2008

18-61

The results of the sensitivity analysis on a pre-tax basis of the IRR for iron ore price, capital cost and operating cost and pre-tax NPV @ 5% discount rate are presented in graphical form in Figures 18-19 and 18-20, respectively. The analysis shows that the project is most sensitive to revenue, that being iron ore price, ore grade and weight recoveries.

Figure 18-27: IRR Bloom Lake Iron Ore Project - Sensitivity Analysis (before tax) - IRR%

Bloom Lake- 8 m tpy Concentrate FS Sensivity Analysis (pre-tax) Internal Rate of Return (IRR)

20

30

40

50

60

70

80

90

-20% -15% -10% -5% 0% 5% 10% 15% 20%

Change (%)

IRR

(%)

OPEX CAPEX SALES


November 2008

18-62

Bloom Lake- 8 m tpy Concentrate PFS Sensitivity Analys is (pre-tax)

Net Present Value (NPV) @ 5% Discount rate

1000

1500

2000

2500

3000

3500

4000

4500

5000

5500

-20% -15% -10% -5% 0% 5% 10% 15% 20%

Change (%)

NPV

(M$)

OPEX CAPEX SALES

Figure 18-28: NPV Bloom Lake Iron Ore Project – Sensitivity Analysis (before tax)

NPV @ 5% Discount Rate

18.11.4 Discussion

The results of the financial analysis indicate that the Bloom Lake project has a potential before-tax IRR of 58.9% and a NPV of US$ 6 813 M and US$ 3 847 M at a discount rate of 0% and 5% respectively. On an after-tax basis, the IRR is 48.6% and the NPV amounts to US$ 4 493 M and US$ 2 520 M at a discount rate of 0% and 5% respectively. These strong financial indicators show the excellent economic viability of the project and the sustainability of a good profitability in the advent of iron price fluctuation


November 2008

19-1

19. PROJECT IMPLEMENTATION

The master schedule for the Bloom Lake project starts in April 2006 and ends with start-up activities on September 30 2009. A simplified version of the master schedule is shown in Figure 19-1. The total duration is 41 months or three (3) years. The schedule was prepared according to these major activities:

• Environmental permitting • Test work program • Basic engineering, detailed engineering and construction of the concentrator and the

infrastructure.

Figure 19-1: Schedule for Bloom Lake Project

The critical path of the project follows the line of these groups of activities; environmental permitting, construction and commissioning.

An eight-month schedule was developed from project notification to environmental authorization for the project. This is a short schedule considering that collection of environmental data had to be completed in the spring of 2006 and public audiences needed to be held for a large mining project. However, since the project is in a mining friendly area and the socioeconomic conditions are favorable, a short environmental schedule was forecasted.

CLM received the certificate of authorization in April 2008. Until the certificate was obtained, no construction activity was underway. However, by mobilizing the contractor and preparing the access to the site in the winter of 2007-2008, it was possible to start the civil / foundation work for the concentrator in the spring of 2008, taking advantage of better conditions. Priorities were given to concentrator and crusher foundations to minimize concrete pours in winter.

Construction of the magnetic separation plant is scheduled for 2011 with start-up the following year.

Project schedules for the Port and Railway facilities, which were developed by the consultants CEMA+ and Cantech, are in accordance with the BBA master schedule. In the case of the railway, the schedule is contingent on permit approval From Newfoundland and Labrador. An Environmental Review Report was submitted to the Newfoundland and Labrador Government on August 29, 2008. The authorisation was obtained on October 27, 2008, which will allow the Construction of the railway to proceed.

Apr-Jun Jan-Mar Apr-Jun Jul-Sept Oct-Dec Apr-Jun Jul-Sept Oct-Dec Jan-Mar Apr-Jun Jul-Sept

Environmental Impact Study

Public Consultations, Replies toQuestions, project Autorisation

Laboratory Testwork

Basic Engineering

Detailed Engineering & Purchasing

Construction

Commissioning

2009Jul-Sept Oct-Dec

2007 20082006Jan-Mar


November 2008

20-1

20. OTHER RELEVANT DATA AND INFORMATION No other data or information are relevant for the review of the Bloom Lake Project.


November 2008

21-1

21. INTERPRETATION AND CONCLUSIONS

The Bloom Lake Iron Ore Project is financially and technically feasible with an estimated capital cost of $446.9 M for the mine and the mill and $108.9 M for the maintenance shops, railway, port and loadout facilities. The total of US$555.8 M, excluding US$17.1 M for working capital, results in an IRR of 58.9% and a simple payback period of 2.6 years before taxes

The level of accuracy of the capital and operating cost estimates is ± 15% and the capital cost estimate includes a 7% contingency on direct costs. Costs of leased mine equipment are included in operating costs. Costs for construction and operation of the railway and for capital expenditures and operation of the Pointe Noire stockpiling facilities will be covered by operating leases and so are also included in the operating costs.

Based on an estimate of mineral resources produced by Watts, Griffis and McOuat Limited in compliance with the National Instrument 43-101, the mineral reserves are evaluated at 580 million tonnes and will be sufficient for over 27 years of operation at a production rate of 8,000,000 tonnes of concentrate per year.

All the testwork results to date are consistent with good iron recovery using conventional spirals and are also in line with the results obtained from the other iron ore producers in the area. Testwork is currently underway to optimize the magnetic separation flowsheet to increase iron recovery.

At this point, the concentrator is under construction. The environmental permit for the rail spur line was approved on October 27, 2008. Construction of the railway can now proceed.


November 2008

22-1

22. RECOMMENDATIONS

The recommendations presented in feasibility study dated April 2006 have been implemented and included the following:

• Environmental permitting work (Environmental Impact Studies) is completed and submitted to the concerned government environmental agencies. (Done).

• A request is made to Hydro-Quebec to begin detailed studies to provide power to the site. (Done).

• New drill core samples for the confirmation test work and bulk samples for pilot plant test work have been collected. (Done).

• Hydrogeology and geotechnical studies have been initiated. (Done).

• Preliminary talks with the public common rail carrier have been initiated. (Done)

• Detailed topographic mapping of the project area is completed. (Done).

• Detailed engineering activities for final flowsheet and design criteria selection have started. (Done).

The status of new recommendations presented in the feasibility study for 7-million tonnes of concentrate per year dated May 2007 are as follows:

• Start the environmental restoration studies to replace the lakes and river sections affected by the tailings ponds in the summer of 2007 (underway).

• Finalize the agreements with the mining fleet equipment suppliers (near completion).

• Finalize the leasing agreements with the railway and port operators (underway).

• Perform exploration drilling to expand the reserves on the west side of the pit (underway).

• Perform definition drilling to confirm detailed mining plan for the first 3 years of initial mining (underway)


November 2008

23-1

23. REFERENCES/SOURCES OF INFORMATION

BBA, Bloom Lake Iron Ore Project – Bankable Feasibility Study, April 2006

Roche, Bloom Lake Project, Feasibility Study – Environmental Aspects, March 2006

Journeaux, Bédard, Feasibility and Cost Evaluation for Tailings Impoundment, Bloom

Lake Project, February 17, 2006

WGM, A technical Review and Mineral Resource Estimate for the Bloom Lake Iron

Deposit, Labrador Trough, Quebec, May 26, 2006

H.E. Neal & Associates, Preliminary Evaluation, Queco Property, Quebec Cobalt &

Exploration Limited, For Republic Steel Corporation, March 1976

SGS Lakefield Research Limited, Proposed Grinding System for the Bloom Lake Circuit

Based on Small-Scale Data, Prepared for BBA on Behalf of CLM, January 26, 2006

SGS Lakefield Research Limited, Beneficiation of Iron Ore Samples from the Bloom

Lake Property, Prepared for WGM on behalf of CLM, February 23, 2006

SGS Lakefield Research, An Investigation into the Water Quality and Treatability of

Bloom Lake Tailings, Final Report, March 9, 2007

SGS Lakefield Research, Preliminary Grinding Circuit Study for the Bloom Lake Circuit

Based on Small-Scale Data, February 16, 2007

E. Palumbo, Sizing of Pan Filters for 7-million tpy Concentrate, February 5, 2007

J. Dinsdale, SGS Thickener and Filter Sizing Testwork, October 6, 2006

J. Dinsdale, Tailings Thickener Sizing for 7 Mtpy Production Rate, February 22, 2007

E. Palumbo, Metallurgical and Grindability Test Results on Drill Core Samples, 21 March

2007

SGS Lakefield Reasearch Limited, Grindability Characteristics and Beneficiation Testing

of Iron Ore Samples from the Bloom Lake Property, prepared by Breton Banville &

Associates Inc., project 11478-001 – Draft Report, May 9, 2007


November 2008

23-2

Feasibility Study for 8 Mtpy Iron Ore Concentrate, Bloom Lake Project, report prepared

by BBA inc. for Consolidated Thompson Iron Mines Ltd., September 2008.


November 2008

25-1

25. CERTIFICATES


November 2008

25-2

CERTIFICATE OF QUALIFIED PERSON

I, Patrice Live, Eng., do hereby certify that :

1. I am currently employed as Manager – Mining in the consulting firm:

BBA Inc. (BBA)

630 René-Lévesque blvd. W.

Suite 2500

Montréal, Québec

Canada H3B1S6

2. I graduated from Université Laval of Québec, Canada with a B. Sc. in Mining in 1976.

3. I am in good standing as a member of the Order of Engineers of Québec (#38991).

4. I have practiced my profession continuously since my graduation.

5. I have read the definition of “qualified person” set out in the National Instrument 43-101 (“NI 43-101”) and certify that as a result of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6. I am responsible for the preparation of Section 18.1 and Section 18.11 of the Technical Report entitled “Bloom Lake Technical Report” based on my knowledge and extensive experience in mining studies and operation on iron ore projects worldwide. I have visited the property on November 2, 2008.

7. As of the date of this certificate I am not aware of any changes in fact or circumstances with respect to the subject matter of this report which materially affects the content of the report or the conclusion reached.

8. I am independent of the issuer applying all of the tests in Section 1.5 of National Instrument 43-101.

9. I have read national Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

10. I consent to the filing of the Technical Report with any stock exchange or any regulatory authority and any publication by them, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.


November 2008

25-4


I, Enzo Palumbo, do hereby certify that:

1. I am currently employed as Metallurgist in the consulting firm:

BBA Inc. (BBA)

630 René-Lévesque Blvd. W

Suite 2500

Montréal, Québec

Canada H3B1S6

2. I graduated from McGill University of Montréal with a B.Eng. in Metallurgy in 1981, and M.Eng. in 1986.

3. I am a member of the Canadian Institute of Mining, Metallurgy, and Petroleum and a member of The Minerals, Metals & Materials Society (TMS) of the American Institute of Mining, Metallurgical, and Petroleum Engineers, Inc.



6. I am responsible for or was involved in the preparation of this 43-101 F1 Technical Report entitled Bloom Lake Technical Report. I have reviewed all sections of the technical report. I have not visited the project site.



9. I have read national Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.


November 2008

25-6


I, René Scherrer, residing at 13, rue Nicolas, Sept-Iles, Québec, G4R 5K3, do hereby certify

that:

1. I am currently Vice President – Exploration and Mine Operation for:

Consolidated Thompson Iron Mines Ltd (CLM)

1155 University Street

Suite 508

Montréal, Québec

Canada H3B 3A7

2. I graduated from École Polytechnique de Montréal, Québec, Canada with a B. Sc. in Mining, in 1987.

3. I am in good standing as a member of the Order of Engineers of Québec (#44587).



6. I am responsible for the preparation and review of Sections 7 to 14 and Section 17 of the Technical Report entitled “Bloom Lake Technical Report” based on my knowledge and extensive experience in mining, in engineering and operation departments, on iron, gold and nickel open pit projects over North and South America.

7. I have personally visited the property on May 16, 2006 and on many occasions since that date.


9. I am working for Consolidated Thompson and therefore not independent of the issuer as defined in Section 1.4 of National Instrument 43-101.


November 2008

25-8


I, André Allaire, Eng., do hereby certify that :

1. I am currently employed as Manager – Mining and Metallurgy and partner in the consulting firm:

BBA Inc. (BBA)

630 boul. René-Lévesque Blvd. W

Suite 2500

Montréal, Québec

Canada H3B1S6

2. I graduated from McGill University of Montréal with a B. Eng in Metallurgy in 1982, a M.Eng. in 1986 and a Ph.D. in 1991.

3. I am in good standing as a member of the Order of Engineers of Québec (#38480) and a member of the Canadian Institute of Mining Metallurgy and Petroleum.



6. I am responsible for or was in involved in the preparation of this 43-101 F1 Technical Report entitled Bloom Lake Technical Report based on my knowledge on iron ore process design. I am aware of the metallurgical testworks performed on ores collected from the Bloom Lake property. I have reviewed all sections of the report. I visited the project site with Richard Quesnel, Eng., President and Chief Operating Officer of Consolidated Iron Ore Mines Ltd on December 7-8, 2005 and on way after subsequent occasions since that date..




November 2008

Appendix A

List of CLM Mining Claims

#1363096.2

CONSOLIDATED THOMPSON IRON MINES LIMITED

DESCRIPTION OF BLOOM LAKE PROPERTY

(Townships of Normanville and Lislois and Moisie River Basin, Registration Division of Saguenay (240 claims))

(Information initially compiled from printouts from the Public Register of Real Immovable Mining Rights dated August 25, 2006 and

updated in the Ministry of Natural Resources and Wildlife GESTIM website on May 18, 2007)

CLAIM NUMBER

CLAIM MAP REFERENCE

REGISTRATION DATE

(Y-M-D)

EXPIRY DATE (Y-M-D)

AREA (HA)

WORK SURPLUS ($)

RENEWALS

CDC 0098977 23B/14 2005-10-03 2007-10-02 16,61 0.00 0 CDC 0098978 23B/14 2005-10-03 2007-10-02 15,96 0.00 0 CDC 0098979 23B/14 2005-10-03 2007-10-02 15,32 0.00 0 CDC 0098980 23B/14 2005-10-03 2007-10-02 14,68 0.00 0 CDC 0098981 23B/14 2005-10-03 2007-10-02 33,05 0.00 0 CDC 0098982 23B/14 2005-10-03 2007-10-02 52,07 0.00 0 CDC 0098983 23B/14 2005-10-03 2007-10-02 52,08 0.00 0 CDC 0098984 23B/14 2005-10-03 2007-10-02 52,08 0.00 0 CDC 0098985 23B/14 2005-10-03 2007-10-02 52,08 0.00 0 CDC 0098986 23B/14 2005-10-03 2007-10-02 52,07 0.00 0 CDC 0098987 23B/14 2005-10-03 2007-10-02 52,06 0.00 0 CDC 0098988 23B/14 2005-10-03 2007-10-02 52,05 0.00 0 CDC 0098989 23B/14 2005-10-03 2007-10-02 52,04 0.00 0 CDC 0098990 23B/14 2005-10-03 2007-10-02 52,03 0.00 0 CDC 0098991 23B/14 2005-10-03 2007-10-02 52,03 0.00 0 CDC 0098992 23B/14 2005-10-03 2007-10-02 52,03 0.00 0 CDC 0098993 23B/14 2005-10-03 2007-10-02 52,08 0.00 0 CDC 0098994 23B/14 2005-10-03 2007-10-02 51,52 0.00 0 CDC 0098995 23B/14 2005-10-03 2007-10-02 34,37 0.00 0 CDC 0098996 23B/14 2005-10-03 2007-10-02 52,07 0.00 0 CDC 0098997 23B/14 2005-10-03 2007-10-02 52,07 0.00 0 CDC 0098998 23B/14 2005-10-03 2007-10-02 52,06 0.00 0 CDC 0098999 23B/14 2005-10-03 2007-10-02 44,15 0.00 0 CDC 0099000 23B/14 2005-10-03 2007-10-02 14,88 0.00 0 CDC 0099001 23B/14 2005-10-03 2007-10-02 8,16 0.00 0

CDC 0099016 23B/14 2005-09-27 2007-09-26 52,07 0.00 0 CDC 0099017 23B/14 2005-09-27 2007-09-26 52,07 0.00 0 CDC 0099018 23B/14 2005-09-27 2007-09-26 52,06 0.00 0 CDC 0099019 23B/14 2005-09-27 2007-09-26 52,06 0.00 0 CDC 0099020 23B/14 2005-09-27 2007-09-26 52,06 0.00 0

CONSOLIDATED THOMPSON IRON MINES LIMITED FRASER MILNER CASGRAIN LLP Bloom Lake Property May 18, 2007 Page 2

#1363096.2

CLAIM NUMBER

CLAIM MAP REFERENCE

REGISTRATION DATE

(Y-M-D)

EXPIRY DATE (Y-M-D)

AREA (HA)

WORK SURPLUS ($)

RENEWALS

CDC 0099021 23B/14 2005-09-27 2007-09-26 52,05 0.00 0 CDC 0099022 23B/14 2005-09-27 2007-09-26 52,05 0.00 0 CDC 0099023 23B/14 2005-09-27 2007-09-26 52,04 0.00 0 CDC 0099024 23B/14 2005-09-27 2007-09-26 52,04 0.00 0 CDC 0099025 23B/14 2005-09-27 2007-09-26 52,04 0.00 0 CDC 0099026 23B/14 2005-09-27 2007-09-26 52,04 0.00 0 CDC 0099027 23B/14 2005-09-27 2007-09-26 52,04 0.00 0 CDC 0099028 23B/14 2005-09-27 2007-09-26 52,03 0.00 0 CDC 0099029 23B/14 2005-09-27 2007-09-26 52,03 0.00 0 CDC 0099030 23B/14 2005-09-27 2007-09-26 52,03 0.00 0 CDC 0099031 23B/14 2005-09-27 2007-09-26 52,03 0.00 0 CDC 0099032 23B/14 2005-09-27 2007-09-26 52,03 0.00 0 CDC 0099033 23B/14 2005-09-27 2007-09-26 52,03 0.00 0 CDC 0099034 23B/14 2005-09-27 2007-09-26 52,07 0.00 0 CDC 0099035 23B/14 2005-09-27 2007-09-26 52,06 0.00 0 CDC 0099036 23B/14 2005-09-27 2007-09-26 52,06 0.00 0 CDC 0099037 23B/14 2005-09-27 2007-09-26 52,05 0.00 0 CDC 0099038 23B/14 2005-09-27 2007-09-26 52,05 0.00 0 CDC 0099039 23B/14 2005-09-27 2007-09-26 52,04 0.00 0 CDC 0099040 23B/14 2005-09-27 2007-09-26 52,04 0.00 0 CDC 0099041 23B/14 2005-09-27 2007-09-26 43,64 0.00 0 CDC 0099042 23B/14 2005-09-27 2007-09-26 52,07 0.00 0 CDC 0099043 23B/14 2005-09-27 2007-09-26 52,07 0.00 0 CDC 0099044 23B/14 2005-09-27 2007-09-26 52,07 0.00 0 CDC 0099045 23B/14 2005-09-27 2007-09-26 52,07 0.00 0 CDC 0099046 23B/14 2005-09-27 2007-09-26 52,07 0.00 0 CDC 0099047 23B/14 2005-09-27 2007-09-26 52,07 0.00 0

CDC 0099881 23B/14 2005-11-02 2007-11-01 17,12 0.00 0 CDC 0099882 23B/14 2005-11-02 2007-11-01 48,14 0.00 0 CDC 0099883 23B/14 2005-11-02 2007-11-01 52,09 0.00 0 CDC 0099884 23B/14 2005-11-02 2007-11-01 52,09 0.00 0 CDC 0099885 23B/14 2005-11-02 2007-11-01 1,80 0.00 0 CDC 0099886 23B/14 2005-11-02 2007-11-01 17,07 0.00 0 CDC 0099887 23B/14 2005-11-02 2007-11-01 2,99 0.00 0 CDC 0099888 23B/14 2005-11-02 2007-11-01 39,32 0.00 0 CDC 0099889 23B/14 2005-11-02 2007-11-01 52,07 0.00 0 CDC 0099890 23B/14 2005-11-02 2007-11-01 4,48 0.00 0 CDC 0099891 23B/14 2005-11-02 2007-11-01 41,93 0.00 0 CDC 0099892 23B/14 2005-11-02 2007-11-01 52,06 0.00 0 CDC 0099893 23B/14 2005-11-02 2007-11-01 52,06 0.00 0 CDC 0099894 23B/14 2005-11-02 2007-11-01 33,16 0.00 0 CDC 0099895 23B/14 2005-11-02 2007-11-01 29,59 0.00 0


#1363096.2

CLAIM NUMBER

CLAIM MAP REFERENCE

REGISTRATION DATE

(Y-M-D)

EXPIRY DATE (Y-M-D)

AREA (HA)

WORK SURPLUS ($)

RENEWALS

CDC 0099896 23B/14 2005-11-02 2007-11-01 26,03 0.00 0 CDC 0099897 23B/14 2005-11-02 2007-11-01 22,63 0.00 0 CDC 0099898 23B/14 2005-11-02 2007-11-01 44,25 0.00 0 CDC 0099899 23B/14 2005-11-02 2007-11-01 52,05 0.00 0 CDC 0099900 23B/14 2005-11-02 2007-11-01 52,05 0.00 0 CDC 0099901 23B/14 2005-11-02 2007-11-01 52,05 0.00 0 CDC 0099902 23B/14 2005-11-02 2007-11-01 52,04 0.00 0 CDC 0099903 23B/14 2005-11-02 2007-11-01 52,04 0.00 0 CDC 0099904 23B/14 2005-11-02 2007-11-01 52,04 0.00 0 CDC 0099905 23B/14 2005-11-02 2007-11-01 52,04 0.00 0 CDC 0099906 23B/14 2005-11-02 2007-11-01 52,04 0.00 0 CDC 0099907 23B/14 2005-11-02 2007-11-01 52,04 0.00 0 CDC 0099908 23B/14 2005-11-02 2007-11-01 52,04 0.00 0 CDC 0099909 23B/14 2005-11-02 2007-11-01 52,04 0.00 0 CDC 0099910 23B/14 2005-11-02 2007-11-01 52,03 0.00 0 CDC 0099911 23B/14 2005-11-02 2007-11-01 52,03 0.00 0 CDC 0099912 23B/14 2005-11-02 2007-11-01 52,03 0.00 0 CDC 0099913 23B/14 2005-11-02 2007-11-01 52,03 0.00 0 CDC 0099914 23B/14 2005-11-02 2007-11-01 52,03 0.00 0 CDC 0099915 23B/14 2005-11-02 2007-11-01 52,03 0.00 0 CDC 0099916 23B/14 2005-11-02 2007-11-01 52,03 0.00 0 CDC 0099917 23B/14 2005-11-02 2007-11-01 52,03 0.00 0 CDC 0099918 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099919 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099920 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099921 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099922 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099923 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099924 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099925 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099926 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099927 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099928 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099929 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099930 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099931 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099932 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099933 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099934 23B/14 2005-11-02 2007-11-01 52,02 0.00 0 CDC 0099935 23B/14 2005-11-02 2007-11-01 52,01 0.00 0 CDC 0099936 23B/14 2005-11-02 2007-11-01 52,01 0.00 0 CDC 0099937 23B/14 2005-11-02 2007-11-01 52,01 0.00 0 CDC 0099938 23B/14 2005-11-02 2007-11-01 52,01 0.00 0


#1363096.2

CLAIM NUMBER

CLAIM MAP REFERENCE

REGISTRATION DATE

(Y-M-D)

EXPIRY DATE (Y-M-D)

AREA (HA)

WORK SURPLUS ($)

RENEWALS



#1363096.2

CLAIM NUMBER

CLAIM MAP REFERENCE

REGISTRATION DATE

(Y-M-D)

EXPIRY DATE (Y-M-D)

AREA (HA)

WORK SURPLUS ($)

RENEWALS

CDC 0099982 23B/14 2005-11-02 2007-11-01 12,55 0.00 0 CDC 0099983 23B/14 2005-11-02 2007-11-01 9,37 0.00 0

CDC 1133830 23B/14 2005-11-11 2007-07-18 52,06 166,206.49 0 CDC 1133831 23B/14 2005-11-11 2007-07-18 52,06 166,206.49 0 CDC 1133832 23B/14 2005-11-11 2007-07-18 52,06 166,206.49 0 CDC 1133833 23B/14 2005-11-11 2007-07-18 52,06 166,206.49 0 CDC 1133834 23B/14 2005-11-11 2007-07-18 52,06 166,206.49 0 CDC 1133835 23B/14 2005-11-11 2007-07-18 52,05 166,174.57 0 CDC 1133836 23B/14 2005-11-11 2007-07-18 52,05 166,174.57 0 CDC 1133837 23B/14 2005-11-11 2007-07-18 52,05 166,174.57 0 CDC 1133838 23B/14 2005-11-11 2007-07-18 52,05 166,174.57 0 CDC 1133839 23B/14 2005-11-11 2007-07-18 52,05 166,174.57 0 CDC 1133840 23B/14 2005-11-11 2007-07-18 52,05 166,174.57 0 CDC 1133841 23B/14 2005-11-11 2007-07-18 52,04 166,142.64 0 CDC 1133842 23B/14 2005-11-11 2007-07-18 52,04 166,142.64 0 CDC 1133843 23B/14 2005-11-11 2007-07-18 52,04 166,142.64 0

CDC 1133844 23B/14 2005-11-11 2007-08-28 52,02 55.75 0 CDC 1133845 23B/14 2005-11-11 2007-08-28 52,02 55.75 0 CDC 1133846 23B/14 2005-11-11 2007-08-28 52,01 55.74 0 CDC 1133847 23B/14 2005-11-11 2007-08-28 52,01 55.74 0

CDC 2082920 23B/14 2007-05-08 2009-05-07 5,12 0.00 0 CDC 2082921 23B/14 2007-05-08 2009-05-07 13,4 0.00 0 CDC 2082922 23B/14 2007-05-08 2009-05-07 50,53 0.00 0 CDC 2082923 23B/14 2007-05-08 2009-05-07 52,08 0.00 0 CDC 2082924 23B/14 2007-05-08 2009-05-07 48,79 0.00 0 CDC 2082925 23B/14 2007-05-08 2009-05-07 1,72 0.00 0 CDC 2082926 23B/14 2007-05-08 2009-05-07 52,04 0.00 0 CDC 2082927 23B/14 2007-05-08 2009-05-07 52,04 0.00 0 CDC 2082928 23B/14 2007-05-08 2009-05-07 52,03 0.00 0 CDC 2082929 23B/14 2007-05-08 2009-05-07 52,03 0.00 0 CDC 2082930 23B/14 2007-05-08 2009-05-07 52,03 0.00 0 CDC 2082931 23B/14 2007-05-08 2009-05-07 52,03 0.00 0 CDC 2082932 23B/14 2007-05-08 2009-05-07 52,03 0.00 0 CDC2082933 23B/14 2007-05-08 2009-05-07 52,03 0.00 0 CDC 2082934 23B/14 2007-05-08 2009-05-07 52,02 0.00 0 CDC 2082935 23B/14 2007-05-08 2009-05-07 52,02 0.00 0 CDC 2082936 23B/14 2007-05-08 2009-05-07 52,02 0.00 0 CDC 2082937 23B/14 2007-05-08 2009-05-07 52,02 0.00 0

Application for renewal filed with the MNRW on April 11, 2007 and currently pending (file no. 668088).


#1363096.2

CLAIM NUMBER

CLAIM MAP REFERENCE

REGISTRATION DATE

(Y-M-D)

EXPIRY DATE (Y-M-D)

AREA (HA)

WORK SURPLUS ($)

RENEWALS



#1363096.2

CLAIM NUMBER

CLAIM MAP REFERENCE

REGISTRATION DATE

(Y-M-D)

EXPIRY DATE (Y-M-D)

AREA (HA)

WORK SURPLUS ($)

RENEWALS

CDC 2082981 23B/14 2007-05-08 2009-05-07 45,78 0.00 0 (240 mining claims)

Documents

43-101_bloom_lake_nov_08[1]