URSA MAJOR MINERALS INCORPORATED FEASIBILITY...

URSA MAJOR MINERALS INCORPORATED

FEASIBILITY STUDY

FOR THE

SHAKESPEARE NICKEL DEPOSIT,

NEAR ESPANOLA, ONTARIO

JANUARY, 2006

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EXECUTIVE SUMMARY

INTRODUCTION AND TERMS OF REFERENCE

Micon International Limited (Micon) has been retained by Dr. Richard H. Sutcliffe, President and CEO of URSA Major Minerals Incorporated (URSA), to prepare a feasibility study concerning URSA’s Shakepeare Project, a deposit comprising nickel-copper mineralization with associated cobalt, platinum, palladium and gold. The deposit has potential for development as an open pit mining operation.

The Shakespeare property is located just north of Agnew Lake, near the village of Webbwood, Ontario. Webbwood lies just west of Espanola on Highway 17, and approximately one hour by road from Sudbury. See Figure I.

The climate in the area is characterized by moderately long, cold winters and shorter, warm summers and is typical of continental conditions.

The project is a joint venture between URSA and Falconbridge Limited (Falconbridge). The joint venture agreement, dated June 16, 2000, gave URSA the right to acquire up to a 75% interest in the Shakespeare Property, which interest URSA has earned. The 28 leased and patented claims which make up the core of the Shakespeare joint venture are thus 75% legally owned by URSA and 25% legally owned by Falconbridge. URSA’s 75% legal ownership is registered on title.

URSA has a further beneficial ownership resulting from its expenditure: Falconbridge chose not to participate in the feasibility study and its interest in the joint venture has been diluted. As of January, 2006, URSA’s beneficial interest has increased to 86%. Upon URSA’s interest reaching 90%, Falconbridge’s interest can be converted to a 1.5% net smelter return (NSR) royalty.

North American Palladium (NAP) has entered into an agreement with URSA whereby, upon completion of the feasibility study described herein, it has 180 days to exercise an option to acquire a 60% undivided interest in URSA’s interest in the Shakespeare Property, and become the operator, by making aggregate payments of $1.5 million to URSA and securing the project financing for commercial production. URSA remains the operator of the feasibility program.

GEOLOGY AND MINERAL RESOURCES

The project area is located located along the southern margin of the Superior Province of the Canadian Shield. The Shakespeare deposit is hosted within a north-dipping differentiated gabbroic sill, approximately 80 m thick, situated along the north contact of a Nipissing Suite mafic intrusive with quartzites of the Mississagi Formation. The bulk of the mineralization occurs as a broad conformable zone of magmatic sulphides located approximately 30 m to 80 m below the upper contact of the intrusion with the quartzites. The style of magmatic sulphides in the mineralized zone progresses downward from sporadically disseminated, to scattered multi-centimetre sized blebby composite to locally net-textured pyrrhotite-chalcopyrite-pentlandite grains, to more evenly distributed, heavily disseminated magmatic sulphides. The overall dip of the deposit is steeply northwards.

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Figure I Project Location Plan

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The zone of nickel, copper, cobalt and platinum group metal-bearing sulphide mineralisation plunges 30° to the northeast. It remains open along strike and at depth, and the potential exists for an economically viable zone of underground mineable mineralization. See Figure II.

EXPLORATION

Five separate diamond drill programs had been completed on the Shakepeare property by Falconbridge over the period between 1942 and 1986. These programs amount to 47 holes totalling 6,655 m.

Work by URSA since 2000 has involved digital compilation, geological mapping, sampling, geophysics and diamond drilling. Drilling at the Shakespeare project that has been completed by URSA is summarized in Table I.

Table I URSA Diamond Drill Holes

Year Number of Holes Metreage 2002 9 4,904.90 Winter 2003 18 3,263.00 2003/2004 28 5,950.33 Summer 2004 10 3,648.90 2005 19 2,443.10 Total 84 20,210.23

The results of the diamond drilling programs carried out by URSA were positive and encouraging. Most of the drill holes intersected a wide interval of sulphide mineralization and collectively defined a single layer of nickel-copper-gold-platinum group metal-bearing sulphide mineralization divided into two adjacent zones known as East and West. Typical widths are approximately 40 m and grades for nickel, copper and precious metals in the East Zone are generally higher than the average grade of the original Shakespeare, or West, deposit discovered by Falconbridge.

Drill core logging, sampling and sample preparation throughout the URSA exploration program have been carried out within normal industry standards, with periodic independent review. Similarly, independent review of the assaying, data verification, quality assurance and quality control procedures such as the use of blanks, standard and duplicate samples, have been fully described and documented.

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Figure II Shakespeare Property Geological Map

(For Legend – see Figure 3.2)

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MINERAL RESOURCE ESTIMATION

A Gemcom database was constructed, the data validated and all identified errors corrected. The data in the assay table included values for nickel, copper, cobalt, gold, platinum and palladium.

A geological domain model was constructed to control grade interpolation. Domain boundaries were determined by lithology, mineralization style and grade boundary interpretation from visual inspection of drill hole sections, taking account of faults interpreted from stratigraphic offsets and logged fault gouge in drill-holes. The geological domains were extrapolated 25 m in either direction from the end of drilling. The gap between the East and West Zones has received relatively little drilling and was not modelled. As a result, a two-part geological domain model, with an aggregate length of approximately 1,435 m in total, has been generated.

Four separate domains, called “Disseminated” (West and East) and “Blebby” (West and East) were established to represent the two styles of sulphide mineralization in the two deposits after a statistical analysis of all assays from the entire mineralized body indicted separate populations of data. The polylines from each section and domain were wireframed into 3-dimensional solids using Gemcom. The resulting domain solids were used for grade interpolation purposes. Figure III shows the resulting domains in a three-dimensional isometric view.

Figure III Three-Dimensional View (looking South) showing Geological Domains and Drill Hole Traces

Grade capping was utilized where appropriate to ensure that erratic high values did not bias the grade estimation. Grade capping values used are set out in Table II..

Semivariograms were created and analyzed for the nickel, copper, cobalt, gold, platinum and palladium data from the constrained assay composite extraction files for both Blebby and Disseminated domains.

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Table II Grade Capping Values

Element East Blebby Domain

East Disseminated Domain

West Blebby Domain

West Disseminated Domain

Nickel 1.10% 1.05% No capping No capping Copper No capping 1.40% No capping 0.92% Cobalt 0.11% 0.107% No capping 0.10% Gold 0.420g/t 0.861g/t 0.355g/t No capping Platinum 0.725g/t No capping No capping 0.758g/t Palladium 0.943g/t 1.000g/t No capping 0.799g/t

The specific gravities used for the mineral resource estimate were obtained from measurements on a total of 257 samples, representing all lithologies, which were subjected to wet/dry bulk specific gravity determination tests.

A block model framework was created with 380 columns (X), 200 rows (Y) and 120 levels (Z), for a total of 9,120,000 blocks, each measuring 5 x 5 x 5 metres. Inverse distance squared (ID2) interpolation was used for the West Blebby domain while Ordinary Kriging (OK) interpolation was used for the East Blebby and both East and West Disseminated domains. Two interpolation passes were used to determine the Indicated and Inferred classifications and to interpolate nickel, copper, cobalt, gold, platinum and palladium grades into each block.

Resources

Under the CIM definitions, a mineral resource must be potentially economic in that it must be “in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction”. The bulk of the resource is expected to be amenable to open pit mining although the deeper part of the resource may be mined by underground methods. The resource may therefore be divided between these two mining methods, and different criteria must therefore be applied in assessing the prospects of economic extraction. Micon has used an internal open pit NSR cut-off equal to Cdn$24.23/t within a Whittle-optimized pit shell for the reporting of the mineral resources in the Shakespeare West and East Zones. Mineralization outside the pit limit was evaluated at a Cdn$50.00/t cut-off to determine the underground potentially mineable portion of the mineral resource.

Mineral resources were classified as indicated or inferred based on search ellipse distances required for interpolation of grade into a block and the ranges from the nickel variograms.

The resources within and beneath the optimized pit shell are shown in Table IIIa and IIIb, respectively. The mineral resources presented are current as of September 18, 2005.

MINING AND MINERAL RESERVES

Mining will be carried out with standard open pit methods. Drilling will be performed on all ore and waste by conventional in-the-hole hammer rigs with single pass drilling capability. Blasting operations will use an emulsion-ANFO blend and a downhole delay initiation system. A hydraulic shovel and a front-end-loader will be used to load haulage trucks of 91 tonne capacity.

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Table IIIa Mineral Resources within Optimised Pit Shell above Cdn$24.23/t NSR Cut-off

Category Tonnes Ni (%)

Cu (%)

Co (%)

Au (g/t)

Pt (g/t)

Pd (g/t)

NSR Cdn$/t

Indicated East 9,460,000 0.37 0.38 0.02 0.204 0.357 0.393 61.25 West 2,970,000 0.29 0.33 0.02 0.181 0.333 0.361 50.90 Total 12,430,000 0.35 0.37 0.02 0.199 0.351 0.386 58.78

Inferred East 220,000 0.32 0.24 0.02 0.127 0.225 0.208 48.95 West 30,000 0.32 0.38 0.02 0.171 0.307 0.340 54.56 Total 250,000 0.32 0.26 0.02 0.132 0.234 0.222 49.54

Table IIIb Mineral Resources beneath Optimised Pit Shell above Cdn$50.00/t NSR Cut-off


Cu (%)

Co (%)

Au (g/t)

Pt (g/t)

Pd (g/t)

NSR Cdn$/t

Indicated East 1,763,000 0.37 0.41 0.03 0.219 0.363 0.388 62.68 West 69,000 0.35 0.43 0.02 0.176 0.327 0.361 60.17 Total 1,832,000 0.37 0.41 0.03 0.218 0.361 0.387 62.59


The pits will be developed using double-benching with 8.5 m catch berms and a 75° bench face angle. This results in an inter-ramp slope of 55°. Pre-split blasting techniques will be required for the final 20-m high bench faces.

OPEN PIT OPTIMIZATION

The Shakespeare deposit is amenable to development as an open pit mine. To determine the economic limits of open pit mining, optimized open pit shells were developed using Whittle 4X pit optimization software.

For a polymetallic deposit, such as at the Shakespeare project, the Net Smelter Return (NSR) value of the resource is used to determine the net value of each mining block, and may be used in the same way as grades in a single-metal deposit in determining economic cut-offs.

The NSR for each block model was calculated utilizing a number of parameters including a provisional forecast of metal prices, estimated concentrate recoveries, smelter treatment charges, smelter payables, refining charges and exchange rate between US and Canadian dollars. These parameters are shown in Table IV.

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Table IV Parameters for Calculation of Net Smelter Return

Metal, unit

Metal Prices US$ per lb or oz

% Recoveryto Concentrate

Smelter Payables,%

Refining ChargesCdn$ per lb or oz

Ni, lb 5.158 76.0 92 0.64 Cu, lb 1.259 95.5 89 0.31 Co, lb 21.332 71.0 50 3.03 Au, oz 437.26 38.0 85 18.24 Pt, oz 780.81 75.0 85 18.24 Pd, oz 233.65 42.0 85 18.24

An estimated smelter treatment charge of Cdn$212/t was applied which, at an ore:concentrate ratio of 30.4:1, results in a smelter treatment charge of Cdn$6.97 per ore tonne milled. An exchange rate of US$0.8224/Cdn$1.00 was used.

Pit shells were optimized over an NSR value range of 10% to 110% of the value determined using the parameters set out above, in increments of 1%. The resulting optimization produced 74 nested pit shells, from which shell number 64 provided the maximum operating revenue and was selected for design purposes.

MINERAL RESERVES

All of the material designated as ore in the pit design was derived from Indicated Resources. That material, having been demonstrated to be economic in this feasibility study, is thus classified as a Probable Mineral Reserve using the CIM standards. The Mineral Reserve Statement for the Shakespeare Project is shown in Table V below.

The mineral reserves are determined by applying the Cdn$11.75/t NSR internal cut-off value to the modeled mineralization contained within the final pit design.

Table V Mineral Reserve Statement

Classification Tonnes %Ni %Cu %Co g/t Au g/t Pt g/t Pd Probable Mineral Reserve 11,226,000 0.33 0.35 .02 0.19 0.33 0.37

The mineral reserve estimate presented in the table above is current as of December 17, 2005.

PRODUCTION SCHEDULE

Ore will be mined at a rate of 4,500 t/d for approximately seven years (see Table VI). The average stripping ratio for the project is 5.15 tonnes of waste per tonne of ore. However, 10 million tonnes have been scheduled for pre-stripping and the ratio is 5.7:1 during the period Year 1 to Year 4. Subsequently, the ratio will fall, when operations will be recovering ore from the deepest benches of the East Pit.

The West Pit will be mined out by early Year 5, after which it will be used to store potentially acid generating mine rock.

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Table VI Production Schedule

Year0 Year1 Year2 Year3 Year4 Year5 Year6 Year7 Year8 Year9 Total 2006 2007 2008 2009 2010 2011 2012 2013 2014 2015

PRODUCTION Waste Mined ( 000 tonnes ) 10,000 10,403 10,403 8,212 9,672 7,117 1,825 180 - 57,813Ore Mined ( 000 tonnes ) - 1,643 1,643 1,643 1,643 1,643 1,643 1,371 - 11,226

Waste/Ore Ratio - 6.33 6.33 5.00 5.89 4.33 1.11 0.13 - 5.15Total Mined ( 000 tonnes ) 10,000 12,045 12,045 9,855 11,315 8,760 3,468 1,551 - 69,039Ore Grade Nickel ( % ) - 0.362 0.387 0.331 0.266 0.282 0.338 0.344 - 0.330 Copper ( % ) - 0.345 0.402 0.369 0.298 0.304 0.361 0.376 - 0.350 Cobalt ( % ) - 0.024 0.025 0.023 0.021 0.020 0.023 0.023 - 0.023 Platinum ( grams / tonne ) - 0.335 0.370 0.370 0.303 0.273 0.332 0.342 - 0.332 Palladium ( grams / tonne ) - 0.373 0.424 0.416 0.325 0.290 0.362 0.376 - 0.366 Gold ( grams / tonne ) - 0.179 0.205 0.203 0.164 0.153 0.195 0.209 - 0.186

OPEN PIT MINING OPERATIONS

Open pit operations will be carried out on two 12-hour shifts, seven days per week for 50 working weeks per year. The rotation schedule will have employees working 7 days on, 7 days off. All equipment will be diesel powered.

Grade control will be overseen by the mine geologists. Cuttings from blastholes in or near ore will be sampled for grade control purposes. The results of these samples will be used to determine which portion of the mineralization is above the internal cut-off grade of Cdn$ 11.75/tonne NSR. Ore blasting limits will be set accordingly.

METALLURGY AND PROCESS SELECTION

Three phases of metallurgical testing have been conducted by SGS Lakefield (SGS) during the development of the Shakespeare project. Work started in 2003 and culminated in the 2005 program which developed data for design of an on-site milling facility.

Preliminary flowsheet development included mineralogy, bulk flotation and initial nickel-copper separation.

The objective of the 2005 metallurgical test program was to develop an optimized process route to maximize nickel, copper and PGM recovery to a single concentrate with a target combined copper and nickel content of between 15 and 20%.

The 2005 feasibility study testwork program comprised both flowsheet development and variability test programs, including grindability work, mineralogical investigations and bench scale flotation tests.Three locked cycle flotation tests completed the program by producing a range of concentrate recovery and grades, shown in Table VII.

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Table VII Feasibility Study Recovery Estimates

Product 15% (Cu+Ni)

Product 18% (Cu+Ni)

Product 20% (Cu+Ni)

Element Feed Grade (% or g/t)

Recovery(%)

Grade (% or g/t)

Recovery(%)

Grade (% or g/t)

Recovery (%)

Grade (% or g/t)

Copper (Cu) 0.42 96.1 8.7 95.9 10.6 95.6 11.9 Nickel (Ni) 0.37 79.8 6.3 76.4 7.4 73.5 8.1 Platinum (Pt) 0.34 81.0 5.9 74.8 6.7 69.5 7.0 Palladium (Pd) 0.41 49.0 4.3 42.4 4.6 36.2 4.4 Gold (Au) 0.21 43.8 2.0 38.4 2.1 33.4 2.1

The process selected for the feasibility study is based on the interpretation of the results from the historic and 2005 metallurgical testwork programs and comprises primary crushing, semi-autogenous grinding (SAG) with pebble crushing, ball milling, flotation to produce a single bulk concentrate, concentrate dewatering, sulphide scalping from tailings and tailings disposal. Table VIII shows principal design criteria.

Table VIII Summary of Key Process Design Criteria

Parameter Units Value Source Primary Crushing Design throughput t/h 313 Calculation Crusher utilization % 60 Micon Grinding Nominal throughput t/h 204 Calculation Circuit utilization % 92 Micon SAG mill circuit product size (80% passing) µm 850 Testwork/Micon SAG mill estimated unit power consumption kWh/t 8.60 Micon/Calculation Grinding circuit product size (80% passing) µm 85 Testwork/Micon Ball mill estimated unit power consumption kWh/t 9.52 Micon/Calculation Flotation Equipment utilization % 92 Micon Primary rougher retention time min 5 Testwork Secondary rougher retention time min 35 Testwork First cleaner retention time min 14 Testwork Second cleaner retention time min 9 Testwork Third cleaner retention time min 5 Testwork Fourth cleaner retention time min 5 Testwork Scavenger concentrate regrind mill feed rate dry t/h Calculation Regrind mill circuit product size (80% passing) µm Testwork/Micon Regrind mill estimated unit power consumption kWh/t Micon/Calculation Tailings sulphide rejection min 12 Testwork Concentrate dewatering Equipment utilisation % 92 Micon Concentrate thickener underflow density wt% solids 65 Micon Filter cake water content wt% water 7 Micon Tailings (design and mass balance purposes) Nominal production rate dry t/h 175 Calculation Nominal solids feed density wt% solids 32 Calculation

Crushing, milling and flotation flowsheets are shown in Figures IV(a), (b) and (c) respectively, while Figure V shows the Shakepseare Project site layout.

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Figure IV (a) Flowsheet - Crushing & Stockpile Area

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Figure IV (b) Flowsheet - Grinding Area

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Figure IV (c) Flowsheet - Flotation Area

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Figure V Site Layout Plan

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PROCESS DESCRIPTION

The blasted ore from the open pit is hauled by mine trucks to the primary crusher located about 1,700 m from the middle point of the West and East pits. The 91-t capacity trucks dump the ore into a 135-t hopper from which an apron feeder is used to feed the primary 36-in by 48-in (0.91 by 1.22m) jaw crusher. The crusher, which is equipped with a dust collector, is covered with a winterized building. An operator’s cabin and hydraulic rock breaker are located at close proximity to the crusher.

The ore is crushed to minus 150 mm and transported on a short transverse 1.22-m wide conveyor belt to a 0.91-m wide stacking conveyor, which discharges onto a 4,500-t capacity stockpile.

Two reciprocating feeders located in a tunnel at the base of the stockpile are used to reclaim the crushed ore from the stockpile. The ore is then transported by a 0.91-m wide conveyor to the SAG mill, which is situated in the main mill building. A belt scale and standby feed hopper are installed on the SAG mill feed conveyor.

The grinding circuit comprises a 24 ft (7.31 m) high by 10 ft (3.05 m) long SAG mill, and a 15.5 ft (4.72 m) high by 21 ft (6.60 m) long ball mill. The SAG and ball mills are fitted with 4,000-HP (2,985-kW) and 3,000-HP (2,220-kW) motors, respectively.

The product from the SAG mill discharges onto a 1.22-m by 2.44-m double deck vibrating screen. Material larger than 15 mm is conveyed to a 5½-ft (1.68-m) diameter short head cone pebble crusher. The product from the crusher is recycled back to the SAG mill. The undersize product from the vibrating screen discharges into the mill discharge pump box where it is joined by the ball mill discharge stream and dilution process water.

The mill discharge products are pumped to a cyclone cluster, from which the overflow sizing 80% passing 85 µm, is routed to the flotation circuit. The cyclone cluster underflow feeds the ball mill.

The flotation circuit comprises two conditioners in series, a primary rougher stage (2 x 20-m3 tank cells), a secondary rougher stage (6 x 50-m3 tank cells), four cleaner stages and a tailings sulphide rejection stage (2 x 50-m3 tank cells). The cleaner circuits consist of 5 x 5-m3 primary conventional cells, 2 x 5-m3 secondary conventional cells, 3 x 1.5-m3 tertiary conventional cells and 2 x 1.5-m3 quaternary conventional cells.

The secondary rougher concentrate and is fed to the regrind mill circuit. The discharge from the 6.5-ft (1.98-m) high by 10-ft (3.05-m) long regrind mill, with 150-kW motor, is pumped to a cyclone cluster. The cyclone underflow returns to the mill while the overflow stream, sizing 80% passing 40 µm, feeds the primary cleaner cells.

The tailings product from the scavenger cells is combined with the primary cleaner tailings and pumped to the tailings sulphide rejection flotation cells, from which the separate concentrate and tailings products are pumped to their separate tailings disposal systems.

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The quaternary cleaner concentrate is the final product and is pumped to the concentrate thickener and filter section for dewatering.

The overflow from the 5-m diameter thickener is pumped to the process water tank while the underflow is pumped to the 5.2-m diameter concentrate stock tank. The thickened concentrate is pumped at a controlled rate from the stock tank and fed to the Larox filter. The filtrate product from the filter is recovered and pumped to the process water tank while the concentrate filter cake, containing about 7% moisture, is stored in a stock pile located on the ground floor of the mill building, under the Larox filter.

The sulphide tailings stream is pumped to lined disposal area situated within the tailings and effluent treatment catchment area.

The non-sulphide tailings are pumped to a high-compression tailings thickener where flocculant is added to aid settling. The thickener overflow stream is pumped to the settling pond while the underflow, containing 75% solids by weight, is pumped to the tailings co-disposal area, where it is mixed with mine rock from the open pit mine. Seepage from the co-disposal area is directed to the settling pond. The overflow from the settling pond gravitates to the recirculation pond where line is added for pH adjustment. The overflow from the recirculation pond is directed to the polishing pond where it is either reclaimed to the process water tank at the plant site or discharged to the environment.

PROCESS PLANT AND SITE

The main criteria considered during the site layout development for the process plant and surface infrastructure were:

• The minimum haulage distance from the pit to the crusher and to the mine rock storage.

• Minimal site visibility for cottagers on the edge of Agnew Lake.

• Minimal noise emissions from the plant site.

The general site layout is located on an area at the northeast corner of the West pit limit, at a distance averaging 1,700 m from the mid-point between the West and East pits. The total site area to be cleaned and levelled for construction is approximately 40 – 45 ha. The surface site plan has been designed to have a minimum footprint. Facilities comprise:

• A main mill/concentrator building with an adjacent building housing mechanical and electrical maintenance workshops below second floor offices, conference room, lunchroom and washrooms.

• Pre-fabricated trailers for the office complex, changehouse and laboratory.

• Mobile equipment maintenance shops and warehouse in a combined sructure,

• The main electrical substation located adjacent to the mill building.

Sewage generated at the mine site would be collected in septic holding tanks and treated in the septic tank and filtration bed system. All non-toxic garbage from the operation would be disposed of using a waste disposal contractor removing materials to an appropriate disposal site.

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PROCESS PLANT PRODUCTION SCHEDULE

Table IX presents the process plant production schedule.

Table IX Process Plant Production Schedule


Ore Milled ( 000 tonnes ) - 1,643 1,643 1,643 1,643 1,643 1,643 1,371 0 11,226 Mill Recovery

Nickel ( % ) 76.0 76.0 76.0 76.0 76.0 76.0 76.0 76.0 76.0 Copper ( % ) 95.5 95.5 95.5 95.5 95.5 95.5 95.5 95.5 95.5 Cobalt ( % ) 71.0 71.0 71.0 71.0 71.0 71.0 71.0 71.0 71.0 Platinum ( % ) 75.0 75.0 75.0 75.0 75.0 75.0 75.0 75.0 75.0 Palladium ( % ) 42.0 42.0 42.0 42.0 42.0 42.0 42.0 42.0 42.0 Gold ( % ) 38.0 38.0 38.0 38.0 38.0 38.0 38.0 38.0 38.0

Recovered Metal

Nickel (000 lbs) - 9,964 10,647 9,114 7,318 7,763 9,299 7,907 - 62,012 Copper (000 lbs) - 11,928 13,914 12,768 10,313 10,510 12,480 10,849 - 82,762 Cobalt 000 lbs) - 625 633 588 528 524 596 494 - 3,989 Platinum (000 oz) - 13.27 14.67 14.65 11.99 10.80 13.16 11.31 - 89.85 Palladium (000 oz) - 8.27 9.40 9.22 7.20 6.44 8.04 6.97 - 55.53 Gold (000 oz) - 3.59 4.11 4.08 3.30 3.06 3.91 3.51 - 25.56

Concentrate Grade ( % Ni ) - 8.19 7.80 7.50 7.47 7.65 7.69 7.59 - 7.71Concentrate Grade ( % Cu ) - 9.81 10.20 10.50 10.53 10.35 10.31 10.41 - 10.29Concentrate Produced (000 dmt) - 55.167 61.895 55.142 44.431 46.048 54.881 47.264 - 364.829Mass Pull 0.00% 3.36% 3.77% 3.36% 2.71% 2.80% 3.34% 3.45% 0.00% 3.25%Payability of Metal

Nickel ( % ) 92.0 92.0 92.0 92.0 92.0 92.0 92.0 92.0 92.0 92.0 Copper ( % ) 89.0 89.0 89.0 89.0 89.0 89.0 89.0 89.0 89.0 89.0 Cobalt ( % ) 50.0 50.0 50.0 50.0 50.0 50.0 50.0 50.0 50.0 50.0 Platinum ( % ) 85.0 85.0 85.0 85.0 85.0 85.0 85.0 85.0 85.0 85.0 Palladium ( % ) 85.0 85.0 85.0 85.0 85.0 85.0 85.0 85.0 85.0 85.0 Gold ( % ) 85.0 85.0 85.0 85.0 85.0 85.0 85.0 85.0 85.0 85.0

Payable Metal

Nickel ( 000 lbs ) - 9,167 9,796 8,385 6,733 7,142 8,555 7,275 - 57,051 Copper ( 000 lbs ) - 10,616 12,384 11,364 9,179 9,354 11,107 9,655 - 73,658 Cobalt ( 000 lbs ) - 313 317 294 264 262 298 247 - 1,994 Platinum (000 oz) - 11.28 12.47 12.45 10.19 9.18 11.19 9.61 - 76.37 Palladium (000 oz) - 7.03 7.99 7.84 6.12 5.48 6.83 5.92 - 47.20 Gold ( 000 oz ) - 3.05 3.49 3.47 2.80 2.60 3.32 2.98 - 21.73

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MINE ROCK AND TAILINGS DISPOSAL

The proposed deposition area is a large valley to the northeast of the pits. Potentially acid generating material (mine rock and tailings) will be disposed of within a sub-aqueous facility fully contained within a larger co-disposal area (CDA).

Non-sulphide process plant tailings will be thickened and pumped to the CDA. The main pipeline from the thickening plant will branch out to the east and west sections of the CDA where spigot lines will be used to deposit and integrate the thickened tailings and mine rock. Thickener overflow will be pumped by centrifugal pumps (one service, one standby) out to the settling/polishing pond before returning to the process plant.

A second tailings slurry pump and pipeline from the process plant will transport the pyrrhotite (sulphide) concentrate slurry directly to the CDA. Based on the geochemistry and disposal design, this stream does not require thickening and will therefore bypass the thickening plant.

WATER MANAGEMENT

Site water management will be undertaken by means of six ponds including one in the CDA that will permanently maintain the acid-generating materials in a submerged state.

All water collected on the property will be passed through a settling pond before treatment and recycling to the process plant.

Seven dams and dykes will be required for the water management system. All will be conventional zoned water retaining dams founded on overburden or grouted bedrock.

ENVIRONMENTAL MANAGEMENT AND PERMITTING

Various environmental baseline studies were completed for the Shakespeare Project in 2004 and 2005. N.A.R. Environmental Consultants Inc. (NAR) undertook several formal and informal desktop reviews to provide input on environmental permitting and present and future environmental management issues for the Preliminary Feasibility Study. In August 2004, several specific technical tasks were implemented by NAR to determine baseline conditions both within the physical limits of the project and the zone of potential impact, notably the receiving water environment (Agnew Lake).

Knight Piésold was retained early in 2005 to provide environmental permitting and management services. The field program was expanded in early to mid 2005 to include the collection of benthic macroinvertebrate community and sediment quality data from three stations in Agnew Lake, installation of several groundwater monitoring wells and collection and testing of groundwater samples, acid rock drainage testing of waste rock and tailings samples and the installation of an on site weather station. The surface water quality and quantity monitoring program continued. In early fall 2005 terrestrial habitat and heritage studies were completed for the project site. An environmental baseline report was prepared for the project and in support of the feasibility study.

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The results of the surface water, groundwater and sediment sampling programs are considered as background values prior to the pre-production and production phases of the Shakespeare project.

No areas of nesting, breeding or significant bird habitat were identified by NAR.

A field inspection was conducted by NAR in August 2005 on a series of ponds located north of the open pits. These ponds are not fish habitat as defined by the Fisheries Act. Also NAR reported that there was no evidence of undisturbed wetlands which supported unique plant assemblages or rare, threatened or endangered wildlife habitat.

The scope of work has been developed to complete air and noise studies as part of the permitting phase of the project.

HERITAGE STUDY

Stage 1 and Stage 2 archaeological and heritage impact assessments were completed for the project site by Horizon Archaeology in early fall 2005. It was concluded that there are no concerns related to the destruction of cultural materials by the continued development of this project.

CAPITAL EXPENDITURES

The total pre-production capital expenditure for the Shakespeare property, at a production rate of 4,500 t/d, is estimated at Cdn$118.7 million (including contingencies, expressed in 2005 constant money terms). This estimate is summarized in Table X

Table X Total Initial Capital Cost Estimate

Cdn$000 Mining Equipment 18,916 Pre-Stripping 10,784 Process Plant direct costs 48,978 Infrastructure direct costs 8,819 Tailings Dams (CDA) 6,672 High Compression Thickener (HCT) Plant 3,506

Total Direct Costs 97,675 Owner’s Costs 4,643 EPCM 7,965 Contingency 8,190

Total Capital Costs 118,473

The contingency was calculated as a specific percentage of each of the direct cost items. The provision of $8.19 million represents an average of 12% of the direct cost estimate for the process plant, infrastructure, the tailings co-disposal area and HCT plant.

Table VII shows a breakdown of direct capital costs, before contingency, for the process plant area only.

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Table XI Process Plant Direct Capital Cost Estimate

Direct Capital Cost - before contingency Total (Cdn$) Plant Site Preparation 2,901,670 Primary Crushing Area 3,242,453 Reclaim Tunnel Area 1,317,648 Pebble Crushing Area 1,606,678 Concentrator Area 27,077,318 Global Costs (Piping & Electricity) 12,832,028 Process Plant Direct Total 48,977,795

OPERATING COSTS

The costs summarized in Table VIII reflect the proposed overall mining rate of 35,000 t/d, and an ore production rate of 4,500 t/d.

Table XII Operating Cost Summary

Area Payroll Complement

Total Cost (Cdn$ per year)

Unit Cost (Cdn$ per tonne ore)

Mining labour 94 6,758,028 4.11 Plant labour 44 3,286,404 2.00 G&A labour 14 1,050,124 0.64

Sub-total labour 152 11,094,556 6.75 Mining consumables 10,909,485 6.64 Plant consumables, etc 13,817,549 8.41 G&A and other 1,149,000 0.70

Total 36,970,590 22.51 ECONOMIC ANALYSIS

The economic analysis takes as its base case the assumption of a reversion of metal prices to their long-term (in this case, 10-year) historical median Canadian dollar prices, expressed in 2005 money terms. Current price levels, whether lying above or below the 10-year median price, are assumed to regress exponentially toward the median, with a ‘decay’ half-life of five years. The resulting average prices over the life of the project, expressed in 2005 US dollars, are given in Table IX.

Table XIII Average Metal Price Forecasts

(Life of Mine, 2005 constant dollars)

Metal Unit Price Nickel (US$/lb) 5.48 Copper (US$/lb) 1.34 Cobalt (US$/lb) 20.05 Platinum (US$/oz) 805.30 Palladium (US$/oz) 225.20 Gold (US$/oz) 438.30

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The base exchange rate for the economic analysis is the average for 2005, of US$0.8224/Cdn$1.00. No escalation has been applied to the model, which is presented in 2005 constant money terms.

Table XIV shows a summary of NSR calcualtion and the annual cash inflows to the project.

Table XIV NSR Calculation Schedule


METAL PRICES Nickel ( US $ / pound ) 6.14 5.91 5.72 5.56 5.42 5.30 5.20 5.12 5.05 5.48 Copper ( US $ / pound ) 1.49 1.44 1.39 1.35 1.32 1.29 1.27 1.25 1.23 1.34 Cobalt ( US $ / pound ) 17.86 18.54 19.17 19.73 20.23 20.67 21.07 21.42 21.73 20.05 Platinum ( US $ / ounce ) 852.74 836.67 822.93 811.16 801.04 792.34 784.84 778.37 772.78 805.3 Palladium ( US$ / ounce ) 210.24 215.08 219.38 223.19 226.57 229.55 232.17 234.48 236.51 225.2 Gold ( US $ / ounce ) 440.27 439.62 439.07 438.58 438.16 437.79 437.47 437.19 436.95 438.3

Exchange Rate (US$/Cdn$) 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224

Metal Prices – Canadian $ Nickel (Cdn$/pound ) 7.46 7.19 6.95 6.76 6.59 6.45 6.33 6.23 6.14 6.67 Copper (Cdn$/pound ) 1.81 1.75 1.69 1.65 1.61 1.57 1.55 1.52 1.50 1.62 Cobalt (Cdn$/pound ) 21.71 22.55 23.31 23.99 24.59 25.14 25.62 26.04 26.42 24.38 Platinum (Cdn$/ounce) 1,036.91 1,017.38 1,000.67 986.35 974.05 963.47 954.35 946.48 939.69 979.17 Palladium (Cdn$/ounce) 255.65 261.53 266.76 271.40 275.50 279.12 282.32 285.13 287.60 273.87 Gold ( Cdn. $ / ounce ) 535.36 534.58 533.90 533.30 532.79 532.34 531.95 531.61 531.32 532.96

Gross Revenue from Sales Nickel - 65,869 68,112 56,657 44,374 46,063 54,141 45,289 - 380,504 Copper - 18,554 20,967 18,716 14,759 14,728 17,173 14,693 - 119,591 Cobalt - 7,049 7,380 7,055 6,497 6,581 7,637 6,431 - 48,630 Platinum - 11,476 12,474 12,280 9,926 8,846 10,678 9,098 - 74,779 Palladium - 1,838 2,132 2,127 1,686 1,528 1,929 1,689 - 12,928 Gold - 1,633 1,865 1,850 1,492 1,385 1,768 1,586 - 11,580

GROSS SALES REVENUE - 106,419 112,930 98,685 78,734 79,131 93,326 78,787 - 648,011Smelting - 11,695 13,122 11,690 9,419 9,762 11,635 10,020 - 77,344

Refining – Nickel - 5,867 6,269 5,366 4,309 4,571 5,475 4,656 - 36,513Refining – Copper - 3,291 3,839 3,523 2,845 2,900 3,443 2,993 - 22,834Refining – Cobalt - 947 959 891 800 793 903 748 - 6,043Refining – Platinum - 206 227 227 186 167 204 175 - 1,393Refining – Palladium - 128 146 143 112 100 125 108 - 861Refining – Gold - 56 64 63 51 47 61 54 - 396

Falconbridge Royalty - 1,263 1,325 1,152 915 912 1,072 900 - 7,539NET REVENUE (NSR) - 82,965 86,979 75,630 60,096 59,878 70,408 59,131 - 495,088

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Table XV shows the schedule of cash flows and presents the computation of Net Present Value (NPV) and Internal Rate of Return (IRR).

Table XV Annual Cash Flows and Discounted Cash Flow Valuation

(Thousand 2005 constant Canadian dollars)


NET REVENUE (NSR) - 82,965 86,979 75,630 60,096 59,878 70,408 59,131 - 495,088Waste Mining (consumables) - 11,218 11,218 8,856 10,431 7,676 1,968 194 - 51,561 Ore Mining (consumables) - 1,771 1,771 1,771 1,771 1,771 1,771 1,479 - 12,106 Mining Labour 4,328 4,749 4,749 4,749 5,089 4,478 2,817 1,884 - 32,845 Fleet Maintenance Labour 907 996 996 996 1,067 939 591 395 - 6,886 Mine Eng/Geology, incl camp 923 1,013 1,013 1,013 1,085 955 601 402 - 7,005 Crushing, Milling & Flotation - 16,567 16,567 16,567 16,567 16,567 16,567 13,831 - 113,232 Concentrate Transportation - 552 619 551 444 460 549 473 - 3,648 General and Administration 1,100 2,199 2,199 2,199 2,199 2,199 2,199 302 - 14,596 Total Operating Costs 7,257 39,065 39,132 36,703 38,654 35,045 27,063 18,959 - 241,878 OPERATING PROFIT (7,257) 43,900 47,847 38,927 21,442 24,834 43,345 40,172 - 253,210 Less: Federal/Provincial Taxes 70 255 605 535 1,906 7,881 9,014 15,682 14,560 266 50,773 PROFIT AFTER TAX (70) (7,512) 43,296 47,312 37,021 13,561 15,820 27,664 25,612 (266) 202,436 Less: Initial Capital Expendit:-

Mining Equipment 11,938 6,619 360.1 18,916 Pre-Stripping 10,784 10,784 Process Plant direct costs 14,693 34,244 48,978 Infrastructure direct costs 2,646 6,173 8,819 Tailings Dams 2,002 4,670 6,672 HiComprThicknr (HCT) 1,052 2,454 3,506 Owners Costs 1,393 3,250 4,643 EPCM 2,389 5,575 7,965 Contingency 2,457 5.733 8,190

Total Initial Capital Expend. 26,632 84,862 6,619 360 118,473 Sustaining/Repl.Capital Mining Equipment 133 - 286 - 419 Mine Services 75 75 75 75 300 Tailings & Water Mgmnt 165 165 165 165 165 825 Contingency/(Resid value) - 62 62 62 62 62 - (3,836) - (3,526) Permits, Recl & Closure 200 - 1,500 - 1,700 Change in Working Cap. (2,200) (5,000) 26,100 1,500 (2,800) (4,200) 100 3,400 (16,900) - -

ANNUAL CASH FLOW (24,702) (87,374) 10,275 45,150 39,386 17,459 15,206 24,264 44,847 (266) 84,246 %Disc. Pre-tax Aftertax Internal Rate of Return 20.0% 14.5% Net Present value 0% 135,019 84,246 CDN $000, Real 2005 terms 5.0% 78,312 43,108 10.0% 41,187 16,242 15.0% 16,516 (1,516) 20.0% (54) (13,331)

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The cash flow projection reflects a maximum cash outflow of Cdn$112 million at the end of the construction period when pre-stripping of the open pit is complete but no milling of ore has yet taken place. Year 2 is the first year of full production, and is cash positive despite an increase in working capital. Undiscounted payback occurs at the end of Year 5, leaving almost three full years of production ‘tail’.

The base case cash flow provides a pre-tax IRR of 20.0%, or 14.5% after tax. Net Present Value (NPV) of the project at a 10% (real) discount rate is Cdn$41.2 million pre-tax, or Cdn$16.2 million after tax.

SENSITIVITY STUDY

As shown in Figure VI, the NPV is most sensitive to metal prices, grade and recovery factors. It is less sensitive to operating costs and reserve tonnage, and is relatively insensitive to throughput and project capital costs.

Figure VI Sensitivity Study Results

Variances of around 15% from the base would be required in the technical parameters before project returns fall below an acceptable level. The precision of the underlying estimates lies inside this range and thus the sensitivity study results suggest the impact of project-specific technical risk on the project cash flow is manageable.

Sensitivity of NPV(10%)

-40,000

-20,000

0

20,000

40,000

60,000

NPV

(10%

) Cdn

$ 00

0 (2

005

mon

ey te

rms)

Capital cost estimate 27,299 25,282 23,173 20,958 18,649 16,242 13,663 10,902 8,059 5,077 1,957

Operating cost estimate 36,634 32,907 29,014 24,949 20,689 16,242 11,469 6,477 1,217 -4,419 -10,385

Revenue drivers (Grade, Recoveries,US$ Prices, Exchange rate)

-35,143 -25,126 -15,182 -4,909 5,547 16,242 27,146 38,540 50,464 62,859 75,621

Throughput 10,966 12,156 12,626 13,656 14,738 16,242 16,606 17,047 18,742 18,670 18,611

Reserve tonnage -3,121 928 5,176 7,791 11,977 16,242 17,815 19,978 22,240 25,260 27,444

80% 84% 87% 91% 96% 100% 105% 109% 114% 120% 125%

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Metal price sensitivity was further tested by evaluating the project using the average metal prices for 2005 held constant for the life of mine. Nickel and copper prices are thus 23% and 22% higher on average over the life of mine, and the gross revenue increase is around 17%. Under this assumption, the IRR rises to 23.9% and the NPV at a discount rate of 10% per year rises by Cdn$38.7 million to Cdn$54.8 million after tax, demonstrating considerable upside potential over the base case.

RECOMMENDATIONS AND CONCLUSIONS

The Shakespeare project contains an economic mineral reserve and is worthy of continued development through detailed engineering and construction to produce 4,500 t/d of mined ore and subsequent metal concentrate for sale. Environmental studies indicate that no significant negative impact from the project will be encountered. The co-disposal concept for tailings and mine rock will mitigate effects from these disposal systems.

It is unlikely that ore grade, metal recovery or operating cost can be significantly improved. The main economic improvement that may be contemplated is the reduction in capital cost by the selection of used equipment and buildings, and Micon recommends an immediate search for suitable items.

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TABLE OF CONTENTS

Page

EXECUTIVE SUMMARY i

1.0 INTRODUCTION AND TERMS OF REFERENCE.................................................. 1 1.1 Project Description ............................................................................................. 1 1.2 Project Ownership .............................................................................................. 1 1.3 Work Carried Out ............................................................................................... 3 1.4 Preparation of the Feasibility Study ................................................................... 3 1.5 Responsibility ..................................................................................................... 5 1.6 Basis of the Feasibility Study ............................................................................. 5

1.6.1 Technical and Economic Parameters ...................................................... 5 1.7 Disclaimer........................................................................................................... 7

2.0 PROPERTY DESCRIPTION AND LOCATION........................................................ 9 2.1 Access and Infrastructure ................................................................................. 11 2.2 Physiography And Climate............................................................................... 11

3.0 GEOLOGY AND MINERAL RESOURCES............................................................. 13 3.1 Regional Geology............................................................................................. 13 3.2 Local Geology .................................................................................................. 14 3.3 Deposit Types ................................................................................................... 15 3.4 Mineralization................................................................................................... 16 3.5 Exploration ....................................................................................................... 17

3.5.1 Exploration Following the 2003 Resource Estimate............................. 18 3.5.2 Exploration Following the 2004 Resource Estimate............................. 18

3.6 Drilling ............................................................................................................. 20 3.6.1 Falconbridge Drilling............................................................................ 20 3.6.2 URSA Drilling Programs ...................................................................... 22 3.6.3 Other Drilling........................................................................................ 24 3.6.4 Summary ............................................................................................... 24 3.6.5 Database Checks ................................................................................... 25 3.6.6 Use of Falconbridge Data ..................................................................... 25

3.7 Mineral Resources ............................................................................................ 25 3.7.1 Mineral Resource Estimation Methodology ......................................... 25

3.8 Resource Classification .................................................................................... 37 3.9 Mineral Resources ............................................................................................ 37 3.10 Confirmation Of Estimation ............................................................................. 39

4.0 MINING AND MINERAL RESERVES..................................................................... 40 4.1 Mining Method................................................................................................. 40 4.2 Pit Slope Analysis............................................................................................. 40

4.2.1 Review of Wardrop Preliminary Design............................................... 40 4.2.2 Compressive Strength Testing .............................................................. 41 4.2.3 Geotechnical Field Investigation .......................................................... 42

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4.2.4 Pit Wall Design ..................................................................................... 45 4.2.5 Pit Water Inflow.................................................................................... 45

4.3 Open Pit design................................................................................................. 46 4.3.1 Net Smelter Return Values ................................................................... 46 4.3.2 Model Preparation................................................................................. 46 4.3.3 Optimization Parameters....................................................................... 47 4.3.4 Generation of Optimized Pit Shells ...................................................... 47 4.3.5 Open Pit Design .................................................................................... 48

4.4 Mineral Reserves .............................................................................................. 50 4.4.1 Mining Dilution & Recovery ................................................................ 50 4.4.2 Statement of Mineral Reserves ............................................................. 50

4.5 Open Pit Production Schedule.......................................................................... 51 4.6 Open Pit Mining Operations............................................................................. 51

4.6.1 Drilling and Blasting............................................................................. 52 4.6.2 Load and Haul ....................................................................................... 52 4.6.3 Grade Control........................................................................................ 52

4.7 Mine Support Facilities .................................................................................... 52 4.7.1 Logistical support.................................................................................. 52 4.7.2 Dewatering ............................................................................................ 52 4.7.3 Mine Dry ............................................................................................... 52

4.8 Personnel .......................................................................................................... 53 4.9 Mining Operating costs .................................................................................... 53 4.10 Mining Capital Costs ........................................................................................ 54

5.0 METALLURGY............................................................................................................ 55 5.1 2003 TestWork ................................................................................................. 55 5.2 2004 Testwork .................................................................................................. 56

5.2.1 2004 Locked Cycle Tests ...................................................................... 59 5.2.2 Bulk Concentrate Tests ......................................................................... 59

5.3 2005 Testwork .................................................................................................. 60 5.3.1 Metallurgical Samples........................................................................... 61 5.3.2 Characterization .................................................................................... 62 5.3.3 Grindability Testwork ........................................................................... 64 5.3.4 Flotation Testwork ................................................................................ 65 5.3.5 Locked Cycle Tests ............................................................................... 70 5.3.6 Variability testwork............................................................................... 71

5.4 Metallurgical Recovery Estimates.................................................................... 76 5.5 Process Selection .............................................................................................. 77

5.5.1 Grinding Mill Sizing ............................................................................. 77

6.0 PROCESS PLANT........................................................................................................ 78 6.1 Process Design Criteria .................................................................................... 78

6.1.1 Design Criteria ...................................................................................... 78 6.2 Process Description .......................................................................................... 79

6.2.1 Crushing and Storage Facility............................................................... 80 6.2.2 Grinding ................................................................................................ 80 6.2.3 Flotation Circuit .................................................................................... 80

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6.2.4 Concentrate Thickening and Filtering................................................... 81 6.2.5 Tailings Disposal and Effluent Treatment ............................................ 81 6.2.6 Services, Utilities, Process and Fresh Water......................................... 81 6.2.7 Reagents ................................................................................................ 82 6.2.8 On-Stream Analysis .............................................................................. 82

6.3 Equipment List ................................................................................................. 82 6.4 Mass Balance.................................................................................................... 83 6.5 Process Plant Production Schedule .................................................................. 85

7.0 INFRASTRUCTURE ................................................................................................... 86 7.1 Plant And Site Layout ...................................................................................... 86 7.2 Roads ................................................................................................................ 87

7.2.1 Access Road.......................................................................................... 87 7.2.2 Site Roads ............................................................................................. 87 7.2.3 Plant Gate and Site Fencing .................................................................. 87 7.2.4 Parking Area ......................................................................................... 88

7.3 Crushing Facility and Processing Plant Layout................................................ 88 7.3.1 Crusher .................................................................................................. 88 7.3.2 Concentrator.......................................................................................... 88 7.3.3 Office Complex..................................................................................... 88 7.3.4 Laboratory Building.............................................................................. 89 7.3.5 Mine Equipment Maintenance Building ............................................... 89 7.3.6 Warehouse............................................................................................. 89 7.3.7 Truck Scale ........................................................................................... 89

7.4 Services............................................................................................................. 89 7.4.1 Fuel Storage and Fuelling Station......................................................... 89 7.4.2 Mine Explosive Storage ........................................................................ 90 7.4.3 Bulk Explosive Plant............................................................................. 90 7.4.4 Water Systems....................................................................................... 90 7.4.5 Potable Water Treatment ...................................................................... 91 7.4.6 Heating, Ventilation and Air Conditioning........................................... 91 7.4.7 Plant Mobile Equipment ....................................................................... 91

7.5 Electrical Power................................................................................................ 92 7.5.1 Power Requirement & Supply .............................................................. 92 7.5.2 Main Electrical Room ........................................................................... 92 7.5.3 Main Electrical Room ........................................................................... 93 7.5.4 On-site reticulation................................................................................ 93

7.6 Automation & Instrumentation......................................................................... 94 7.6.1 PLC Controls......................................................................................... 94 7.6.2 Instrumentation ..................................................................................... 94

7.7 Waste Disposal ................................................................................................. 95 7.7.1 Sanitary Waste Water............................................................................ 95 7.7.2 Garbage Disposal .................................................................................. 95

7.8 Telecommunications and Computer Networking............................................. 95 7.8.1 Telecommunications ............................................................................. 95 7.8.2 Computers and Networking .................................................................. 95

xxviii

8.0 MINE ROCK AND TAILINGS MANAGEMENT ................................................... 96 8.1 Introduction ...................................................................................................... 96 8.2 Co-disposal Concept......................................................................................... 96 8.3 Laboratory Evaluation ...................................................................................... 98

8.3.1 Laboratory Summary ............................................................................ 98 8.3.2 Additional Laboratory Testwork........................................................... 98 8.3.3 Laboratory Conclusion.......................................................................... 99

8.4 Geochemistry.................................................................................................... 99 8.4.1 Principal Lithologies ............................................................................. 99

8.5 Physical Properties of the Tailings ................................................................. 101 8.6 Geotechnical Considerations .......................................................................... 101 8.7 Site Precipitation and Evaporation ................................................................. 101 8.8 Seismic Risk ................................................................................................... 102 8.9 Closure............................................................................................................ 102 8.10 Design Drivers................................................................................................ 102 8.11 Site Selection for the Co-disposal Facility ..................................................... 103 8.12 Operating Data................................................................................................ 103

8.12.1 Filling Plan.......................................................................................... 104 8.13 Environmental Planning ................................................................................. 106

9.0 SITE WATER MANAGEMENT .............................................................................. 107 9.1 Water Management ........................................................................................ 107 9.2 Dams and Dykes............................................................................................. 107 9.3 Flow Model .................................................................................................... 108

10.0 ENVIRONMENTAL PERMITTING AND MANAGEMENT .............................. 109 10.1 Summary of Baseline Studies......................................................................... 110

10.1.1 Surface Water...................................................................................... 110 10.1.2 Groundwater........................................................................................ 112 10.1.3 Sediment.............................................................................................. 112

10.2 Terrestrial Habitat........................................................................................... 113 10.3 Aquatic Habitat............................................................................................... 113 10.4 Air and Noise.................................................................................................. 114 10.5 Mine Rock and Tailings Characterization ...................................................... 114

10.5.1 Mine Rock........................................................................................... 114 10.5.2 Tailings................................................................................................ 115

10.6 Meteorology ................................................................................................... 116 10.7 Heritage Study ................................................................................................ 116

11.0 CAPITAL EXPENDITURES .................................................................................... 117 11.1 Basis of Estimate ............................................................................................ 117

11.1.1 Currency Base Date and Exchange Rate............................................. 117 11.1.2 Labor Hourly Rate .............................................................................. 117 11.1.3 Civil Work, Concrete Quantities and Unit Costs ................................ 117 11.1.4 Structural Buildings Quantities and Unit Costs .................................. 117 11.1.5 Process Equipment .............................................................................. 118 11.1.6 Piping and Pipelines............................................................................ 118

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11.2 Project Capital Costs Summary...................................................................... 118 11.3 Mining Capital Costs ...................................................................................... 119 11.4 Process Plant Capital Costs ............................................................................ 120 11.5 Infrastructure Capital Costs ............................................................................ 121 11.6 Tailings Management Area Capital Costs ...................................................... 121 11.7 Capital Cost Summary Schedule .................................................................... 121

12.0 OPERATING COSTS ................................................................................................ 123 12.1 Mining Operating Costs ................................................................................. 123 12.2 Ore Processing Costs ...................................................................................... 123 12.3 General and Administration Operating Costs................................................. 124 12.4 Summary of Operating Costs ......................................................................... 124

13.0 ECONOMIC ANALYSIS........................................................................................... 125 13.1 Macro-Economic Assumptions ...................................................................... 125

13.1.1 Metal Price Forecasts .......................................................................... 125 13.1.2 Dollar Exchange Rate ......................................................................... 125 13.1.3 Escalation and Money terms ............................................................... 125

13.2 Base Case Model ............................................................................................ 125 13.2.1 Working Capital .................................................................................. 125 13.2.2 Taxation .............................................................................................. 125 13.2.3 Discounted Cash Flow Valuation ....................................................... 126

13.3 Sensitivity Study............................................................................................. 128

14.0 PROJECT IMPLEMENTATION SCHEDULE...................................................... 130 14.1 General ........................................................................................................... 130 14.2 Schedule of Activities..................................................................................... 130

14.2.1 Engineering ......................................................................................... 130 14.2.2 Procurement ........................................................................................ 130 14.2.3 Mine Pre-Stripping.............................................................................. 130 14.2.4 Construction ........................................................................................ 130

14.3 Priority Activities ........................................................................................... 131

15.0 CONCLUSIONS ......................................................................................................... 133

16.0 REFERENCES............................................................................................................ 134

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LIST OF TABLES

Page

Table I URSA Diamond Drill Holes......................................................................... iii Table II Grade Capping Values...................................................................................vi Table IIIa Mineral Resources within Optimised Pit Shell above Cdn$24.23/t

NSR Cut-off...................................................................................................vii Table IIIb Mineral Resources beneath Optimised Pit Shell above Cdn$50.00/t

NSR Cut-off...................................................................................................vii Table IV Parameters for Calculation of Net Smelter Return ..................................... viii Table V Mineral Reserve Statement......................................................................... viii Table VI Production Schedule ......................................................................................ix Table VII Feasibility Study Recovery Estimates ............................................................x Table VIII Summary of Key Process Design Criteria ......................................................x Table IX Process Plant Production Schedule.............................................................xvii Table X Total Initial Capital Cost Estimate ..............................................................xix Table XI Process Plant Direct Capital Cost Estimate ..................................................xx Table XII Operating Cost Summary .............................................................................xx Table XIII Average Metal Price Forecasts (Life of Mine, 2005 constant dollars).........xx Table XIV NSR Calculation Schedule ..........................................................................xxi Table XV Annual Cash Flows and Discounted Cash Flow Valuation

(Thousand 2005 constant Canadian dollars)................................................xxii Table.1.4.1 Feasibility Study Team...................................................................................4 Table 1.8.1 List of Abbreviations ......................................................................................7 Table 2.0.1 List of Claims .................................................................................................9 Table 3.6.1 Pre- 2002 Diamond Drill Holes....................................................................20 Table 3.6.2 URSA Diamond Drill Holes since 2002.......................................................22 Table 3.7.1 Block Model Rock Codes.............................................................................30 Table.3.7.2 Grade Capping Values..................................................................................31 Table 3.7.3 Block Model Grade Interpolation Parameters Kriging (Indicated)

and ID2 (Indicated First Pass) .......................................................................33 Table 3.7.4 Block Model Grade Interpolation Parameters Kriging (Indicated)

and ID2 (Indicated First Pass) .......................................................................34 Table 3.7.5 Block Model Grade Interpolation Parameters Kriging (Inferred) and

ID2 (Inferred).................................................................................................35 Table.3.7.6 Bulk Specific Gravity ...................................................................................36 Table 3.8.1 Grade Block Coding by Confidence Category .............................................37 Table 3.9.1 Parameters Used in Pit Optimization (Cdn$ per tonne) ...............................38

xxxi

Table 3.9.2 Shakespeare Open Pit Mineral Resources (Cut-off Cdn$24.23/t NSR) ..............................................................................................................38

Table 3.9.3 Shakespeare Underground Mineral Resources (Cut-off Cdn$24.23/t NSR) ..............................................................................................................38

Table 3.10.1 Comparison of Weighted Average Grade of Database to Total Block Model Average Grade .........................................................................39

Table 4.2.1 Summary of Point Load Strength Results (Wardrop)...................................41 Table 4.2.2 Empirical Rock Mass Properties (Wardrop).................................................41 Table 4.2.3 Orientation of Structural Discontinuities......................................................41 Table 4.2.4 Compressive Strengths by Lithology from Point Load Data........................42 Table 4.2.5 Summary of Geotechnical Drill Holes .........................................................42 Table 4.2.6 Hydraulic Conductivity Estimates................................................................43 Table 4.2.7 Summary of Discontinuity Sets for Surface Mapping and Optical

Televiewer Logging.......................................................................................44 Table 4.2.8 Shakespeare Project Rock Mass Classification ............................................44 Table 4.2.9 Wall Slope Parameters..................................................................................45 Table 4.3.1 Parameters for Calculation of NSR ..............................................................46 Table 4.3.2 Whittle Pit Optimization Parameters ............................................................47 Table 4.4.1 Mineral Reserve Statement (Cut-off Cdn$11.75/t NSR)..............................50 Table 4.5.1 Open Pit Production Schedule ......................................................................51 Table 4.8.1 Mining Personnel..........................................................................................53 Table 4.9.1 Mining Operating Cost (Cdn$/t rock mined)................................................54 Table 4.10.1 Principal Mining Equipment Items...............................................................54 Table 5.1.1 2003 Sample Assays .....................................................................................55 Table 5.1.2 Comparison Rougher Flotation, Tests F2 and F9 (2003) .............................56 Table 5.1.3 Cleaning Flotation Tests (2003) ...................................................................56 Table 5.2.1 2004 Composite Sample Assays...................................................................57 Table 5.2.2 2004 Tests - Summary of Results .................................................................58 Table 5.2.3 Locked Cycle Test Results (2004)................................................................59 Table 5.2.4 Bulk Concentrate Test F-29 (2004) ..............................................................60 Table 5.2.5 Predicted Optimum Metal Recoveries, 2004 Testwork................................60 Table 5.3.1 Metallurgical Composite Sample (2005)......................................................61 Table 5.3.2 Grindability Samples (2005)........................................................................61 Table 5.3.3 Metallurgical Variability Samples (2005) ....................................................62 Table 5.3.4 Metallurgical Composite (2005) Feed Main Element Analysis ...................63 Table 5.3.5 Metallurgical Composite (2005) Feed Major and Minor Element

Analysis .........................................................................................................63 Table 5.3.6 Metallurgical Composite (2005) Summary of Qualitative X-ray

Diffraction Analysis.......................................................................................63

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Table 5.3.7 Summary of Results from SGS Rapid Mineral Scan....................................64 Table 5.3.8 Bond Ball Mill Work Indices .......................................................................64 Table 5.3.9 Summary of MinnovEX SPI Testwork Results............................................65 Table 5.3.10 MacPherson Testwork Results .....................................................................65 Table 5.3.11 Summary of Variability Grinding Testwork Results....................................66 Table 5.3.12 Rougher Flotation Test Results (Stage 1).....................................................67 Table 5.3.13 Rougher Flotation Test Results (Stage 2).....................................................68 Table 5.3.14 Summary of Locked Cycle Flotation Test Conditions .................................70 Table 5.3.15 Locked Cycle Flotation Test Results............................................................71 Table 5.3.16 Full Element Analysis of Shakespeare Flotation Concentrate .....................72 Table 5.3.17 Variability Samples Head Grades.................................................................73 Table 5.3.18 Summary of Variability Flotation Test Results ............................................74 Table 5.4.1 Feasibility Study Recovery Estimates ..........................................................76 Table 5.5.1 Comparison of Mill Circuit Selections .........................................................77 Table 6.1.1 Process Design Basis ....................................................................................78 Table 6.1.2 Summary of Key Process Design Criteria ....................................................79 Table 6.2.1 List of Flowsheet Drawings..........................................................................79 Table 6.4.1 Process Mass and Water Balance .................................................................83 Table 6.5.1 Process Plant Production Schedule...............................................................85 Table 7.0.1 List of Infrastructural Drawings ...................................................................86 Table 7.5.1 List of Electical and Automation Drawings .................................................92 Table 7.5.2 Estimated Total Power Requirement ............................................................93 Table 11.2.1 Total Initial Capital Cost Estimate .............................................................119 Table 11.3.1 Mining Equipment Capital (Cdn$, 2005 terms) .........................................120 Table 11.4.1 Process Plant Direct Capital Cost Estimate (Cdn$, 2005 terms)................120 Table 11.5.1 Infrastructure Direct Capital Costs .............................................................121 Table 11.6.1 Tailings Area Direct Capital Costs .............................................................121 Table 11.7.1 Capital Cost Summary Schedule ................................................................122 Table 12.1.1 Mining Consumables Cost/tonne moved (Ore+Waste)..............................123 Table 12.1.2 Mine Operating Costs, including Labour ...................................................123 Table 12.2.1 Process Operating Costs .............................................................................124 Table 12.3.1 General and Administrative Operating Costs .............................................124 Table 12.4.1 Summary of Operating Costs......................................................................124 Table 13.1.1 Average Metal Price Forecasts (Life of Mine, 2005 constant

dollars) .........................................................................................................125 Table 13.2.1 NSR Calculation Schedule .........................................................................126 Table 13.2.2 Annual Cash Flows and Discounted Cash Flow Valuation

(Thousand 2005 constant Canadian dollars)................................................127

xxxiii

LIST OF FIGURES

Page

Figure I Project Location Plan..........................................................................................ii Figure II Shakespeare Property Geological Map..............................................................iv Figure III Three-Dimensional View (looking South) showing Geological

Domains and Drill Hole Traces ...........................................................................v Figure IV (a) Flowsheet - Crushing & Stockpile Area ............................................................xi Figure IV (b) Flowsheet - Grinding Area................................................................................xii Figure IV (c) Flowsheet - Flotation Area.............................................................................. xiii Figure V Site Layout Plan...............................................................................................xiv Figure VI Sensitivity Study Results .............................................................................. xxiii Figure 1.1 Project Location Plan..........................................................................................2 Figure 2.1 Shakespeare Project Claim Configuration ........................................................10 Figure 2.2 Shakespeare Project Location and Access Plan................................................12 Figure 3.1 Shakespeare Property Geological Map.............................................................19 Figure 3.2 Cumulative Distribution Curve, Nickel (Whole Database) ..............................27 Figure 3.3 Cumulative Distribution Curve, Nickel Blebby Domain..................................28 Figure 3.4 Cumulative Distribution Curve, Nickel Disseminated Domain .......................28 Figure 3.5 Example Section Showing/Domain Interpretation (Section 1+00 E,

Looking East).....................................................................................................29 Figure 3.6 3-Dimensional View of East Zone (Left Side) and West Zone (Right

Side) Geological Domains and Drill Hole Traces .............................................30 Figure 4.1 NPV of Nested Pit Shells..................................................................................48 Figure 4.2 Plan showing Ultimate Pit ................................................................................49 Figure 5.1 Testing Flow Sheet ...........................................................................................58 Figure 5.2 Copper to Pyrrhotite Flotation Selectivity ........................................................68 Figure 5.3 Nickel to Pyrrhotite Flotation Selectivity .........................................................69 Figure 5.4 Locked Cycle Flotation Test Results ................................................................71 Figure 5.5 Variability Samples Comparison of Nickel and Copper Feed Analyses ..........72 Figure 5.6 Variability Flotation Test Metal Recoveries vs Concentrate Grade .................75 Figure 5.7 Variability Tests – Head Grade versus Recovery .............................................76 Figure 8.1 Layout of Project Site showing Co-Disposal Area .........................................105 Figure 10.1 Plan Showing Water Quality Monitoring Stations .........................................111 Figure 13.1 Sensitivity Study Results ................................................................................128 Figure 14.1 Project Schedule..............................................................................................132

xxxiv

LIST OF APPENDICES

VOLUME II

Appendix 1 Kallio, Eric A., P.Geo., Technical Report for the Shakespeare Property, Shakespeare Township, Ontario, NTS 41I/5, for URSA Major Minerals Incorporated, November 28, 2002. (Volume 1).

Appendix 2 (a) Mineral Resource Estimation supporting data (b) P&E Mining Consultants Inc. Open Pit design and Capital Cost estimate; Supporting Schedules for Operating Cost estimates (Mining, Processing, G&A)

Appendix 3 Met-Chem Canada Inc., Shakespeare Project – Capital Cost Estimate – Ore Processing and Infrastructure, Flowsheets, Mechanical Drawings, Electrical and Automation Drawings.

VOLUME III

Appendix 4 Golder Associates Ltd., Feasibility Level Pit Slope Design and Water Inflow Study Shakespeare Project Webbwood, Ontario, Canada, January 2006.

VOLUME IV

Appendix 5 Golder Paste Technology Ltd., Trade-off Study Tailings and Mine Rock Disposal, November 11, 2005.

Appendix 6 Golder Paste Technology Ltd., Feasibility Study Tailings and Mine Rock Disposal, Volume 1, December 21, 2005.

Appendix 7 Golder Associates Ltd., Feasibility Study Tailings and Mine Rock Disposal, Volume 2, Co-disposal and Water Management, January 7, 2006.

VOLUME V

Appendix 8 SGS Lakefield Research Limited, Proposed Grinding System for the Shakespeare Deposit Based on Small Scale Data, Project 10044-139 – Report 1, February 3, 2006.

Appendix 9 SGS Lakefield Research Limited, An Update on Metallurgical Variability Testing of the Shakespeare Ores, Project 10616-003 – Report 2, February 3, 2006.

Appendix 10 SGS Lakefield Research Limited, An Update on Metallurgical Flowsheet Development for the Shakespeare Ores, Project 10616-003 – Report 1, February 3, 2006.

xxxv

VOLUME VI

Appendix 11 SGS Lakefield Research Limited, Acid Rock Drainage and Metal Leaching Characterization of Waste Rock – Shakespeare Property, Interim Report 5, November 14, 2005.

Appendix 12 SGS Lakefield Research Limited, Acid Rock Drainage and Metal Leaching Characterization of Tailings – Shakespeare Project, Interim Report 3, November 25, 2005.

VOLUME VII

Appendix 13 Knight Piésold Consulting, Environmental Baseline Report for Feasibility Study (Ref NB101-00222/1-2), December, 2005. (Volume 1).

VOLUME VIII

Appendix 13 continued

Knight Piésold Consulting, Environmental Baseline Report for Feasibility Study (Ref NB101-00222/1-2), December, 2005. Appendices (Volume 2)

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1.0 INTRODUCTION AND TERMS OF REFERENCE

1.1 PROJECT DESCRIPTION

Micon International Limited (Micon) has been retained by Dr. Richard H. Sutcliffe, President and CEO of URSA Major Minerals Incorporated (URSA), to prepare a feasibility study concerning URSA’s Shakespeare project, a deposit comprising nickel-copper mineralization with associated cobalt, platinum, palladium and gold. The deposit has potential for development as an open pit mining operation. Geological exploration and diamond drilling programs have been carried out, resulting in a revised mineral resource estimate. This resource estimate, together with geotechnical information, has been used in the construction of an open pit mine design and production schedule. Metallurgical testwork has been used to provide parameters for the design of a crushing, milling and flotation concentrator for the production of a saleable bulk sulphide concentrate, and in the design of a co-disposal area for inert mine rock and non-sulphidic tailings. Baseline environmental studies have also been undertaken.

The Shakespeare project is located 1 km north of the north shore of Agnew Lake, and 10 km north of the village of Webbwood, Ontario. Webbwood lies 20 km west of Espanola on Highway 17, approximately 80 km west of Sudbury.

1.2 PROJECT OWNERSHIP

The project is a joint venture with Falconbridge Limited (Falconbridge). The 28 leased and patented claims which make up the core of the Shakespeare joint venture are 75% legally owned by URSA and 25% legally owned by Falconbridge. The joint venture agreement, dated June 16, 2000, gave URSA the right to acquire up to a 75% interest in the Shakespeare Property. URSA is now fully earned into the project. The agreement also provides that if URSA acquires an interest in a property which is contiguous with the joint venture property, Falconbridge will have 30 days in which to elect to make the contiguous property part of the joint venture, subject to certain conditions. Three additional claims, totalling 36 units, have been added. These claims constitute the Falconbridge joint venture. Other claims staked by URSA in the immediate area are not subject to the joint venture.

Upon earn in by URSA, Falconbridge chose not to participate in the feasibility study. Consequently, Falconbridge has been diluted and it is reported to Micon that URSA now has an acknowledged 86% beneficial interest. Upon URSA reaching 90%, Falconbridge’s interest can be converted to a 1.5% net smelter return (NSR) royalty.

North American Palladium (NAP) has entered into an agreement with URSA whereby, upon completion of the feasibility study described herein, it has 180 days to exercise an option to acquire a 60% undivided interest in URSA’s interest in the Shakespeare Property, and become the operator, by making aggregate payments of Cdn$1.5 million to URSA and securing the project financing for commercial production. URSA intends that the Cdn$1.5 million will be used for expenditures to be incurred in connection with the Shakespeare full feasibility program.

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Figure 1.1 Project Location Plan

3

Upon completion of the feasibility study, NAP may elect to form a joint venture and, on arranging financing for commercial production, NAP shall be entitled to an undivided 60% interest in URSA’s interest and become operator of the property. URSA was the operator of the feasibility program.

1.3 WORK CARRIED OUT

The Shakespeare nickel-copper deposit is hosted within gabbroic rocks situated along the north contact between the mafic intrusive body which crosses the property and quartzites of the Mississagi Formation.

The bulk of the mineralization occurs as a broad conformable zone of magmatic sulphides located approximately 30 m to 80 m below the upper contact of the intrusion with the quartzites. The overall dip of the deposit is steeply northwards. The total strike length of mineralization exceeds 1,500 m.

Exploration work by URSA since 2000 has involved digital compilation, geological mapping, sampling, geophysics and diamond drilling. Most of the drill holes intersected a wide interval of sulphide mineralization and collectively defined a single layer of nickel-copper-gold-PGM-bearing sulphide mineralization divided into two adjacent zones known as East and West. The East Zone was discovered in 2002. The West Zone is the original Shakespeare deposit discovered by Falconbridge in the early 1940s. Typical widths are approximately 40 m and grades for nickel, copper and precious metals in the East Zone are generally higher than the average grade of the original Shakespeare, or West, deposit.

URSA has completed several phases of drilling on the East Zone (Summer/Fall, 2002; November, 2002 to March, 2003; May, 2003 to February 2004; February to September, 2004 and March to June, 2005 programs), as well as the infill program on the West Zone (May, 2003 to February 2004; February to September, 2004; and March to June, 2005 programs). Approximately 1,200 m of strike length has been drilled off at a nominal 60-m or 30-m spacing and an additional 375 m has been drilled at 120-m spacing. Within this 1,575 m length, two zones of mineralization (the East and West) have been identified with an aggregate strike length of approximately 1,435 m. The East and West Zones are separated by about 140 m of sparsely mineralized ground believed to be the result of an oblique fault offset. Approximately the last 300 m of strike length is at depth and unlikely to be mined by open pit methods. It does appear to have sufficient grade to justify further study for mining by underground mining methods late in the project’s life.

Micon has completed two previous mineral resource estimates for the Shakespeare project, the initial estimate in 2003 and a second one released in 2004 after the completion of the 2003/2004 drilling program. With the completion of the 2005 drilling, the mineral resource estimate has been updated and forms the basis of the present feasibility study.

1.4 PREPARATION OF THE FEASIBILITY STUDY

In March, 2005, URSA initiated the study reported herein using a combination of URSA personnel Micon and other specialist consultants. URSA provided metallurgical samples and site support for the various consultants. See Table 1.1.1, below.

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Micon assisted URSA in the preparation of an internal study in August, 2005 which showed that an on-site concentrator provided the most attractive processing option for the project and provided the management of URSA with the information needed to support the preparation of a full feasibility study.

Micon has been retained by URSA to prepare this feasibility study based on the establishment, at the Shakespeare property, of an open pit mining operation with on-site production of a saleable sulphide concentrate.

Micon and URSA conducted this study in conjunction with a number of other consultants responsible for specific technical areas, as shown in Table.1.4.1.

Table.1.4.1 Feasibility Study Team

Mineral resource estimate, mine plan, mine equipment and mine facilities.

Micon International Limited (Micon) P&E Mining Consultants Inc. (P&E)

Open Pit geotechnical studies Golder Associates Ltd. (Golder) Metallurgical testing SGS Lakefield Research (SGS)

Micon Process engineering Micon

Met-Chem Canada Inc. (Met-Chem) Mine rock and tailings disposal Golder Environmental baseline and geotechnical studies

Knight Piésold Group (Knight Piésold)

Environmental management plan and permitting

Golder

Infrastructure and plant design, capital expenditures and operating costs

Met-Chem Micon

Economic evaluation Micon

P&E, Golder, SGS and Knight Piésold were retained directly by URSA for the purposes of this feasibility study. Met-Chem was retained by Micon.

The principal supporting documents for this study are provided as Appendices 1 through 13 to this report.

Site visits have been carried out as follows:

March 19-21, 2003: B. Terrence Hennessey, P.Geo., of Micon in the company of Harold Tracanelli of URSA.

June 7, 2004: Eugene Puritch, P.Eng., of P&E, hosted by H. Tracanelli and Malcolm Buck.

March 21, 2005: Richard Gowans, P.Eng., of Micon, Francois Biron, Eng., Project Manager, Met-Chem, Steve Aiken, P.Eng., Manager Environmental Services, Knight Piésold, Matt Parfitt, P.Eng., Project Manager, Knight Piésold, Dr. Alan Cameron, P.Eng., Senior Mining Consultant and Principal, Golder and Irwin Wislesky, P.Eng.,

5

Senior Geotechnical Engineer and Associate of Golder, in the company of Dr. Richard Sutcliffe, P.Geo. President and CEO of URSA, J.P. Chauvin, P.Eng., President & CEO, Patricia Mining Corp., M. Buck and Gus Garisto, URSA Investor Relations.

September 13, 2005: Ian Ward, P.Eng., President of Micon and Study Manager, Lionel Poulin, Ing., and Daniel Houde of Met-Chem, Frank Palkovits and Irwin Wislesky of Golder and Steve Aiken of Knight Piésold, accompanied by Dr. Sutcliffe and J.P. Chauvin.

Several Golder and Knight Piésold field personnel worked intermittently on the project site during 2004 and 2005.

1.5 RESPONSIBILITY

The mineral resource estimate contained herein was prepared under the supervision and direction of B. Terrence Hennessey, P.Geo., with the assistance of Eugene Puritch, P.Eng., of P&E who operated the Gemcom software. The open pit designs and production schedules, as well as the capital and operating cost estimates for the mining aspects of the project, were prepared by Eugene Puritch. Met-Chem prepared the process plant and infrastructure designs and the associated capital cost estimates under the direction of Lionel Poulin, P.Eng. Golder was responsible for the design and costing of the co-disposal area for mill tailings and mine rock, and for geotechnical work in relation to the open pit. Knight Piésold, under the direction of Steve Aiken, P.Eng., conducted the environmental baseline studies and geotechnical work not relating to the open pit. Operating cost estimates (other than for mining) were prepared by Micon, as were the metal price, revenue and cash flow projections.

1.6 BASIS OF THE FEASIBILITY STUDY

1.6.1 Technical and Economic Parameters

Ore will be mined at 4,500 t/d of ore from an open pit. The mining operation will use an excavator/loader truck configuration, hauling ore directly to the processing plant.

All concentrate produced by the onsite processing plant will be shipped to a custom smelter/refiner in Canada or overseas for further processing.

A 1.5% royalty paid to Faconbridge, based on gross revenue, is provided for in the financial evaluation.

Concentrate sales terms used are based on indicative quotes.

All capital, operating costs were estimated and economic evaluations were conducted using Canadian dollars of mid-2005 value.

Economic analysis was carried out by means of discounted cash flow analysis expressed in constant dollar terms.

Table 1.6.1 provides the principal criteria for the Shakespeare project.

6

Table 1.6.1 Principal Criteria for the Shakespeare Project

Item Unit Quantity Life of Mine Production Ore Tonnes 11,226,250 Waste Tonnes 57,812,620 Total Tonnes 69,038,870 Strip ratio Waste/Ore (t/t) 5.15 Average nickel grade % 0.330 Average copper grade % 0.350 Average cobalt grade % 0.023 Average platinum grade Grams per tonne 0.332 Average palladium grade Grams per tonne 0.366 Average gold grade Grams per tonne 0.186 Contained nickel Thousand pounds 81,595 Contained copper Thousand pounds 86,662 Contained cobalt Thousand pounds 5,618 Contained platinum Thousand troy ounces 119.80 Contained palladium Thousand troy ounces 132.23 Contained gold Thousand troy ounces 67.27 Plant Recovery Nickel recovery % 76.0 Copper recovery % 95.5 Cobalt recovery % 71.0 Platinum recovery % 75.0 Palladium recovery % 42.0 Gold recovery % 38.0 Annual Production (payable metal) Average annual production, ore milled Tonnes 1,642,500 Average annual production, nickel in concentrate Thousand pounds 8,150 Average annual production, copper in concentrate Thousand pounds 10,523 Average annual production, cobalt in concentrate Thousand pounds 285 Average annual production, platinum in concentrate Thousand troy ounces 10.91 Average annual production, palladium in concentrate Thousand troy ounces 6.74 Average annual production, gold in concentrate Thousand troy ounces 3.10 Costs per lb nickel NSR value of nickel C$ per pound nickel 6.67 Smelter costs C$ per pound nickel (2.55) By-product credits C$ per pound nickel 4.69 Royalty payable to Falconbridge C$ per pound nickel (0.13) NSR (incl. by-product credits) C$ per pound nickel 8.68 Operating costs C$ per pound nickel (4.24) Operating Profit before taxes C$ per pound nickel 4.44 Taxes C$ per pound nickel (0.89) Profit after tax C$ per pound nickel 3.55 Capital costs Mining C$ 000 29,700 Milling C$ 000 48,978 Infrastructure C$ 000 18,997 Contingency 8,190 Sub-total Direct Capital Costs 105,865 Indirect (EPCM, etc) 12,607 Total initial capital C$ 000 118,473 Working capital (max) C$ 000 20,400 Ongoing incl Reclamation C$ 000 3,553 Residual value of plant C$ 000 (3,836) Total Project capital C$ 000 138,590

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1.7 DISCLAIMER

Micon has reviewed and analyzed data provided by URSA, its consultants and previous operators of the property, and has drawn its own conclusions therefrom, augmented by its direct field examination. Micon has not carried out any independent exploration work, drilled any holes or carried out any sampling and assaying. However, the nickel-copper-precious metal-bearing sulphide mineralization is visible in the local rocks, both in core and outcrop, and has been substantiated by the previous exploration history of Falconbridge on the property. Micon has reviewed the exploration data and undertaken the estimation of a mineral resource for both the East and West Zones in conjunction with P&E.

While exercising all reasonable diligence in checking, confirming and testing it, Micon has relied upon the data presented by URSA and the previous operators in formulating its opinion.

The various agreements under which URSA holds title to the mineral lands for this project have not been investigated or confirmed by Micon and Micon offers no opinion as to the validity of the mineral title claimed.

All currency amounts are stated in Canadian or US dollars, as specified, with costs typically expressed in Canadian dollars and commodity prices in US dollars. Quantities are generally stated in SI units, the Canadian and international practice, including metric tonnes (t) and kilograms (kg) for mass, kilometres (km) or metres (m) for distance, hectares (ha) for area, grams (g) and grams per metric tonne (g/t) for precious metal grades (e.g., g/t Au). See Table 1.8.1 for a list of abbreviations.

PGM grades may be expressed in parts per billion (ppb) or grams per metric tonne (g/t) and quantities may also be reported in troy ounces (ounces, oz), a common practice in the mining industry. Early exploration work on the property was conducted using the imperial system of measurement and the exploration grid and certain early exploration results may be expressed in feet (ft) although these were converted to metric measurements for resource estimation.

Micon is pleased to acknowledge the helpful cooperation of URSA’s management and field staff, all of whom made any and all data requested available and responded openly and helpfully to all questions, queries and requests for material.

Table 1.8.1 shows a list of the abbreviations used in this study.

Table 1.8.1 List of Abbreviations

Acid base accounting ABAAcid rock drainage ARDCentimetre(s) cmAmmonium Nitrate – Fuel Oil ANFO Centimetres per second cm/sCo-disposal area CDACubic metre(s) m3

Cubic metres per day m3/dCubic metres per second m3/sCubic metres per year m3/yDay(s) dDays per week d/w

8

Degree(s) o

Degrees Celsius oCDollar(s), Canadian and US $, Cdn$ and US$ Engineering, procurement and construction management EPCM Environmental design flood EDFEnvironmental Protection Agency (US Federal) EPAFoot(feet) ftHectare(s) haHertz HzHorse power HPHour(s) hHydraulic conductivity KInductively coupled plasma-optical emission spectrometry/mass spectrometry ICP-OES/MS Inch(es) inLitre(s) lKilogram(s) kgKilograms per day kg/dKilometre(s) kmKilovolt(s) kVKilowatt hours per tonne kWh/t KiloNewton per cubic metre kN/m3

Megapascal(s) MPaMegavolt ampere(s) MVA Metal Mining Effluent Regulations (Federal) MMER Metre(s) mMicron µmMillion years MaMilligram(s) mgMillimetre(s) mmMinistry of Natural Resources (Ontario) MNRMinistry of Northern Development and Mines (Ontario) MNDM Minute(s) minMunicipal Industrial Strategy for Abatement (Ontario) MISA Net acid generating NAGNet present value NPVNet smelter return NSRNot available/applicable n.a.Peak ground acceleration (acceleration due to gravity) gPercent(age) %Platinum group metals PGMPotentially Acid Generating PAGProbable maximum precipitation PMPProvincial Sediment Quality Guidelines (Ontario) PSQG Provincial Water Quality Objectives (Ontario) PWQO Rock Mass Quality RMQRock Quality Designation RQDSecond sSpecific gravity SGUniaxial compressive strength UCSUS gallons per minute USgpm Volt(s) VWeight percent wt%x-ray diffraction XRDYears(s) yGolder Associates, Golder Paste Technology Ltd. Golder Horizon Archeology .. Horizon Archeology Knight Piésold Ltd. Knight Piésold Micon International Limited Micon NAR Environmental NARNorth American Palladium Ltd. NAPTestmark Laboratories Ltd. Testmark SGS Lakefield Research Limited SGSURSA Major Minerals Incorporated URSA

9

2.0 PROPERTY DESCRIPTION AND LOCATION

The Shakespeare property is located in Shakespeare Township, immediately north and east of Agnew Lake, (See Figure 1.1.1) The property is approximately 70 km west-southwest of Sudbury, Ontario. The closest towns are Webbwood, which is 9 km southwest of the property, and Espanola, which is 11 km southeast. The property is situated on N.T.S. 41I/5 near Latitude 46°21'00"N and Longitude 81°49'47"W.

In 2005, URSA had total land holdings in the Sudbury area of 771 staked claim units and 28 leased and patented claims totalling 799 claim units (12,784 ha), located in Shakespeare, Baldwin, Porter and Hyman Townships. Of this total, the 28 leased and patented claims in Shakespeare Township plus 1 staked claim in Baldwin Township, 1 staked claim in Dunlop Township and 8 staked claims in Shakespeare Township totalling 100 claim units (1,600 ha or 4,000 acres) formed part of a joint venture agreement with Falconbridge Limited, (Falconbridge). See Table 2.0.1.

This report and feasibility study concern only the Shakespeare deposit on the original Falconbridge option claims and the surrounding ‘area of interest’. In the event that URSA acquires any contiguous property within the ‘area of interest’, Falconbridge may, within 30 days of acquisition by URSA, elect to make the contiguous property part of the joint venture.

The NAP option on the Shakespeare project is limited to these areas. The regional claims of URSA, along with the Falconbridge option claims and those newly acquired ones subject to the Falconbridge joint venture and included within the NAP option, are shown in Figure 2.1.

Table 2.0.1 List of Claims

Township Number of Claims

Number of Units

Type

Shakespeare 21 21 Patented Shakespeare 7 7 Leases Shakespeare 8 52 Staked Baldwin 1 12 Staked Dunlop 1 8 Staked Total 38 100

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Figure 2.1 Shakespeare Project Claim Configuration

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2.1 ACCESS AND INFRASTRUCTURE

Access road to the site will be from the northeast via a secondary road branching north from the Trans Canada Highway # 17 approximately 7.5 km east of Nairn Center. An existing logging road connects to the west side of the secondary road, approximately 13 km from Highway 17 and allows access to the property. For much of its length, this existing logging road is considered to be suitable for site access without upgrading. However, the portion branching south to the Shakespeare Project site will require upgrading over 3.4 km to the property gate location.

Figure 2.2 shows the location of the Shakespeare Project and route of the project access road.

Power will be provided to the site by a new 7.5 km, 44 kV overhead line from the existing 115 kV grid line located northeast of Webbwood. Substations with 10 MVA capacity will be built at the connection with the grid (115 kV/44 kV) and at the end of the 44 kV line adjacent to the Shakespeare processing plant (44 kV/4,160 V).

2.2 PHYSIOGRAPHY AND CLIMATE

The topography on the property is rugged with abrupt ridges and valleys ranging from 272 m to 381 m and generally averaging about 300 m above sea level. The average topographic relief is about 90 m feet and bedrock outcrops are common. Much of the general area is covered by timber resources which consist of second growth birch, poplar, oak, maple, jackpine and spruce.

The principal drainage channel is the Spanish River. The Spanish River and its tributaries drain the major part of the property. The part of the river near the property has been dammed for hydroelectric power generation and has resulted in the creation of Agnew Lake. Numerous private cottages and several commercial tourist operators are located on Agnew Lake.

Climate is typical of continental conditions with moderately long, cold winters and shorter, warm summers. Winter temperatures may drop below minus 20oC for extended periods and, in summer, maximum daily temperatures may exceed 25oC for extended periods. From December through March, daily mean temperatures typically are below zero oC.

Precipitation is moderate. The wettest months are between May and October but rainfall is generally distributed evenly through the year. Estimated average annual precipitation is 899 mm with 657 mm falling as rain and the balance (242 mm water equivalent) as snow.

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Figure 2.2 Shakespeare Project Location and Access Plan

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3.0 GEOLOGY AND MINERAL RESOURCES

The following descriptions of the regional and local geological settings for the Shakespeare Project were given by Eric. A. Kallio, P.Geo., in his Technical Report entitled “Technical Report for the Shakespeare Property, Shakespeare Township, Ontario, NTS 41I/5, for URSA Major Minerals Incorporated, Volume 1 of 2”, dated November 28, 2002 and filed with SEDAR on December 2, 2002 (Appendix 1).

The Shakespeare copper-nickel deposit is hosted by sulphide mineralization associated with gabbroic rocks of the Nipissing Intrusive Suites. Rocks of the East Bull Lake Mafic Intrusive are also present at Agnew Lake. Mineralization typically occurs as fine disseminations of magmatic pyrrhotite, chalcopyrite and pentlandite with some of the sulphides exhibiting magmatic blebby textures.

3.1 REGIONAL GEOLOGY

“The Dunlop-Shakespeare-Baldwin-Porter Township area [is] located along the southern margin of the Superior Province of the Canadian Shield and has had a prolonged evolutionary history involving the interaction between three structural provinces including the Superior, Southern and Grenville.

“The bedrock underlying the area is dominated by rocks of Precambrian age, including Early Precambrian (Archean) felsic plutonic rocks of the Superior Province and by Middle Precambrian (Proterozoic) supracrustal rocks of the Huronian Supergroup of the Southern Province. These rocks have been cut by mafic intrusions of several ages including the East Bull Lake Suite, Nipissing Suite and possibly the Sudbury Igneous Complex.

“The rocks of the Southern Province unconformably overly the Archean basement rocks. The Southern Province forms a discontinuous belt extending 750 miles (1,200 km) west from Quebec to central Minnesota along the southern margin of the Superior Province. The western portion of the Southern Province comprises a passive margin supracrustal sequence of the Marquette Range Supergroup, whereas in central Ontario the Southern Province is defined by the distribution of the Huronian Supergroup succession which is part of a basin forming rift margin.

“The Huronian Supergroup consists of a thick sequence (12,000 m) of clastic metasedimentary rocks. The Huronian rocks include sandstone, conglomerate, siltstone and greywacke, which were derived from the Archean granitoid terrains to the north. Also, mafic to intermediate metavolcanics, including flows and pyroclastic rocks are intercalated with the metasedimentary units in the basal part of the Huronian Supergroup succession.

“The East Bull Lake Suite is part of a major magmatic episode that occurred at 2480 – 2470 Ma in Central Ontario contemporaneous with rifting of the Archean Superior Province Protocontinent and the formation of the Huronian Rift Zone, now represented by the Southern Province. The intrusions typically occur near the boundary between the Archean Superior Province and the Early Proterozoic Southern Province, and

14

generally appear to have been emplaced as large sills. Magmatism is also manifested in the form of mafic dykes, and as bimodal continental flood basalt sequences (Huronian Volcanics). The most prominent intrusions of the East Bull Lake suite surrounding the project include the: East Bull Lake, Agnew, and May Township Intrusions.

“The Nipissing Suite was emplaced at roughly 2.2 Ma and form a trend extending from Sault Ste. Marie through the Sudbury Region to the Cobalt and Gowganda Regions. The intrusions are located dominantly within the Huronian Supergroup, but are also localized along the Archean- Proterozoic unconformity. The intrusions consist dominantly of gabbros with lesser diabase and granophyre, which range in thickness from a few hundred meters to over a thousand meters and typically outcrop at the present erosional levels as open ring structures, ring dikes, cone sheets, dykes and undulatory sills. (Hriskevich, 1952, 1968). The Nipissing Intrusions have traditionally been described as undulatory sheets consisting of a series of basins and arches connected by limbs ( Hriskevitch, 1968). The basinal portions of the sills consist of quartz diabase overlain by Hypersthene gabbro, and are overlain by vari-textured gabbro with pegmatoidal patches. The arches consist of vari-textured gabbro overlain by quartz diorite, granodiorite, granophyre and aplitic granitoids.

“The west limit of the Sudbury Igneous Complex is centered close to Sudbury and was emplaced at approximately 1.85 Ma. The Sudbury Igneous Complex occurs along the contact between the Superior and the Southern Province and consists of a thick composite mafic- felsic intrusion forming an elliptical ring having a major east-northeast trending axis that is 60 kilometers in length and a minor axis of 27 kilometers.

“The present outcrop distribution of the Huronian Supergroup does not reflect the size and shape of the original depositional system, but has rather been determined by syn- and post-Huronian folding, faulting and erosion. The most prominent faulting is syn-depositional normal faulting along the east-northeast trending Murray Fault system which is considered to have controlled the accumulation and preservation of most of the Huronian Supergroup in Central Ontario.”

Uranium-lead (U-Pb) age determinations on zircon from the gabbroic rocks hosting the Shakespeare deposit confirm that the host rocks of the Shakespeare deposit belong to the Nipissing Suite.

3.2 LOCAL GEOLOGY

“The area surrounding the Shakespeare property is underlain predominantly by units of the Huronian-aged Mississagi quartzite and gabbroic intrusives, which trend approximately northnortheast and dip moderate to steeply north.

“The Mississagi quartzites dominate the north and south limit of the land package and are typically whitish, medium grained and uniform.

“The gabbroic intrusive rocks occur predominantly in the south to central portion of the Shakespeare option and cover an area up to 1,600 feet [488 m] wide and over one

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mile long in strike length. The ultimate east and west limits for the intrusive, cannot be confirmed at this time, as detailed ground surveys have not yet been completed, but indications from airborne geophysics indicate the unit may extend for several miles east of the option ground, and onto lands owned 100 % by URSA Major.

“The intrusive rock is mainly dark grey and fine grained and consists of dominantly gabbro, however, a range of lithologies from quartz diorite to pyroxentie can be found in various parts of the property. The north and south limits of the intrusion are bounded by the Mississagi quartzite. Thin units of quartzite also occur locally within the overall limits of the intrusion. The contacts between the gabbro and the quartzites is locally sheared and altered.

“The gabbroic intrusives have been interpreted by the Ontario Geological Survey (OGS) (Card, 1976) as Nipissing Diabase, but others suggest that some may be part of the Agnew Intrusion, (Thompson, 1986 and Vogel, 1996) or even the Sudbury Igneous Complex. The possibility exists that intrusions of all three ages occur on the property.

“One of the major structures underlying the area is [the] Porter Syncline. The main axis of the syncline is located north of the Shakespeare property and trends in a northeasterly direction. All rocks within the area including the mafic intrusions appear to have been folded into a series of tight to moderately open, upright, complex folds with axes trending roughly parallel to the above syncline. Recent mapping at the Shakespeare property, suggests that there may also be a major northeast trending anticline located on the Shakespeare joint venture lands, which trends parallel to the Porter syncline. The axis of the projected fold is just south of the Shakespeare deposit and the central part of the fold is defined by a prominent quartzite lens.

“There are three main faults crossing the area which appear to be splays of the Hunter Lake Fault. The strike of the faults is generally northeast – southwest and dip steep. Several more northerly trending cross faults have also been identified in recent mapping.”

3.3 DEPOSIT TYPES

The deposit type descriptions for the Shakespeare Project are set out Kallio (Appendix 1). The relevant portions of these descriptions are reproduced below.

“Within the area of interest, numerous occurrences of copper and nickel sulphides along with platinum group elements have been identified through past work.

“Copper, nickel sulphide mineralization containing platinum group elements, (PGE’s) and gold is typically associated with gabbroic rocks of the : East Bull Lake, Nipissing and Sudbury Complex Intrusive Suites. Mineralization also occurs to a lesser degree in Huronian metavolcanic and metasedimentary rocks. Mineralization typically occurs as fine disseminations of magmatic pyrrhotite, chalcopyrite and pentlandite with some of the sulphides exhibiting magmatic blebby textures.

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“Recent deposit models for the area include the Nipissing model, (Lightfoot, P. G. and Naldrett, A. J., 1996) and the Sudbury Offset model proposed by Lewis, C. L. (1949) and Sutcliffe, R. H, (2002).

“The Nipissing model is based on the recognition of close spatial relationships of nickel, copper and platinum group elements with the Nipissing Diabase, (Lightfoot, 1996). Significant observations regarding mineralization according to the Nipissing model are described in Lightfoot, (1996) and include the following:

“1) Magmatic nickel, copper and PGE mineralization is spatially associated with Nipissing intrusions [which] lie on a trend which extends from Whitefish Falls to River Valley.

“2) The sulphides occur as fine disseminations of magmatic pyrrhotite (50 - 75%) with lesser chalcopyrite and pentlandite.

“3) The disseminated sulphides tend to be localized in the interior of the sills, (100 - 300 m above the base), within coarse grained gabbro-norites and hypersthene rich gabbros.

“4) Sulphides occurring locally as basal concentration can carry 1 - 15 % copper, 2.5 - 6.3 ppm Pt, 17 - 53 [ppm] Pd and 1 - 6 ppm Au, (ie Wanapetei Intrusion).

“The Sudbury offset model considers that mineralization to be related to radial dykes extending outwards from the Sudbury Igneous Complex. Key characteristics for mineralization related to offset dykes are discussed in Dressler, (1991). Economic mineralization within offset dykes is typically spatially associated with inclusion rich quartz-diorite and local thinning. Sulphides often form blebs in the quartz Norite matrix. Past studies indicate a possible zoning in ore composition characterized by an increase in chalcopyrite content with increasing distance from discontinuities in the dyke, Cochrane, (1984). According to Lewis, (1949) some of the rocks surrounding the Shakespeare deposit contain similarities in texture, composition and mineral content to rocks of the Worthington and Copper Cliff offset dykes.”

The most recent work completed or commissioned by URSA, including radiometric age dating, appears to eliminate the possibility of a Sudbury offset model as the intrusive hosting the mineralization is some 400 million years too old and defines the Shakespeare deposit as being hosted in Nipissing-aged intrusions.

3.4 MINERALIZATION

The mineralization styles seen at the Shakespeare Project are described in Kallio (Appendix 1) as provided below.

“The most significant mineralization identified at the Shakespeare property to date is contained in the Shakespeare deposit, located in the central portion of the URSA Major - Falconbridge option ground.

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“The deposit is hosted within gabbroic rocks situated along the north contact between the mafic intrusive body which crosses the joint venture property and quartzites of the Mississagi Formation.

“The bulk of the mineralization occurs as a broad northeast trending zone of magmatic sulphides located between 25 and 150 feet [8 – 46 m] south of the intrusive and quartzite contact. The style of magmatic sulphides in the mineralized zone progresses downward from sporadically disseminated, to scattered multi-centimeter sized blebby composite pyrrhotite-chalcopyrite grains, to more evenly distributed, heavily disseminated to locally net-textured magmatic sulphides. Chalcopyrite-rich cusps in the blebby mineralization indicate a north-facing up direction, consistent with the geologic interpretation. Based on observations, to date the style of mineralization appears to conform best with the Nipissing model as described by Lightfoot, (1993).

“The surface expression of the deposit is marked by a distinct gossan which has formed as a result of oxidation of disseminated sulphides, largely pyrrhotite.

“Past work has indicated that the overall dip for the deposit to be from vertical to very steep northwards. More recent observations, however, including those from mapping and new diamond drilling suggest that in section the zone may be arcuate in shape and the dip in some areas, shallower than original projections.

“The total strike length of mineralization defined from past work is approximately 1,500 feet and situated between 500 W and 1500 W of the property grid.”

Exploration work completed by URSA has resulted in the discovery of a new zone of mineralization very similar to and along strike from, the original Shakespeare deposit. This new zone is now referred to as the East Shakespeare Deposit or East Zone and the original Shakespeare deposit discovered by Falconbridge is called the West Shakespeare Deposit or West Zone.

The main mineralized zone is characterized by an upper part of “blebby” style mineralization and a lower part of “disseminated” style mineralization.

3.5 EXPLORATION

The results of previous surface exploration at the Shakespeare Project are described in Kallio (2002a). A summary of the previous exploration drilling on the East Zone by URSA, taken from this report, is provided below.

The winter 2002/2003 exploration program consisted of diamond drilling on the discovery at the East Zone and early stage geophysical programs on adjacent claims controlled 100% by URSA. The geophysical program was conducted on the along-strike continuation of the lithologies hosting the Shakespeare deposit but is not germane to the present study.

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3.5.1 Exploration Following the 2003 Resource Estimate

During the summer 2003 field season, after completion of the East Zone mineral resource estimate, URSA commenced further exploration activity at the Shakespeare project consisting of a trenching and mapping program and further drilling. The drilling activities in this program will be discussed below. Channel sampling was conducted in areas of exposed outcrop or where stripping of thin soil and moss cover could expose subcrop over the deposit. The sampling protocols were conducted in such a way that the samples were continuous and could be entered into the database as pseudo drill holes.

The recent trenching concentrated on increasing knowledge over the East Zone and searching along strike to the east. The work has created considerable new surface exposure on the mineralized structure and has extended the area of interest to the southeast of the current East Zone mineral resource. Figure 3.1 shows the current geological interpretation of the Shakespeare property, with a provisional open pit outline as at July, 2005. The mineralization is shown to extend into the folded area seen to the right of the larger of the two proposed open pit lobes. Diamond drilling appears to indicate that the main lens of potentially economic mineralization in the East Zone plunges to depth.

3.5.2 Exploration Following the 2004 Resource Estimate

After the completion of the second mineral resource estimate, released in the fall of 2004, URSA continued with exploration and definition drilling activities through late 2004 and 2005. These programs consisted primarily of diamond drilling for exploration, mineral resource estimation and geotechnical purposes, much of it to allow for completion of a feasibility study. This drilling is discussed in Section 3.6 below. Little other exploration work was completed at the Shakespeare deposit.

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Figure 3.1 Shakespeare Property Geological Map

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3.6 DRILLING

3.6.1 Falconbridge Drilling

Kallio (Appendix 1) described the pre-2002 Falconbridge drilling programs at Shakespeare as follows:

“Five separate diamond drill programs have been completed on the property including 1942, 1948, 1951, 1985 and 1986. These programs amount to 47 holes totalling 21,833 feet [6,655 m] and are summarized in Table 6 [see Table 3.6.1].

Table 3.6.1 Pre- 2002 Diamond Drill Holes

Program Number of Holes Footage 1942 12 2,687 1948 3 4,462 1951 12 6,000 1985 16 3,380 1986 4 5,304 TOTAL 47 21,833

“Drill holes completed in 1942 included twelve short holes, totalling 2,687 feet [819 m]on the Shakespeare deposit. These holes ranged in length from 40 to 445 feet [12 – 136 m]. They were drilled for assessment work and to define the geology and grade of the deposit.

“Drill holes completed in 1948 included three holes, totalling 4,462 feet [1,360 m]. These holes, number 13, 14 and 15 were drilled to a depth of 1,050, 1,862 and 1,550 feet [320, 568 and 472 m] respectively. The drilling was used for assessment work requirements and to investigate the possibilities of enrichment with depth. The holes were drilled on section 1300 W, approximately midway along the strike of the ore zone.

“Drilling in 1951 included twelve short holes, numbered 16-27, totalling 6,000 feet [1,892 m]. The length of the holes range from 300 to 630 feet [91-192 m]and designed for the purpose of checking the width and grade of mineralization to a 500 foot [152 m] depth. The holes were positioned at roughly 200 foot [61-m] centres between 200 W and 2000 W.

“In 1985, sixteen holes totalling 3,380 feet [1,030 m] were drilled. These holes were drilled to test the near surface resource and to evaluate the precious metal (Au, Pt and Pd) potential of the zone. Holes from the program were completed between 200 W and 20000 W and designed to provide coverage on 100 foot [30.5-m] centres across the Shakespeare deposit at depths less than 100 feet [30.5 m] from surface.

“In 1986, a further 4 holes totalling 5,304 feet [1,617 m] were drilled to test the deposit at depth and along strike to the southwest. Two of the holes were drilled on 2900W,

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one on 2300W and the other on 1800 W. All of the holes were drilled to total depths of 1,200 to 1,500 feet [355 to 457 m] and designed to test the deposit at a depth below surface of approximately 500 feet [152 m].

“It is thought that most of the hole collars from previous drilling were initially spotted using an exploration grid for reference. The casings for almost all of these holes along with the baseline for the Falconbridge grid were surveyed in 1985 by D. White an Ontario Land Surveyor. Data from the survey is plotted on a map and listed in a chart compiled at the time.

“Most of the downhole surveying from the historic holes has been conducted either by acid or Tropari testing. The frequency of tests seems rather low in many cases, with only one test at the top and bottom of each hole being recorded. The magnitude of deviations for azimuth and dip, however is also quite low, suggesting more tests might not significantly improve the accuracy of hole locations.

“The name of the drill companies and the core size drilled are unknown for sure for much of the early drilling. Notes on some of the 1985 and 86 Falconbridge drill logs, however, indicate that this drilling was conducted by Triangle Drilling of Lively using BQ sized equipment.

“Very little of the historic drill core from past programs at the property has been retained. A search for historic drill core which might be available was carried out by G. T. Shore in June of 2002 at the Falconbridge core repository.

“Most of the drill core logging from historic programs was either done at the project site or at the Lockerby Core Shack owned by Falconbridge. Initial compilation of information from historic logging and surveying was to hard copy drill logs, which are all quite similar in format. The logs are subdivided into separate sections for header and descriptive information. The header area contains fields for: hole name, date drilled, identity of logger, property name, and grid location, azimuth and starting dip. The descriptive section is subdivided into two: one area containing intervals for various lithologic units and the other the name and description of the interval. There does not appear to be a standard coding system for lithologies, but the terminology is relatively consistent with time.

“Work to compile and verify the historic logging information to a Gemcom Computer database, has been ongoing by personnel at URSA.

“Results of the historic diamond drill data indicate a continuous zone of sulphide and precious metal mineralization extending over a total strike length of 1,800 feet [549 m] between 200 W and 2000 W to a depth of approximately 250 feet [76 m] with very few holes testing below the 250 foot [76-m] level. The center of the zone is usually close to the baseline or slightly north of this and the dip variable, from shallow to steep north. Possible explanations for the variability in dip are faulting or that the overall shape of the zone is arcuate with a slight curve to the north. If the zone is in fact accurate in shape, then it is possible that the variations in dip observed on sections are simply a function of where the various drill holes intersected this.

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“The width of most intersections ranges between 75 and 125 feet [23 and 38 m], with the longest intersection of 261 feet [79.6 m] being recorded in hole # 1 and the shortest of 3.0 feet [0.9 m] being in hole 85-4. The range of grades intersected for nickel is : 0.09 % to 0.49 %, copper : 0.09 to 0.61 %, gold : 0.11 to 0.30 g/t, platinum : 0.15 to 0.57 g/t and palladium : 0.17 to 0.57 g/t.

3.6.2 URSA Drilling Programs

The remaining drilling at the Shakespeare project has been completed by URSA since 2002 and is summarized in Table 3.6.2 below.

Table 3.6.2 URSA Diamond Drill Holes since 2002

Program Dates Number of Holes Metreage

2002 Summer/Fall 9 4,904.90 Winter 2003 Nov-02 to Mar-03 18 3,263.00 2003/2004 May-03 to Feb-04 28 5,950.33

Summer 2004 Feb-04 to Sep-04 10 3,648.90 2005 Mar-05 to Jun-05 19 2,443.10 Total 84 20,210.23

Drill Programs Summer 2002

URSA Major carried out drilling in 2002 at the Shakespeare project in areas to the east and west of the main zone of mineralization identified through past work. Total drilling completed to the end of October, 2002 was 1,495 metres in 9 holes. Two of the holes which were drilled, UR-03-01 and UR-03-02, were located approximately 1,200 feet (366 metres) south-west of the Shakespeare deposit to test a new area of the property and seven were drilled east of 5+00 W (the limit of the 1985 Falconbridge resource), to test for an easterly extension of the Shakespeare deposit. Results of the diamond drilling carried out east of the Shakespeare deposit were very good. Several wide intervals of mineralization were intersected, at distances of up to 1,400 feet (427 metres) beyond the east limit of the resource defined by Falconbridge in 1985. Typical widths of Nickel and copper mineralization was intersected with better grades and widths than in drilling at the original Shakespeare deposit. Grades for gold and platinum group elements were also found to be marginally higher. Results from both diamond drill-holes in the south-west were disappointing, however, intersecting mainly barren gabbro and/or quartzite with no significant assays.

Drill Program Winter 2002/2003

Between November, 2002 and March, 2003, URSA completed 3,263 m of drilling in 18 holes. This program concentrated on an area to the east of the Shakespeare deposit outlined by Falconbridge.

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Drill Program Summer 2003 to Winter 2004

During the summer, fall and early winter of 2003/2004 a program of additional diamond drilling was performed by URSA on the Shakespeare Project. This work was carried out from May 21, 2003 through to February 12, 2004 with the primary goal of further expanding and defining the limits of the Shakespeare East Zone. During this period 4,006 m in 15 diamond drill holes was drilled on the East Zone, exploring along strike from Line 9+00E through to Line 17+00E. Additionally 2,057 m in 13 diamond drill holes were completed on the original Falconbridge Shakespeare deposit (West Zone) and were essentially confined to the areas between Line 6+00W through to Line 20+00W.

Drill Programs Late 2004 and 2005

Between February 16, 2004, and September 15, 2004 URSA completed a further 10 NQ-sized, inclined drill holes for a total of 3,649 m. The bulk of this drilling was performed on East Zone mineralization and potential strike extensions around the fold nose seen in Figure 9.1 (to the right of the known mineralization). Seven of the 10 drill holes (UR-03-59 through to UR-03-65) were targeted on known mineralization in the deposit for a total of 3,332 m. While several of these holes were drilled prior to the release of the 2004 mineral resource estimate their assay results were not available for use in the estimate as the database had to be “frozen” prior to commencement of work.

Drill Program Late 2004 to 2005

During the period from March 12, 2005 to June 24th, 2005, URSA completed an additional 18 NQ-sized drill holes (15 inclined and 1 vertical) and an HQ-sized vertical drill hole at the Shakespeare Project. The 19 drill holes represented a total of 2,443 m of drilling. Fifteen of the holes were drilled principally to provide locally needed information on mineralization contacts in the deposit and to fill in sampling and grade information for the geological model and mineral resource estimate. The other four holes were drilled for geotechnical purposes related to mine design for the feasibility study. These were not assayed but were logged for geological and geotechnical information.

Discussion

URSA has found the results of the 2003, 2004 and 2005 diamond drilling programs to be positive and encouraging. Most infill holes have intersected a wide interval of sulphide mineralization and collectively defined a single layer of nickel-copper-gold-PGE-bearing sulphide mineralization divided into 2 adjacent, near-surface zones known as East and West. Typical widths are approximately 40 m and grades for nickel, copper and precious metals in the East Zone are generally higher than the average grade of the original Shakespeare West deposit.

Infill drilling of the West zone has provided confirmation for most of the Falconbridge drilling. One early Falconbridge hole, the coordinates of which are suspect, has not been used in the mineral resource estimate.

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3.6.3 Other Drilling

In addition to the exploration drilling described above, 17 vertical diamond drill holes ranging from 10 m to approximately 50 m length, were drilled for hydrogeological study purposes and for geotechnical purposes and have been identified as “water monitoring wells”. They are not included in the drill hole database for the estimation of mineral resources and reserves. These holes are located in the peripheral areas surrounding the East and West Zones and the anticipated optimized pit shell areas.

3.6.4 Summary

Most drilling has intersected a wide interval of sulphide mineralization and collectively defined a single layer of nickel-copper-gold-PGM-bearing sulphide mineralization divided into two adjacent near surface zones known as East and West. Typical widths are approximately 40 m and grades for nickel, copper and precious metals in the East Zone are generally higher than the average grade of the original Shakespeare West deposit.

The mineralization is contained within an approximately 80-m thick differentiated sill, herein defined as the Shakespeare sill, that has intruded between quartzite in the northwest (hanging wall) and Nipissing gabbro in the southeast (footwall). From northwest to southeast the Shakespeare sill grades downward from biotite-quartz diorite, to biotite-quartz gabbro, to mineralized gabbro and pyroxenite, downward into unmineralized gabbro. Magmatic sulphides in the mineralized zone progress downward from sporadically disseminated, to scattered multi-centimetre sized composite blebs and grains of pyrrhotite-chalcopyrite mineralization, to more evenly distributed, heavily disseminated to locally net-textured magmatic sulphides. Strong mineralization starts at or near the contact between the quartz gabbro and the gabbro-pyroxenite-mela gabbro and persists through most of the pyroxenite-mela gabbro. Locally, the contact between the mineralized mela gabbro and the underlying footwall Nipissing type gabbroic rocks, has been found to be markedly sharp and chilled against the lower rocks, or the visible contact relationship between these two rock types can be described in terms of being a lower admixed-cooked zone. Intrusive contacts and zone of sulphide mineralization dip north at 50º to 60º.

Recent drilling around the discovery holes for the East Zone, near Line 3+00E and over to Line 5+00E, has identified a significant thickening of the zone and a development of a “belly” which made up for any losses of mineralization elsewhere in definition drilling. Infill drilling of the West Zone has helped firm up geological modelling interpretations and provided confirmation for most of the Falconbridge drilling.

The zone of nickel, copper, cobalt and PGM bearing sulphide mineralization appears to plunge gently to the northeast where, based on the early 2004 drilling, it is approximately 150 m below surface on Line 17+00E. Later, more widely spaced, 2004 drilling has extended the deposit out to Line 29+00E where it continues to plunge. While likely not mineable from an open pit, a reasonably high grade core of mineralization has been discovered here. It remains open along strike at depth, and the potential exists for an economically viable zone of underground mineable mineralization here, should an open pit be developed on the main East and West Zones from which a ramp would be required in order to access to lower levels.

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3.6.5 Database Checks

Surface drill hole collar locations were picked up by a professional surveying crew under the direction of an Ontario Land Surveyor.

Sperry Sun was contracted to conduct gyroscopic down-hole surveys on most of the 2002/2003 holes as a check of the Tropari/EZ shot/Flexit readings. The Sperry Sun instrument is unaffected by magnetism however, azimuth readings by the others systems can be affected by magnetic rocks. URSA felt it necessary to confirm that the mafic intrusives and pyrrhotite of the host rocks were not significantly affecting readings. No significant problems were encountered.

URSA staff thoroughly checked the database for compliance with required data entry format prior to submission to P&E for use in modelling. Assay entry was spot-checked from original assay certificates, as were from/to intervals and lithologic codes from the original drill logs.

Gemcom has utilities for checking database integrity such as missing entries, crossed from/to intervals and improper coding of lithologies or other descriptive elements. These utilities were used to ensure database integrity.

3.6.6 Use of Falconbridge Data

In order to confirm the validity of the previous Falconbridge drilling, URSA decided to drill on four sections, spaced equally along the known length of the West Zone. It was intended that these holes, interspersed within the Falconbridge drilling, would provide confirmatory evidence of the Falconbridge assay results and allow for the lithologic codes used by them to be converted to the URSA system.

Additionally, once modelling was complete, all assays from each domain were extracted and then separated into Falconbridge and URSA populations. The population statistics for these data were calculated and compared.

3.7 MINERAL RESOURCES

3.7.1 Mineral Resource Estimation Methodology

3.7.1.1 Database

The drill hole data used for the resource estimation were provided by URSA in the form of Microsoft Excel files. A Gemcom database was constructed containing 47 Falconbridge holes, 28 trench segments (as pseudo drill holes) and 84 URSA holes (UR series), of which only 6 were not used for modeling purposes. Holes UR-03-01, UR-03-02, UR-03-47, UR-03-56, UR-03-57 and UR-03-58 were exploration holes not in the vicinity of the Shakespeare Deposit. The database was validated and corrected in Gemcom until it was found to be error free. The data in the assay table included assays for nickel, copper, cobalt, gold, platinum and palladium.

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Topographic surface data were obtained from an airborne LIDAR (Light Detection and Ranging) survey completed by Mosaic Mapping Systems Inc. The data were provided as an AutoCAD polyline file that was converted into a Gemcom 3D topographic surface for modelling. Drill hole collars were surveyed by Paul H. Torrance Surveying Ltd. Down hole surveys were conducted using an “EZ Shot” electronic instrument. Some down hole check surveys were conducted by Sperry Sun Drilling Services of Canada using a gyroscopic instrument. All coordinates were UTM, based on the NAD 83 datum. The decision was made to express all quantities in the resource estimate in metric units, although the original Falconbridge exploration grid, established in Imperial units, was retained,. Surface drilling equipment in Canada is now largely metric and some of the 2003 drill program utilized metric rods. The older Falconbridge drilling data was in imperial units and was converted to metric.

3.7.1.2 Domain Interpretation

A geological domain model was constructed to control grade interpolation. Domain boundaries were determined by lithology, domain mineralization style and grade boundary interpretation from visual inspection of drill hole sections, as well as modelled faults interpreted from stratigraphic offsets and logged fault gouge in holes. There were 38 drill sections at two different interval spacings. Drilling in the west area of the deposit was at nominal 30-m (100-ft) spaced intervals with twenty three sections from Lines 0+00 W to 22+00 W. The east area of the deposit was drilled at nominal 60-m (200-ft) spaced intervals with fifteen sections from 1+00 E to 29+00 E. A total of 5,100 ft (approximately 1,555 m) of strike length was drilled off during the program. However, a gap in the mineralization between the East and West Zones (sections 0+00 to 4+00 W) has received relatively little drilling. It shows limited mineralization and this area was not modelled. This gap appears to be caused by the mineralization transitioning from the hanging wall to the footwall of a group of faults and at this time appears to represent a local absence of mineral resources. The geological domains were extrapolated 25 m in either direction from the end of drilling and terminated. This resulted in a two-part geological domain model approximately 1,435 m in total aggregate length.

The domains were physically created by computer screen digitizing on drill hole sections in Gemcom based on input from URSA with Micon and P&E. The outlines were influenced by lithology, structure and a Cdn$30 in situ metal value which approximately selects continuous mineralization boundaries.

The Canadian dollar in situ values for nickel and copper were derived from the 24-month trailing average of published spot commodity prices on the Kitco Base Metals website (www.kitcometals.com), for the period September 1, 2003 to August 31, 2005. The metals prices used were US$6.32 per pound (lb) for nickel and US$1.31/lb for copper. The remaining commodity prices were derived from the October 2004 pre-feasibility study and are as follows: US$15/lb for cobalt, US$400/oz for gold, US$800/oz for platinum and US$200/oz for palladium. The exchange rate used was Cdn$1.00 = US$0.80.

On each section, polyline interpretations were digitized in three dimensions, from drill hole to drill hole, but never extrapolated more than 25 m into untested territory. Polylines were not typically projected to surface, unless justified by trenching or surface mapping since a flat-

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dipping fault at shallow elevation has locally offset the zone. A statistical analysis of all assays from the full mineralized domain indicated separate populations of data. Accordingly, four separate domains, called “Disseminated” (West and East) and “Blebby” (West and East) were established to represent the two styles of sulphide mineralization and the two areas of strike.

An analysis performed for the entire mineralized width of the East Zone (both sulphide domains) in the first resource estimate yielded probability plots similar to that seen in Figure 3.1, where the cumulative distribution curve shows a distinct inflection at 0.5% nickel indicating that two lognormally distributed populations of data are present.

Separation of the data by blebby and disseminated domains resulted in probability plots such as those seen in Figures 3.2 and 3.3. The blebby domain appears to be close to a single, lognormally-distributed population of data. However, the disseminated domain still contains two populations.

It was considered possible that these two populations occupy the same location in space and cannot be separated by domaining. Therefore two separate sulphide domains were used for resource estimation because the blebby domain successfully separated one population of data. The polylines from each section and domain were wireframed into 3-dimensional solids using Gemcom. The resulting domain solids were used for grade interpolation purposes.

Further work in this area, including more detailed examination and description of sulphide textures, was recommended however, no criteria for sorting the two populations in the disseminated domain have yet been identified. Probability plots which include the new URSA and the old Falconbridge drilling show similar results to those seen previously and the two sulphide texture domains continue to be used.

Figure 3.2 Cumulative Distribution Curve, Nickel (Whole Database)

Shakespeare

Ni (%)

0.01 0.1 1

CumulativePercent

0.1

1

10

30

50

70

90

99

99.9

Nickel vs Cumulative Distribution

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Figure 3.3 Cumulative Distribution Curve, Nickel

Blebby Domain

Figure 3.4 Cumulative Distribution Curve, Nickel

Disseminated Domain

Shakespeare

Ni (%)

0.01 0.1 1

CumulativePercent

1

10

30

50

70

90

99

Nickel vs Percent Ni

Shakespeare

Ni (%)

0.01 0.1 1

CumulativePercent

1

10

30

50

70

90

99

99.9

Nickel vs Percent Ni

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Figure 3.5 shows an example drill section (1+00 E) from the first resource estimate with the blebby and disseminated domains interpreted on it. Figure 3.6 is a 3-dimensional isometric view of the blebby and disseminated domains showing the topographic surface and drill hole traces.

Figure 3.5 Example Section Showing/Domain Interpretation

(Section 1+00 E, Looking East)

BlebbyDomain

DisseminatedDomain

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Figure 3.6 3-Dimensional View of East Zone (Left Side) and West Zone (Right Side)

Geological Domains and Drill Hole Traces

Green = blebby domain, Red = disseminated domain, Brown = surface

3.7.1.3 Rock Type Determination

The rock types used for the resource model were coded from the mineralized domain solids as well as surface topography. The overburden/bedrock contact was used to limit upward extension of the domains. The list of rock codes used is set out in Table 3.7.1.

Table 3.7.1 Block Model Rock Codes

Rock Code Description

10 East Blebby Sulphide Domain 11 West Blebby Sulphide Domain 20 East Disseminated Sulphide Domain 21 West Disseminated Sulphide Domain

3.7.1.4 Composites

Length-weighted assay composites were generated for the portion of each of the 91 drill holes and 17 surface trenches that fell within the constraints of the blebby sulphide and disseminated sulphide domains. These composites were calculated for nickel, copper, cobalt, gold, platinum and palladium. The composites were compiled over 2.5-m lengths starting at the first point of intersection between drill hole and hanging wall of the 3-dimensional zonal constraints. The compositing process was halted upon exit from the footwall of the aforementioned constraint. Un-assayed intervals were treated as nulls and not utilized in the composite calculation. Any

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calculated composites less than 1.0 m in length, were discarded so as not to introduce any short sample bias in the interpolation process.

The composite data was then transferred to Gemcom extraction files as X, Y, Z, value files for the grade interpolation.

Grade Capping

Grade capping was investigated on the raw assay values in the database, prior to compositing to ensure that the possible influence of erratic high values did not bias the grade estimates.

Extraction files were created for domain-constrained data for all six metals of interest. From the extraction files, normal histograms, log histograms and probability plots were generated. In addition, sample means, standard deviations and coefficients of variation were calculated. These plots were used to identify the point at which lognormal populations broke down into scattered outlier values. The grade at these points was selected as the top cutting value.

The resulting capping (top-cut) values used are listed in Table 3.7.2. The population statistics and graphs for the data described above can be seen in Appendix 2a.

Table.3.7.2 Grade Capping Values

Element Unit East Blebby

Domain East

Disseminated Domain

West Blebby Domain

West DisseminatedDomain

Nickel % 1.10 1.05 No capping No capping Copper % No capping 1.40 No capping 0.92 Cobalt % 0.11 0.107 No capping 0.10 Gold g/t 0.420 0.861 0.355 No capping Platinum g/t 0.725 No capping No capping 0.758 Palladium g/t 0.943 1.000 No capping 0.799

The capping values chosen have resulted in less than 1% of the data being capped.

3.7.1.5 Variography

Semivariograms (hereafter referred to as variograms) were created and analyzed for the nickel, copper, cobalt, gold, platinum and palladium data from the constrained assay composite extraction files for both blebby and disseminated domains. In past resource estimates, cobalt variograms could not be modelled, and it was suspected that this was due to the relatively low grades and the very limited variability in the data set. However, the additional data provided by the 2005 drilling has allowed the modelling of cobalt variograms although they still show limited variability of data and relatively low grades.

The two blebby sulphide domains are of limited size and the relatively small amount of contained data did not yield any discernable pattern in interpretation of the resulting

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variograms during the first two resource estimation attempts. However, the 2004/2005 drilling has provided enough data to allow for the modelling of variograms in the East Blebby domain.

The East Disseminated domain gave reasonable variograms for both the down-hole and omnidirectional in-the-plane-of-dip interpretations. Coherent along-strike and down-dip variograms could not be modelled. Nugget values were modelled on the omnidirectional variograms and forced on the across dip ones. All four variogram types could be modelled for the West Disseminated and East Blebby domains. The resulting variograms are provided in Appendices 2l and 2 m.

The variograms were modelled with one or two spherical structures as required however, two structures were typically used for the omnidirectional variograms and one structure was normally used for the discreet variograms. The range of the first structure in the East Disseminated zone omnivariograms varied from 12 m to 20 m for all elements modelled. Second structure ranges of 99 m to 108 m were obtained for all elements in the dip plane. Across dip ranges varied from 13 m to 23 m.

The West Disseminated domain had somewhat shorter ranges of 44 m to 85 m along strike and 36 m to 106 m down dip. The ranges in the across-dip or down-hole variograms varied from 8 m to 29 m. Nugget values and nugget to sill ratios were generally very low.

The East Blebby domain also had somewhat shorter ranges of 51 m to 83 m along strike and 23 m to 48 m down dip. The ranges in the across-dip or down-hole variograms varied from 11 m to 28 m.

Nugget values and nugget to sill ratios were generally very low for all variograms.

The ranges and parameters used for grade interpolation at Shakespeare are set out in Tables 3.7.3, 3.7.4 and 3.7.5, below. Disseminated Zone and East Blebby Zone mineralization was interpolated using Ordinary Kriging with kriging parameters as set out in the tables. The West Blebby Zone mineralization was interpolated in two passes using Inverse Distance Squared (ID2) grade interpolation because no variograms could be modelled for it. The first interpolation pass (Indicated 1) was made using a search ellipse distance of two-thirds of the full range of variogram structures modelled. Any blocks not filled were interpolated in a second pass (Indicated 2) which used the full range. In all zones inferred mineralization was interpolated in a final pass using ranges double that used in the previous pass, so as to fill the model. The ranges for the West Blebby Zone ID2 grade interpolation were determined after a review of the typical ranges seen in the variograms from the other zones.

The search ellipsoid ranges used for grade interpolation, as established by the variography, were sufficient to code a large majority of the constrained mineralization within the proposed open pit as indicated mineral resources. The remainder of the block model was filled using the process described above and was classified as inferred mineral resources. The majority of this mineralization occurs as deep mineralization down plunge to the east and would most likely be mined by underground methods.

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Table 3.7.3 Block Model Grade Interpolation Parameters

Kriging (Indicated) and ID2 (Indicated First Pass)

Profile Dip Strike Dip Model Model Max No. Min No.

Direction Along Strike Down Dip Across Dip Nugget TypeRange 1 Gamma Range 2 Gamma 2/3 Range 1 Gamma Range 2 Gamma 2/3 Range Gamma 2/3 Gamma per of

( º ) ( º ) ( º ) (m) (H) (m) (H) Range (m) (H) (m) (H) Range (m) (H) Range (H) Hole SamplesWest Bl-Ni 327 57 -60 40 N/A - N/A 40 N/A - N/A 10 N/A N/A N/A 2 3West Bl-Cu 327 57 -60 40 N/A - N/A 40 N/A - N/A 10 N/A N/A N/A 2 3

ID2 West Bl-Co 327 57 -60 40 N/A - N/A 40 N/A - N/A 10 N/A N/A N/A 2 3First West Bl-Au 327 57 -60 40 N/A - N/A 40 N/A - N/A 10 N/A N/A N/A 2 3Pass West Bl-Pt 327 57 -60 40 N/A - N/A 40 N/A - N/A 10 N/A N/A N/A 2 3

West Bl-Pd 327 57 -60 40 N/A - N/A 40 N/A - N/A 10 N/A N/A N/A 2 3

East Dis-Ni 327 57 -60 17 0.25 104 0.93 42 see along Strike parameters 19 0.43 8 0.08 Spherical 2 3East Dis-Cu 327 57 -60 20 0.51 108 1.03 28 see along Strike parameters 16 0.35 9 0.06 Spherical 2 3

Kriging East Dis-Co 327 57 -60 18 0.23 99 0.45 24 see along Strike parameters 23 0.16 10 0.04 Spherical 2 3East Dis-Au 327 57 -60 14 0.31 108 0.93 38 see along Strike parameters 14 0.31 6 0.09 Spherical 2 3East Dis-Pt 327 57 -60 14 0.26 105 0.89 26 see along Strike parameters 13 0.29 5 0.07 Spherical 2 3East Dis-Pd 327 57 -60 12 0.26 106 0.96 41 see along Strike parameters 13 0.26 6 0.05 Spherical 2 3

West Dis-Ni 327 57 -60 85 0.70 - - 31 93 0.77 - - 39 25 0.60 7 0.23 Spherical 2 3West Dis-Cu 327 57 -60 67 0.48 - - 23 106 0.74 - - 44 29 0.55 11 0.17 Spherical 2 3

Kriging West Dis-Co 327 57 -60 44 0.20 - - 16 36 0.12 90 0.22 25 8 0.14 2 0.06 Spherical 2 3West Dis-Au 327 57 -60 70 0.46 - - 24 94 0.74 - - 39 16 0.56 12 0.16 Spherical 2 3West Dis-Pt 327 57 -60 71 0.45 - - 29 78 0.57 - - 33 16 0.19 1 0.12 Spherical 2 3West Dis-Pd 327 57 -60 68 0.38 - - 17 82 0.65 - - 32 16 0.25 0 0.18 Spherical 2 3

East Bl-Ni 327 57 -60 17 0.19 80 0.57 29 44 0.76 - - 21 21 0.86 10 0.06 Spherical 2 3East Bl-Cu 327 57 -60 81 0.71 - - 36 47 1.15 - - 23 21 1.38 10 0.05 Spherical 2 3

Kriging East Bl-Co 327 57 -60 51 0.37 - - 22 23 0.47 - - 10 11 0.34 5 0.06 Spherical 2 3East Bl-Au 327 57 -60 79 1.07 - - 37 23 1.74 - - 10 21 2.00 10 0.05 Spherical 2 3East Bl-Pt 327 57 -60 83 0.64 - - 29 48 0.85 - - 19 28 1.44 13 0.19 Spherical 2 3East Bl-Pd 327 57 -60 82 0.69 - - 39 23 0.87 - - 11 26 1.72 12 0.05 Spherical 2 3

Bl - blebby sulphide, Dis - disseminated sulphide

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Kriging (Indicated) and ID2 (Indicated First Pass)



( º ) ( º ) ( º ) (m) (H) (m) (H) Range (m) (H) (m) (H) Range (m) (H) Range (H) Hole SamplesWest Bl-Ni 327 57 -60 80 N/A - - 45 N/A - - 20 N/A - N/A N/A 2 3West Bl-Cu 327 57 -60 80 N/A - - 45 N/A - - 20 N/A - N/A N/A 2 3

ID2 West Bl-Co 327 57 -60 80 N/A - - 45 N/A - - 20 N/A - N/A N/A 2 3Second West Bl-Au 327 57 -60 80 N/A - - 45 N/A - - 20 N/A - N/A N/A 2 3

Pass West Bl-Pt 327 57 -60 80 N/A - - 45 N/A - - 20 N/A - N/A N/A 2 3West Bl-Pd 327 57 -60 80 N/A - - 45 N/A - - 20 N/A - N/A N/A 2 3


35


Kriging (Inferred) and ID2 (Inferred)



( º ) ( º ) ( º ) (m) (H) (m) (H) Range (m) (H) (m) (H) Range (m) (H) Range (H) Hole SamplesWest Bl-Ni 327 57 -60 160 N/A - - 90 N/A - - 20 N/A - N/A N/A 3 3West Bl-Cu 327 57 -60 160 N/A - - 90 N/A - - 20 N/A - N/A N/A 3 3West Bl-Co 327 57 -60 160 N/A - - 90 N/A - - 20 N/A - N/A N/A 3 3

ID2 West Bl-Au 327 57 -60 160 N/A - - 90 N/A - - 20 N/A - N/A N/A 3 3West Bl-Pt 327 57 -60 160 N/A - - 90 N/A - - 20 N/A - N/A N/A 3 3West Bl-Pd 327 57 -60 160 N/A - - 90 N/A - - 20 N/A - N/A N/A 3 3

East Dis-Ni 327 57 -60 34 0.25 209 0.93 see along Strike parameters 38 0.43 N/A Spherical 3 3East Dis-Cu 327 57 -60 40 0.51 216 1.03 see along Strike parameters 33 0.35 N/A Spherical 3 3East Dis-Co 327 57 -60 35 0.23 199 0.45 see along Strike parameters 47 0.16 N/A Spherical 3 3East Dis-Au 327 57 -60 28 0.31 216 0.93 see along Strike parameters 27 0.31 N/A Spherical 3 3East Dis-Pt 327 57 -60 28 0.26 210 0.89 see along Strike parameters 26 0.29 N/A Spherical 3 3East Dis-Pd 327 57 -60 25 0.26 211 0.96 see along Strike parameters 27 0.26 N/A Spherical 3 3

KrigingWest Dis-Ni 327 57 -60 170 0.70 - - 185 0.77 - - 50 0.60 0.23 Spherical 3 3West Dis-Cu 327 57 -60 133 0.48 - - 211 0.74 - - 58 0.55 0.17 Spherical 3 3West Dis-Co 327 57 -60 87 0.20 - - 71 0.12 180 0.22 16 0.14 0.06 Spherical 3 3West Dis-Au 327 57 -60 140 0.46 - - 188 0.74 - - 32 0.56 0.16 Spherical 3 3West Dis-Pt 327 57 -60 142 0.45 - - 157 0.57 - - 32 0.19 0.12 Spherical 3 3West Dis-Pd 327 57 -60 136 0.38 - - 165 0.65 - - 32 0.25 0.18 Spherical 3 3

KrigingEast Bl-Ni 327 57 -60 33 0.19 159 0.57 89 0.76 - - 42 0.86 0.06 Spherical 3 3East Bl-Cu 327 57 -60 162 0.71 - - 94 1.15 - - 42 1.38 0.05 Spherical 3 3East Bl-Co 327 57 -60 101 0.37 - - 46 0.47 - - 23 0.34 0.06 Spherical 3 3East Bl-Au 327 57 -60 158 1.07 - - 46 1.74 - - 41 2.00 0.05 Spherical 3 3East Bl-Pt 327 57 -60 166 0.64 - - 96 0.85 - - 56 1.44 0.19 Spherical 3 3East Bl-Pd 327 57 -60 165 0.69 - - 46 0.87 - - 52 1.72 0.05 Spherical 3 3


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3.7.1.6 Bulk Specific Gravity

The specific gravities used for the resource model were obtained from measurements taken from test work performed by URSA personnel on drill hole numbers UR-03-23, UR-03-26, UR-03-30 and UR-03-36. Representative samples from all lithologies were taken and subjected to a wet/dry bulk specific gravity determination test. A total of 257 samples were analyzed. The results are set out in Table 3.7.6 below.

Table.3.7.6 Bulk Specific Gravity

Lithology Lithology Code Bulk Specific Gravity

Quartzites 1a 2.67 Biotite quartz diorite 4d 2.78 Quartz gabbro 4c 2.91 Rock fragment phase 4f 3.00 Melagabbro 4b 3.02 Nipissing gabbro 3a 2.97 Mafic dyke 6a 3.08

The mineralization is contained within lithological units 4b and 4f.

3.7.1.7 Block Modelling

A block model framework was created with 9,120,000 blocks that were 5 m in the X direction, 5 m in the Y direction and 5 m in the Z direction. There were 380 columns (X), 200 rows (Y) and 120 levels (Z). The model was rotated 33 degrees in order to align it with the drill sections. The coordinates for the block model are in UTM units.

A percent block model was set up to accurately represent the volume and subsequent tonnage that was occupied by each block and partial block inside the constraining domains. As a result, the domain boundaries were properly represented by the percent model’s ability to measure infinitely variable inclusion percentages.

The nickel, copper, cobalt, gold, platinum and palladium composites were extracted from the Microsoft Access database composite table into 24 separate files for the East and West Blebby and East and West Disseminated domains.

Inverse distance squared interpolation was used for the West Blebby domain while Ordinary Kriging interpolation was used for the East Blebby and both East and West Disseminated domains. Two interpolation passes were used to determine the Indicated and Inferred classifications and to interpolate nickel, copper, cobalt, gold, platinum and palladium grades into each block. Contained metal values in Canadian dollars were also calculated for each block using the commodity price and exchange rate assumptions set out in Section 4.3.1. The contained metal values were used to estimate a simplified NSR for the Whittle pit optimization process. The open pittable mineralization table was reported from the Whittle runs.

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The interpolation parameters utilized for nickel, copper, cobalt, gold, platinum and palladium by domain are described in Section 3.7.1.5.

3.8 RESOURCE CLASSIFICATION

For the purposes of this resource estimate, confidence category classifications were derived from the Indicated and Inferred search ranges and interpolation parameters for nickel. Any grade block coded by the Indicated search ellipsoids for the nickel grade interpolation run was classified as Indicated and all other blocks were classified as Inferred. All of the grade blocks inside the constraining domains were coded as shown in Table 3.8.1 below.

Table 3.8.1 Grade Block Coding by Confidence Category

Domain Number Percent

East Blebby Domain Indicated Grade Blocks 11,497 14.9 East Blebby Domain Inferred Grade Blocks 7,686 10.0 West Blebby Domain Indicated Grade Blocks 2,074 2.7 West Blebby Domain Inferred Grade Blocks 607 0.8 East Disseminated Domain Indicated Grade Blocks 42,199 54.7 East Disseminated Domain Inferred Grade Blocks 1,665 2.2 West Disseminated Domain Indicated Grade Blocks 10,030 13.0 West Disseminated Domain Inferred Grade Blocks 1,339 1.7 Total Grade Blocks 77,097 100.0

3.9 MINERAL RESOURCES

The mineral resource estimates used as the basis for this feasibility study were classified according to the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council November, 2004.

Under the CIM definitions, a mineral resource must be potentially economic in that it must be “in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction.” The Shakespeare deposit plunges toward the northeast so it was necessary to determine the limit within which there are reasonable prospects for economic extraction by open pit and by underground methods, Micon has applied Whittle pit optimization software and reports mineral resources from within the resulting open pit shell having NSR value above a cut-off of Cdn$24.23/t and resources outside (beneath) the shell with potential for underground mining based on having an NSR value above a cut-off of Cdn$50.00/t.

The resulting mineral resource estimate is based upon a review of the deposit geometry, local topography and recent experience of actual mining and processing costs on other projects in the Sudbury and other parts of Northern Ontario. The parameters used to determine the cut-offs within the Whittle software are shown in Table 3.9.1.

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Table 3.9.1 Parameters Used in Pit Optimization (Cdn$ per tonne)

Ore mining cost, $/t 3.06 Waste mining cost, $/t 1.89 Haulage to plant, $/t 0.34 Processing costs, $/t milled 11.00 G & A costs, $/tmilled 1.73 Smelter Treatment charges, $/t concentrate

212

Concentration ratio 19:1 Total cash operating costs, $/t milled 43.65 Payable metal, % of total

Ni Cu Co Au Pt Pd

69.0 83.7 35.5 28.9 55.3 34.0

Pit Slopes, degrees 55

A cut-off NSR value of Cdn$24.23/t was applied to mineralization within the pit shell to determine the open pit mineable portion of the mineral resource as shown in Table 3.9.2 below

Table 3.9.2 Shakespeare Open Pit Mineral Resources (Cut-off Cdn$24.23/t NSR)


Cu (%)

Co (%)

Au (g/t)

Pt (g/t)

Pd (g/t)

NSR Cdn$/t

Indicated East 9,460,000 0.37 0.38 0.02 0.204 0.357 0.393 61.25 West 2,970,000 0.29 0.33 0.02 0.181 0.333 0.361 50.90 Total 12,430,000 0.35 0.37 0.02 0.199 0.351 0.386 58.78


A Cdn$50.00/t cut-off was applied to mineralization beneath the pit to determine the underground mineable portion of the mineral resource as shown in Table 3.9.3 below.

Table 3.9.3 Shakespeare Underground Mineral Resources (Cut-off Cdn$24.23/t NSR)


Cu (%)

Co (%)

Au (g/t)

Pt (g/t)

Pd (g/t)

NSR Cdn$/t

Indicated East 1,763,000 0.37 0.41 0.03 0.219 0.363 0.388 62.68 West 69,000 0.35 0.43 0.02 0.176 0.327 0.361 60.17 Total 1,832,000 0.37 0.41 0.03 0.218 0.361 0.387 62.59


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Most of the resource blocks within the pit were coded as indicated resources and the mineralization shows good continuity from hole to hole and section to section. Sectional and plan views of the block model, showing copper and nickel grades and Canadian dollar NSR values are included in Appendices 2n through 2u.

3.10 CONFIRMATION OF ESTIMATION

The use of a Whittle-optimized pit renders meaningless the reporting of sensitivity of the mineral resource estimate to cut-off grade, unless the pit were completely re-optimized using different assumptions. As such, no analysis of the block model at other contained metal value cut-offs has been performed.

However, as a test of the reasonableness of the estimate, the block model was queried at a Cdn$0.01 cut-off and all blocks were summed. The values of the interpreted grades for the entire block model were compared to the length weighted average grades of all samples from within the Blebby and Disseminated domains. The results are presented in Table 3.10.1.

Table 3.10.1 Comparison of Weighted Average Grade of Database to

Total Block Model Average Grade

Category Ni (%)

Cu (%)

Co (%)

Au (g/t)

Pt (g/t)

Pd (g/t)

Database - Blebby Domains 0.269 0.239 0.022 0.117 0.209 0.210 Block Model – Blebby Domains 0.220 0.232 0.022 0.116 0.206 0.208 Database - Disseminated Domains 0.332 0.372 0.023 0.196 0.348 0.389 Block Model - Disseminated Domains 0.300 0.341 0.021 0.186 0.322 0.354

The comparison above shows that the estimation is not significantly biased relative to the data used to create it.

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4.0 MINING AND MINERAL RESERVES

4.1 MINING METHOD

Mining will be carried out using conventional open pit methods. Drilling will be performed on all ore and waste with conventional in-the-hole hammer rigs with single pass drilling capability. Blasting operations will use an emulsion-ANFO blend and a downhole delay initiation system. A hydraulic face-shovel and a front-end-loader will be used to load rigid haulage trucks of 91-t capacity.

Haulage trucks will exit from the adjacent East and West open pits onto a common haul road approximately 1.7 km west of the plant site. Inert mine rock will be hauled to a storage area located 2.5 km from the exit of the pits to the north east, where it will be co-mingled with hydraulically placed mill tailings. Potentially acid generating (PAG) mine rock will be stored adjacent to the West pit for future placement within that pit once mining operations have ceased.

4.2 PIT SLOPE ANALYSIS

URSA had previously engaged Wardrop Engineering Inc. (Wardrop) to conduct a geotechnical stability analysis, the results of which were provided in the report, Shakespeare Project Preliminary Open Pit Design, dated November, 2004.

Golder Associates (Golder) was retained by URSA to carry out a pit slope design study as part of the overall open pit design. Golder’s findings were presented in the Report on Feasibility Study Level Pit Slope Design and Water Inflow Study, Shakespeare Project, Webbwood, Ontario, Canada, dated January 10, 2006. (See Appendix 4).

Golder’s analysis included a review of the existing data which comprised:

• Review of the Wardrop geotechnical stability analysis and preliminary pit slope design.

• Review and analysis of existing data inlcuding point load data, discontinuity measurements and large discrete features.

• Assessment of regional geology.

The regional and local geology has been discussed in Section 3 of this report and will not be referred to further here.

4.2.1 Review of Wardrop Preliminary Design

Point tests had been performed by URSA on 1,076 representative samples of core from six drill holes the results of which are shown in Table 4.2.1.

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Table 4.2.1 Summary of Point Load Strength Results (Wardrop)

Drill Hole Average UCS (MPa)

Maximum UCS(MPa)

Minimum UCS (MPa)

Standard Deviation

Population

U-03-23 241 332 49 67 242 U-03-26 240 332 42 75 106 U-03-30 242 332 8 73 230 U-03-36 234 332 34 85 198 U-03-38 211 332 26 87 150 U-03-61 281 332 45 63 76 Rock mass properties were estimated by Wardrop using derived relationships based on rock mass quality and compressive strength as shown in Table 4.2.2. An average unit weight of 27 kN/m3 was used for the Wardrop stability analysis.

Table 4.2.2 Empirical Rock Mass Properties (Wardrop)

Rock Mass Property Minimum Average Maximum Rock mass rating 75 80 84 Cohesion (MPa) 10.6 13.6 16.7 Friction angle 40.0 42.5 44.5 Compressive strength (MPa) 45 60 77

Surface mapping by URSA of the mineralized zones, which are related to an east-west trending shear zone, indicated a series of transverse fault zones that strike perpendicularly to the trend of the mineralization.

The orientation of 569 structural discontinuities were mapped and plotted as a series of stereonet projections. The general orientation of structural discontinuities is summarized in Table 4.2.3.

Table 4.2.3 Orientation of Structural Discontinuities

Structure Dip (degrees) Dip Direction (degrees) Bedding 70 246 Foliation 76 251

75 251 Jointing 80 337

4.2.2 Compressive Strength Testing

Golder assessed the compressive strength for each of the lithologies occurring in the Shakespeare deposit with the results shown in Table 4.2.4.

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Table 4.2.4 Compressive Strengths by Lithology from Point Load Data

Lithology Code

Lithology Average UCS(MPa)

Maximum UCS (MPa)

Minimum UCS (MPa)

Standard Deviation

Population

1a Quartzite 187 332 8 70 266 1b Metasediments 112 115 109 4 2 3a Nipissing gabbro 280 332 49 60 295 4b Massive melagabbro 294 332 106 50 60 4c Quartz gabbro 293 332 160 42 114 4d Biotite-quartz gabbro 215 332 42 69 298 4f Rock-fragmented

melagabbro 223 332

45 87 12

5a Rheomorphic breccia 60 60 60 1 6a Diabase dyke 250 332 58 97 22 6b Fine-grained mafic

intrusive 156 217 34 52 12

3a, 4b, c, d, f

All gabbro 258 332

42 70 779

4.2.3 Geotechnical Field Investigation

The geotechnical field studies comprised the following programs:

• Drilling of six NQ-sized diamond drill holes. Data from a seventh hole, which had been previously drilled was also used.

• Collection of rock mass properties and logging of all seven holes.

• Logging of six of the holes with an optical televiewer in order to identify discontinuities in the pit area.

• Hydraulic conductivity testing consisting of single packer tests conducted on 14 intervals in 6 of the holes.

Details of the geotechnical drill holes are summarized in Table 4.2.5.

Table 4.2.5 Summary of Geotechnical Drill Holes

UTM Coordinates NAD 83 Hole

Number East North Elevation (m)

Inclination (degrees)

Azimuth (degrees)

Length (m)

U-03-42 435977 5133188 329 -45 329 142.0 U-03-73 435900 5133271 319 -48 331 158.4 U-03-76 436412 5133619 295 -74 159 215.0 U-03-77 436367 5133531 324 -45 146 127.0 U-03-78 436460 5133657 325 -55 322 156.0 U-03-79 436608 5133649 322 -50 147 174.0 U-03-80 436727 5133713 330 -59 063 115.0

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The detailed geotechnical logging provided the following data:

• Depth. • Lithology and features of interest. • Feature depth. • Run number. • Total and solid core recovery. • Rock Quality Designation (RQD). • Fracture frequency. • Discontinuity dip with respect to core axis. • Type and surface description of discontinuity. • Rock strength and weathering indices. • Comments.

The optical televiewer logging of intervals in six holes (all except U-03-77) showed that few open discontinuities were present and joints and foliations were generally tight or healed. In contrast to the previous logging, the televiewer data indicate that the drilling process and/or the handling of core open up discontinuities and show poorer Rock Mass Quality (RMQ).

Packer testing was undertaken by a contractor under the supervision of Golder and used conventional wireline packer testing equipment. The results, summarized in Table 4.2.6, show that relatively low hydraulic conductivities were encountered, generally in the range of 10-5 cm/s and 10-6 cm/s, and that hydraulic conductivity decreases with depth.

Table 4.2.6 Hydraulic Conductivity Estimates

Depth (m) Hole Number Pit From To Average

Hydraulic Conductivity

(cm/s) U-03-70 West 20 56.2 38.1 9.02E-06 U-03-71 West 20 38.1 29.05 2.56E-06 U-03-73 West 10 150 80 2.92E-06 U-03-76 East 118.3 215 166.65 8.06E-08 U-03-76 East 76.2 118.3 97.25 2.02E-06 U-03-76 East 43.3 76.2 59.75 5.90E-06 U-03-78 East 70.3 102 86.15 4.88E-06 U-03-78 East 38.6 70.3 54.45 1.86E-06 U-03-78 East 8.3 38.6 23.45 2.47E-06 U-03-78 East 102.3 155 128.65 8.83E-07 U-03-79 East 94.6 138 116.15 3.54E-06 U-03-79 East 52.3 138 95.15 2.64E-06 U-03-79 East 13.3 138 75.65 1.66E-06 U-03-79 East 137.5 175 156.25 3.39E-06

Golder compared the data obtained using the optical televiewer and the earlier discontinuity data collected from the surface mapping program. Golder observed that the majority of the data from the surface mapping program was collected along the perimeters of the pits at relatively high elevation and which it noted tends to skew the data towards the orientation of

44

the dominant wall and from the exposed lithology. The data from the optical televiewer was collected from subsurface, in-situ discontinuities and across geological, structural and alteration domains. The optical televiewer data also provides information on whether the discontinuity is open, tight or healed. The Golder analysis, therefore, focused on data collected using the optical televiewer. The data from both programs are summarized in Table 4.2.7.

Table 4.2.7 Summary of Discontinuity Sets for Surface Mapping and Optical Televiewer Logging

Televiewer Logging Surface Mapping Feature Type Set Set Type Set Set Type

JC1 Major JM2 and JM3 Major/minor Moderate to steep dip NNW JC2 Minor JM5 Minor Shallow dip SE JC3 Minor None n.a. Moderate dip SE JC4 Minor JM4 Minor Moderate to steep dip SSE JC5 Minor None n.a. Shallow to moderate dip NW JC6 Minor JM1 Minor Moderate to steep dip ENE JC7 Minor None n.a. Moderate to steep dip SW

Golder assessed the point load tests undertaken by URSA in order to provide an estimate of the compressive strengths of each lithology. Both UCS and International Society for Rock Mechanics (ISRM) tests were completed for six boreholes in the geotechnical program in 2005.

A comparison of the results of the point load tests and the UCS testing indicated good agreement between the two data sets.

Golder developed rock mass classifications based on the 2005 geotechnical drill core data, summarized in Table 4.2.8.

Table 4.2.8 Shakespeare Project Rock Mass Classification

Pit Domain Main Lithology

Q1 Min Q Ave Q Max RMREquivalent 2

Minimum RMREquivalent

Average RMREquivalent

Maximum East pit hanging wall

Quartzite 0.2 10.0 58.5 28.3 59.2 80.6

East pit footwall

Gabbro 1.4 12.3 200.0 47.0 60.1 91.7

West pit hanging wall

Quartzite 0.3 13.3 39.1 33.5 63.6 77.0

West pit footwall

Gabbro 0.5 17.1 387.8 37.1 56.5 97.6

1 Barton’s Q System 2 Bieniawski/Laubsher’s RMR system

Golder concluded that, on the basis of the calculated Q values from the core logged during the 2005 geotechnical program, the rock quality in both the East and West pits can be characterized as good. In view of the open discontinuities logged with the televiewer, Golder concludes that the Q rating should be in the range of good to very good.

45

4.2.4 Pit Wall Design

Golder carried out kinematic stability and limit equilibrium assessments in order to evaluate the effect of geological structure on pit wall stability.

The resulting wall slope parameters are summarized in Table 4.2.9.

Table 4.2.9 Wall Slope Parameters

Pit Sector Bench Height (m)

Catch Berm Width (m)

Bench Face Angle (degrees)

Inter-ramp Angle (degrees)

East pit hanging wall 20 8.5 75 55 East pit footwall 20 8.5 75 55 West pit hanging wall 20 10.5 80 55 West pit footwall 20 8.5 75 55

The West pit hanging wall has potential for bench scale planer failures and Golder recommends that careful wall control blasting will be required in order to minimize backbreak and damage to the walls.

Golder recommended double-benching with 8.5 m catch berms and a 75° bench face angle. This results in an inter-ramp slope of 55°. Pre-splitting of the final 20 m high bench face is required.

4.2.5 Pit Water Inflow

Golder’s assessment concludes that the majority of inflow into the open pits will result from precipitation and surface runoff.

Continuous groundwater inflow from the bedrock aquifer was estimated from bedrock hydraulic test data and determined as follows, based on 10% and 50% poor rock quality, respectively:

East pit: 170-450 m3/d.

West pit: 50-134 m3/d.

The surface water runoff comprises direct precipitation into the pits and the contribution from the up-gradient catchment area. For both pits, it was estimated that 51% of the surface water inflow would come from direct precipitation. Maximum inflow is expected to occur in April due to a combination of rainfall and snowmelt.

Short term pumping requirements were assessed on the basis of the Timmins 100-y regional storm as a result of which the April volume may be expected to occur within a 12-h period.

The pumping capacities for the 2-y return period 24-h storm and the Timmins 100-y regional storm conditions for each pit are as follows:

46

East pit: 3,270 USgpm and 11,900 Usgpm (or approx. 18,000 and 65,000 m3/d respectively

West pit: 1,670 USgpm and 6,100 USgpm (or approx. 9,000 and 33,000 m3/d respectively

Further evaluation will be required in order to determine the optimum pumping and water storage capacity for each pit to handle anticipated precipitation events.

4.3 OPEN PIT DESIGN

The Shakespeare deposit is amenable to development as a open pit mine. In order to determine the economic limits of open pit mining, optimized open pit shells were developed using Whittle 4X pit optimization software.

4.3.1 Net Smelter Return Values

For a polymetallic deposit such as at the Shakespeare Project, the NSR value of the resource is used to determine the net value of each mining block.

The NSR for each block model was calculated utilizing a number of parameters on the basis of the anticipated concentrate recovery, smelter payables and applicable refining charges, as shown in Table 4.3.1.

Table 4.3.1 Parameters for Calculation of NSR

Metal Unit Metal Prices $US per unit

Concentrate Recovery,

%

Smelter Payables, %

Refining ChargesCdn$ per unit

Ni lb 5.158 76.0 92 0.64 Cu lb 1.259 95.5 89 0.31 Co lb 21.332 71.0 50 3.03 Au oz 437.26 38.0 85 18.24 Pt oz 780.81 75.0 85 18.24 Pd oz 233.65 42.0 85 18.24

In addition, for the purposes of pit design, the following assumptions were made:

• A concentration ratio of 30.4:1.

• A smelter treatment charge of Cdn$212/t or Cdn$6.97/t of ore milled.

• A currency exchange rate of US$0.8224:Cdn$1.00 was used.

4.3.2 Model Preparation

In the resource estimation process, four domains were created to represent the East Blebby, East Disseminated, West Blebby and West Disseminated mineralized zones. In some places, these domains are less than one block model block width (5 m). All blocks were processed and

47

combined into homogenous blocks with a single NSR value. In the event that a block was only partially contained within a domain, the volume portion outside of the domains was assigned a background NSR value of Cdn$3 per tonne. This background grade was based on the average NSR value existing within the deposit outside the ore zones.

The block homogenization procedure is a critical step prior to export to Whittle 4X in order to ensure that the model is representative of actual operating conditions while taking account of the dimensions of the selected mining unit (5 m x 5 m x 5 m), dilution, mining losses and operational ore/waste misclassification. The homogenized model was exported to Whittle via Gemcom’s Whittle 4X export utility for use in the pit optimization process.

4.3.3 Optimization Parameters

The operating costs (in Cdn$) and parameters used in the Whittle 4X optimization are given in Table 4.3.2.

These preliminary cost estimates, used for pit optimization, necessarily differ somewhat from the final unit costs used in the financial analysis section of this report, the latter being derived in part from the results of the pit optimization and design process.

Table 4.3.2 Whittle Pit Optimization Parameters

Waste mining cost, Cdn$/t 1.68 Ore mining cost, Cdn$/t 1.68 Ore processing cost, Cdn$/t 10.41 General and Administration cost, Cdn$/t 1.34 Process production rate (ore tonnes per year) 1,642,500 Pit slopes 55o

4.3.4 Generation of Optimized Pit Shells

The preceding operating costs and parameters were entered into Whittle 4X for the optimization runs. A structure arc file was generated from the pit slope requirements and was adjusted to keep the calculated slope error to less than one degree. Due to the large number of blocks in the model (9.12 million), re-blocking was necessary in order to significantly reduce optimization processing time.

Pit shells were optimized over an NSR factor range of 10% to 110% of the value determined by the base NSR parameters, in increments of 1%. The resulting optimization produced 74 nested pit shells. The net present value (NPV) of each of the nested pit shells is presented in the chart at . The shell with the highest NPV, number 64, was selected for design purposes.

48

Figure 4.1 NPV of Nested Pit Shells

4.3.5 Open Pit Design

For each of the East and West pits, the optimized pit shell was exported from Whittle 4X into Gemcom and a 3-dimensional surface was created and subsequently contoured on 10-m intervals that matched the block model vertical spacing. These contours were then used as a guide for the creation of toe and crest lines with Gemcom’s pit design tool utility.

The East pit will be mined in two phases. Phase I uses the upper portion of the permanent ramp and accesses higher grade ore at a lower waste to ore ratio at an early stage. Phase II features a push-back to the north and northeast and allows access to the full economic depth of the pit. The benches were doubled to a final height of 20 m with catch berm widths of 8.5 m and a batter (bench face) angle of 75o. This configuration results in an inter-ramp slope of 55°. In some cases, bench heights were reduced to 10 m to accommodate localized pit bottom access, as well as to minimize mined waste tonnage during the mining of flat ore zones.

Pit main access ramps have a 10% gradient and have drive-on access to all berms for scaling and loose rock removal. Ramp width is 24 m to facilitate two-way truck traffic at all points along the open pit haulage roadway. Final pit bottom access ramps were designed at a gradient of 15% and a width of 15 m to accommodate one-way traffic. In each pit, the last bench will be excavated primarily with a front end loader or back-hoe due to the steep temporary access ramp on broken ore and minimal working space.

Consideration was given at all times in the design process to issues regarding existing topography, haulage roads, mine rock storage locations and drainage areas. During the design process, consideration was also given to providing adequate operational space requirements for equipment in areas approaching final design pit walls. Pit design data prepared by P&E are presented in Appendix 2b. The final pit ouline is shown in Figure 4.2.

$

$20,000

$40,000

$60,000

$80,000

$100,000

$120,000

$140,000

$160,000

1 6 11 16 21 26 31 36 41 46 51 56 61 66 71

Pit Shell Number

NPV

(7%

) (C

DN

$ 00

0)

49

Figure 4.2 Plan showing Ultimate Pit

49

50

4.4 MINERAL RESERVES

4.4.1 Mining Dilution & Recovery

Conversion of mineral resources into reserves takes into account dilution and losses occuring during blasting and loading operations. Dilution tonnage was assigned to the mineable ore zones at a rate of 10%, equivalent to the amount of dilution contributed by one half of a drill pattern burden spacing on both the footwall and hangingwall contacts of the ore zones. Therefore, at the drill pattern burden of 5.7 m and an average ore zone width of approximately 56 m, the assigned dilution was calculated at (5.7m/56m) x 100 = 10.2%. Additionally, a mining recovery rate of 97% was applied to the diluted tonnage to arrive at an overall effective dilution rate of 6.9%.

Dilution grade was added to the dilution tonnage as determined by the average metal assay values for the mineralization between the ore economic contact limits and the edge of the resource estimate constraining domain envelope. The resulting diluting grades were as follows: Ni = 0.102%, Cu = 0.096%, Co = 0.012%, Au = 0.050 g/t, Pt = 0.098 g/t and Pd = 0.105 g/t. This dilution mineralization was consolidated on a weighted average basis of 6.9% with the undiluted resource to arrive at a final diluted ore grade.

4.4.2 Statement of Mineral Reserves

The mineral reserves were determined by applying the Cdn$11.75/t NSR internal cut-off value described in section 4.3.3 to the modeled mineralization contained within the final pit design.

The cut-off value (Cdn$11.75/t) is the sum of the mill processing costs and general and adminstration costs. The value of material hauled from the pit must exceed this cutoff value in order to be profitably processed. Material only slightly below the cut-off may be stored on a low-grade stockpile for processing later, at the end of the mine life, if storage space permits. The remainder, having little or no value, will be sent to the mine rock storage facility.

All of the material designated as ore in the pit design was derived from Indicated Resources, is demonstrated to be economic in this feasibility study, and is classified as a Probable Mineral Reserve using the CIM standards.

The mineral reserve estimate presented in Table 4.4.1 is current as of December 17, 2005.

Table 4.4.1 Mineral Reserve Statement (Cut-off Cdn$11.75/t NSR)

Classification Tonnes %Ni %Cu %Co g/t Au g/t Pt g/t Pd

Probable Mineral Reserve 11,226,000 0.33 0.35 0.02 0.19 0.33 0.37

51

4.5 OPEN PIT PRODUCTION SCHEDULE

Ore will be mined from the East and West pits at an average rate of 4,500 t/d for 6.8 years. Haulage trucks will dump directly to the crusher or to a nearby stockpile. A front-end-loader will move material from the stockpile to the crusher as required. The average stripping ratio for the project is 5.1 t of waste per tonne of ore. However, 10 million t of waste have been scheduled as pre-stripping and the ratio is 5.7:1 during the Year 1 to Year 4 period. This leaves a somewhat lower amount of waste to be stripped in Year 5 (5.2:1 ratio). Subsequently, the ratio will continue to fall, when operations will be recovering ore from the deepest benches of the East pit. In Years 6 and 7, shifts will be reduced in order to maintain a high productivity rate for the mining fleet.

The West pit will be mined out by early Year 5, after which the West pit will be used to store potentially acid generating mine rock.

Table 4.5.1 shows the annual production tonnage and grades mined over the life of the project.

Table 4.5.1 Open Pit Production Schedule


PRODUCTION Waste Mined ( 000 tonnes ) 10,000 10,403 10,403 8,212 9,672 7,117 1,825 180 - 57,813 Ore Mined ( 000 tonnes ) - 1,643 1,643 1,643 1,643 1,643 1,643 1,371 - 11,226

Waste/Ore Ratio - 6.33 6.33 5.00 5.89 4.33 1.11 0.13 - 5.15 Total Mined ( 000 tonnes ) 10,000 12,045 12,045 9,855 11,315 8,760 3,468 1,551 - 69,039 Ore Grade Nickel ( % ) - 0.362 0.387 0.331 0.266 0.282 0.338 0.344 - 0.330 Copper ( % ) - 0.345 0.402 0.369 0.298 0.304 0.361 0.376 - 0.350 Cobalt ( % ) - 0.024 0.025 0.023 0.021 0.020 0.023 0.023 - 0.023 Platinum ( grams / tonne ) - 0.335 0.370 0.370 0.303 0.273 0.332 0.342 - 0.332 Palladium ( grams / tonne ) - 0.373 0.424 0.416 0.325 0.290 0.362 0.376 - 0.366 Gold ( grams / tonne ) - 0.179 0.205 0.203 0.164 0.153 0.195 0.209 - 0.186

4.6 OPEN PIT MINING OPERATIONS

Open pit operations will be carried out on two 12-hour shifts, seven days per week for 50 working weeks per year. The rotation schedule will have employees working 7 days on, 7 days off. The shift change will be a hot change with mine employees transported to the drills and loading equipment and truck operators changing either at the truck shop area or in the open pit area. All equipment will be diesel powered.

52

4.6.1 Drilling and Blasting

Two drill rigs with in-the-hole hammers will be used to drill 9-in (22.9 cm) diameter blast-holes on an approximately 5.7 m by 7.2 m pattern.

Considering the hardness of the rock, blast-holes will be loaded with a blend of emulsion explosives and ANFO delivered to the holes by the explosives supplier. The supplier will erect a small plant on site and use a blending truck between the plant and the pit. The blend will be adjustable, but current design is based on 70% emulsion, 30% ANFO. The product will have a high velocity of detonation and good water resistance. The design powder factor is 0.4 kg of explosives per tonne of rock. After mining commences, fragmentation will be evaluated and the drill and blast parameters may be further refined to optimize results.

4.6.2 Load and Haul

The main loading tool will be a 9.2-m3 hydraulic front shovel. As back up to the primary loader, a 10.5-m3 front-end loader will be available for simultaneous loading from a second face, clean-up around the shovel, clean-up after blasting, etc. A fleet of up to six 91-t payload trucks will haul ore to the crusher and mine rock to the storage area.

4.6.3 Grade Control

Grade control will be overseen by the mine geologists. Cuttings from blastholes in or near ore will be sampled for grade control purposes. The results of these samples will be used to determine which portion of the mineralization is above the Cdn$11.75/t NSR internal cut-off value. Ore blasting limits will be set accordingly.

4.7 MINE SUPPORT FACILITIES

4.7.1 Logistical support

Half-ton pickup trucks will be used by engineering, geology and management personnel for travelling on the site and between the site and other facilities.

4.7.2 Dewatering

Each pit will have a sump to collect seepage and water from precipitation. Initially, each pit will be equipped with a 43-kW submersible pump, while later the two sumps will be equipped with 104-kW pumps. In the final stages, the 43-kW pumps can be used as in-line boosters. The East and West pits will have separate 150-mm pipelines discharging into the settling pond near the waste storage area north of the pit. These pumps will be suspended from flotation modules and controlled to start and stop automatically.

4.7.3 Mine Dry

Separate changing and washing facilities (“mine dry”) for men and women will be located adjacent to the mine superintendent’s office. Employee parking will be available close to this facility.

53

4.8 PERSONNEL

Mining crews will work 12-h shifts around the clock, seven days per week, using a 7 days on, 7 days off rotation. In Years 6 and 7, the schedule will become a single shift 12-h/d, 7 days per week. Mining personnel requirements are listed in Table 4.8.1.

Table 4.8.1 Mining Personnel

Description Years 1 to 5 Mine superintendent 1 Shift boss 4 Production loader operator 6 Production truck driver 20 Driller 8 Blaster 2 Blaster helper 2 Dozer, grader, water truck etc. 10 Stockpile loader operator 4 Service truck driver 4 Labourer 8 Maintenance foreman 2 Mechanics 8 Welders 2 Mechanical technician 1 Tech. Services superintendent 1 Geologists 2 Core storage labourer 2 Mining engineer 1 Technicians 2 Surveyors 2 Computer technician 1 Environmental coordinator 1 Total 94

4.9 MINING OPERATING COSTS

Mine maintenance and technical services are included in the mine operating cost. The labour cost estimate is Cdn$0.67/t of rock moved, based on current rates (2005) for similar operations in Northern Ontario. The consumables cost estimate is Cdn$1.08/t of rock, for a total estimated average mining operating cost of Cdn$1.75/t of rock, at full production and at the average waste to ore ratio. The mining operating cost breakdown is shown in Table 4.9.1.

54

Table 4.9.1 Mining Operating Cost (Cdn$/t rock mined)

Item Cost (Cdn$/t)

Production labour 0.47 Maintenance labour 0.10 Technical services labour 0.10

Sub-total labour 0.67 Fuel 0.44 Drill rods & bits 0.03 Explosives & accessories 0.40 Maintenance supplies 0.21

Sub-total consumables 1.08 Total Mine Operating Cost (Cdn$/t) 1.75

4.10 MINING CAPITAL COSTS

The principal mining equipment items are shown in Table 4.10.1. In order to accommodate the higher stripping ratio, this quantity of equipment will be required during the first five years of production. Haulage truck requirements will diminish towards the end of the mine life.

Table 4.10.1 Principal Mining Equipment Items

Item Quantity

Hydraulic shovel, 9.2-m3 1 Wheel loader, 10.5-m3 1 Haulage truck, 91-t 6 Wheel loader, stockpile, etc. 1 Production drill, 9-in 2 Track dozer, 346-kW 2 Excavator, general 1 Road grader 1 Crawler drill & compressor 1 Water truck 1

URSA intends that the following used equipment items will be purchased from the Lac des Iles Mine of NAP, and rebuilt for use at the Shakespeare operation:

3 Caterpillar 777B 91-t trucks.

1 Komatsu WA800 10.5-m3 wheel loader.

1 Caterpillar model D10 (substituted for 1 346-kW dozer).

The major mining production units will last the life of the operation. The used equipment will be fully rebuilt by the Ontario dealer so that it will be reliable for the required eight years. As a result, the sustaining capital requirements are low. The capital cost schedule is summarized in Section 11, below.

55

5.0 METALLURGY

Three phases of metallurgical testing have been conducted by SGS-Lakefield (SGS) during the development of the Shakespeare project. These testwork programs are termed; 2003 testwork, 2004 testwork and 2005 testwork. The 2003 testwork comprised preliminary flowsheet development including mineralogy, bulk flotation and an initial assessment of nickel-copper separation.

The metallurgical test work in 2004 was designed primarily on the premise that all production would be toll-milled at the Strathcona concentrator of Falconbridge. Accordingly, testing was focused on providing satisfactory proof that the metallurgical characteristics of the Shakespeare deposit would allow its treatment in that concentrator, either alone or mixed with Falconbridge ore. This testwork program included a series of locked cycle flotation tests.

The 2005 testwork program is an extension of the 2003 work and was designed to optimize a process for the on-site beneficiation of ore on site, with the production of a combined (Cu/Ni) concentrate.

5.1 2003 TESTWORK

Metallurgical testing of the Shakespeare deposit commenced in 2003, at SGS, with two composite samples of drill core provided by URSA. The assays for the two samples were similar, as shown in Table 5.1.1.

Table 5.1.1 2003 Sample Assays

Sample Reference Cu% Ni% Fe% Co% Au g/t Pt g/t Pd g/t U3Met1 0.59 0.53 12.3 0.03 0.25 0.57 0.55 U3Met2 0.49 0.44 11.1 0.03 0.25 0.45 0.53

Mineralogical studies and flotation tests were conducted. The mineralogical study concluded that the two samples were very similar. Both pentlandite and pyrrhotite grains were found to have a mean size of between 20-30 µm, and chalcopyrite and pentlandite were both found to have a high level of surface exposure (low mineral locking). Approximately 10% of the nickel is in solid solution in the pyrrhotite.

Flotation tests at different grind sizes led SGS to recommend 82 µm as the optimum (K80) grind size for nickel, but the flotation test results with sample Met2 were similar at grind sizes of 126 and 82 µm. SGS noted that finer grinding seemed to improve platinum group metal recovery but did not improve nickel recovery.

Initial tests (referenced F1 to F6) produced low-grade bulk concentrates due to excessive pyrrhotite recovery. Later tests, including the addition of lime, improved pyrrhotite rejection and, hence, the grade of the final concentrate, without significant loss of nickel recovery. However, as seen in Table 5.1.2, for staged recovery after 7 minutes roughing, there appears to be little advantage to the more selective flowsheet at the roughing level, and PGM recovery is also adversely affected.

56

Table 5.1.2 Comparison Rougher Flotation, Tests F2 and F9 (2003)

Grade (%) Recovery (%) Test Sample Product Wt. (%) Cu Ni % Cu+Ni Cu Ni Pt Pd

F-2 U3Met2 7 min. Ro. 8.6 7.74 5.56 13.30 96.7 86.8 85.7 61.6 F-9 U3Met2 7 min. Ro. 6.1 7.36 5.23 12.59 96.5 84.1 70.5 48.4 Feed 100.0 0.48 0.39

Subsequent tests employed cleaning of the rougher concentrate to obtain a final bulk concentrate, with the target of >20% combined Cu+Ni. Copper recovery was maintained during cleaning but nickel recovery fell steeply as the grade increased. Attempts to improve performance by regrinding the rougher concentrate and the use of pyrrhotite depressants were not successful. Test F-10 indicated the best conditions. A copper-nickel separation in tests F-13 and F-14 yielded copper concentrates with up to 25% Cu, but with unacceptably high nickel content. Summary results are shown in Table 5.1.3.

Table 5.1.3 Cleaning Flotation Tests (2003)

Grade Recovery % Test Sample Product Cu % Ni % Cu Ni Pt Pd

Bulk cl 2 Conc 13.50 8.80 93.5 68.8 68.7 27.6 Ro conc 6.23 4.44 96.2 77.6 77.3 49.9

F-10 U3Met2

Feed 0.47 0.41 Cu conc 25.10 3.94 90.0 16.8 9.5 10.3 Cu-Ni cl Conc 7.16 5.05 96.5 81.0 81.6 53.0

F-14 U3Met2

Feed 0.48 0.40

At the conclusion of the 2003 work, SGS recommended a further mineralogical study to investigate nickel losses in cleaning and the development of a less selective flotation upgrading scheme.

5.2 2004 TESTWORK

The metallurgical testwork in 2004, also conducted at the SGS laboratory, was designed primarily on the premise that all production would be toll-milled at the Strathcona concentrator of Falconbridge.

Accordingly, testing focused on providing satisfactory proof that the metallurgical characteristics of the Shakespeare deposit would allow its treatment in that concentrator, either alone or mixed with Falconbridge ore, as well as to yield (i) acceptable concentrates for the Falconbridge smelter, and (ii) economic recovery levels of copper, nickel, cobalt and platinum group metals.

The primary grind size and flotation process were selected to mimic those of the Strathcona concentrator. Samples from drill cores of the Shakespeare deposit were collected and combined to form a number of composites. Five composites of the two main geological zones, (Blebby and Disseminated), were prepared and assayed, then combined to form main composites. Within each group of five zonal composites the content of each metal is extremely

57

consistent, but there is some variation between the blebby and disseminated zone samples. The average content of the zone and main composites are shown in Table 5.2.1.

Table 5.2.1 2004 Composite Sample Assays

Composite Cu % Ni % Fe % S % Au g/t Pt g/t Pd g/t Blebby 1-5 av. 0.39 0.34 11.1 2.00 0.20 0.37 0.40 Disseminated 1-5 av. 0.42 0.50 14.3 4.28 0.34 0.48 0.50 Master Comp 1-5 av. 0.43 0.39 11.5 2.46 0.21 0.33 0.42

The significant difference between the Shakespeare ore and average feed to the Strathcona concentrator is the lower nickel content of the former at 0.38% Ni, compared with approximately 1.5% Ni at Strathcona. The first phase of the 2004 program examined the response of the Shakespeare ore following the processing scheme employed at the Strathcona mill and using the same reagent scheme. The sample was ground to an average of 140 µm (80% passing size, also employed for all following tests), prior to flotation with reagent additions of lime, xanthate, copper sulphate and frother added at various stages. After recovery of most of the copper and nickel to a bulk concentrate, additional nickel recovery was obtained by extended flotation of middlings products, regrinding of this product and re-flotation. The bulk concentrate was re-treated to separate copper and nickel concentrates. The flowsheet is illustrated in Figure 5.1. Copper and nickel recoveries by the above scheme were acceptable but the low nickel content of the feed prevented attainment of a suitable concentrate grade.

In follow up work, the flow sheet and conditions were tested on a sample of Strathcona ore. Recoveries were acceptable but, even with this sample, nickel concentrate showed a low final grade. The test was repeated with different blends of the two ore samples, and results are summarized in Table 5.2.2. For comparison, summary results only are shown for copper and nickel recovery. Nickel recovery in similar concentrate grades appears to be strongly dependent on the feed content, and all tests show difficulty in producing a concentrate grade consistently at or above 10% Ni.

Subsequent testing examined the possibility of increasing the concentrate grade and maintaining acceptable recovery by additional copper cleaning. Results were marginally improved but nickel concentrate grade versus recovery was still seen to be sensitive. The results for Test F-20 are presented in Table 5.2.2.

58

Figure 5.1 Testing Flow Sheet

Table 5.2.2

2004 Tests - Summary of Results

Grade (%) Recovery % Test Blend Shakespeare

and Strathcona

Product Cu Ni Cu Ni

Cu conc 19.00 1.91 86.8 4.5 Ni conc 0.28 4.98 10.0 91.8 F-4 0:100 Feed 0.78 1.53 Cu conc 11.70 2.44 88.5 12.2 Ni conc 0.28 4.11 8.4 81.7 F-5 60:40 Feed 0.56 0.85 Cu conc 17.00 1.92 87.0 5.9 Ni conc 0.30 4.47 9.9 89.1 F-6 40:60 Feed 0.64 1.06 Cu conc 16.90 2.31 85.9 6.4 Ni conc 0.29 4.33 11.2 89.7 F-7 20:80 Feed 0.70 1.29 Cu conc 21.20 1.05 85.7 4.5 Ni conc 0.78 6.36 8.8 74.9 F-20 100:0 Feed 0.42 0.40

Ball Mill

SecondaryRougher

Feed Ore

PrimaryRougher

PoR ejectionRougher

Cu/NiSeparation

Scavenger

Regrind

1

7

65

4

32

Product Legend1. Cu/Ni Sep Conc2. Cu/Ni Sep Tails3. 2° Ro Cl Conc4. Po Ro Conc5. Po Scav Con6. Po Scav Tail7. Ro/Scav Tails

2° RoCleaner

PoRejection

Scavenger

59

5.2.1 2004 Locked Cycle Tests

Locked cycle tests were conducted, together with analysis of cobalt and PGMs, on combined concentrates. Four or five cycles were employed with conditions similar to those established in the most recent 2004 batch tests. A full locked cycle test was completed for each of the Blebby, Disseminated and master composites.

Copper and nickel performances in these tests were not significantly different from those in the better batch tests. It appeared that while a concentrate grade of 10% Ni was obtainable, a nickel recovery of 65-68% would ensue. Results shown in Table 5.2.3 are compared by adding the lower grade products from the secondary recovery steps, reducing the concentrate grade to a level in the order of 5-7%, but maximizing recovery of nickel and the minor payable elements.

Table 5.2.3 Locked Cycle Test Results (2004)

Grade Recovery % Test Blend Shakespeare

Product Cu % Ni % Cu Ni

Cu conc1 14.00 1.20 86.8 6.2 Ni conc 0.38 5.21 6.4 78.5 LCT-1 Blebby Feed 0.39 0.48 Cu conc 20.90 1.35 84.8 6.1 Ni conc 0.99 7.89 7.7 67.9 LCT-2 Disseminated Feed 0.37 0.33 Cu conc 18.90 1.47 87.2 7.4 Ni conc 1.04 9.22 7.3 70.3 LCT-3 Master Comp1 Feed 0.42 0.38

Cobalt and PGM Recovery Test Blend Shakespeare

Product Co % Au % Pt % Pd %

Cu conc 6.1 19.7 10.1 13.6 Ni conc 76.4 24.0 71.4 33.2 LCT-1 Blebby Total 82.5 43.7 81.5 46.8 Cu conc 6.8 9.3 13.4 7.3 Ni conc 58.3 17.0 47.7 22.5 LCT-2 Disseminated Total 65.1 26.3 61.1 29.8 Cu conc 8.7 13.8 20.3 12.6 Ni conc 62.5 20.0 42.9 26.4 LCT-3 Master Comp1 Total 71.2 33.8 63.2 39.0

1 Test LCT-1 appeared to be unstable and with possibly too much collector added. Copper concentrate grade varied from above 18% to 13%, 14% being the weighted average of the last two cycles, but is not likely the optimum from this sample.

5.2.2 Bulk Concentrate Tests

Additional tests were completed within the 2004 test program in order to examine process options. In test F-29, a bulk concentrate was produced with correspondingly improved copper and nickel recoveries. For a combined Cu+Ni grade of 16.9%, prior to final cleaning, the respective recoveries were 92.8% and 76.3%. Summary results are shown in Table 5.2.4.

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Table 5.2.4 Bulk Concentrate Test F-29 (2004)

Grade Recovery % Test Blend Shakespeare

Product Cu % Ni % Cu+Ni% Cu Ni

Bulk cl. conc 11.60 8.60 20.20 89.8 72.5 Bulk ro. conc 9.64 7.27 16.91 92.8 76.3 F-29 Composite Feed 0.44 0.40

The conclusions from the 2004 testwork program were:

• Production of a nickel concentrate containing about 10% Ni from Shakespeare ore affects recovery and a lower grade concentrate may be a better target.

• Retreatment of the primary copper concentrate will improve nickel rejection and hence

overall nickel recovery.

• Predicted optimum metal recoveries for project economic evaluation, based on a copper concentrate of about 18% Cu and a nickel concentrate of about 7% Ni, are shown in Table 5.2.5.

Table 5.2.5

Predicted Optimum Metal Recoveries, 2004 Testwork

Copper 94% Nickel 75% Cobalt 71% Gold 34% Platinum 65% Palladium 40%

5.3 2005 TESTWORK

The 2005 metallurgical test program was also conducted at SGS. As a result of studies conducted by Micon on behalf of URSA, an on-site milling facility was selected as a more attractive option than processing via the Strathcona mill.

The objective of the 2005 testwork was to develop a flowsheet and process design criteria for an on-site beneficiation facility. The testing program was to based on production of a single concentrate with a target combined copper and nickel content of between 15 and 20%. and maximum recovery of nickel, copper and PGM’s.

The 2005 feasibility study testwork program comprised both flowsheet development and variability test programs, including grindability work, mineralogical investigations and bench scale flotation tests.

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5.3.1 Metallurgical Samples

The samples used in this test program were selected and prepared on-site by URSA personnel. The samples included a single metallurgical composite and five grindability samples. The metallurgical composite sample was made up from 200-m of quarter NQ core weighing approximately 240 kg, details of this sample are included in Table 5.3.1.

Table 5.3.1 Metallurgical Composite Sample (2005)

Interval (m) Assay Hole No. Location From To

Length (m) Ni (%) Cu (%)

Weight (kg)

U-03-55 11+00E / 1+16N 123.1 144.4 21.3 0.37 0.39 26 U-03-59 17+00E / 7+65N 337.2 384.9 47.7 0.37 0.43 57 U-03-60 21+00E / 8+74N 383.8 444.9 61.1 0.44 0.42 73 U-03-61 25+15E / 8+65N 413.3 463.6 50.3 0.39 0.39 60 U-03-63 21+00E / 10+49N 433.7 451.1 17.4 0.34 0.41 21

The composite metallurgical sample included both Blebby and Disseminated styles of mineralization.

Samples for grindability testing were selected to include a variety of typical mineralization that occurs at Shakespeare. These samples, measuring 5 to 15 cm (2 to 6 in) in diameter and weighing a minimum of 20 kg each, included the scattered and interconnected blebby styles of pyrrhotite (po) and chalcopyrite (cpy) mineralization in Shakespeare quartz gabbro, and disseminated styles of pyrrhotite and chalcopyrite mineralization in the Shakespeare rock fragment bearing and the massive equigranular melagabbro.

A description of the five grindability samples is presented in Table 5.3.2.

Table 5.3.2 Grindability Samples (2005)

Sample No.

Location UTM Coordinates

Description

98825 0435888E / 5133221N Blebby po and cpy in quartz gabbro (4c) type B1/B2 mineralization 98826 0435882E / 5133275N Disseminated po and cpy in melagabbro (4b) type D1 mineralization 98827 0435976E / 5133282N Disseminated and blebby po and cpy in melagabbro (4f), transitional

slightly altered mineralization types D1/D1S/B1 98828 0436041E / 5133326N Disseminated and blebby po and cpy in melagabbro (4f), transitional

strongly altered mineralization types D1/D1S/B1 98829 0436241E / 5133428N Disseminated po and cpy in melagabbro (4f/4b) slightly altered type

D1 mineralization

A series of SAG mill tests were conducted on a bulk sample consisting of 390 kg of material crushed to minus 150 mm. This bulk sample comprised near-surface mineralization selected from five locations on the property.

The program to determine the metallurgical variability was performed on a set of 19 samples. The first eight samples, identified as SVMLO-1 to SVMLO-8, were near-surface samples, the

62

remaining 11 samples, identified as SVMDC-1 to SVMDC-11, were split diamond drill core. Table 5.3.3 shows the location of these samples.

Table 5.3.3 Metallurgical Variability Samples (2005)

Location UTM Coordinates

Sample No.

UTM - N UTM - E

Description

SVMLO-1 5133221 0435888 Near surface sample, -150 mm Lump Ore SVMLO-2 5133275 0435882 Near surface sample, -150 mm Lump Ore SVMLO-3 5133282 0435976 Near surface sample, -150 mm Lump Ore SVMLO-4 5133326 0436041 Near surface sample, -150 mm Lump Ore SVMLO-5 5133428 0436241 Near surface sample, -150 mm Lump Ore SVMLO-6 5233221 0435888 Near surface sample, -150 mm Lump Ore SVMLO-7 5133377 0436109 Near surface sample, -150 mm Lump Ore SVMLO-8 5133600 0436538 Near surface sample, -150 mm Lump Ore Sample No. Hole No. From, m To, m Description

SVMDC-1 U-03-66 1.18 11.04 Diamond drill core SVMDC-2 U-03-66 11.04 20.9 Diamond drill core SVMDC-3 U-03-69

U-03-70 U-03-71

33.8 4.75 25.7

68 11.17 34.7

Diamond drill core

SVMDC-4 U-03-69 U-03-70 U-03-71

68 11.17 34.7

102.2 17.6 43.7

Diamond drill core

SVMDC-5 U-03-74 U-03-75 U-03-76

90.5 89.4

180.2

103.65 117.5

185.35

Diamond drill core

SVMDC-6 U-03-74 U-03-75 U-03-76

103.65 117.5

185.35

116.8 145.6 190.5

Diamond drill core

SVMDC-7 U-03-82 U-03-79

90 62.2

127.1 68.55

Diamond drill core

SVMDC-8 U-03-82 U-03-79

127.1 68.55

164.2 74.9

Diamond drill core

SVMDC-9 U-03-83 49.5 90.3 Diamond drill core SVMDC-10 U-03-36

U-03-36 U-03-36 U-03-36 U-03-36 U-03-40

76.46 81.3

85.25 89.25 93.25 49.69

78.46 82.3

88.25 92.25 97.25 77.55

Diamond drill core

SVMDC-11 U-03-29 U-03-33 U-03-52

137.94 226.25

207

145.58 242.3

221

Diamond drill core

5.3.2 Characterization

The principal economic element assays for the metallurgical composite sample are reported in Table 5.3.4 and the minor elements in Table 5.3.5.

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Table 5.3.4 Metallurgical Composite (2005) Feed Main Element Analysis

Cu (%) Ni (%) S (%) Fe (%) Co (%) Pt (g/t) Pd (g/t) Au (g/t) 0.42 0.37 2.45 11.79 0.02 0.34 0.41 0.21

Table 5.3.5

Metallurgical Composite (2005) Feed Major and Minor Element Analysis

Element Units Value Element Units Value Element Units Value SiO2 % 49.4 Ag g/t < 3 Mn g/t 1,100 Al2 O3 % 13.3 Al g/t 66,000 Mo g/t < 8 Fe2 O3 % 17.5 As g/t 130 Na g/t 12,000 MgO % 7.14 Ba g/t 170 Ni g/t 3,500 CaO % 6.89 Be g/t < 0.3 P g/t 240 Na2O % 1.69 Bi g/t < 20 Pb g/t < 25 K2O % 0.77 Ca g/t 43,000 Sb g/t < 10 TiO2 % 0.73 Cd g/t < 4 Se g/t < 30 P2O5 % 0.07 Co g/t 240 Sn g/t < 20 MnO % 0.17 Cr g/t 170 Sr g/t 130 Cr2O3 % 0.04 Cu g/t 3,800 Ti g/t 3,700 V2O5 % 0.05 Fe g/t 120,000 Tl g/t < 30 LOI % 2.82 K g/t 7,800 V g/t 220 Li g/t < 8 Y g/t 11 Mg g/t 42,000 Zn g/t 100

Table 5.3.6 presents the results of a qualitative x-ray diffraction (XRD) analysis performed on a sample of the metallurgical composite. The relative proportions of crystalline mineral assemblages shown in Table 5.3.6 are based on relative peak heights. The results do not reflect the presence of non-crystalline/amorphous compounds.

Table 5.3.6 Metallurgical Composite (2005) Summary of Qualitative X-ray Diffraction Analysis

Major Moderate Minor Trace Chlorite Actinolite Calcite Rhodochrosite Quartz Mica Pyrite Montmorillonite Plagioclase feldspar Mordenite Magnetite Pyrrhotite

A rapid mineral scan data report on the metallurgical composite, provided by SGS, presented semi-qualitative indicative information on ore mineralogy. A summary of these results is included in Table 5.3.7. These results, although not fully quantitative, indicate fairly good liberation of sulphide minerals at a grind size of 80% passing 75 µm and no significant evidence of problematic pentlandite-pyrrhotite flames.

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The results from mineralogical testwork completed during the 2005 pre-fesibility testwork program tended to support previous work, which indicated that chalcopyrite and pentlandite have a relatively low level of mineral locking and that approximately 10% of the nickel is in solid solution in the pyrrhotite.

Table 5.3.7 Summary of Results from SGS Rapid Mineral Scan

Liberation, % Mineral Abundance, % Liberated Attached Locked

Arsenopyrite 0 100 0 0 Pyrite 6 0 0 0 Chalcopyrite 12 74 10 16 Pyrrhotite 80 82 7 11 Pentlandite 3 621 0 0 1 The pentlandite liberation is provided from previous work.

5.3.3 Grindability Testwork

The abrasion index of the metallurgical composite, determined by using the standard Bond abrasion test procedure, was 0.2909. This material would be considered moderately abrasive.

The metric Bond rod mill index of the metallurgical composite, determined by SGS, was 16.9 kWh/t. The Bond ball mill work indices for the various grindability samples and the metallurgical composite are listed in Table 5.3.8. It is noted that the two samples of disseminated material, 98826 and 98829, have relatively high work indices.

Table 5.3.8 Bond Ball Mill Work Indices

80% Passing Size (µm) Sample

BWI –metric (kWh/t) Feed (F80) Product (P80)

Metallurgical composite 13.5 2,170 119 Sample No. 98825 14.9 2,200 120 Sample No. 98826 15.7 2,161 122 Sample No. 98827 13.5 2,234 116 Sample No. 98828 11.0 2,125 120 Sample No. 98829 15.9 2,307 119

As a comparison, the metric Bond ball and rod indices for the “Shakespeare Composite” sample, determined during the 2004 testwork program, were 12.6 and 15.9 kWh/t, respectively.

The five grindability samples listed in Table 5.3.9 were submitted to MinnovEX Technologies Inc. (MinnovEX) for a preliminary design of the Shakespeare grinding circuit based on CEET® (Comminution Economic Evaluation Tool) technology. The goal was to achieve an average tonnage throughput of 204 t/h at an average P80 of 80 µm in an advanced grinding design and performance predicting tool that uses the SPI® energy relationship and Bond’s third theory of comminution to model the energy performance of SAG/ball mill circuits.

65

The samples submitted to MinnovEX were too fine for a full SPI® test and, therefore, the SPI®-Lite test was used with an accuracy estimated at +/-15%. The circuit studied contains a SAG mill (with crusher) and a ball mill. A summary of results is presented in Table 5.3.9.

Table 5.3.9 Summary of MinnovEX SPI Testwork Results

Sample No. CI SPI WI F80 F50 PCCL T80 P80 t/h

kWh/t (SAG)

kWh/t (Ball)

98825 10 165 14.9 150 68 28 2240 85 205 8.75 11.54 98826 10 152 15.7 150 68 27 2293 85 194 8.40 12.19 98827 10 158 13.5 150 68 28 2267 75 210 8.56 11.27 98828 10 116 11.0 150 68 22 2465 70 224 7.33 10.54 98829 10 187 15.9 150 68 31 2170 85 193 9.30 12.26

CI = Crusher Index, SPI = SAG Power Index, W I= Bond Ball Work Index, PCCL = % Pebble Circulating load, F = Feed Size (mm), T = Transition Size (µm ), P =Mill Circuit product Size (µm).

A MacPherson test was performed by SGS using the 390-kg bulk SAG mill sample. Table 5.3.10 presents the results from the MacPherson test. The test resulted in a MacPherson Autogenous Work Index (Ai) of 15 kWh/t. The Bond rod and ball work index test results for this bulk sample were 16.3 and 13.8 kWh/t, respectively.

Table 5.3.10 MacPherson Testwork Results

Feed (kg/h) Percentile

of Hardness F80

(µm) P80

(µm) Gross Work

Index (kWh/t)

Correlated Work Index

(kWh/t)

Gross Specific Energy Input

(kWh/t) 6.7 81 22,225 174 18.0 15.0 12.4

The 19 variability samples were subjected to standard Bond ball mill tests and JK Drop-Weight Tests. The latter is generally used to confirm the breakage response of an ore and to evaluate ore variability. The test determines energy vs. breakage distribution functions. Table 5.3.11 presents the results from the variability grinding tests.

Details of this testwork are provided in Appendix 8.

5.3.4 Flotation Testwork

The objective of the feasibility study flotation testwork program was the development of a robust, simple mineral processing scheme, employing principally flotation, for the production of a single high-grade copper-nickel bulk concentrate, containing approximately 20% combined Cu+Ni.

The objective of the rougher flotation development program was to exploit the differential flotation kinetics between the PGM/Cu/Ni-bearing sulphides and the pyrrhotite.

66

Table 5.3.11 Summary of Variability Grinding Testwork Results

Drop Weight Test Parameters

Sample Reference

Lithology

Ore Density (g/m3)

A b A x b ta DWI kWh/t (Ball)

SAG composite SS 3.05 100 0.22 22.0 0.23 12.4 13.8 SAG composite SS 3.02 100 0.23 23.0 - 11.7 SVMLO-1 SS 3.10 84.3 0.34 28.7 - 9.5 14.5 SVMLO-2 SS 3.06 81.8 0.34 27.8 - 9.9 15.8 SVMLO-3 SS 3.01 60.6 0.42 25.5 - 10.5 13.3 SVMLO-4 SS 2.97 70.9 0.37 26.2 - 10.2 12.3 SVMLO-5 SS 3.00 80.8 0.28 22.6 - 11.7 14.9 SVMLO-6 SS 3.22 72.9 0.47 34.3 - 8.5 14.4 SVMLO-7 SS 3.00 50.9 0.60 30.5 - 8.8 13.8 SVMLO-8 SS 3.00 74.9 0.33 24.7 - 10.8 14.1 SVMDC-1 DissemS 3.00 74.9 0.33 24.7 - 10.8 14.5 SVMDC-2 Dissem 3.01 66.5 0.37 24.6 - 11.0 13.1 SVMDC-3 Blebby 2.99 66.6 0.37 24.6 - 10.8 13.0 SVMDC-4 Dissem 3.04 59.8 0.40 23.9 - 11.5 12.5 SVMDC-5 Blebby 3.03 52.4 0.52 27.2 - 9.9 12.7 SVMDC-6 Dissem 3.02 74.7 0.32 23.9 - 11.2 13.0 SVMDC-7 Blebby 3.03 76.1 0.28 21.3 - 12.9 11.9 SVMDC-8 Dissem 2.97 63.2 0.41 25.9 - 10.1 13.9 SVMDC-9 Blebby 3.04 100 0.21 21.0 - 13.1 12.2 SVMDC-10 Dissem2 3.03 82.4 0.28 23.1 - 11.8 13.4 SVMDC-11 Dissem3 3.00 100 0.20 20.0 - 13.6 13.0 Disseminated (average) 3.01 66.1 0.38 24.6 - 11.0 13.1 Blebby (average) 3.02 73.8 0.35 23.5 - 11.7 12.5 Average (Total) 3.03 75.9 0.35 25.0 0.23 11.0 13.5 Std. Dev. 0.05 15.1 0.10 3.4 - 1.3 1.0 Relative Std. Dev. 1.8 19.9 29.4 13.4 - 12.3 7.4 Minimum 2.97 50.9 0.20 20.0 - 8.5 11.9 10rd Percentile 2.99 59.8 0.22 21.3 - 9.5 12.3 25rd Percentile 3.00 66.5 0.28 23.0 - 10.1 12.9 Median 3.02 74.9 0.34 24.6 - 10.8 13.3 75th Percentile 3.04 82.4 0.40 26.2 - 11.7 14.2 90th Percentile 3.06 100 0.47 28.7 - 12.9 14.6 Maximum 3.22 100 0.60 34.3 - 13.6 15.8 Lithology Reference:

SS = Surface Sample Blebby = exhibit predominantly the blebby styles of po and cpy mineralization Dissem = exhibit predominantly the disseminated style of po and cpy mineralization DissemS = represent the typical disseminated style of po and cpy mineralization in the upper part of the Shakespeare project Dissem2 = exhibit pred. the disseminated styles of po and cpy with only the occasional scattered blebs of po and cpy mineralization. Dissem3 = exhibit pred. the diss. styles of po and cpy, the minor localized occurrences of the blebby styles of po and cpy mineral.

67

The concept was to produce a primary fast-floating concentrate, which was expected to contain minimal pyrrhotite, and would be cleaned by silicate rejection. This would then be followed by the production of a secondary slow-floating concentrate, which was expected to contain substantially more pyrrhotite. This concentrate would then be cleaned by establishing suitable pyrrhotite rejection chemistry.

The main phases of the bench scale flotation program included investigations into:

• Rougher flotation kinetics. • Pyrrhotite rejection in the rougher circuit. • Effect of primary grind fineness. • Two stage rougher circuit. • Regrinding. • Cleaner flotation kinetics.

5.3.4.1 Flotation Testwork Results

Table 5.3.12 summarizes the results from the first stage of rougher flotation testwork. Tests F1 to F6 looked at rougher kinetics, tests F7 to F9 investigated pyrrhotite rejection and tests F11 to F13 considered different primary grind sizes. Tests F11 to F13 used primary grinds of 80% passing (P80) 157, 104 and 60 µm, respectively, all other tests used a primary grind (P80) size of 80 µm. All the results are based on a 15-min concentrate recovery.

Table 5.3.12 Rougher Flotation Test Results (Stage 1)

Conc Grade (%) Recovery (%) Test No.

Weight (%) Cu Ni Cu+Ni Cu Ni Pt Pd

F1 9.62 3.97 3.23 7.19 97.01 85.76 82.60 50.77 F2 8.46 4.27 3.64 7.90 96.10 85.50 80.14 52.28 F3 8.94 4.27 3.41 7.68 96.77 84.83 76.29 50.64 F4 9.60 4.00 3.28 7.28 97.03 87.44 84.82 56.38 F5 9.74 3.92 3.24 7.16 92.97 86.18 89.69 58.65 F6 10.59 3.68 3.00 6.68 93.96 86.36 90.20 60.67 F7 17.34 2.27 1.91 4.18 97.35 86.98 90.43 62.34 F8 12.66 3.14 2.51 5.65 96.81 84.85 76.83 54.14 F9 9.08 4.60 3.48 8.08 96.23 82.26 63.08 47.11 F11 7.26 4.86 3.57 8.44 82.64 71.78 64.91 36.16 F12 8.91 4.03 3.13 7.15 83.11 75.36 76.53 41.65 F13 9.30 4.06 3.19 7.25 97.42 82.35 69.06 48.59

A number of rougher flotation tests investigating the two stage rougher circuit were performed during the testwork program. These results are summarized in Table 5.3.13. Cleaner kinetic flotation tests have been completed but results were pending as of January, 2006.

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Table 5.3.13 Rougher Flotation Test Results (Stage 2)

Primary Rougher Concentrate Secondary Rougher Concentrate Recovery Concentrate Grade Recovery Concentrate Grade

Test Number (%Cu) (%Ni) (%Cu) (%Ni) (%Cu) (%Ni)

(%Cu) (%Ni)

F14 91.55 51.06 10.59 5.25 4.86 29.99 0.32 1.74 F15 92.33 55.64 10.52 6.00 4.29 27.04 0.23 1.40 F16 94.10 61.90 8.59 5.16 2.19 20.83 0.14 1.24 F17 92.69 63.49 8.39 5.11 3.19 19.19 0.20 1.09 F18 93.54 59.59 8.65 4.97 3.44 23.17 0.18 1.07 F20 96.75 84.62 3.33 2.58 2.91 14.36 0.17 0.74 Primary Rougher Cleaner Concentrate Secondary Rougher Cleaner Concentrate F14 86.26 22.19 21.85 4.99 3.22 10.93 3.32 10.00 F15 85.10 9.94 25.14 2.78 2.41 6.76 2.98 7.91 F16 83.17 6.19 28.40 1.93 1.56 7.13 1.67 6.98 F17 88.83 28.84 22.40 6.47 1.86 5.75 2.14 5.89 F18 90.02 16.51 21.90 3.62 1.88 4.30 3.25 6.68 F20 86.16 13.63 26.60 3.72 1.80 5.09 1.10 2.76

5.3.4.2 Discussion of the Initial Flotation Test Results

These rougher tests indicated that relatively good copper recoveries can be expected, irrespective of the details of the final flowsheet selected. Figure 5.2 plots copper recovery versus pyrrhotite (Po) recovery for these initial rougher tests and some data from the 2004 testwork program.

Figure 5.2 Copper to Pyrrhotite Flotation Selectivity

Copper/Po Selectivity

0

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60 70 80 90 100

Copper recovery

Po r

ecov

ery

Red: 2004study

69

As historical work confirmed, nickel is less responsive to flotation, relative to copper. Approximately 8% of the nickel is estimated to be present in non-sulphide form and since this is considered to be unrecoverable into any form of sulphide concentrate the ultimate recovery limit is around 92%. Of this 92%, some nickel is in solid solution in pyrrhotite. This is estimated to be about 4.5% of the total nickel in the composite sample, leaving approximately 87.5% of the remaining nickel in the sample as either as fine pentlandite (Pn) inclusions in pyrrhotite, or as discrete pentlandite particles.

Figure 5.3 presents the pyrrhotite (Po) verses nickel recovery results from the preliminary rougher testwork performed in 2005. The blue (thin) line indicates that nickel is recovered as discrete pentlandite mineralization up to about 75% recovery. Higher recovery than 75% appears to be closely associated with the recovery of pyrrhotite (dotted line). Extrapolating the pyrrhotite recovery to 0% recovery gives us an indication of the nickel content in pyrrhotite, which approximates to17% at the 80 µm grind. Of this 17%, about 4.5% is solid solution nickel while the remainder (12.5%) is as micro-inclusions in pyrrhotite.

Mineralogy determined that Pn is mainly liberated in the primary rougher concentrate, while the Pn in the secondary rougher concentrate (between the arrows) was 50% liberated, and 50% locked with pyrrhotite. Both these observations fit well with the above model.

Figure 5.3 Nickel to Pyrrhotite Flotation Selectivity

Nickel/Po Selectivity

0

10

20

30

40

50

60

70

80

90

100

0 20 40 60 80 100Nickel recovery

Po re

cove

ry

pre-F20F26F25

70

5.3.5 Locked Cycle Tests

A series of four locked cycle tests targeting a range of concentrate qualities were performed by SGS using the 2005 metallurgical composite sample. The test conditions for these tests are summarized in Table 5.3.14.

Table 5.3.14 Summary of Locked Cycle Flotation Test Conditions

Units F38 F40 F43 F47 Primary grind (k80) µm 82 82 82 82 Rougher time min 16 16 16 16 Cleaner 1 time min 5.5 5 4.5 5.5 Cleaner 2 time min 3.5 3 2.5 2.5 Cleaner 3 time min 2 2 2 2 Cleaner 4 time min 2 2 1.5 1.5 Ca(OH)2 g/t 290 315 390 315 SIPX reagent g/t 30.5 25 25 30 Aero 3477 reagent g/t 20 20 20 22.5 CMC reagent g/t 20 20 20 15 MIBC reagent g/t 35 32.5 32.5 31.5 Float feed pH - 9.5 9.5 9.5 9.5

and Figure 5.4 present the results from the locked cycle tests. SGS concluded that tests F43 was unstable and the results were not indicative of probable plant operation. Results of test F47 were used for metallurgical predictions in the plant design and economic analysis. The regression curves shown in Figure 5.4 have the following equations and regression coefficients:

%Cu recovery = -0.0028x2 + 0.0032x + 96.704 R2=0.8473

%Ni recovery = -0.0665x2 + 1.069x + 78.684 R2=0.9998

%Pd recovery = -0.1064x2 + 1.4267x + 83.544 R2= 0.999

%Pt recovery = -0.1884x2 + 4.0202x + 31.106 R2 = 0.9964

%Au recovery = -0.1348x2 + 2.6519x + 34.309 R2 = 1.0

A full element scan of a final concentrate produced during locked cycle testing is presented in Table 5.3.16.

71

Table 5.3.15 Locked Cycle Flotation Test Results

Concentrate Grade (%) Recovery (%) Test No. Wt% Cu Ni Cu+Ni S Cu Ni S

F38 5.78 6.51 5.10 11.62 27.80 96.33 82.12 72.69 F40 3.60 10.08 7.22 17.30 25.65 96.15 77.38 46.88 F43 2.37 15.21 8.24 23.45 27.17 95.28 67.22 29.32 F47 3.45 11.67 7.59 19.27 29.38 95.47 74.49 46.59

Concentrate Grade (g/t) Recovery (%) Test No. Wt% Pt Pd Au - Pt Pd Au

F38 5.78 4.42 3.45 1.81 - 85.71 52.26 46.92 F40 3.60 6.05 4.07 2.35 - 76.82 45.17 39.87 F43 2.37 6.88 3.20 2.32 - 58.62 21.99 22.36 F47 3.45 7.10 4.11 2.08 - 71.10 37.65 35.32

Figure 5.4 Locked Cycle Flotation Test Results

5.3.6 Variability testwork

The head grades of the metallurgical variability samples are shown in Figure 5.5 and Table 5.3.17.

The variation in ore hardness and grindability between the variability samples are presented in Table 5.3.11, above. The Bond ball mill work index for the 19 variability samples ranged between 11.9 and 15.8 kWh/t. The average work indices for disseminated and blebby type ores was 13.1 and 12.5 kWh/t, respectively.

0

10

20

30

40

50

60

70

80

90

100

10.00 12.00 14.00 16.00 18.00 20.00 22.00 24.00 26.00

Ni + Cu Grade (%)

Met

al R

ecov

ery

(%)

CuNiPtPdAu

72

Figure 5.5 Variability Samples Comparison of Nickel and Copper Feed Analyses

-

0.10

0.20

0.30

0.40

0.50

0.60

0.70

0.80

SVMLO-4

SVMDC-11

SVMLO-7

SVMLO-3

SVMDC-6

Met Compo

site

SVMLO-5

SVMDC-9

SVMLO-8

SVMDC-4

SVMDC-5

SVMDC-8

SVMDC-3

SVMLO-1

SVMDC-1

SVMDC-10

SVMDC-7

SVMDC-2

SVMLO-2

SVMLO-6

Perc

ent

NiCu

Table 5.3.16

Full Element Analysis of Shakespeare Flotation Concentrate

Element Units Value Element Units Value Solution XRF ICP

Cu % 10.9 P g/t <200 Fe % 29.1 Bi g/t <200 Ni % 7.4 Cd g/t <20 Pb % 0.01 Co % 0.41 Mo g/t 138 Al g/t 10,389 Zn g/t 331 Ca g/t 7,993 As % 0.1 Cr g/t 152 Sb % 0.003 Mg g/t 4,138 U % <0.002 Mn g/t 126

Fire Assay Si g/t 37.6 Pt g/t 6.7 Ti g/t 552 Pd g/t 3.7 V g/t 71.3 Au g/t 1.7 Na g/t 2,207

Carbon/Sulphur K g/t 1,079 C(t) % 0.3 ICP-MS S % 25.3 Ga g/t 4.7

Powder XRF Ge g/t 2.5 Cl g/t 50.4 In g/t 0.8 F % <0.005 Te g/t 82

Cold Vapour AA Tl g/t 19.4 Hg g/t 0.6

73

Table 5.3.17

Variability Samples Head Grades

No Drill Ref. Pt (g/t) Pd (g/t) Cu (%) Ni (%) S (%) V1 SVMLO-1 0.32 0.28 0.44 0.45 4.90 V2 SVMLO-2 0.56 0.68 0.69 0.57 3.19 V3 SVMLO-3 0.30 0.39 0.32 0.32 1.67 V4 SVMLO-4 0.38 0.46 0.37 0.27 1.67 V5 SVMLO-5 0.35 0.43 0.43 0.39 2.22 V6 SVMLO-6 0.32 0.25 0.68 0.58 7.55 V7 SVMLO-7 0.40 0.48 0.49 0.31 1.80 V8 SVMLO-8 0.37 0.44 0.45 0.42 2.28 V9 SVMDC-1 0.46 0.57 0.51 0.45 2.40 V10 SVMDC-2 0.51 0.60 0.60 0.51 2.72 V11 SVMDC-3 0.38 0.48 0.52 0.44 2.47 V12 SVMDC-4 0.43 0.51 0.47 0.43 2.46 V13 SVMDC-5 0.42 0.47 0.50 0.43 2.74 V14 SVMDC-6 0.32 0.41 0.41 0.34 1.97 V15 SVMDC-7 0.49 0.62 0.49 0.48 3.10 V16 SVMDC-8 0.38 0.52 0.55 0.43 2.51 V17 SVMDC-9 0.43 0.52 0.44 0.40 2.09 V18 SVMDC-10 0.45 0.57 0.56 0.47 2.58 V19 SVMDC-11 0.28 0.30 0.32 0.28 1.51

The variability flotation testing was undertaken by SGS in October and November, 2005. This work comprised 38 batch cleaner flotation tests using 2-kg feed samples and was supported by mineralogical studies using QEMSCANTM.

Variability flotation tests were conducted on all 19 variability samples, eight surface lump ore samples (V1 to V8) and 11 drill core samples (V9 to V19). It was noted that the results obtained for the oxidized, surface material were poor and therefore, as they are not truly representative of the future mill feed, these results are not reported here. The drill core samples were representative of typical fresh mineralization.

The standard flotation test protocol followed for the batch variability flotation work was based on the locked cycle tests, which corresponds to the final process flowsheet adapted for the feasibility study. While a consistent grind time was used in preparing the flotation feed samples, there was some variability is product size between the different samples. This is shown in the table of results, Table 5.3.18, which presents the estimated metal recoveries for a target concentrate grade of 18% combined copper plus nickel.

The flotation test results showing the individual metal recoveries against Ni+Cu% in the concentrate are presented in Figure 5.6.

74

Table 5.3.18 Summary of Variability Flotation Test Results

Sample Feed k80 (µm)

Ni Recovery (%)

Cu Recovery (%)

Pt Recovery (%)

Pd Recovery

(%) V9 81 63.33 90.85 47.94 22.91 V10 88 62.70 84.95 71.34 26.64 V11 79 59.30 89.79 65.11 25.78 V12 73 64.75 89.19 59.98 26.28 V13 75 69.03 93.35 72.26 37.11 V14 84 71.11 91.98 66.19 32.37 V15 70 64.06 93.05 67.51 24.06 V16 84 72.60 93.88 66.03 38.69 V17 73 73.35 94.09 71.64 46.95 V18 82 69.15 89.47 61.73 35.81 V19 78 62.25 93.61 61.37 38.20

75

Figure 5.6 Variability Flotation Test Metal Recoveries vs Concentrate Grade

Variability Tests - Ni Recovery

0102030405060708090

100

0.0 5.0 10.0 15.0 20.0 25.0 30.0

Ni +Cu Concentrate Grade (%)

Ni R

ecov

ery

(%)

V9V10V11V12V13V14V15V16V17V18V19

Variability Tests - Cu Recovery

80828486889092949698

100

0.0 5.0 10.0 15.0 20.0 25.0 30.0


Cu

Rec

over

y (%

)

V9V10V11V12V13V14V15V16V17V18V19

Variability Tests - Pt Recovery

0102030405060708090

100

0.0 5.0 10.0 15.0 20.0 25.0 30.0


Pt R

ecov

ery

(%)

V9V10V11V12V13V14V15V16V17V18V19

Variability Tests - Pd Recovery

0

10

20

30

40

50

60

70

0.0 5.0 10.0 15.0 20.0 25.0 30.0


Pd R

ecov

ery

(%)

V9V10V11V12V13V14V15V16V17V18V19

76

The data, as plotted in Figure 5.7, show that there is no correlation between head grade and recovery for any of the principal metals, based on the samples tested in the variability program.

Figure 5.7 Variability Tests – Head Grade versus Recovery

0

10

20

30

40

50

60

70

80

90

100

0.20 0.25 0.30 0.35 0.40 0.45 0.50 0.55 0.60 0.65

Variability Sample Head Grade (% or g/t)

Rec

over

y (%

)

NiCuPtPd

The recoveries from the batch variability tests are generally 3 to 10% lower than for locked cycle tests conducted using the metallurgical composite. This is not unexpected due to the lack of supplemental recovery gleaned by re-processing the middlings streams, as per locked cycle testing (eee Appendix 10).

5.4 METALLURGICAL RECOVERY ESTIMATES

The recovery estimates for a range of concentrate product qualities, based on the regression curves shown in Figure 5.4, are shown in Table 5.4.1. These estimated results are based on the metallurgical composite sample and do not take into account variations in head grade. Note that variability tests indicated that there is no correlation between head grade and recovery.

Table 5.4.1 Feasibility Study Recovery Estimates

Product 15% (Cu+Ni)

Product 18% (Cu+Ni)

Product 20% (Cu+Ni)

Element Feed Grade (% or g/t)

Recovery(%)

Grade (% or g/t)

Recovery(%)

Grade (% or g/t)

Recovery (%)

Grade (% or g/t)

Copper (Cu) 0.42 96.1 8.7 95.9 10.6 95.6 11.9 Nickel (Ni) 0.37 79.8 6.3 76.4 7.4 73.5 8.1 Platinum (Pt) 0.34 81.0 5.9 74.8 6.7 69.5 7.0 Palladium (Pd) 0.41 49.0 4.3 42.4 4.6 36.2 4.4 Gold (Au) 0.21 43.8 2.0 38.4 2.1 33.4 2.1

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5.5 PROCESS SELECTION

The process selected for the feasibility study is based on the interpretation of the results from the historic and 2005 metallurgical testwork programs.

The basic process selected comprises primary crushing, SAG milling with pebble crushing, ball milling, flotation to produce a single bulk concentrate, concentrate dewatering and tailings disposal.

5.5.1 Grinding Mill Sizing

A number of methods were used to estimate the grinding circuit requirements. Micon used the Bond ball and rod mill indices and applied typical industry scale-up factors to calculate the SAG mill size. In addition, the MacPherson Autogenous work index, the MinnovEX SAG Power Index (SPI) and JK SimMet simulation using JK Drop-Weight test breakage parameters, were also considered. A comparison is presented in Table 5.5.1.

Table 5.5.1 Comparison of Mill Circuit Selections

Mill Parameter Micon Minnovex JK SimMet F80 (mm) 150 150 150 P80 (mm) 0.75 2.287 0.735 Diameter (ft) 24 21 26 Length (ft) 10 10.5 9 HP1 3000 2526 3681 Pebble diameter (mm) 15 16 6.3 Pebble CL 45% 27% 37% kWh/t 9.60 8.80 13.46

SAG Mill

Pebble crusher Yes Yes Yes F80 (mm) 850 2287 735 P80 (mm) 80 80 80 Dia (ft) 15 15 16 L (ft) 19 27 22 HP1 3000 3323 2420

Ball Mill

kWh/t 9.50 11.578 8.85 Total Total kWh/t 19.10 20.38 22.31 Total HP 6000 5849 6101

1 The HP for Micon selection is motor selected. For Minnovex and JK it represents gross power.

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6.0 PROCESS PLANT

The process plant flowsheet and design criteria are based mainly on the results from the metallurgical testwork program conducted in 2005, which is discussed in Section 5.3 of this feasibility study. When certain specific results from the 2005 test campaign were unavailable, results from the 2003 and 2004 testwork program have been used in the design.

6.1 PROCESS DESIGN CRITERIA

The process design basis and design criteria are summarized in Table 6.1.1 and Table 6.1.2, respectively. Complete process design criteria used to establish mass balances and preliminary equipment sizing are included in Appendix 3.

6.1.1 Design Criteria

Table 6.1.1 summarizes the process design basis. The operation is designed to treat 4,500 dry t/d of nickel-bearing Shakespeare ore from the open pit mine on a 24-h/d, 7-d/w basis. The utilization factors used for the calculation of the nominal hourly flow rates are 60% for the primary crusher and 92% for the remainder of the process facilities. A summary of the key criteria used as the basis for the design of the unit operations is shown in Table 6.1.2.

Table 6.1.1 Process Design Basis

Parameter Units Value Source Operating time d/y 365 Micon Operating time h/d 24 Micon Primary crusher operating criteria d/w 6 Micon Primary crusher utilization % 60 Micon Leach and plant operating criteria d/w 7 Micon Leach and plant utilization % 92 Micon Throughput Nominal annual throughput thousand t 1,643 Micon Design daily throughput t 4,500 Micon Run-of-Mine Ore Characteristics Maximum rock size mm 1,250 Micon Ore specific gravity 3.1 Micon Ore moisture wt % 3.0 Micon Average feed grade – Ni % 0.34 Indicated resource Average feed grade – Cu % 0.36 Indicated resource Average feed grade – Co % 0.02 Indicated resource Average feed grade – Pt g/t 0.34 Indicated resource Average feed grade – Pd g/t 0.38 Indicated resource Average feed grade – Au g/t 0.19 Indicated resource Metallurgical Efficiency (plant design only) Total Ni plus Cu concentrate grade % 20 Micon Total nickel recovery % 73.5 Testwork Total copper recovery % 95.6 Testwork Concentration ratio 34 Calculation Final concentrate grade – Ni % 8.4 Calculation Final concentrate grade – Cu % 11.6 Calculation Final concentrate production dry t/d 134 Calculation Total Ni plus Cu concentrate grade % 15 Micon Total nickel recovery % 79.8 Testwork Total copper recovery % 96.1 Testwork Concentration ratio 24 Calculation Final concentrate grade – Ni % 6.3 Calculation Final concentrate grade – Cu % 8.7 Calculation Final concentrate production dry t/d 194 Calculation

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Table 6.1.2

Summary of Key Process Design Criteria

Parameter Units Value Source Primary Crushing Design throughput t/h 313 Calculation Crusher utilization % 60 Micon Grinding Nominal throughput t/h 204 Calculation Circuit utilization % 92 Micon SAG mill circuit product size (80% passing) µm 850 Testwork/Micon SAG mill estimated unit power consumption kWh/t 8.60 Micon/Calculation Grinding circuit product size (80% passing) µm 85 Testwork/Micon Ball mill estimated unit power consumption kWh/t 9.52 Micon/Calculation Flotation Equipment utilization % 92 Micon Primary rougher retention time min 5 Testwork Secondary rougher retention time min 35 Testwork First cleaner retention time min 14 Testwork Second cleaner retention time min 9 Testwork Third cleaner retention time min 5 Testwork Fourth cleaner retention time min 5 Testwork Scavenger concentrate regrind mill feed rate dry t/h Calculation Regrind mill circuit product size (80% passing) µm Testwork/Micon Regrind mill estimated unit power consumption kWh/t Micon/Calculation Tailings sulphide rejection min 12 Testwork Concentrate dewatering Equipment utilisation % 92 Micon Concentrate thickener underflow density wt% solids 65 Micon Filter cake water content wt% water 7 Micon Tailings (design and mass balance purposes) Nominal production rate dry t/h 175 Calculation Nominal solids feed density wt% solids 32 Calculation

6.2 PROCESS DESCRIPTION

The process selected for the beneficiation of the Shakespeare ore is based on design criteria summarized in Section 6.1. The process as described below is presented in the flowsheets listed in Table 6.2.1, which are included at the end of this section.

Table 6.2.1 List of Flowsheet Drawings

Number Rev. Description 25047-0201-F E Crushing Area Flowsheet 25047-0202-F E Grinding Area Flowsheet 25047-0203-F E Flotation Area Flowsheet 25047-0204-F E Concentrate Dewatering Area Flowsheet 25047-0205-F D Tailings Disposal and Effluent Treatment Flowsheet 25047-0206-F E Make-Up Water and Services Flowsheet 25047-0207-F E Reagents Sheet 1 of 2 Flowsheet 25047-0208-F E Reagents Sheet 2 of 2 Flowsheet 25047-0209-F E On-Line Analyser System Flowsheet

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6.2.1 Crushing and Storage Facility.

Refer to Flowsheet 25047-0201-F

The blasted ore from the open pit is hauled by mine trucks to the primary crusher located about 1,700 m from the middle point of the West and East pits. The 91-t capacity trucks dump the ore into a 135-t hopper from which an apron feeder is used to feed the primary 36-in by 48-in (0.91 by 1.22m) jaw crusher. The crusher, which is equipped with a dust collector, is covered with a winterized building. An operator’s cabin is located at close proximity to the crusher.

The ore is crushed to minus 150 mm and transported on a short transverse 1.22-m wide conveyor belt to a 0.91-m wide stacking conveyor, which discharges onto a 4,500-t capacity stockpile.

Two reciprocating feeders located in a tunnel at the base of the stockpile are used to reclaim the crushed ore from the stockpile. The ore is then transported by a 0.91-m wide conveyor to the SAG mill, which is situated in the main mill building. A belt scale and standby feed hopper are installed on the SAG mill feed conveyor.

6.2.2 Grinding


The grinding circuit comprises a 24 ft (7.31 m) high by 10 ft (3.05 m) long SAG mill, and a 15.5 ft (4.72 m) high by 21 ft (6.60 m) long ball mill. The SAG and ball mills are fitted with 4,000-HP (2,985-kW) and 3,000-HP (2,220-kW) motors, respectively.

The product from the SAG mill discharges onto a 1.22-m by 2.44-m double deck vibrating screen. Material larger than 15 mm is conveyed to a 5½-ft (1.68-m) diameter short head cone pebble crusher. The product from the crusher is recycled back to the SAG mill. The undersize product from the vibrating screen discharges into the mill discharge pump box where it is joined by the ball mill discharge stream and dilution process water.

The mill discharge products are pumped to a cyclone cluster, from which the overflow sizing 80% passing 85 µm, is routed to the flotation circuit. The cyclone cluster underflow feeds the ball mill.

6.2.3 Flotation Circuit


The flotation circuit comprises two conditioners in series, a primary rougher stage (2 x 20-m3 tank cells), a secondary rougher stage (6 x 53-m3 tank cells), four cleaner stages (using conventional cells, in banks), and a tailings sulphide rejection stage (2 x 53-m3 tank cells). The cleaner circuit consists of 5 x 5-m3 first cleaner cells, 2 x 5-m3 second cleaner cells, 3 x 1.6-m3 third cleaner cells and 2 x 1.6-m3 fourth cleaner cells.

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The secondary rougher concentrate and is fed to the regrind mill circuit. The discharge from the 6.5-ft (1.98-m) high by 10-ft (3.05-m) long regrind mill, with 150-kW motor, is pumped to a cyclone cluster. The cyclone underflow returns to the mill while the overflow stream, sizing 80% passing 40 µm, feeds the first cleaner cells.

The tailings product from the scavenger cells is combined with the first cleaner tailings and pumped to the tailings sulphide rejection flotation cells, from which the separate concentrate and tailings products are pumped to their separate tailings disposal systems.

The fourth cleaner concentrate is the final product and is pumped to the concentrate thickener and filter section for dewatering. The routing of the intermediate flotation streams within the flotation circuit is shown on the flowsheet.

6.2.4 Concentrate Thickening and Filtering

Refer to flowsheet.25047-0204-F

The overflow from the 6 m diameter thickener is pumped to the process water tank while the underflow is pumped to the 5.1 m diameter concentrate stock tank. The thickened concentrate is pumped at a controlled rate from the stock tank and fed to the Larox filter. The filtrate product from the filter is recovered and pumped to the process water tank while the concentrate filter cake, containing about 7% moisture, is stored in a stock pile located on the ground floor of the mill building, under the Larox filter.

6.2.5 Tailings Disposal and Effluent Treatment

Refer to flowsheet 25047-0205-F

The sulphide tailings stream is pumped to lined disposal area situated within the tailings and effluent treatment catchment area.

The non-sulphide tailings are pumped to a high-compression tailings thickener where flocculant is added to aid settling. The thickener overflow stream is pumped to the settling pond while the underflow, containing 75% solids by weight, is pumped to the tailings co-disposal area, where it is mixed with mine rock from the open pit mine. Seepage from the co-disposal area is directed to the settling pond. The overflow from the settling pond gravitates to the recirculation pond where lime is added for pH adjustment. The overflow from the recirculation pond is directed to the polishing pond where it is either reclaimed to the process water tank at the plant site or discharged to the environment.

6.2.6 Services, Utilities, Process and Fresh Water

Refer to flowsheets 25047-0206-F.

A flotation low-pressure air blower, plant air compressor and instrument air compressor with ancillaries have been included within the scope of concentrator building services.

The process, fresh/fire and potable water systems are represented in the flowsheet.

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6.2.7 Reagents

Refer to flowsheets 25047-0205-F, 25047-0207-F, and 25047-0208-F

Four reagents are used in the Ni/Cu flotation circuit; these are:

• Aero 3477 as a collector; consumption 10.5 g/t of float feed or 47 kg/d. • Isopropyl Xanthate as a collector; consumption 28 g/t of float feed or 126 kg/d. • MIBC as a frother; 30 g/t of float feed or 135 kg/d. • Hydrated lime for pH control; 290 g/t (100% CaO) of float feed or 1,305 kg/d.

A flocculent will be used for thickening the concentrate. A dosing rate of 20 g/t of dry concentrate has been allowed for.

Five reagents are used in the tailings sulphide flotation circuit; these are:

• Aero 3477 as a collector; consumption 15 g/t of tailings or 65 kg/d. • Sodium Isopropyl Xanthate as a collector; consumption 45 g/t of float feed or 194 kg/d. • MIBC as a frother; consumption 17.5 g/t of float feed or 76 kg/d. • Copper sulphate as a sulphide activator; consumption 150 g/t of float feed or 648 kg/d. • Sulphuric acid for pH control; consumption 250 g/t of float feed or 1,080 kg/d.

Suitable storage, make-up systems and dosing facilities have been included within the scope of the process plant and infrastructure design.

6.2.8 On-Stream Analysis

Refer to flowsheets 25047-0209-F.

The following process streams are sampled and analysed for nickel, copper and iron using a Courier on-line analyser:

• Flotation feed. • Secondary rougher tailings. • Secondary rougher concentrate. • Primary cleaner tailings. • Final concentrate. • Primary rougher concentrate.

6.3 EQUIPMENT LIST

The detailed mechanical equipment list developed by Met-Chem for the 4,500 t/d concentrator is included in Appendix 3. The equipment selection and sizing was based on the flowsheets, design criteria and the mass balance.

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6.4 MASS BALANCE

The mass balance for the 4,500 t/d processing plant is provided in Table 6.4.1 (cont’d over).

Table 6.4.1 Process Mass and Water Balance

Running Time

Solids Solution Total

d/w h/d t/d t/h SG t/d SG m3/h t/h m3/h % solids SG Primary crusher Primary crusher feed 7.00 14.40 4,500 313 3.10 139 1.00 9.7 322 - 97 - Primary crusher product 7.00 14.40 4,500 313 3.10 139 1.00 9.7 322 - 97 - Grinding SAG mill feed 7.00 22.08 4,500 204 3.10 139 1.00 6.3 210 - 97 - SAG mill circulating load 7.00 22.08 1,350 61 3.10 238 1.00 10.8 72 31 85 2.36 SAG mill feed water 7.00 22.08 - - 3.10 1,573 1.00 71.2 71 71 0 1.00 SAG mill product 7.00 22.08 5,850 265 3.10 1,950 1.00 88.3 353 174 75 2.03 SAG mill pebble crusher feed 7.00 22.08 1,350 61 3.10 238 1.00 10.8 72 31 85 2.36 SAG mill screen spray water 7.00 22.08 - - 3.10 221 1.00 10.0 10 10 0 1.00 SAG mill screen underflow 7.00 22.08 4,500 204 3.10 1,933 1.00 87.5 291 - 70 - Ball mill circulation load 7.00 22.08 11,250 510 3.10 3,360 1.00 152.2 662 317 77 2.09 Ball mill feed water 7.00 22.08 - - 3.10 390 1.00 17.6 18 18 0 1.00 Ball mill product 7.00 22.08 11,250 510 3.10 3,750 1.00 169.8 679 334 75 2.03 Cyclone feed 7.00 22.08 15,750 713 3.10 12,886 1.00 583.6 1,297 814 55 1.59 Cyclone feed dilution water 7.00 22.08 - - 3.10 7,204 1.00 326.3 326 326 0 1.00 Cyclone underflow 7.00 22.08 11,250 510 3.10 3,360 1.00 152.2 662 317 77 2.09 Cyclone overflow 7.00 22.08 4,500 204 3.10 9,526 1.00 431.4 635 497 32 1.28 Rougher Flotation Feed to Flotation 7.00 22.08 4,500 204 3.10 9,526 1.00 431.4 635 497 32 1.28 Flotation feed 7.00 22.08 4,500 204 3.10 9,526 1.00 431.4 635 497 32 1.28 Primary roughr concentrate 7.00 22.08 180 8 3.20 420 1.00 19.0 27 22 30 1.26 Rougher conc spray water 7.00 22.08 - - 3.10 110 1.00 5.0 5 5 0 1.00 Secondary rougher feed 7.00 22.08 4,320 196 3.10 9,106 1.00 412.4 608 476 32 1.28 2ndary rougher concentrate 7.00 22.08 360 16 3.15 840 1.00 38.0 54 43 30 1.26 Secondary rougher conc spray water 7.00 22.08 - - 3.10 221 1.00 10.0 10 10 0 1.00

Rougher flotation tailings 7.00 22.08 3,960 179 3.10 8,266 1.00 374.4 554 432 32 1.28 Concentrate regrind Regrind mill circuit feed 7.00 22.08 360 16 3.15 1,061 1.00 48.0 64 53 25 1.21 Regrind mill circulation load 7.00 22.08 540 24 3.15 144 1.00 6.5 31 14 79 2.17 Regrind mill feed water 7.00 22.08 - - 3.15 36 1.00 1.7 2 2 0 1.00 Regind mill product 7.00 22.08 540 24 3.15 180 1.00 8.2 33 16 75 2.05 Cyclone feed 7.00 22.08 900 41 3.15 1,241 1.00 56.2 97 69 42 1.40 Cyclone feed dilution water 7.00 22.08 - - 3.15 - 1.00 - - - 0 1.00 Cyclone underflow 7.00 22.08 540 24 3.15 144 1.00 6.5 31 14 79 2.17 Cyclone overflow 7.00 22.08 360 16 3.15 1,097 1.00 49.7 66 55 25 1.20 Cleaner flotation Cleaner 1 feed 7.00 22.08 719 33 3.20 2,165 1.00 98.1 131 108 25 1.21 Cleaner 1 concentrate 7.00 22.08 360 16 3.30 926 1.00 41.9 58 47 28 1.24 Cleaner 1 conc spray water 7.00 22.08 - - 3.10 66 1.00 3.0 3 3 0 1.00 Cleaner 1 tailings 7.00 22.08 359 16 3.10 1,240 1.00 56.1 72 61 22 1.18 Cleaner 2 feed 7.00 22.08 472 21 3.30 1,329 1.00 60.2 82 67 26 1.22 Cleaner 2 concentrate 7.00 22.08 293 13 3.30 791 1.00 35.8 49 40 27 1.23 Cleaner 2 conc spray water 7.00 22.08 - - 3.10 44 1.00 2.0 2 2 0 1.00

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Running Time

Solids Solution Total

d/w h/d t/d t/h SG t/d SG m3/h t/h m3/h % solids SG Cleaner 2 tailings 7.00 22.08 179 8 3.10 538 1.00 24.4 32 27 25 1.20 Cleaner 3 feed 7.00 22.08 337 15 3.30 1,012 1.00 45.8 61 50 25 1.21 Cleaner 3 concentrate 7.00 22.08 225 10 3.30 675 1.00 30.6 41 34 25 1.21 Cleaner 3 conc spray water 7.00 22.08 - - 3.10 44 1.00 2.0 2 2 0 1.00 Cleaner 3 tailings 7.00 22.08 112 5 3.10 337 1.00 15.2 20 17 25 1.20 Cleaner 4 feed 7.00 22.08 225 10 3.30 719 1.00 32.6 43 36 24 1.20 Cleaner 4 concentrate 7.00 22.08 181 8 3.30 543 1.00 24.6 33 27 25 1.21 Cleaner 4 conc spray water 7.00 22.08 - - 3.10 22 1.00 1.0 1 1 0 1.00 Cleaner 4 tailings 7.00 22.08 44 2.0 3.10 177 1.00 8.0 10 9 20 1.16 Concentrate dewatering Concentrate 7.00 22.08 181 8.2 3.30 565 1.00 25.6 34 28 24 1.20 Conc thickener underflow 7.00 22.08 181 8.2 3.30 97.4 1.00 4.4 12.6 6.9 65 1.83 Conc thickener overflow 7.00 22.08 - - 3.30 467 1.00 21.2 21.2 21.2 0 1.00 Filter cake 7.00 14.40 181 12.6 3.30 14 1.00 0.9 13.5 4.8 93 2.84 Filtrate 7.00 14.40 - - 3.30 84 1.00 5.8 5.8 5.8 0 1.00 Tailings Rougher flotation tailings 7.00 22.08 3,960 179 3.10 8,266 1.00 374 554 432 32 1.28 Cleaner 1 tailings 7.00 22.08 359 16 3.10 1,240 1.00 56 72 61 22 1.18 Total tailings 7.00 22.08 4,319 196 3.10 9,506 1.00 431 626 494 31 1.27 Sulphide scavenger flotation conc 7.00 22.08 450 20 3.30 1,350 1.00 61.1 82 67 25 1.21

Sulphide tails conc spray water 7.00 22.08 - - 3.30 44 1.00 2.0 2 2 0 1.00 Sulphide tailings 7.00 22.08 450 20 3.30 1,394 1.00 63.1 84 69 24 1.20 Non sulphide tailings 7.00 22.08 3,869 175 3.10 8,156 1.00 369.4 545 426 32 1.28 Tailings thickener u/flow 7.00 22.08 3,869 175.2 3.10 1,289.7 1.00 58.4 233.6 114.9 75 2.03 Tailings thickener overflow 7.00 22.08 - - 3.10 6,866 1.00 311.0 311.0 311.0 0 1.00 Internal Plant Water Balance Water Balance - IN SAG mill feed 7.00 22.08 139 1.00 6.3 SAG mill feed water 7.00 22.08 1,573 1.00 71.2 SAG mill screen spray water 7.00 22.08 221 1.00 10.0 Ball mill feed water 7.00 22.08 390 1.00 17.6 Cyclone feed dilution water 7.00 22.08 7,204 1.00 326.3 Rougher conc spray water 7.00 22.08 110 1.00 5.0 2ndary rougher conc spray water 7.00 22.08 221 1.00 10.0 Cyclone feed dilution water 7.00 22.08 - 1.00 - Regrind mill feed water 7.00 22.08 36 1.00 1.7 Cleaner 1 conc spray water 7.00 22.08 66 1.00 3.0 Cleaner 2 conc spray water 7.00 22.08 44 1.00 2.0 Cleaner 3 conc spray water 7.00 22.08 44 1.00 2.0 Cleaner 4 conc spray water 7.00 22.08 22 1.00 1.0 Sulphide tails conc spray water 7.00 22.08 44 1.00 2.0 Total 10,114 Water Balance - OUT Sulphide tailings 7.00 22.08 1,394 1.00 63.1 Tailings thickener u/flow 7.00 22.08 1,290 1.00 58.4 Tailings thickener overflow 7.00 22.08 6,866 1.00 311.0 Conc thickener overflow 7.00 22.08 467 1.00 21.2 Filter cake 7.00 14.40 14 1.00 0.9 Filtrate 7.00 14.40 84 1.00 5.8 Total 10,114

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6.5 PROCESS PLANT PRODUCTION SCHEDULE

Table 6.5.1 shows the annual process plant production schedule over the life of the project.

Table 6.5.1 Process Plant Production Schedule


Ore Milled ( 000 tonnes ) - 1,643 1,643 1,643 1,643 1,643 1,643 1,371 - 11,226 Mill Recovery


Recovered Metal

Nickel ( 000 lbs ) - 9,964 10,647 9,114 7,318 7,763 9,299 7,907 - 62,012 Copper ( 000 lbs ) - 11,928 13,914 12,768 10,313 10,510 12,480 10,849 - 82,762 Cobalt ( 000 lbs ) - 625 633 588 528 524 596 494 - 3,989 Platinum ( 000 oz ) - 13.27 14.67 14.65 11.99 10.80 13.16 11.31 - 89.85 Palladium ( 000 oz ) - 8.27 9.40 9.22 7.20 6.44 8.04 6.97 - 55.53 Gold ( 000 oz ) - 3.59 4.11 4.08 3.30 3.06 3.91 3.51 - 25.56

Concentrate Grade ( % Ni ) - 8.19 7.80 7.50 7.47 7.65 7.69 7.59 - 7.71 Concentrate Grade ( % Cu ) - 9.81 10.20 10.50 10.53 10.35 10.31 10.41 - 10.29 Concentrate Produced (000 dmt) - 55.167 61.895 55.142 44.431 46.048 54.881 47.264 - 364.829Mass Pull (Conc t / Milled t) 0.00% 3.36% 3.77% 3.36% 2.71% 2.80% 3.34% 3.45% 0.00% 3.25%Payability of Metal


Payable Metal

Nickel ( 000 lbs ) - 9,167 9,796 8,385 6,733 7,142 8,555 7,275 - 57,051 Copper ( 000 lbs ) - 10,616 12,384 11,364 9,179 9,354 11,107 9,655 - 73,658 Cobalt ( 000 lbs ) - 313 317 294 264 262 298 247 - 1,994 Platinum ( 000 oz ) - 11.28 12.47 12.45 10.19 9.18 11.19 9.61 - 76.37 Palladium ( 000 oz ) - 7.03 7.99 7.84 6.12 5.48 6.83 5.92 - 47.20 Gold ( 000 oz ) - 3.05 3.49 3.47 2.80 2.60 3.32 2.98 - 21.73

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7.0 INFRASTRUCTURE

The site and plant general layout arrangements are presented in the drawings listed in Table 7.0.1, and which are included at the end of this section.

Table 7.0.1 List of Infrastructural Drawings

Number Rev. Description 25047-0001-L C Project Location Plan 25047-0002-L E General Site Layout 25047-0003-L E Surface Facilities Layout 25047-0010-L C Primary Crushing and Stockpile General Layout 25047-0011-L C Primary Crushing Plan 25047-0012-L C Primary Crushing Sections 25047-0013-L B Stockpile Reclaim Tunnel Plan and Sections 25047-0015-L B Pebble Crushing Plan and Elevations 25047-0020-L D Concentrator Ground Floor 25047-0021-L D Concentrator Operating Floor 25047-0022-L C Concentrator Sections A and B 25047-0023-L C Concentrator Sections C, D, E and F 25047-0030-L C Laboratory, Dry and Offices Complex Plan 25047-0041-L C Mine Equipment Maintenance Building Plans 25047-0042-L C Mine Equipment Maintenance Building Elevation

and Sections 25047-0050-L B Fuel Farm Plan and Sections 25047-0070-L B Water Supply Plan and Sections 25047-0080-L B Effluent Treatment Plant Plan and Sections

7.1 PLANT AND SITE LAYOUT

The main criteria considered during the site layout development for the mill and other surface infrastructure were:

• The minimum haulage distance from the pit to the crusher and to the waste dump.

• Minimize site visibility for cottagers on the edge of Agnew Lake.

• Minimal noise emissions from the plant site.

The general site layout is located on an area at the northeast corner of the East pit limit, at a distance averaging 1,700 m from the mid-point between the West and East pit. The site area is relatively level and easily accessible from the existing exploration road built by URSA. The total site area to be cleared and levelled for construction is approximately 40 – 45 ha.

The plant site area is covered by trees, mainly pine and spruce. Tree-cutting is required and the plant site will be cleaned and grubbed. Based on test pits excavated in 2005, the overburden thickness is estimated to be approximately 1.5 m and about 40% of the surface area is exposed bedrock. The estimated amount of rock to be blasted and used as fill in lower depressions is approximately 30,000 m3.

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7.2 ROADS

7.2.1 Access Road

Access road to the site will be from the northeast via a secondary road branching north from the Trans Canada Highway # 17 approximately 7.5 km east of Nairn Center. An existing logging road connects to the west side of the secondary road, approximately 13 km from Highway 17 and allows access to the property. This existing logging road is considered to be suitable for site access without upgrading. The portion branching south to the Shakespeare Project site will require upgrading over 3.4 km to the property gate location. The upgrade is planned using mine waste fill and road surfacing consisting of 0-150 mm borrow pit run material. Mine waste is assumed non acid generating and in sufficient quantities to complete road work in time. The site preparation for road consists of clearing of trees twice the width of the final road. No grubbing is included. Nine drainage culverts are planned for the upgrade.

MineSight software was used to select the most economical route and estimate cut and fill quantities. Rock cut quantities were kept to a minimum respecting 8% slope maximum.

7.2.2 Site Roads

The site roads are required to access different parts of the property. The site roads include the road from the gate to the plant site facilities, the fresh water intake road, the mine haulage road, the mine explosive storage road, the bulk explosives plant road, the tailings thickening plant road, the pyrrhotite tailings disposal area road, the effluent treatment plant road and the mine dewatering line road.

All roads are constructed from mine waste covered by 0-150 mm of borrow pit run material. For each road an estimated number of drainage culverts are included. MineSight software was used to select the most economical route and estimate cut and fill quantities. Rock cut quantities were kept to a minimum respecting maximum slope of 8% for mine haulage trucks and 8 to 10% for other roads.

The main haulage road from the pit to the crusher and the waste dump is located in the extension of west open pit ramp on the hanging wall side of the pit. The amount of waste to be removed for this haulage road is estimated at 33,500 m3. The waste will be utilized for road and site construction. The estimated amount of filling material required for the final road and site construction is 484,000 m3.

Alternatives have been discussed to reduce the amount of waste displacement providing from the haulage road and reduce the excavation costs. Further investigation during the detailed engineering phase of the project may result in additional savings in excavation and fill requirements, while facilitating access to the crusher and the waste dump from both pits.

7.2.3 Plant Gate and Site Fencing

The plant gate is located northwest of the waste rock/tailings co-disposal area. The gate is located 1.7 km north of the main site facilities. It consists of a remote operated gate complete with camera and radio communication.

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Fencing is restricted to the main gate, the main sub-station, the propane tanks and explosive storage areas. The explosive magazine area will be enclosed by a fence with a locked gate allowing access to the area by authorized employees only.

7.2.4 Parking Area

The employee parking is located northwest of the plant facilities. The parking area (6000 m2) has a capacity of 100 vehicles. It will serve as temporary laydown for construction before being transferred to operations. The parking has 30 power outlets for winter.

The parking is constructed from mine waste covered by 150 mm of pit run material.

7.3 CRUSHING FACILITY AND PROCESSING PLANT LAYOUT

The surface site plan has been designed to have a minimum footprint. The site access road follows essentially the same path as the URSA exploration road. It has been assumed that this road will be upgraded using mine rock material produced during pre-stripping of the pit.

7.3.1 Crusher

The primary crusher area is of a compact design. The site for the crusher installation was selected by using natural topography of the land to benefit from the vertical elevation between the dumping point and the output of the crusher.

7.3.2 Concentrator

The concentrator building is a conventional ore processing type building. The dimensions are 42 m by 63 m. The concentrator building houses the grinding, flotation, reagent and filtration areas, the compressors as well as tailings pumps and pumpboxes. The electrical rooms (two floors of 12 m by 21 m) are located on the west side. The concentrate loadout is located at the east end of the building. The loadout is designed for a Caterpillar 988 front-end loader and a drive through for concentrate trucks.

Mechanical and electrical maintenance shops are located adjacent and south of the concentrate loadout. Offices, conference room, lunchroom and washrooms are provided for on the second floor of the maintenance shops.

7.3.3 Office Complex

The office complex is composed of 17 Atco type trailers assembled in modules. Each trailer has dimensions of 3.65 m by 18.3 m (12 ft x 60 ft). The complex has five modules, one for each department. The mine building housing the mining group is composed of two trailers put together giving an area of 7.4 m by 18.3 m (135 m2). The change house building is composed of five trailers put together giving an area of 18.5 m by 18.3 m (340 m2). The mill building housing the mill personnel is similar to the mine building. The administration building is composed of five trailers similar to the change house building. The laboratory building is detailed in the following section.

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The complex is installed on wooden supports, called cages, approximately 1.2 m above ground. Each building/module is installed with skirting. Each building/module has connecting heated corridors except for the laboratory, which is isolated.

7.3.4 Laboratory Building

The laboratory building has a surface area of 200 m2 and is composed of three trailers. Two simple wooden structures are added to the laboratory building. One is for the electrical room, where all electrical equipment is installed. The other one houses the precision scales and has a concrete floor. The laboratory building is installed on wooden supports, approximately 1.2 m above ground. It is also installed with skirting.

7.3.5 Mine Equipment Maintenance Building

The mine equipment maintenance building is a standard structural steel, siding and concrete foundations building normally constructed for similar plant sites. The building includes a wash bay, three major equipment maintenance bays, the warehouse and offices. The warehouse is described in the following section.

The wash bay is equipped with a pressure washing system. The maintenance bays are equipped with a 15 tonne overhead crane and a dedicated air compressor. Each bay has an area of 205 m2. An oil/water separator is included. The mine equipment maintenance building is designed to accommodate the Caterpillar 777F or Komatsu HD785-5 models of 91t mine trucks.. It is also planned to service a Caterpillar 992G wheel loader as well as smaller equipment in the facilities.

7.3.6 Warehouse

The warehouse is an integral part of the mine equipment maintenance building, separated from the maintenance bays by a concrete block wall. The surface area is 300 m2. The warehouse floor is elevated 1.2 m above the surrounding ground elevation, obviating the need for a ramp down to a loading dock and thus eliminating potential water and ice problems.

7.3.7 Truck Scale

A truck scale is located approximately 90 m northeast of the concentrator loadout. It is designed to weigh the concentrate trucks and any other delivery trucks, if required. The type selected for the project is a Balance Industrielles Montreal (BIM) type DURABIM 10100-100-A, capacity of 100 tonnes.

7.4 SERVICES

7.4.1 Fuel Storage and Fuelling Station

The fuel storage is located north east of the plant facilities on the north side of the site access road. The base will be constructed with pit-run material. The tank will be surrounded by containment berms sufficient for the full capacity of the tanks plus an allowance for water and

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snow. The berms will be constructed with pit-run material and lined with a leak-proof membrane.

The fuelling station is a Chamco fuelling station all-inclusive package located on the south dyke of the fuel storage reservoir. All equipment is contained in a 3.6 m by 13.7 m container. The fuel trucks will move up a ramp to the fuel dispensing equipment part of the fuelling station. All potential leaks from trucks or the fuelling station are directed to the fuel storage reservoir via sloping ground.

7.4.2 Mine Explosive Storage

Cartridge explosives to be used for pre-shearing and secondary blasting will be stored in a magazine provided by the explosive supplier. The magazine for detonators, also to be provided by the supplier, will be located nearby. The location of the magazine approximately 500 m to the west of the West Pit is based on a maximum quantity of 30,000 kg as per the requirements of NRCan explosive quantities/distances regulations.

7.4.3 Bulk Explosive Plant

The bulk explosive supplier will erect a bulk explosive mixing plant on the Shakespeare property. The required services (power, water, septic installation) will be provided by URSA.

7.4.4 Water Systems

The water systems include: the fresh water intake system, the plant collection pond system, the reclaim water system, and the mine dewatering system.

The fresh water intake system provides fresh water to the mill fresh water and fire protection water tank. The fresh water intake is from Agnew Lake just east of Long Bay. A 1.4 km road gives access to a fresh water barge. The fresh water barge is a Chamco intake barge package. The barge dimensions are 9.8 m by 4.6 m. The barge package includes two 75 hp vertical turbine pumps capable of supplying the mill and potable water requirements. The barge is equipped with a de-icing system to prevent ice formation around the barge during winter. This is done with a 15 hp pump, pumping water in a peripheral perforated piping system around the barge. The barge is also equipped with a 24 m by 1.5 m walkway that can withstand a 7 m water level variation expected in Agnew Lake. The Chamco intake barge package is all-inclusive, only requiring power supply to the barge and provision for the return piping to the mill.

The fresh water tank near the concentrator building has a total capacity of 1,113,000 liters. The capacity for fire protection is 625,000 liters.

The plant collection pond is located south of the concentrator building. It is designed to collect all surface drainage water coming from plant facilities. The plant collection pond system is mainly a pumphouse, pumps and piping system to pump water back to the mill tailings pump box. The pumphouse includes a standard vertical turbine pump placed in a galvanized pipe well with a galvanized pipe intake. The pumphouse’s foundations are a concrete slab on pit run material backfill on the upstream side of the collection pond dam. The pump and building

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are included in a Chamco all inclusive pump package reducing site work. The piping is 100 mm (4”) diameter HDPE pipe running to the mill building for a distance of 340 m.

The polishing pond is located east of the waste rock/tailings co-disposal area. It is designed to hold all water coming from the settling pond and the precipitation pond. The reclaim water system is mainly a pumphouse, pumps and piping system to pump water back to the mill building after being treated and decanted. The pumphouse includes a standard vertical turbine pumps in a galvanized pipe well with a galvanized pipe intake. The pumphouse foundations are a concrete slab on pit run material backfill on the upstream side of the polishing pond dam. The pumps and building are part of a Chamco all-inclusive pump package reducing site work. The piping is 250 mm (10”) diameter HDPE pipe running to the mill building for a distance of 2,650 m.

The mine dewatering system takes mine water from open pits and transports it to the settling pond east of waste rock/tailings co-disposal area. The system is made of two HDPE lines 150 mm (6”) in diameter running for 1,025 m from the pits to the pond.

7.4.5 Potable Water Treatment

The potable water treatment system is provided by a Chamco water treatment all-inclusive package. The potable water treatment system fits in a standard 2.4 m by 6.1 m container. The package reduces the site work needed for installation and provides the required water treatment equipment to satisfy Ontario regulations.

The container is installed in the concentrator building near line 10 on the second floor. The potable water is stored in tank 520-TAK-02 just below the treatment container. The water is then distributed via pumps to the different buildings.

7.4.6 Heating, Ventilation and Air Conditioning

Exhaust fans are installed for each building to insure proper air change and temperature control in the summer.

The major buildings and facilities are heated by propane direct fired units, namely, the crusher and pebble crusher buildings, the reclaim tunnel, the concentrator including the loadout area, the main electrical room, the mine equipment maintenance building including the wash bay, the garage area and the warehouse.

The secondary installations are heated by electrical baseboard heaters, namely the laboratory, the change house, and the administration offices. The offices in the concentrator and maintenance building are also heated with baseboard heaters and have air conditioning.

7.4.7 Plant Mobile Equipment

Plant mobile equipment includes four crew cab 4x4 pick-up trucks, one fork lift 6000# capacity with heated cab and one bob-cat loader with cab, bucket, forks and quick coupler.

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7.5 ELECTRICAL POWER

The electrical and automation drawings listed in Table 7.5.1 are included at the end of this section.

Table 7.5.1 List of Electical and Automation Drawings

Number Rev. Description A1-25047-7001-01-E B Main Single Line Diagram Sheet 1 of 2 A1-25047-7001-02-E C Main Single Line Diagram Sheet 2 of 2 A1-25047-7380-E B 44-4.16kV Substation Layout A1-25047-8050-Y B Global Control System Architecture

7.5.1 Power Requirement & Supply

The estimated power demand for the plant is shown in Table 7.5.2. The estimate assumes utilisation factor of 0.8 and a power factor of 0.9. The plant will be fed by a new 44 kV overhead electrical power line to be built by a local contractor over a distance of 7.5 km as shown on the General Site layout drawing (# 25047-0001-L Rev. C).

The grid connection of this electrical power line will be at the existing 115 kV power line located south of the Project site northeast of the town of Webbwood. A 115 kV/44 kV Substation/Switching Station of 10 MVA capacity will be constructed at that location. A 44 kV/4,160 V Substation of 10 MVA capacity will be constructed at the end of the 44 kV line adjacent to the concentrator building.

A 1250 kW, 4,160 volt, emergency diesel generator will be located near the main electrical room and will feed the main busbar at 4,160 volt to give the emergency power to critical loads in the concentrator and elsewhere on the site.

7.5.2 Main Electrical Room

The main electrical room, located near the main loads, will contain the following equipment: medium voltage switchgear, including breakers to distribute power to the different areas of the plant; ball mill starters and accessories for 3,000 hp motor; SAG mill starters and accessories for 4,000 hp motor; power transformers 3,000/4,000 kVA 4.16 kV / 600 volt; low voltage switchgear 600 volt with one (1) main breaker 5,000 amp, eight (8) breakers 800 amp; four (4) Motor Control Centres (MCC) for the concentrator motors.

Two other MCC’s are located respectively in the central building and in the mine equipment maintenance building.

Variable frequency drives are planned for the concentrate thickener underflow pumps.

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Table 7.5.2 Estimated Total Power Requirement

Description Operating

kW Operating

HP Comments Crushing area 566 Grinding area 7,780 Flotation 1,257 Regrind area 239 Nickel concentrate 102 Talilings pump (at concentrator) 290 Talilings thickener area 300 By Golder Effluent Treatment 21 Process water 250 Fresh water 407 Potable water 9 Reagents 632 On-line analyser 12 Plant site collection pond 10 Roof exhaust fan 17 Emergency generator 0 Sub-total Milling 8,870 11,891

Open Pit Pumps 295 396 Operating 24 hours/day, 2 operating units Bulk Explosive Plant 200 Surface Garage & Mtce 55 Estim. Operating 16 hours/day Heating General 100 Estim. Operating 24 hours/day. 8 months/year Lighting General 250 Estim. Others General 458 614 Estim. 5% of above. Operating time 24 hours/day. TOTAL 10,229 12,901 Equivalent MVA Requirement 9.09 Assuming Utilization Factor 0.8, Power Factor 0.9 Available capacity (MVA) 10.00 Approx HydroOne 115 kV powerline avail capacity I ( load line demand at 44 kV ) 119 Utilizing existing line with new substation 115-44 kV 7.5.3 Main Electrical Room

The main electrical room, located near the main loads, will contain the following equipment: medium voltage switchgear, including breakers to distribute power to the different areas of the plant; ball mill starters and accessories for 3,000 hp motor; SAG mill starters and accessories for 4,000 hp motor; power transformers 3,000/4,000 kVA 4.16 kV / 600 volt; low voltage switchgear 600 volt with one (1) main breaker 5,000 amp, eight (8) breakers 800 amp; four (4) Motor Control Centres (MCC) for the concentrator motors.

Two other MCC’s are located respectively in the central building and in the mine equipment maintenance building.

Variable frequency drives are planned for the concentrate thickener underflow pumps.

7.5.4 On-site reticulation

The distribution of power around the Shakespeare project site will require the following lines:

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• A 4,160 volt line between the effluent treatment and the transfer water area, with isolated cable installed on special tray;

• A 4,160 volt line between the Plant Access Road and the reclaim water pumphouse;

• A 4,160 volt overhead line between the main substation and the pumphouse station to feed the water station;

• A 4,160 volt overhead line between the main substation and the open pit area with electrical room in a container;

• An extension of the open pit 4,160 volt line to the bulk explosive plant;

• A tap on the 4,160 volt overhead line to the crushing plant area; and

• A tap on the 4,160 volt overhead line to the collecting pond area;

• A 4,160 volt overhead line to the parking area will supply electrical power during the construction;

• A 4,160 volt overhead line to the tailing thickener area;

• A 4,160 volt overhead line to the effluent treatment

• A 600 volt line to the service building electrical room to feed the office building and surrounding buildings.

• Extension of the line from the tailings thickener area to the remote-controlled gate

The concentrator will be supplied from the switchgear located in the main electrical room.

7.6 AUTOMATION & INSTRUMENTATION

7.6.1 PLC Controls

The conceptual control system architecture for the new ore processing complex was designed based on the utilisation of Programmable Logic Controllers (PLC) in each key process area. The areas will be interconnected in Network via an Ethernet communication fibre optic system for the concentrator area and by modem for the others areas. The Ethernet communication system is fast, reliable and is becoming the standard in the industry and all PLC manufacturers support the protocol.

The control system is designed with three operator’s workstations in the central control room to supervise and control the plant operation and for power management. The Engineer station is located in the concentrator electrical room for the system programming and the maintenance debugging. The operator and the engineer stations are connected to the PLC network system by Ethernet fibre and are part of the same network.

7.6.2 Instrumentation

The instrument list was developed using the process flowsheet and co-ordination meeting with the process group and verification with potential supplier.

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The instruments will be wired to the analogue input/output module located in different electrical rooms by standard 4-20 ma signals. For complex instruments, a communication link could be added for remote calibration and diagnostic.

7.7 WASTE DISPOSAL

7.7.1 Sanitary Waste Water

The sanitary waste water system consists of collecting all sanitary water from each building via an underground piping and manholes system. The system discharges in a concrete septic tank located south of the crushed ore stockpile. The septic tank, after the proper retention time, discharges in a leaching bed made of perforated pipes. The septic tank and leaching bed design is based on Ontario Water Resources, REG. 525/98 and Ontario Building Code Act, REG. 403/97.

7.7.2 Garbage Disposal

Garbage disposal will be handled by a local contractor removing materials to an appropriate disposal site. No capital cost allowance is included.

7.8 TELECOMMUNICATIONS AND COMPUTER NETWORKING

Telephone and internet services infrastructure are available in the area but not directly to the property. The nearest telecommunications systems are located at the Agnew Lake Lodge.

7.8.1 Telecommunications

A cabled telephone trunk line from the main telephone line at the Agnew Lake Lodge would be laid to provide connection to the main Bell telephone network.

A small telephone switch would be located in the office building, with telephone extensions for individual office spaces connected to this switch. The switch and telephone cabling would allow employees to access telephone and internet services as required.

7.8.2 Computers and Networking

The corporate computer systems of the mine would be based on Microsoft.NET Enterprise Servers, ethernet LAN and Windows XP-based PC’s. Software would comprise an industry standard office applications package, accounting, stores, and technical systems for mine planning and production control.

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8.0 MINE ROCK AND TAILINGS MANAGEMENT

8.1 INTRODUCTION

URSA requested Golder Paste Technology Ltd. and Golder Associates Ltd. (Golder) to perform a feasibility study and produce a capital and operating cost estimate (+/-15%) for the evaluation and design of a tailings thickening and mine rock deposition plan for the Shakespeare Project located along Agnew Lake, north of Espanola.

The main design criteria were provided by URSA, Met-Chem and Micon, and which formed the basis for the analysis and design of the thickening plant.

The original concept for tailings disposal was conventional slurry deposition in a valley site about 7 km north east of the mine. URSA retained Golder to review that approach and to recommend a cost effective, fully functional concept. This included the possibility of co-disposing thickened tailings and mine rock in a site close to the mine in order to reduce the area of disturbance and environmental impact. Golder then recommended the co-disposal concept as the preferred approach in its trade-off study (see Appendix 5).

Golder prepared a subsequent report as part of the Shakespeare feasibility study, presented in two volumes, Volume 1, Tailings Preparation, dated December 21, 2005, and Volume 2, Co-disposal and Water Management, dated January 9, 2006. These are provided in Appendices 6 and 7 to this report.

8.2 CO-DISPOSAL CONCEPT

The co-disposal of tailings and mine rock on surface is not common in Ontario although it is carried out underground and there are many examples where mine rock has been used to contain tailings. In the case of the Shakespeare project, co-disposal of the tailings and mine rock on surface has a number of advantages over conventional slurry disposal of tailings for which the selected site was 7 km from the mine. Generally, the advantages of co-disposal include:

• Ease of management of disposal in a single facility.

• Reduced area of disturbance.

• Reduced issues relating to seepage and evaporation losses.

• Reduced infrastructure requirements (e.g., roads, pipelines, pumps et cetera).

• Better control of acid generation though efficient mixing (tighter matrix).

• Reduced water management issues (single point of discharge).

• No ponded water on top of the co-disposed material (hence less risk).

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• Reduced closure issues resulting from smaller area, single water management system, single treatment plant, ease of monitoring.

Acid generating pyrrhotite tailings and acid generating mine rock will be disposed of in the low central valley section of the co-disposal area in the early years and kept permanently submerged behind a water retaining dam. In the later years of operation they will be disposed of in the mined out West open pit. The non-acid generating mine rock and tailings will be placed initially in the upper reaches of the valley and eventually over the central region, covering and encapsulating the acid generating materials.

Golder provides a discussion of the dam and dyke construction based on its assessment that the topography of each location is suitable for conventional, zoned, water retaining embankment dams with central water retaining cores, founded on stiff clay or grouted bedrock. This is a proven approach for dam construction at other mines in the region.

The principal features of a zoned embankment dam are a barrier core to retain water enclosed by a supporting structure or shell constructed of compacted granular material. Where cold winters are experienced a frost protection cap is required and, at Shakespeare, Golder has recommended a 2-m frost cap.

Before placement of the core, overburden is excavated to bedrock (or a suitable water-retaining soil). The surface is cleaned and slush-grouted and curtain grouting is carried out. On a soil foundation, all organic and unsuitable (e.g., frost impacted) material is removed and the foundation compacted before placing the core.

The zones of the embankments proposed for the Shakespeare project are:

1. Core (clayey soil).

2. Upstream shell (random fill or select mine rock).

3. Downstream shell (select granular, or select mine rock).

4. Upstream transition (processed or select granular).

5. Downstream filter (processed or select granular).

6. Downstream transition.drain (processed or select granular).

7. Blanket or finger drains under downstream shell (processed or select granular).

8. Frost protection cap (granular, non frost susceptible material.

9. Upstream erosion protection (processed granular or mine rock).

10. Downstream erosion protection (processed granular or mine rock – can be finer).

11. Toe drain (processed granular or mine rock).

12. Road surface on dam crest (sand and gravel).

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8.3 LABORATORY EVALUATION

In the initial laboratory program, several tests were performed on the Shakespeare project combined tailings, which consist of pyrrhotite and pyrrhotite-free tailings, and the sulphur reduced or pyrrhotite-free tailings samples. Once these tests were completed, only approximately 2 kg of material remained in the Golder PasteTec laboratory. The remaining sample was shipped to SGS Lakefield (Lakefield) for additional geochemical testing. The remaining pyrrhotite-free tailings underwent segregation testing.

The summary from the laboratory work in the Golder trade-off study (Appendix 5) will be reviewed briefly as follows. An additional laboratory-scale test was completed for the purposes of the Golder work for the feasibility study.

8.3.1 Laboratory Summary

The following are the main results of the trade-off study testwork:

• Both tailings samples had over 40% of the material passing 20 µm, which is well over the minimum required 15% passing 20 µm; therefore, pipeline transport problems are not expected.

• The pyrrhotite-free tailings have an SG of 2.97, the pyrrhotite concentrate has an SG of 3.67, while the combined sample has an SG of 3.10.

• Based on the mineralogical and chemical results, there does not seem to be any significant component that would affect either the thickening or pipeline transport of the tailings.

• Both tailings samples showed sensitivity to water additions, that is, the slump changed significantly with relatively small amounts of water addition.

• As the solids content increased, each sample displayed the typical exponential increase in yield stress but the results indicate for the most part it should be possible to pump with centrifugal pumps at 71 wt% solids.

• Approximately 10 wt% of the total water available separated within 2 hours, but once deposited, most of the ‘bleed’ water tended to evaporate before pooling occurred.

• Flocculant screening and settling tests determined that Magnafloc 338 at a dose of 24 g/t was the optimum flocculant dosage.

8.3.2 Additional Laboratory Testwork

8.3.2.1 Segregation Testing

Additional testing was carried out to determine the non-segregating limit for the pyrrhotite free sample. A series of segregation tests involving a range of solids contents from 42 to 75 wt% solids revealed that non-segregating slurry could be expected as low as 55 wt% solids.

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Generally, a paste consistency starts at approximately a 250-mm slump (10 in) which, in this case, corresponds to a solids content of approximately 73 wt%.

Approximately 71 wt% solids is the expected underflow density, which is well within the range of solids contents which were determined to be non-segregating as well as below the range of paste densities, which is anticipated to be the appropriate consistency which will allow centrifugal pumping and proper mixing of the tailings and mine rock in the deposition area. Segregation is determined by profiling a sample which has settled for a period of 15 minutes. This involves laser particle diffraction of the upper and lower portion of a test vessel to determine if segregation has occurred. The results are compared to the original particle size distribution, and a ratio is developed representing the amount of coarse particles (P80 of the original sample) remaining in suspension.

At 71 wt% solids, approximately 13 wt% of available water was released over a 24-h period. Dessication testing performed under standard conditions, (20ºC, 70% relative humidity) revealed approximately 12 wt% of the water released would naturally evaporate over the same time period.

8.3.3 Laboratory Conclusion

The laboratory results indicate that the Shakespeare tailings are suitable as a non-segregating, thickened slurry, are amenable to centrifugal pumping and will not likely cause problems from a thickening/transport perspective.

8.4 GEOCHEMISTRY

Preliminary testing was undertaken by NAR on approximately 130 samples of mine rock in order to identify the general geochemical characteristics of the principal rock types. Subsequently, more detailed testing was carried out on a selected number of mine rock samples by SGS under the direction of Knight Piésold. As of the time of writing of the Golder report, January 2006, testing of two tailings samples had been in progress for approximately 20 weeks and further testwork on additional tailings samples had been initiated at SGS.

The geochemical characterization program is discussed also in Section 10.5 of this report.

8.4.1 Principal Lithologies

The four principal lithologies identified by URSA are:

• Hanging wall quartzite.

• Quartz gabbro.

• Footwall gabbro.

• Mineralized rock comprising Disseminated ore (the majority) and Blebby ore (occurring on the hanging wall side of both the East and West Zones).

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The ARD rock, amounting to some 6% of the total, is identified as the quartz gabbro and, possibly, the footwall gabbro. The majority of the non-ARD rock is the hanging wall quartzite.

The characteristics of each type are discussed briefly below.

8.4.1.1 Hanging Wall Quartzite

The hanging wall quartzites generally have limited sulphide available for oxidation but, also, no effective neutralization capacity. Kinetic testing of this material, classified as non-ARD, has produced acidic pH values below the Federal MMER limits and the Ontario MISA limits. Sulphate concentrations, which are assumed to relate directly to oxidation of sulphide, are stable to decreasing. Golder recommended that kinetic testing should continue in order to confirm the preliminary conclusions that the longevity and magnitude of ARD will not pose a long term issue.

8.4.1.2 Gabbros and Ore

The quartz gabbro and, possibly, the footwall gabbro and disseminated sulphide mineralization, are considered ARD material in view of the sulphide content. However, these rock types also have higher neutralization capacity than the quartzite.

The disseminated sulphide mineralization will go to the mill for processing.

Kinetic testing will demonstrate the anticipated timing of the onset of ARD from these rock types. As of January, 2006, the neutralization capacity had not been exhausted in the ongoing kinetic tests so that, although the sulphate content is stable to increasing (indicating that oxidation of sulphide is occurring), metal concentrations are generally low. The exception is aluminum which was low but had decreased from the elevated concentrations of 0.1-0.25 mg/l.

It is expected that the sulphide mineralized gabbros will be acid-generating, however, the footwall gabbro is not expected to be acid generating.

8.4.1.3 Tailings

Tailings samples tested as of January, 2006 are:

• Sample F19: combined flotation tailings.

• Sample F30: rougher tailings (sulphur-reduced rougher flotation tailings).

The F19 tailings have significantly higher sulphide content than the F30 material, but the same neutralization capacity. Testing indicates that F19 material will be strongly acid-generating. One type of leach testing which, at dilute water to rock ratios, produced pH values lower than the MMER limits and nickel and zinc concentrations higher than the respective MMER limits.

As of January, 2006, kinetic testing on the F19 sample had not generated similar metal concentrations and it is assumed that the available neutralization capacity had not been

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exhausted. Decreasing pH, alkalinity and increased sulphate during kinetic testing indicate that neutralization capacity was being used although there was no detectable acidity and metals concentrations were below MMER limits as of January, 2006. Thiosalts detected in the supernatant decant water from the F19 tailings degraded from a level of 25 mg/l to below than 10 mg/l in less than one day.

The ARD potential of the F30 material is low, because of the lower sulphide content. Kinetic testing has demonstrated no significant ARD potential. Testing showed stable and neutral pH, no detectable acidity, low sulphate and all metals below the MMER standards.

As of January, 2006, geochemical testing of the pyrrhotite tailings, had not been undertaken but is planned on seven samples. These high sulphur tailings are anticipated to generate significant acidity and high metal dissolution. As noted above, it is planned that pyrrhotite tailings will be stored under water in the lower central portion of the CDA.

8.5 PHYSICAL PROPERTIES OF THE TAILINGS

The tailings for co-disposal are a coarse grind with approximately 46% passing the 200 mesh sieve (silt sizes). The specific gravity of the non-acid generating tailings is 3.10 while the specific gravity of the pyrrhotite tailings is 3.80.

8.6 GEOTECHNICAL CONSIDERATIONS

Knight Piésold carried out the geotechnical investigations for the CDA through a series of bore holes and test pits. Five holes were drilled in the area of the co-disposal facility of which three are close to proposed dam sites. The latter three holes were drilled for groundwater monitoring well installations and were also logged for geotechnical data.

The CDA is located in a valley comprising silt and soft clay overlying bedrock in the lower part with the sloping sides made up of exposed bedrock. A test pit dug by URSA indicated the sediment to be competent clay.

The holes completed for monitoring wells indicate that the dam sites are located on very shallow overburden overlying bedrock. Golder assumes that the dams and dykes will be founded essentially on bedrock. This is also the case for the aging pond and the plant site collection pond.

8.7 SITE PRECIPITATION AND EVAPORATION

The mean annual precipitation for the site is 876 mm and the mean annual lake evaporation is 520 mm. This is based on records from meteorological stations at Massey, Ontario and Amos, Quebec. Snowfall represents 26% of the total precipitation that falls between November and March. The precipitation for the 100 year wet return is estimated to be 1,203 mm and for the 100-y dry return it is 625 mm. The peak daily rainfall is estimated to be 104 mm for the 100-y return event. A historical storm for the region is the Timmins storm that produced 193 mm of rain in 12 hours. For this study the 100 year storm has been assumed for the environmental design flood (EDF) because of the relatively short life of mine. Any storm in excess of the

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EDF will be allowed to spill over the emergency spillway that will be designed to handle the routed probable maximum flood (PMF).

8.8 SEISMIC RISK

The seismic (earthquake) risk in the region is relatively low.

As of January, 2006, a seismic risk assessment had not been carried out for the Shakespeare project and was recommended by Golder for the final design stage.

Studies that were carried out for the closure of the Elliot Lake uranium facilities are relevant at this stage of development of the Shakespeare project. The 1-in-10,000-y event at Elliot Lake was as a 5.5 magnitude event 22 km from the site, or a 7.0 magnitude event 80 km from the site. The magnitude 5.8 event at Saguenay, is also commonly used for dynamic stability assessment of dams in Eastern Canada. It would describe the 10,000-y seismic response spectrum within the frequency range of about 2 Hz with a peak ground acceleration of 0.07g.

8.9 CLOSURE

The closure objectives for the tailings and mine rock facilities are to:

• Design for closure. • Promote progressive closure. • Prevent the migration of contaminants in surface and groundwater. • Inhibit surficial erosion. • Prevent (wind blown) dusting. • Enhance the visual aspects of the site.

The relatively small volume of acid-generating pyrrhotite tailings and acid generating mine rock will be disposed of separately and kept permanently submerged in the bottom of the co-disposal area (placed in the early years of operation) or in the mined out west open pit (in the later years of operation) in order to inhibit oxidation.

It is anticipated that the majority of the tailings and mine rock will not actively oxidize and release contaminants (metals). These areas will be graded and vegetated. Erosion protection against wind and runoff will be provided on exposed surfaces and will utilize clean mine rock separated and stockpiled for this purpose during operation.

The water treatment plant, located between the polishing pond and the settling pond, will continue to operate until it is no longer required.

8.10 DESIGN DRIVERS

The key design driver for the CDA is the geochemistry of the tailings and mine rock. Based on the test results that were available as of January, 2006, about 6% of the mine rock is sulphide-bearing gabbro that is potentially acid generating. The remaining mine rock is quartzite, which has a low sulphide content but shows some potential to generate acid. About 90% of the tailings are sulphur reduced and show no signs of acid generation or metal leaching and about

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10% of the tailings have a high pyrrhotite content and will be acid generating. The two tailings streams are produced simultaneously in the milling process. Co-mixing the non-acid generating tailings with the mine rock will reduce the potential for the mine rock to oxidize.

Other design drivers include:

• aging the tailings water for at least 30 days and treating before recycling to the mill.

• minimizing visual impacts.

• reducing the foot print and the environmental disturbance.

8.11 SITE SELECTION FOR THE CO-DISPOSAL FACILITY

The selected site for the co-disposal area (CDA) is about 1 km from the open pit mine and about 800 m from the processing plant, or mill. It has good topographic containment that enhances stability. It is currently heavily wooded. The flat base area is underlain with clay that will inhibit seepage. Selective placement of the co-disposed materials will be required to inhibit instability on the clay. The sides are predominantly rock outcrop and there is space for ponds downstream of the disposal area. The tailings thickening plant is located on the south side of the co-disposal area about 500 m north of the mill.

The main dam for the CDA will be located on the western, downstream side of the area with the settling pond and polishing ponds on the western side of the main dam. The dam for the external seepage collection pond is located on the northeastern, upstream side of the CDA.

8.12 OPERATING DATA

The basic operating data for a mine life of approximately 7 years, are:

Ore 11.227 million t

Annual production rate 1,642,500 t/y

Mill availability 92%

Tailings/ore ratio 0.96

Tailings discharge slurry density (before thickening) 31% solids by mass

Acid generating (pyrrhotite) portion of the tails 10% of total

Mine rock 57.813 million t

Potentially acid generating mine rock 6% of total

Water content of ore to mill 3% of total mass

Water content of concentrate 7% of concentrate mass

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Fresh make-up water required in mill 0.050 m3/t ore milled

The acid generating and non-acid generating mine rock will be hauled by truck to the co-disposal area and end dumped. The acid generating tailings will be pumped in slurry form to the co-disposal site and deposited underwater. The non-acid generating tailings will be thickened and pumped (via a pipeline) to the co-disposal area for co-mingling with the non-acid generating mine rock. The co-mingling process is discussed in the Volume 1 of the Golder report (Appendix 6).

Golder estimates that the total volume required for disposal of all tailings and mine waste rock is 31.679 million m3.

The layout of the CDA is shown in Figure 8.1.

8.12.1 Filling Plan

The CDA has been divided into two areas in order to facilitate deposition of tailings and mine rock. The low central valley section will be used primarily for acid generating rock and high sulphur tailings. A dam will be constructed across the western end of the CDA in order to flood the valley and create a pond over the acid generating material in order to prevent oxidation and acid generation. The pyrrhotite tailings will be deposited subaqueously in slurry form. The remaining area will be used for the non-acid generating rock and pyrrhotite-reduced tailings. The non-acid generating tailings will be thickened and disposed of with the non-acid generating rock.

Golder notes that, in the initial one or two years of operation, alternative deposition strategies for the non-acid generating materials will be evaluated in order to determine the most efficient method of mixing tailings and rock. A combination of rock dumping, spigotting, dozer pushing and other methods will be tested.

The acid-generating materials will be placed in the flooded basin with the mine rock placed at the west end near the main dam, and the slurried tailings placed at the east end. In order to facilitate placement, the rock will be end-dumped from a platform about 1 m above the operating water level. This will be kept about 2 m below the final water level.

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Figure 8.1 Layout of Project Site showing Co-Disposal Area

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In addition to the CDA facility, the mined out West pit will also be used for storage of approximately 0.67 million m3 of acid-generating material. The West pit will be used as follows. At the beginning of Year 4, acid-generating rock will be stored temporarily near the West pit and will be pushed into the West pit after Year 5. From the end of Year 5, both pyrrhotite tailings and acid-generating rock will be disposed of entirely in the West pit which will be permanently flooded. During Years 4 and 5, only pyrrhotite tailings will be deposited in the central section of the CDA.

8.13 ENVIRONMENTAL PLANNING

In its January 9, 2006 report, Golder noted the following items to be addressed as the Shakespeare project is moved forward:

• Conduct a specific assessment of the seismic risk for the Shakespeare site.

• Carry out subsurface investigations and evaluate borrow sources for construction of dams and the foundation of the thickening plant.

• Finalize the dam/dyke design and stability analyses.

• Prepare construction specifications and contract documents for dams, borrow sources and construction access roads.

• Complete geochemical characterization testwork including:

• Elemental and more detailed acid base accounting (ABA) analyses on existing samples.

• Perform additional leach tests.

• Carry out planned humidity cell testing to assess the geochemical characteristics of the co-mingled mine rock and sulphur-reduced tailings.

• Continue test program underway to assess the characteristics of the combined flotation, sulphur reduced and pyrrhotite tailings.

The elemental analyses, ABA and leach testing are required in order to comply with the Mining Act of Ontario. Part of this work is already scheduled (at January, 2006).

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9.0 SITE WATER MANAGEMENT

The site water management plan has been prepared by Golder and presented in its report which is provided in Appendix 7.

9.1 WATER MANAGEMENT

Six ponds are required for the water management on the Shakespeare property:

• Pond in the co-disposal area behind the main dam that will cover the acid generating materials.

• Settling pond immediately downstream of the co-disposal facility.

• Polishing pond for water that is recycled to the mill and/or discharged to the environment.

• Mill (plant) site collection pond.

• External pond at the east end of the co-disposal facility that will collect seepage from the co-disposal area in the later years of operation.

• Aging pond near the mill that provides time to age the process water before it is recycled to the mill.

All the water that is collected on the property is passed through the settling pond before being treated. The treatment plant, consisting of a lime storage and dosing system, is located between the settling pond and the polishing pond. It has to be operational through the winter months because water has to be continually recycled to the mill.

The aging pond is sized to retain process water for thirty days before recycling it to the mill. The critical period is just before freeze up in the fall when there must be enough water in the pond to run the mill through the winter months when there is no runoff.

For all the other ponds, the critical period is during the spring runoff at the end of April or early May when the winter snow melt runs off and accumulates in the ponds. These ponds will be drawn down to their lowest levels in the fall.

9.2 DAMS AND DYKES

A total of seven dams and dykes are required to manage the water on site:

• Main dam at the west end of the co-disposal area to permanently maintain the acid generating materials in a flooded/submerged state.

• Settling pond dam.

• Polishing pond dam.

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• Mill (plant) site collection pond.

• Two dams around the aging pond near the mill.

• Later in the mine life a seepage collection pond dam is required at the east end of the co-disposal facility.

The topographies at all the dam and dyke locations are amenable for conventional, zoned water retaining dams founded on overburden or grouted bedrock. The main elements of a zoned embankment dam are a barrier to retain water and supporting shells to support the barrier. Internal drains, transition and filter zones minimize the effects of cracking and inhibit the migration (piping) of fine soil particles under the influence of seeping water. A dam is normally defined as a barrier across a water course and a dyke infills a topographic low.

9.3 FLOW MODEL

A site wide deterministic flow (water balance) model has been developed for the site to size ponds and water management systems. The input data can be changed to investigate worst case and what if scenarios. The model is simple to use. Sensitivity analyses can be carried out and it can be used by mine personnel over the life of the mine. Ultimately, a probabilistic model might be required to model contaminant transport loadings and to predict the environmental and hydraulic performance over the life of the mine.

There is a large variation in flows on the property depending on the precipitation (100-y wet and dry and mean) and various stages in the life of the mine. For the 100-y dry return at the mid-life of the project, the discharge to the environment is calculated to be 1.3 million m3/yr and for the 100-y wet return it is 2.5 million m3/yr. The mean is about 1.8 million m3/yr.

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10.0 ENVIRONMENTAL PERMITTING AND MANAGEMENT

Various environmental baseline studies were completed for the Shakespeare Project in 2004 and 2005. N.A.R. Environmental Consultants Inc. (NAR) undertook several formal and informal desktop reviews to provide input on environmental permitting and present and future environmental management issues for the preliminary feasibility study. In August, 2004, several specific technical tasks were implemented by NAR to determine baseline conditions both within the physical limits of the project and the zone of potential impact, notably the receiving water environment (Agnew Lake).

Knight Piésold Ltd. was retained by URSA in early 2005 to provide environmental permitting and management services for the Shakespeare project as part of the preparation of the bankable feasibility study and with the understanding that NAR would be retained directly by Knight Piésold to continue with the baseline studies initiated in 2004.

The field program was expanded under the responsibility of Knight Piésold in early- to mid-2005 to include the collection of benthic macroinvertebrate community and sediment quality data from three stations in Agnew Lake, installation of several groundwater monitoring wells and collection and testing of groundwater samples, acid rock drainage testing of mine rock and tailings samples and the installation of an on site weather station. The surface water quality and quantity monitoring program continued. In the early fall of 2005 terrestrial habitat and heritage studies were completed for the project site.

Knight Piésold prepared an Environmental Baseline Report for Feasibility Study on behalf of URSA, dated December 15, 2005. Volume 2 of this report contains the appendices, provided as follows for reference (see Appendix 13).

Appendix A: Surface Water Quality and Quantity, Groundwater Quality and Sediment Quality Results

A1 - Surface Water and Sediment Quality Results

A2 - Flow Monitoring Records

A3 – Groundwater Quality Results

Appendix B: Geotechnical Drilling and Monitoring Well Installation Summary for the Shakespeare Project (Knight Piésold)

Appendix C: Pre-Operative Environmental Baseline Study – Shakespeare Deposit, Espanola, Ontario (NAR)

Appendix D: Terrestrial Assessment – URSA Major Access Corridor (NAR)

Appendix E: Terrestrial Habitat Assessment of Shakespeare Project Site – URSA Major Minerals Incorporated (Maret Tae)

Appendix F: Acid Rock Drainage and Metal Leaching Characterization Interim Reports (SGS)

F1 – Waste Rock Interim Report 5

F2 – Tailings Interim Report 3

Appendix G: Meteorological Data

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Appendix H: Stage One and Two Archeological and Heritage Impact Assessment – Proposed Agnew Lake Mine Development, Agnew Lake, Ontario (Horizon Archeology).

A summary of the results of the baseline studies is provided below.

The results of the surface water, groundwater and sediment sampling programs are considered as background values prior to the pre-production and production phases of the project.

10.1 SUMMARY OF BASELINE STUDIES

10.1.1 Surface Water

Agnew Lake was created when the Spanish River was dammed for electric power generation. The Spanish River flows south from Biscotasi Lake to become Agnew Lake which flows eastwards and then southwest past Espanola to discharge into the North Channel of Lake Huron.

A number of small streams, Order 1 or intermittent, drain the Shakespeare project area.

Monitoring of surface water quality was initiated by NAR in 2004. URSA has monitored surface water quality at stream locations on-site and NAR has collected samples from reference, near-field and far-field stations located on Agnew Lake, the John’s Creek outlet, Spellman’s Cove, Stumpy Bay and Long Bay, see Figure 10.1.

In general, most surface water parameters tested for were well below Provincial Water Quality Objectives (PWQO) and Metal Mining Effluent Regulations (MMER). Some of the surface water samples collected on site (UM-SW-1 through UM-SW-6) exceeded the PWQO for aluminum, ammonia (as N), cadmium, cobalt, copper, iron and zinc. In general, the pH values were slightly acidic for the samples collected from on-site locations, ranging from 4.57 to 7.40. There were single sample exceedances for PWQO for lead (location UM-SW-4), nickel (location UM-SW-3A), silver (location UM-SW-5) and vanadium (location UM-SW-1). The only parameter to exceed MMER was pH where values were slightly acidic as indicated above.

For samples collected from Agnew Lake, including where John’s Creek flows into the lake, only aluminum (location JC-FF-b), ammonia (as N), cadmium (locations UM-AL-REF-b and UM-AL-FF-b), copper and pH (location JC-FF-b) exceeded PWQO limits. For samples collected from the three embayments, only ammonia (as N) exceeded the PWQO limit. There were no exceedances of MMER for any of these samples.

Flow monitoring is carried out at surface water monitoring location UM-SW-3, in Spellman’s Cove. Average recorded flows ranged from zero in August and September, 2004 to approximately 0.16 m3/s in April, 2005. Data collection is ongoing.

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Figure 10.1 Plan Showing Water Quality Monitoring Stations

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10.1.2 Groundwater

A total of 17 groundwater monitoring wells were installed in 2005. Hydraulic conductivities (K) estimated from the rising head tests completed in the wells ranged from 3.4 x 10-6 cm/s in MW-05-01 to 3.1 x 10-4 cm/s in MW-05-07 (average of approximately 7 x 10-5 cm/s). These values are typical for near surface bedrock and the overburden soils encountered at site.

Groundwater samples were collected in August and September, 2005. Some of the groundwater samples (MW-05-01 through MW-05-17) exceeded the PWQO for aluminum, ammonia (as N), arsenic, cadmium, cobalt, copper, iron, total cyanide, tungsten and zinc. It should be noted that for some of the samples the method detection limit for total cyanide was greater than the PWQO. The pH values were slightly acidic for almost half of the samples collected, with lower values ranging from 4.91 to 6.43. There were single sample exceedances for PWQO for lead (MW-05-15), mercury (MW-05-03) and uranium (MW-05-01) and two samples exceeded for nickel (MW-05-01 and MW-05-14) and silver (MW-05-05 and MW-05-15).

A single vertical drill hole (U-03-66) was completed in the area of the bulk sample (between the East and West pits) with the principal objective of confirming the depth to groundwater and the geology of the bulk sample. Groundwater was encountered at a depth of approximately 1.2 m. A monitoring well was not installed and quality sampling was not carried out.

10.1.3 Sediment

Sediment samples were collected by NAR from the three surface water monitoring locations in Agnew Lake in October, 2004 and from locations in Stumpy Bay, Spellman’s Cove and Long Bay in May, 2005. Results were compared to the Provincial Sediment Quality Guidelines (PSQG) which consist of lowest effect and severe effect levels for various parameters.

For some or all of the samples collected from Agnew Lake, the lowest effect levels were exceeded for percent organic, arsenic, cadmium, chromium, copper, lead, manganese, nickel, total kjeldahl nitrogen and zinc. All three far-field samples exceeded the severe effect levels for percent organic while two of the reference and two of the far-field samples exceeded the severe effect level for manganese.

For the samples collected from the embayments (Stumpy Bay, Spellman’s Cove and Long Bay), the lowest effect levels were exceeded for percent organic, arsenic, cadmium, chromium, copper, lead, manganese, nickel, total kjeldahl nitrogen and zinc for some or all of the samples collected. One or more of the samples from each embayment exceeded the severe effect level for manganese while one from Spellman’s Cove and two from Stumpy Bay exceeded the severe effect level for total kjeldahl nitrogen.

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10.2 TERRESTRIAL HABITAT

The local vegetation consists of relatively dense bush comprising coniferous and deciduous bushes and small and large trees. Timber resources are principally second growth birch, poplar, oak, maple, jackpine and spruce.

A terrestrial habitat assessment was completed by Maret Tae, R.P.Bio., in the early fall of 2005. The report concludes that the project site “exhibits vegetation typical of the landscape of the north shore of Lake Huron. Forest habitat is primarily Transitional Boreal/Great Lakes-St. Lawrence Forest. Close to half the site is covered in very shallow soil with bedrock outcrops with Shallow Soil vegetation community dominated by sometimes stunted jack pine and red oak. The site also has a small amount of swamp and marsh habitat, and is adjacent to a large open water marsh to the northwest on Sutherland Creek.” Recommendations were made to conduct vegetation surveys throughout the growing season in order to better document the presence or absence of significant plant species at the project site.

The terrestrial habitat assessment also included mammals, birds and herpetiles. The report concluded that the project site exhibits a variety of wildlife habitat, both upland and wetland, typical of the landscape of the north shore of Lake Huron. The presence of the cliffs and the adjacent Agnew Lake are valuable features for wildlife such as raptors. However, due to the timing and duration of the assessment, it was recommended that additional wildlife surveys, such as amphibian and reptile surveys, waterfowl, raptor and spring songbird surveys, mid-summer mist netting of bats and reporting of on-site wildlife sightings on an ongoing basis be conducted in order to provide more complete inventories of the wildlife in and around the project site.

In 2004, Ministry of Natural Resources (MNR) Values Maps were reviewed for the project site area by NAR. No areas of nesting, breeding or significant bird habitat were identified by NAR.

10.3 AQUATIC HABITAT

NAR provided information on some fish spawning areas, in particular Northern Pike and Walleye, based on its review of MNR Values Maps. However, due to the drawdown of Agnew Lake caused by the operating of the hydropower dam, which can vary water levels in the lake each year by as much as 3 to 6 m, NAR’s assessment concluded that “fisheries resources in Agnew Lake are generally limited by the loss of littoral habitat and the associated benthic productivity.”

Benthic samples were collected from three locations in Agnew Lake by NAR in the spring of 2005, as well as from Spellman’s Cove, Stumpy Bay and Long Bay. NAR’s report stated that “in general, benthic macroinvertebrate communities at both the lake and embayment stations were dominated by worms, midge and clams. Results were typical of unimpacted shield lakes where numbers of organisms and taxa diversity are typically low. Communities in the embayments are also subject to stress through the annual seasonal draw-down” caused by the operation of the hydropower dam.

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A field inspection was conducted by NAR in August, 2005 on a series of ponds located north of the open pits. In its report, NAR stated that “the observational data collected during the site inspection support that ponds downstream of the proposed mine rock and tailings dump are transitory, and do not support a fishery. As such, these ponds are not fish habitat as defined by the Fisheries Act”. Also NAR reported that there was no evidence of undisturbed wetlands which supported unique plant assemblages or rare, threatened or endangered wildlife habitat.

10.4 AIR AND NOISE

No air and noise studies have been completed at the site. A scope of work has been developed to complete these studies as part of the permitting phase of the project.

10.5 MINE ROCK AND TAILINGS CHARACTERIZATION

The program for geochemical characterization of mine rock and tailings was carried out in support of the Shakespeare feasibility study but not for final project permitting. The discussion below addresses the environmental aspects of the geochemical characterization of mine rock and tailings. See Section 8 for discussion of disposal and storage of mine rock and tailings.

The program was based on the Guidelines for the Prediction of Acid Rock Drainage and Metal Leaching for Mines in British Columbia and which were adopted by the Ontario Ministry of Northern Development and Mines (MNDM) by regulation under the Ontario Mining Act.

10.5.1 Mine Rock

Core samples collected during exploration drilling provided the material for testing mine rock. Initial leach and ABA tests were completed by Testmark Laboratories Ltd. (Testmark) in September, 2004, the results of which were included in the Shakespeare pre-feasibility study and the closure plan for advanced exploration for the project.

Further testing, of which the humidity cell (kinetic ABA) testing is ongoing, has been undertaken by SGS and includes:

• US Environmental Protection Agency (EPA) toxicity characteristic leach procedure (TCLP) method 1311 on 16 samples.

• EPA synthetic precipitation leach procedure (SPLP) method 1312 on 16 samples.

• Modified static ABA tests on 10 samples.

• Net Acid Generation (NAG) tests on 10 samples.

• Humidity cell testing (kinetic ABA) on 4 individual samples and 3 composite samples.

As noted in Section 8 above, the rock samples were representative of the principal bedrock units at the Shakespeare site:

• Hanging wall quartzite

• Disseminated sulphide mineralization.

• Footwall gabbro.

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• Quartz gabbro.

Results of the TCLP tests showed all parameters, with the exception of pH, to be within the MMER limits. Results of the SPLP tests showed all parameters to be within the MMER limits.

Modified static ABA and NAG tests for the mine rock have all indicated the potential for acid generation, with the exception of the NAG test on footwall gabbro which indicated a slight potential for acid generation.

As of the time of writing, humidity cell tests (kinetic tests) were in the 30th week of testing of a 40-week testing program. The kinetic test results show a general decrease in pH from 6.75 (week 0) to 4.71 (week 30) for the quartzite material and 9.32 (week 1) to 6.21 (week 30) fo rthe gabbros. Sulphate concentrations have shown a constant to decreasing trend over 30 weeks of testing; however, concentrations are low for both quartzites and gabbros, generally less than 25mg/l. Kinetic test results indicate a potential for acid generation for the mine rock. The quartzites have no neutralization capacity; their generation of acidity is not expected to last for extended periods of time. The gabbros, while not generating acidity thus far in the test program, are expected to do so over an extended period in the future. Final test results were pending as of January, 2006.

The incomplete humidity cell test results indicate that leachate quality is within MMER limits.

10.5.2 Tailings

Geochemical characterization of samples generated from the metallurgical testwork on the Shakespeare ore is either completed or ongoing by SGS. Tests included:

• Mineralogical examination. • Rietveld X-ray diffraction (XRD) analyses. • Whole rock analyses. • ICP-OES/MS strong acid digest trace metal scans. • Modified ABA tests. • NAG tests. • US EPA TCLP method 1311. • US EPA SPLP method 1312. • Humidity cell (kinetic ABA) testing (ongoing). • Supernatant aging tests (ongoing). • Daphnia magna LC50 acute lethality tests (ongoing).

As noted in Section 8, the tailings samples represented combined flotation tailings (sample F19) and rougher tailings (F30, sulphur reduced rougher flotation tailings).

Generally the results from the testing carried out on the sulphide tailings (F19 tailings) indicated that they have a strong potential for acid generation and may leach nickel and zinc. The kinetic test results for the first 30 weeks of testing show a general decrease in pH from 8.03 (week 0) to 6.0 (week 30) and sulphate concentrations have steadily increased from

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6.5 mg/l (week 0) to 75 mg/l (week 30). These kinetic test results indicate that the F19 tailings have a potential for acid generation. Final results were pending as of January, 2006.

Results for the sulphide reduced tailings (F30 tailings) indicate acid generation is unlikely. However, results from the TCLP testing indicated the F30 tailings may have the potential to leach zinc. The kinetic test results for the first 28 weeks of testing show a consistent near-neutral pH and sulphate concentrations have shown a steady decrease from 33 mg/l (week 0 and 1) to 0.56 mg/l (week 28). These kinetic test results indicate acid generation is unlikely for the F30 tailings. Final results were pending as of January, 2006.

10.6 METEOROLOGY

The climate in the Shakespeare Property area is characterized by moderately long, cold winters and shorter, warm summers and is typical of continental conditions. A weather station was installed on site in March, 2005 to monitor local weather conditions including rainfall, temperature and wind. During the period of record, from March 17 through November 16, 2005, temperatures ranged from -12.3°C in March to 33.2°C in July. Total monthly rainfall ranged from 0 mm in March to 89.4 mm in September.

A wind rose plot was developed using the recorded wind speeds and directions. As evidenced by the predominantly east and west wind directions shown on the wind rose plot, the wind data recorded by the weather station may be influenced by the location of the cleared corridor for the access road to the project site. The predominantly east and west directions may also be a result of the fact that data has not currently been collected over the winter months.

Long term rainfall data (1971 through 2000) are available for the meteorological station at Sudbury airport, approximately 85 km northeast of the Shakespeare site. Estimated average annual precipitation is 899 mm with 657 mm falling as rain and the remainder (242 mm water equivalent), falling as snow.

Evaporation data are available from Amos, Quebec, approximately 380 km northeast of the Shakespeare project. The recorded average annual evaporation is 746 mm (1968 through 1992). Knight Piésold reported that the evaporation data from Amos may be used for design purposes.

10.7 HERITAGE STUDY

Stage 1 and Stage 2 archaeological and heritage impact assessments were completed for the project site by Horizon Archaeology in early fall 2005. Based on the studies completed, they concluded that due to the changes in the water level of Agnew Lake and the inaccessibility of the rocky ridges, the likelihood of discovering cultural remains appeared low. Horizon Archeology notes that “the original shoreline would have been a considerable distance further away than is the case today. The lack of appreciable soil deposits also limited the usefulness of this area…the test pit strategy employed by Horizon Archaeology failed to uncover any signs of cultural activity, nor even locate areas where probability modelling would indicate special need be given.” The report stated that, “in the opinion of Horizon Archaeology, there are no concerns related to the destruction of cultural materials by the continued development of this project.”

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11.0 CAPITAL EXPENDITURES

11.1 BASIS OF ESTIMATE

11.1.1 Currency Base Date and Exchange Rate

The currently base date for the cost estimate is the last quarter of 2005. The estimate is expressed in Canadian dollars. No allowances for escalation or currency fluctuation are included.

The exchange rate used is US $0.80/CA $1.00 when quotations were received in US dollars.

11.1.2 Labor Hourly Rate

Labor rate was established as an all-inclusive hourly rate by considering the basic hourly rates given by a local contractor for the tradesman, foreman and superintendent. Additional costs added were the traveling and pension, the direct and indirect supervision, the contractor’s site establishment facilities and some provision for mobilization and demobilization. The calculation method and figures were then confirmed with a qualified local contractor for such a project. The working calendar was defined as regular 8 hours a day, 5 days a week and productivity was kept as 1.0.

11.1.3 Civil Work, Concrete Quantities and Unit Costs

Quantities for site preparation civil work were calculated from plan site layout. Unit costs for the northern part of Ontario were supplied by Micon/URSA for clearing and grubbing. Rock drilling, blasting and removal was quoted by qualified local contractors.

Quantities for excavation, backfill, culverts and manholes were calculated from layouts and test pits results made for URSA in September 2005. Unit costs for excavation and backfill, including equipment and manpower, were quoted by a qualified local contractor. Mine waste fill was estimated at no direct cost to the project. Costs for culverts and concrete manholes were quoted by qualified suppliers.

Quantities for buildings foundations, for slabs on-grade, elevated slabs and equipment foundations were calculated from layouts. Unit costs for concrete including cement, rebar, forms, finishing, freight, equipment and all installation work were quoted by qualified local contractors. It was assumed that: small quantities of concrete would be supplied by a batch plant in Espanola, with bulk delivery by truck from Sudbury for large quantities; 30MPa concrete is used; all concrete work will be carried out during the summer months.

11.1.4 Structural Buildings Quantities and Unit Costs

Quantities for structural steel, steel deck, stairs with handrail, handrails and grating were calculated from layouts. Unit costs including material, freight, equipment and installation work were quoted by qualified local contractors. Steel cost is considered to be subject to sudden variations.

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Quantities for buildings exterior and interior finishes were calculated from layouts. Basic supply costs were quoted by qualified local contractors and overall unit costs were calculated from Met-Chem’s in-house-database. Quantities for louvers, windows and man doors were calculated from layouts. Unit costs were taken from Met-Chem’s in-house database. Quantities for truck doors were calculated from layouts. Unit costs and installation cost were quoted by a qualified supplier.

11.1.5 Process Equipment

All process equipment is new except for the regrind mill, for which a commitment was agreed with North American Palladium (NAP). A process equipment list was derived from flowsheets and is provided in Appendix 3. For the SAG and ball mills, the pebble crusher and the jaw crusher, three quotations were received and the lowest quotation was used for the capital cost estimate except for the pebble crusher. Single source quotes were obtained for 95% of the equipment. The remaining equipment was estimated from Met-Chem’s in-house database of similar projects.

Freight was either quoted by the suppliers or indicated as 5% of the equipment value. Equipment installation hours were either quoted by the suppliers or estimated from Met-Chem’s in-house database for similar projects.

11.1.6 Piping and Pipelines

Quantities for large bore process and water piping were calculated by take-off from flowsheets and layouts. Quantities for service piping and small bore lines were factorized from the take-off. Percentage of total equipment direct cost was also taken into consideration to estimate the overall piping cost. Installation man-hours were estimated from in-house database. The cost includes supply and installation of piping, flanges and couplings, fittings and valves, secondary steel and supports and freight.

Pipelines were estimated as HDPE pipes supplied in 50 foot lengths flanged at both ends. Quantities for mine water pipeline, fresh water line, polishing pond, aging pond and collecting pond reclaim water and the fire water loop were calculated from layouts. Quantities for East and West pits sumps to pond pipelines were provided by P & E Consulting. Unit cost for HDPE pipelines and fittings supply and freight were quoted by a qualified supplier. Installation and bolt-up man hours were estimated from in-house database.

Sanitary waste water pipeline was estimated as PVC DR-35. Quantities were calculated from layouts. Unit costs were quoted by a qualified supplier. Installation man hours were estimated from in-house database.

11.2 PROJECT CAPITAL COSTS SUMMARY

The total pre-production capital expenditure for the Shakespeare property, at a production rate of 4,500 t/d, is estimated at Cdn$118.5 million (including contingencies, expressed in dollars of mid-2005 constant money terms). This estimate is summarized in Table 11.2.1.

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The following sections summarize the capital cost estimate for the mining, plant, tailings and infrastructure sections of the project. The mining capital estimate was prepared by P&E, plant and infrastructure by Met-Chem, and tailings manangement by Golder. Micon provided overall supervision and coordination between the parties.

Table 11.2.1 Total Initial Capital Cost Estimate

Cdn$000 Mining Equipment 18,916 Pre-Stripping 10,784 Process Plant direct costs 48,978 Infrastructure direct costs 8,819 Tailings Dams 6,672 High Compression Thickener (HCT) Plant 3,506

Total Direct Costs 97,675 Owner’s Costs 4,643 EPCM 7,965 Contingency 8,190

Total Initial Capital Costs 118,473

The contingency was calculated as a specific percentage of each of the direct cost items. The provision of $8.19 million represents an average of 12% of the direct cost estimate for the process plant, infrastructure, the tailings co-disposal area and HCT plant.

11.3 MINING CAPITAL COSTS

Mining capital costs are shown in Table 11.3.1. In addition to the equipment costs listed, provision has been for pre-stripping of 10 million tonnes of waste from the open pit ahead of production. The direct cost of consumables for this is estimated to be Cdn$10,784,000 and this cost is reflected in the capital estimate above. Associated labour costs of Cdn$6,158,000 are included in mine operating costs during Year 1.

Certain used equipment has been identified as being available for sale to the project and this equipment, after being rebuilt, would be used in place of similar, new units. The cost saving from this procurement strategy is reflected in the cost estimate. All other equipment is priced new, based on supplier quotations obtained by P&E.

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Table 11.3.1 Mining Equipment Capital (Cdn$, 2005 terms)

Start-up Capital Sustaining

Equipment Type Make/Model Units Year 1 Year 2 Year 3 Years 4-8 Shovel O&K RH90-C 1 2,157,487 Wheel loader (rebuilt) Komatsu WA800 1 1,000,000 Truck, haulage (rebuilt) Cat 777 3 2,850,000 Truck, haulage Cat 777F 3 1,253,882 2,507,764 Wheel loader, stockpile, etc. Cat 988H 1 662,982 Drill, production Reeddrill SKF-12 2 2,714,820 Dozer, track (rebuilt) Cat D10 1 800,000 Dozer, track Cat D9T 1 923,691 Excavator, general Cat 365CL 1 712,103 Grader Cat 16H 1 803,341 Crawler drill & compr. P&E Estimate 1 300,000 Water truck Estimate 1 500,000 Capital spares 5% 542,287 307,267 15,000 Service truck P&E Estimate 1 100,000 Mechanic truck P&E Estimate 1 200,000 Pick-up truck P&E Estimate 7 90,000 120,000 210,000 Lighting tower Frontier PowerT. 4 41,000 41,000 Pump system Flygt 68,600 44,600 208,400 Office equipment Allowance 1 105,000 Survey equipment Allowance 1 40,000 Freight Allowance 10,000 5,000 500 750 Total 11,937,621 6,618,602 360,100 419,150

11.4 PROCESS PLANT CAPITAL COSTS

Direct capital costs include costs of material and equipment, installation and freight charges. Owner’s costs, EPCM, and contingencies are treated as indirect costs. Table 11.4.1 summarizes the direct capital costs (only) for the process plant.

Details of the cost estimate, prepared by Met-Chem, are given in Appendix 3.

Table 11.4.1 Process Plant Direct Capital Cost Estimate (Cdn$, 2005 terms)

Direct Capital Cost - before contingency Total (Cdn$) Plant Site Preparation 2,901,670 Primary Crushing Area 3,242,453 Reclaim Tunnel Area 1,317,648 Pebble Crushing Area 1,606,678 Concentrator Area 27,077,318 Global Costs (Piping & Electricity) 12,832,028 Process Plant Direct Total 48,977,795

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11.5 INFRASTRUCTURE CAPITAL COSTS

Table 11.5.1 shows the direct capital costs of infrastructure associated with establishing the mine and process plant site at Shakespeare Project.

Table 11.5.1 Infrastructure Direct Capital Costs

Direct Capital Cost - before contingency Total (Cdn$) General access road (3.43 km) 224,913 Office complex 1,685,879 Mine equipment maintenance building 2,804,714 Fuel storage and fuelling system 598,330 Water systems 2,048,060 Effluent treatment area 523,801 Bulk Explosives Mixing Bldg (Services) 17,435 Main substation 114,480 Truck scale 160,500 Plant mobile equipment/laboratory equipment 640,628 Infrastructure Total 8,818,740

11.6 TAILINGS MANAGEMENT AREA CAPITAL COSTS

The co-disposal of mine rock and flotation tailings requires dewatering of the non-sulphide tailings ahead of deposition. Table 11.6.1 reflects the capital cost of establishing the tailings dam and associated ponds, pipelines, plus the cost of the high-compression thickener plant.

Table 11.6.1 Tailings Area Direct Capital Costs

Direct Capital Cost - before contingency Total (Cdn$) Tailings dams 6,672,124 High compression thickener (HCT) 3,506,010 Total Tailings Management 10,178,134

11.7 CAPITAL COST SUMMARY SCHEDULE

Table 11.7.1 provides a breakdown of annual expenditure for each of the main project areas, including the provision for indirect capital costs (i.e., owner’s costs, EPCM, permitting and contingencies).

Note that escalation is not provided for, as the economic evaluation has been carried out in constant 2005 dollar terms.

Figure 11.1 shows the construction schedule for the Shakepeare Project.

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Table 11.7.1 Capital Cost Summary Schedule

(Thousand Cdn$, constant 2005)

Year 0 Year 1 Year 2 Year 3 Total Mining equipment 11,938 6,619 360 18,916 Pre-stripping 10,784 10,784 Subtotal – Mining 0 22,722 6,619 360 29,700 Process plant 14,693 34,284 48,978 Infrastructure 2,646 6,173 8,819 Tailings management 3,053 7,125 0 0 10,178 Subtotal – Direct Costs 20,392 70,304 6,619 360 97,675 Owners costs 1,393 3,250 4,643 EPCM 2,389 5,575 7,965 Contingency 2,457 5,733 8,190 Total 26,632 84,862 6,619 360 118,473

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12.0 OPERATING COSTS

12.1 MINING OPERATING COSTS

The mining costs reflect the proposed overall mining rate of 35,000 t/d, and an ore production rate of 4,500 t/d, for the fleet as described under Mining Capital Cost. Table 12.1.1 shows the cost of consumables for operation of the open pit, per tonne of rock moved.

Table 12.1.1 Mining Consumables Cost/tonne moved (Ore+Waste)

Cdn$/t Fuel 0.43 Drill steel and bits 0.03 Explosives 0.40 Other consumables 0.21 Total 1.08

In addition to the consumable costs, labour costs have been allowed for 94 persons as shown in Table 12.1.2.

Table 12.1.2 Mine Operating Costs, including Labour

Payroll

Complement Annual cost

Cdn$ Cost per tonne milled (Cdn$/t)

Ore mining consumables 1,773,900 1.08 Waste mining consumables 9,135,585 5.56 Mining production labour 69 4,749,460 2.89 Mining maintenance labour 13 995,680 0.61 Mining engineering, geology 12 1,012,888 0.62 Total cash mining cost 94 17,667,513 10.76

12.2 ORE PROCESSING COSTS

The estimated average life-of-mine cash operating costs for the on site 4,500 tonnes per day processing option are summarized in Table 12.2.1.

Basis of the estimate is current costs (2005) including freight for all reagents and consumables, plus power provided for at the rate of Cdn$0.10 per kWh. The latter unit cost was obtained as a preliminary estimate from Hydro One.

Details of the operating costs estimate are provided in Appendix 2b.

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Table 12.2.1 Process Operating Costs

Area Payroll

Complement Total Cost

(Cdn$ per year) Unit Cost

(Cdn$/t ore) Plant labour 44 3,286,404 2.00 Reagents 1,427,434 0.87 Media 4,961,123 3.02 Spares and laboratory 1,631,650 0.99 Site power 5,060,245 3.08 Building heating 200,000 0.12 Concentrate shipment 537,097 0.33 Total 17,103,953 10.41

12.3 GENERAL AND ADMINISTRATION OPERATING COSTS

The estimated average life-of-mine General and Administrative operating costs are summarized in Table 12.3.1.

Table 12.3.1 General and Administrative Operating Costs



Unit Cost (Cdn$/t ore)

G&A labour 14 1,050,124 0.64 Other site costs 499,000 0.30 Corporate office 650,000 0.40 Total 2,199,124 1.34

12.4 SUMMARY OF OPERATING COSTS

Table 12.4.1 summarises the cash operating costs of the Shakespeare Project at steady state.

Table 12.4.1 Summary of Operating Costs



Unit Cost (Cdn$/t ore)

Mining labour 94 6,758,028 4.11 Plant labour 44 3,286,404 2.00 G&A labour 14 1,050,124 0.64 Subtotal labour costs 152 11,094,556 6.75 Mining consumables 10,909,485 6.64 Plant consumables, etc 13,817,549 8.41 G&A other 1,149,000 0.70 Total 36,970,590 22.51

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13.0 ECONOMIC ANALYSIS

13.1 MACRO-ECONOMIC ASSUMPTIONS

13.1.1 Metal Price Forecasts

The economic analysis takes as its base case the assumption of a reversion of metal prices to their long-term (in this case, 10-year) historical median Canadian dollar prices, expressed in 2005 money terms. Current price levels, whether lying above or below the 10-year median price, are assumed to regress exponentially toward the median, with a ‘decay’ half-life of five years. The resulting average prices over the life of the project, expressed in 2005 US dollars, are given in Table 13.1.1.

Table 13.1.1 Average Metal Price Forecasts

(Life of Mine, 2005 constant dollars)

Metal Unit Price

Nickel US$/lb 5.48 Copper US$/lb 1.34 Cobalt US$/lb 20.05 Platinum US$/oz 805.30 Palladium US$/oz 225.20 Gold US$/oz 438.30

13.1.2 Dollar Exchange Rate

The base exchange rate for the economic analysis is the average US$/Cdn$ rate of 0.8224 for 2005.

13.1.3 Escalation and Money terms

No escalation has been applied to the model, which is presented in mid-2005 constant money terms.

13.2 BASE CASE MODEL

The following technical and commercial assumptions are contained within the base case economic model.

13.2.1 Working Capital

Working capital computations assume 80% of metal revenue received on delivery, and 20% paid after 120 days. Accounts payable are assumed to amount to 30 days of consumable costs.

13.2.2 Taxation

Canadian Federal and Ontario Provincial tax regimes are assumed to apply.

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13.2.3 Discounted Cash Flow Valuation

Table 13.2.1 shows a summary of NSR calcualtion and the annual revenues accruing to the project. Table 13.2.2 shows a summary of the annual cash flows and the calculation of base case net present value (NPV) and internal rate of return (IRR).

Table 13.2.1 NSR Calculation Schedule


METAL PRICES Nickel ( US $ / pound ) 6.14 5.91 5.72 5.56 5.42 5.30 5.20 5.12 5.05 5.48 Copper ( US $ / pound ) 1.49 1.44 1.39 1.35 1.32 1.29 1.27 1.25 1.23 1.34 Cobalt ( US $ / pound ) 17.86 18.54 19.17 19.73 20.23 20.67 21.07 21.42 21.73 20.05 Platinum ( US $ / ounce ) 852.74 836.67 822.93 811.16 801.04 792.34 784.84 778.37 772.78 805.3 Palladium ( US $ / ounce ) 210.24 215.08 219.38 223.19 226.57 229.55 232.17 234.48 236.51 225.2 Gold ( US $ / ounce ) 440.27 439.62 439.07 438.58 438.16 437.79 437.47 437.19 436.95 438.3 Exchange Rate ( US $ / Cdn. $ ) 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224 0.8224

Nickel ( Cdn. $ / pound ) 7.46 7.19 6.95 6.76 6.59 6.45 6.33 6.23 6.14 6.67 Copper ( Cdn. $ / pound ) 1.81 1.75 1.69 1.65 1.61 1.57 1.55 1.52 1.50 1.62 Cobalt ( Cdn. $ / pound ) 21.71 22.55 23.31 23.99 24.59 25.14 25.62 26.04 26.42 24.38 Platinum ( Cdn. $ / ounce ) 1,036.91 1,017.38 1,000.67 986.35 974.05 963.47 954.35 946.48 939.69 979.17 Palladium ( Cdn. $ / ounce ) 255.65 261.53 266.76 271.40 275.50 279.12 282.32 285.13 287.60 273.87 Gold ( Cdn. $ / ounce ) 535.36 534.58 533.90 533.30 532.79 532.34 531.95 531.61 531.32 532.96Gross Revenue from Metal Sales Nickel - 65,869 68,112 56,657 44,374 46,063 54,141 45,289 - 380,504 Copper - 18,554 20,967 18,716 14,759 14,728 17,173 14,693 - 119,591 Cobalt - 7,049 7,380 7,055 6,497 6,581 7,637 6,431 - 48,630 Platinum - 11,476 12,474 12,280 9,926 8,846 10,678 9,098 - 74,779 Palladium - 1,838 2,132 2,127 1,686 1,528 1,929 1,689 - 12,928 Gold - 1,633 1,865 1,850 1,492 1,385 1,768 1,586 - 11,580GROSS SALES REVENUE - 106,419 112,930 98,685 78,734 79,131 93,326 78,787 - 648,011Smelting - 11,695 13,122 11,690 9,419 9,762 11,635 10,020 - 77,344Refining - Nickel - 5,867 6,269 5,366 4,309 4,571 5,475 4,656 - 36,513Refining - Copper - 3,291 3,839 3,523 2,845 2,900 3,443 2,993 - 22,834Refining - Cobalt - 947 959 891 800 793 903 748 - 6,043Refining - Platinum - 206 227 227 186 167 204 175 - 1,393Refining - Palladium - 128 146 143 112 100 125 108 - 861Refining - Gold - 56 64 63 51 47 61 54 - 396Falconbridge Royalty - 1,263 1,325 1,152 915 912 1,072 900 - 7,539NET REVENUE (NSR) - 82,965 86,979 75,630 60,096 59,878 70,408 59,131 - 495,088

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Table 13.2.2 Annual Cash Flows and Discounted Cash Flow Valuation

(Thousand 2005 constant Canadian dollars)

Year0 Year1 Year2 Year3 Year4 Year5 Year6 Year7 Year8 Year9 Total 2006 2007 2008 2009 2010 2011 2012 2013 2014 2015NET REVENUE (NSR) - 82,965 86,979 75,630 60,096 59,878 70,408 59,131 - 495,088 Waste Mining (consumables) - 11,218 11,218 8,856 10,431 7,676 1,968 194 - 51,561 Ore Mining (consumables) - 1,771 1,771 1,771 1,771 1,771 1,771 1,479 - 12,106 Mining Labour 4,328 4,749 4,749 4,749 5,089 4,478 2,817 1,884 - 32,845 Fleet Maintenance Labour 907 996 996 996 1,067 939 591 395 - 6,886 Mine Eng/Geology, incl camp 923 1,013 1,013 1,013 1,085 955 601 402 - 7,005 Crushing, Milling & Flotation - 16,567 16,567 16,567 16,567 16,567 16,567 13,831 - 113,232 Concentrate Transportation - 552 619 551 444 460 549 473 - 3,648 General and Administration 1,100 2,199 2,199 2,199 2,199 2,199 2,199 302 - 14,596 Total Operating Costs 7,257 39,065 39,132 36,703 38,654 35,045 27,063 18,959 - 241,878 OPERATING PROFIT BEFORE TAX (7,257) 43,900 47,847 38,927 21,442 24,834 43,345 40,172 - 253,210 Less: Federal/Provincial Taxes 70 255 605 535 1,906 7,881 9,014 15,682 14,560 266 50,773 PROFIT AFTER TAX (70) (7,512) 43,296 47,312 37,021 13,561 15,820 27,664 25,612 (266) 202,436 Less: Initial Capital Expendit:- Mining Equipment 11,938 6,619 360.1 18,916 Pre-Stripping 10,784 10,784 Process Plant direct costs 14,693 34,244 48,978 Infrastructure direct costs 2,646 6,173 8,819 Tailings Dams 2,002 4,670 6,672 HiComprThicknr (HCT) 1,052 2,454 3,506 Owners Costs 1,393 3,250 4,643 EPCM 2,389 5,575 7,965 Contingency 2,457 5.733 8,190 Total Initial Capital Expend. 26,632 84,862 6,619 360 118,473 Sustaining/Repl.Capital Mining Equipment 133 - 286 - 419 Mine Services 75 75 75 75 300 Tailings & Water Mgmnt 165 165 165 165 165 825 Contingency/(Resid value) - 62 62 62 62 62 - (3,836) - (3,526)Permits, Recl & Closure 200 - 1,500 - 1,700 Change in Working Cap. (2,200) (5,000) 26,100 1,500 (2,800) (4,200) 100 3,400 (16,900) - - ANNUAL CASH FLOW (24,702) (87,374) 10,275 45,150 39,386 17,459 15,206 24,264 44,847 (266) 84,246 %Disc. Pre-tax Aftertax Internal Rate of Return 20.0% 14.5% Net Present value 0% 135,019 84,246 CDN $000, Real 2005 terms 5.0% 78,312 43,108 10.0% 41,187 16,242 15.0% 16,516 (1,516) 20.0% (54) (13,331)

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The cash flow projection reflects a maximum cash outflow of Cdn$112 million at the end of the construction period when pre-stripping of the open pit is complete but no milling of ore has yet taken place. Year 2 is the first year of full production, and is cash positive despite an increase in working capital. Undiscounted payback occurs at the end of Year 5, leaving almost three full years of production ‘tail’.

The base case cash flow provides a pre-tax IRR of 20.0%, or 14.5% after tax. Net Present Value (NPV) of the project at a 10% (real) discount rate is Cdn$41.2 million pre-tax, or Cdn$16.2 million after tax.

13.3 SENSITIVITY STUDY

Figure 13.1 shows the results of the sensitivity analysis, constructed by varying each factor within the economic model while holding all else equal. It shows that of a given rate of discount, the NPV is most sensitive to metal prices, grade and recovery factors. It is less sensitive to operating costs and reserve tonnage, and relatively insensitive to throughput and project capital costs.

Figure 13.1 Sensitivity Study Results

Variances of around 15% from the base would be required in the technical parameters before project returns fall below an acceptable level. The precision of the underlying estimates lies inside this range and thus the sensitivity study results suggest the impact of project-specific technical risk on the project cash flow is manageable.

Sensitivity of NPV(10%)

-40,000

-20,000

0

20,000

40,000

60,000

NPV

(10%

) Cdn

$ 00

0 (2

005

mon

ey te

rms)

Capital cost estimate 27,299 25,282 23,173 20,958 18,649 16,242 13,663 10,902 8,059 5,077 1,957

Operating cost estimate 36,634 32,907 29,014 24,949 20,689 16,242 11,469 6,477 1,217 -4,419 -10,385

Revenue drivers (Grade, Recoveries,US$ Prices, Exchange rate)

-35,143 -25,126 -15,182 -4,909 5,547 16,242 27,146 38,540 50,464 62,859 75,621

Throughput 10,966 12,156 12,626 13,656 14,738 16,242 16,606 17,047 18,742 18,670 18,611

Reserve tonnage -3,121 928 5,176 7,791 11,977 16,242 17,815 19,978 22,240 25,260 27,444

80% 84% 87% 91% 96% 100% 105% 109% 114% 120% 125%

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Metal price sensitivity was further tested by evaluating the mine with current (i.e., 2005 year-to-date) prices held constant for the life of mine. Nickel and copper prices are thus 23% and 22% higher on average over the life of mine, and the gross revenue increase is around 17%. Under this assumption, the IRR rises to 23.9% and the NPV at a discount rate of 10% per year rises by Cdn$38.7 million to Cdn$54.8 million after tax, demonstrating considerable upside potential over the base case.

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14.0 PROJECT IMPLEMENTATION SCHEDULE

14.1 GENERAL

The project implementation schedule was developed in MS Project format. The main engineering, procurement and construction activities are indicated. The schedule is based on information taken from supplier’s quotes or Met-Chem’s in-house database. The schedule presents the total duration of the construction project considering project design criteria as of November 18, 2005.

It is assumed that:-

(i) all permits required are approved and received prior to the starting date of the schedule;

(ii) construction will be carried out over 5 days/week at regular 8 hours per day;

(iii) all concrete work will be done during the summer months – i.e., no winter allowances are included for civil works namely concrete.

14.2 SCHEDULE OF ACTIVITIES

14.2.1 Engineering

In Engineering, the important activities indicated are the grinding mills specification preparation followed by major equipment specification preparation (i.e. flotation cells, crushers, compressors, etc.). The general mechanical equipment specification preparation follows. It includes all other mechanical equipment. The office complex specification is also indicated. The transformers specification preparation is next, followed by the general electrical and automation specification preparation (i.e. switchgear and motor control center etc.). The last activity is the detailed engineering that takes into account the project design criteria and all specific details from supplier’s shop drawings.

14.2.2 Procurement

The Procurement category repeats the previous activities namely the ones with specification preparation and states the fabrication and expected delivery time to site.

14.2.3 Mine Pre-Stripping

Mine Pre-Stripping indicates the duration required to have enough mine waste rock for construction and to expose sufficient ore in the pit before the start-up of the processing plant.

14.2.4 Construction

The Construction category indicates the duration for site-related activities including site preparation, civil and concrete, tailings dams construction, structural steel, exterior and interior finishes, office complex installation and mechanical, electrical and automation installation.

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Based on the purchase and installation of new equipment for all areas of the project, the estimated duration of the project implementation before start up is twenty-two (22) months.

The project implementation schedule is shown in Figure 14.1.

14.3 PRIORITY ACTIVITIES

The priority activities, which are part of the critical path, are the grinding mills specification preparation, procurement, fabrication, delivery and installation.

The project duration is dependent of the grinding mills selection and can be reduced if suitable used equipment can be found for the Shakespeare Project applications.

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Figure 14.1 Project Schedule

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15.0 CONCLUSIONS

The Shakespeare project contains an economic mineral reserve and is worthy of continued development through detailed engineering and construction to produce 4,500 t/d of mined ore and subsequent metal concentrate for sale. Environmental studies indicate that no significant negative impact from the project will be encountered. The co-disposal concept for tailings and mine rock will mitigate effects from these disposal systems.

It is unlikely that ore grade, metal recovery or operating cost can be significantly improved. The main economic improvement that may be contemplated is the reduction in capital cost by the selection of used equipment and buildings, and Micon recommends an immediate search for suitable items.

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16.0 REFERENCES

Cole, S and Williams, S. R., (2003); Metallurgical Testing of Shakespeare Project Samples for Ursa Major Minerals Incorporated, LR 10616-001 – Progress Report No. 1. SGS Lakefield Research Limited metallurgical report, 14p plus Appendices.

Dressler, B. D., Gupta V. K. and Muir T. L. (1991); The Sudbury Structure in Geology of Ontario, OGS Special Volume 4, Part 1, 1991.

Hennessey, B. T., (2003); A Mineral Resource Estimate for the East Zone of the Shakespeare Deposit, Shakespeare Property, Near Espanola, Ontario, Volumes Volume I and II.

Kallio, Eric A., 2002a; Technical Report for the Shakespeare Property, Shakespeare Township, Ontario, NTS 41I/5, for URSA Major Minerals Incorporated, Volume 1 of 2. Technical Report filed on SEDAR (www.sedar.com) on December 2, 2002.

Kallio, Eric A., 2002b; Technical Report for the Shakespeare Property, Shakespeare Township, Ontario, NTS 41I/5, for URSA Major Minerals Incorporated, Volume 2 of 2, Appendices. Technical Report filed on SEDAR (www.sedar.com) on December 12, 2002.

Lewis, C. L., (1949); A Sulphide-Bearing Quartz Diorite Intrusive in Shakespeare Township, Ontario, unpublished B Sc. Thesis, 29p.

Lightfoot, P. C. and Naldrett, A. J., (1996); Petrology and geochemistry of the Nipissing Gabbro: Exploration strategies for nickel, copper, and platinum group elements in a large igneous province; Ontario Geological Survey, Study 58, 81p.

Perkins, M. J., (2001); Report of Work completed in 2000 at the Shakespeare Project, prepared for URSA Major Minerals Inc., June 2001.

Perkins, M. J. (2002); Summary of Field Work completed in 2001 and 2002 for Shakespeare Project, prepared for URSA Major Minerals Inc., November 2002.

Sutcliffe, R. H., (2002); Annual Report for URSA Major Minerals 2002, June 2002.

Sutcliffe, R. H., Tracanelli H. J., (2003); Winter 2002/2003 Drill Program. Internal URSA Major Mineral Inc. report on exploration activities.

Sutcliffe, R. H., Tracanelli H. J., (2004); Summer and Winter of 2003/2004 Diamond Drilling, Mineral Exploration Program. Internal URSA Major Mineral Inc. report on exploration activities.

Thompson, W. H., (1985); Report on Shakespeare Nickel-Copper and Agnew Lake Regional Projects, Energy, Mines & Resources Canada, Bulletin B-190.

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