17 mineral resource and mineral reserve estimates · 17 mineral resource and mineral reserve...

SCOTT WILSON RPA www.scottwilson.com

Strateco Resources Inc. – Matoush Project Preliminary Assessment –December 17, 2008 7-1

17 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES GENERAL STATEMENT

Scott Wilson RPA updated the Mineral Resource estimate for the Matoush Project

using drill hole data available as of July 25, 2008. A set of cross-sections and plan views

were used to construct three-dimensional wireframe models at a cut-off grade of 0.05%

U3O8. High-grade values were cut to 9% U3O8 prior to compositing. Variogram

parameters were interpreted from two-metre composited values. Block U3O8 grades

within the wireframe models were estimated by ordinary kriging. The Mineral Resources

are contained within three zones: AM-15, MT-22 and MT-34. At a cut-off grade of

0.05% U3O8, Indicated Mineral Resources are estimated to total 250,000 tonnes grading

0.68% U3O8 containing 3.73 million pounds U3O8. Inferred Mineral Resources are

estimated to total 1.3 million tonnes grading 0.44% U3O8 containing 13.07 million

pounds U3O8. There are no Mineral Reserves estimated at Matoush.

RESOURCE DATABASE

Scott Wilson RPA received header, survey, assay, lithology and related data from

Strateco in Microsoft Access format. Data were amalgamated and parsed as required,

converted to ASCII, and imported into Gemcom Software International Inc. (Gemcom)

Resource Evaluation Version 6.13 for modelling. The Matoush Gemcom database

includes 277 diamond core holes. Of these, 217 holes representing 87,000 m of drilling

were used in the modelling process (Table 17-1). The other holes are located outside the

area of the deposit. The wireframe models representing the mineralized zones are

intersected by 117 drill holes.


Strateco Resources Inc. – Matoush Project Preliminary Assessment –December 17, 2008 Page 17-2

TABLE 17-1 GEMCOM DATABASE RECORDS PRIMARY DATA Strateco Resources Inc. - Matoush Project

Table Name Number of Records

Hole-ID 217Survey 2,128U3O8 Values (chemical and gamma-logging) 23,506Geology 1,410Alteration General 1,711Hematite/Limonite Alteration 2,972Fuchsite/Chlorite/Muscovite Alteration 1,004Tourmaline Alteration 350Structure 8,610Radiometric 1,769

The resource modelling used U3O8 values from both chemical analysis and gamma

logging (Table 17-2). A total of 188 of the 217 drill holes contain 5,134 chemical assay

U3O8 values. Sample lengths range from 0.1 m to 5.0 m for a total length of 4,488 m.

The majority (~75%) of the chemical resource assay values range in length from 0.5 m to

1.5 m. The additional 24 holes contain 18,372 eU3O8 values calculated from gamma logs.

These range in “sample length” from 0.7 m to 1.4 m for a total of 13,404 m. Many of

gamma-log values are located outside the resource wireframe models.

TABLE 17-2 SOURCE OF U3O8 VALUES Strateco Resources Inc. - Matoush Project

Drill Hole Series Source of U3O8 Values

AM-01 to AM-23 Chemical analysis MT-06-01 to MT-08-35 Chemical analysis MT-08-43 Chemical analysis MT-08-36 Cameco probe MT-08-037 to MT-08-064 Strateco probe

Scott Wilson RPA ran several validation queries and routines in Excel, Access, and

Gemcom to identify data errors. Details of this process are described in Section 14 of this

report. Several problems were identified and corrected. Scott Wilson RPA also verified

a significant number of data records with original logs. No discrepancies were identified.

Drill hole AM-15, the discovery hole drilled by Uranerz, was excluded from the drill hole

database due to poor location data plus the fact it was drilled at a low angle to the



mineralized lens. In Scott Wilson RPA’s opinion, the Gemcom drill hole database is

valid and is suitable to estimate Mineral Resources at Matoush.

CUT-OFF GRADE

Development of the Matoush deposit would require an underground mining method,

given its depth below surface and geometry. Scott Wilson RPA assumes that part of the

deposit would be amenable to longhole methods while other parts would require narrow

vein mining. The following economic assumptions were used to calculate an economic

breakeven grade:

• U3O8 price: US$55/lb • Processing cost: US$20/t • Underground mining cost: US$55/t • General and Administrative cost: US$15/t • Power cost: US$10/t • U3O8 recovery: 98%

Cut-off grade = Operating Cost/(Metal Price x Recovery)

= $100/t/($55/lb x 98%) = 1.86 lb/t = 0.084% U3O8

Scott Wilson RPA used 60% of the fully costed cut-off grade to estimate a marginal

cut-off grade of 0.05% U3O8 for reporting of Mineral Resources. The material with grade

above the marginal cut-off grade, but below the fully costed cut-off grade, represents

mineralization that has a reasonable prospect of economic extraction assuming sunk

development costs and that this marginal material does not displace higher grade

mineralization in the mill feed. The marginal material is included largely for purposes of

maintaining resource continuity, and does not make up a large proportion of the total

Mineral Resource.

Upon development of the mine operating costs, it was found that the operating costs

were substantially higher than the $100/t used and closer to the $300/t level. With a

uranium market price of US$75/lb and using the same marginal cut-off criteria as above

would increase this marginal cut-off to about 0.10% U3O8. While the economics were run



on the 0.05% U3O8 cut-off, the effect of the higher cut-off grade was also evaluated and

is discussed in the economics section of the report.

GEOLOGICAL INTERPRETATION AND 3D SOLIDS

Scott Wilson RPA created 120 north-northwest–south-southeast oriented vertical

sections spaced 10 m apart and 65 level plans extending from the 0 m to the 650 m

elevation. The sections and plans were used to make an on-screen interpretation of

wireframe models. Composite control intervals were identified for each drill hole using a

0.05% U3O8 cut-off and a 1.5 m minimum true thickness.

The 3D wireframe models were built using 3D wobbly polylines on each cross-

section. Several lower grade intersections were included to facilitate continuity. At

model extremities, polylines were extrapolated 5 m to 25 m beyond the last drill hole

section. Polylines were joined together in 3D using tie lines and the continuity was

checked using the longitudinal section and level-plan views. Three zones make up the

Mineral Resources at Matoush: AM-15, MT-22, and MT-34. Each of the zones is made

up of two or more lenses, all of which strike consistently to the north (~008º) and dip

steeply to the east (Table 17-3 and Figure 17-1).

TABLE 17-3 LENS DIMENSIONS AND DETAILS Strateco Resources Inc. - Matoush Project

Zone Lens Strike Dip Max. Strike Length

Max. Down

Dip Avg. True Thickness Volume

No. Holes Intersecting

Solid (deg.) (deg.) (m) (m) (m) (m3 x1,000)

AM-15 MZ 008 -85 140 80 6.0 44.0 17 AM-15 NZ 008 -90 80 80 2.2 8.7 8 AM-15 SZ 006 -83 110 70 3.5 16.0 16 AM-15 UZ 011 -90 45 40 2.3 3.8 3 MT-22 MT22a 007 -90 580 370 3.0 518.0 41 MT-22 MT22b 007 -85 78 65 3.6 13.8 2 MT-34 MT34a 007 -85 320 150 7.8 234.0 25 MT-34 MT34b 009 -80 130 160 2.7 54.2 5 Total 008 -85 1,010 560 4.4 892.0 117



The MT-22a lens is a large panel measuring 580 m by 370 m with excellent

geometric continuity. Good grade continuity is observed in two southward plunging

ribbons. These are separated by low-grade zones with a gradual decrease in grade. To

maintain continuity and preserve the grade distribution, lower grade holes were included

in the preliminary wireframe (i.e., soft boundary). The lower grade subzones were later

managed during the block modelling process (see ‘Classification’ section below) to

ensure mineable continuity of the Mineral Resource blocks. MT-34b also includes

several low-grade subzones that were managed during the block modelling process.

FIGURE 17-1 3D ISOMETRIC VIEW OF MATOUSH WIREFRAME MODELS (LOOKING WEST)

MT-34a

MT-22b

AM-15

MT-22a

MT-22a (low grade)

MT-34b

MT-34b (low grade)



U3O8 STATISTICS

Uranium values within the wireframes were tagged with domain identifiers and

exported for basic statistical analysis. The sample population includes the low-grade

values within lenses MT-22a and MT-34b. Results on a zone and lens basis are

summarized in Tables 17-4 and 17-5.

Descriptive statistics and examination of probability plots indicated positively skewed

data approaching a log-normal distribution. All zones and lenses have high coefficient of

variation (CV) values. This is especially true for MT-22a and MT-34b where CV values

are exasperated by the low-grade intercepts that have been included to maintain

continuity. MT34a and the MZ lenses appear to have the highest average grades.

TABLE 17-4 DESCRIPTIVE STATISTICS OF U3O8 VALUES BY ZONE Strateco Resources Inc. - Matoush Project

All Zones AM-15 MT-22 MT-34 Length U3O8 Length U3O8 Length U3O8 Length U3O8

No. of Cases 1,098 1,098 374 374 274 274 450 450 Minimum 0.10 0.00 0.20 0.00 0.30 0.00 0.10 0.00 Maximum 3.00 18.40 3.00 16.60 3.00 14.51 2.70 18.40 Median 0.69 0.18 0.60 0.18 0.70 0.09 0.69 0.30 Arithmetic Mean 0.70 0.87 0.68 0.81 0.76 0.60 0.68 1.09 Standard Deviation 0.29 1.94 0.32 1.68 0.29 1.58 0.26 2.30 Coefficient of Variation 0.42 2.23 0.47 2.08 0.38 2.62 0.38 2.10



TABLE 17-5 DESCRIPTIVE STATISTICS OF U3O8 VALUES BY LENS Strateco Resources Inc. - Matoush Project

MZ SZ NZ UZ Length U3O8 Length U3O8 Length U3O8 Length U3O8

No. of Cases 217 217 106 106 37 37 14 14 Minimum 0.25 0.000 0.20 0.001 0.300 0.001 0.400 0.003Maximum 2.40 16.600 3.00 7.050 2.200 3.110 1.000 1.050Median 0.60 0.204 0.60 0.186 0.600 0.080 0.750 0.071Arithmetic Mean 0.66 0.911 0.70 0.734 0.659 0.635 0.779 0.143Standard Deviation 0.29 1.968 0.357 1.268 0.372 0.956 0.208 0.271Coefficient of Variation 0.44 2.160 0.511 1.726 0.563 1.506 0.267 1.887

MT22a MT22b MT34a MT34b Length U3O8 Length U3O8 Length U3O8 Length U3O8

No. of Cases 260 260 14 14 421 421 29 29 Minimum 0.30 0.000 0.694 0.093 0.100 0.000 0.694 0.010Maximum 3.00 14.505 1.000 1.310 2.700 18.397 0.695 4.290Median 0.70 0.080 0.900 0.151 0.694 0.320 0.694 0.060Arithmetic Mean 0.76 0.616 0.855 0.357 0.677 1.145 0.694 0.340Standard Deviation 0.30 1.618 0.152 0.410 0.268 2.352 0.000 1.005Coefficient of Variation 0.39 2.625 0.178 1.149 0.396 2.054 0.001 2.959

CUTTING HIGH-GRADE VALUES

Where the assay distribution is positively skewed or approaches log-normal, erratic

high-grade values can have a disproportionate effect on the average grade of a deposit.

One method of treating these outliers in order to reduce their influence on the average

grade is to cut or cap them at a specific grade level. In the absence of production data to

calibrate the cutting level, inspection of the assay distribution can be used to estimate a

“first pass” cutting level.

Scott Wilson RPA’s interpretation of the frequency distributions of the resource

assays suggests cutting high-grade values to 9% U3O8 (Figure 17-2). Inspecting high-

grade values on vertical sections and level plans confirms that uranium values above this

level are spatially erratic and therefore cutting is appropriate. Scott Wilson RPA

recommends that the cutting level be reassessed as more data become available.



Cutting high-grade U3O8 values to 9% affects 15 values which represent 1.4% of the

resource records and reduces the average grade of the resource assays from 0.875% U3O8

to 0.818% U3O8, a 6% decrease (Table 17-6). In the MT-34a lens, where most of the cut

values are located, the standard deviation is reduced from 2.30 to 1.80 and the coefficient

of variation from 2.10 to 1.80.

FIGURE 17-2 HISTOGRAM OF RESOURCE ASSAY VALUES

0 5 10 15 20U3O8 %

0.00

0.02

0.04

0.06

0.08

0.10

0.12

0.14

0.16

0.18

Proportion per B

ar

0

50

100

150

200

Cou

nt

TABLE 17-6 DESCRIPTIVE STATISTICS OF CUT U3O8 ASSAY VALUESStrateco Resources Inc. - Matoush Project

All Lenses U3O8 (%)

AM-15 U3O8 (%)

MT-22 U3O8 (%)

MT-34 U3O8 (%)

Cutting Level 9% 9% 9% 9% Number of Values Cut 15 3 3 9 Percent of Values Cut 1.40% 0.8% 1.1% 2.0% No. of Cases 1,098 374 274 450 Minimum 0.000 0.000 0.000 0.000 Maximum 9.000 9.000 9.000 9.000 Median 0.176 0.178 0.090 0.297 Arithmetic Mean 0.818 0.774 0.578 1.000 Standard Deviation 1.603 1.458 1.401 1.801 Coefficient of Variation 1.959 1.884 2.422 1.801



COMPOSITING

Sample length intervals within the wireframe models range from ten centimetres to

three metres, and average less than one metre (Table 17-4). Assays within the wireframe

models were composited to two-metre lengths starting at the first mineralized wireframe

boundary from the collar and resetting at each new lens wireframe boundary. Several

shorter composites occur at the bottom of the mineralized zone, immediately above where

the drill hole exits the wireframe. Partial composites less than 60 cm long were removed

from the dataset. Non-assayed intervals were treated as zero grade.

Tables 17-7 and 17-8 summarize statistics of the uncut and cut U3O8 composite

values. The decrease in average composite grade, when compared to raw assay grades, is

mainly due to a sample length bias whereby the geologist logging the core selects shorter

sample lengths for the higher grade intercepts, based on scintillometer response. After

compositing, the CV values are mostly below two.

TABLE 17-7 DESCRIPTIVE STATISTICS OF U3O8 COMPOSITE VALUES BY ZONE

Strateco Resources Inc. - Matoush Project

All Zones AM-15 MT-22 MT-34 Uncut Cut Uncut Cut Uncut Cut Uncut Cut

No. of Cases 428 428 141 141 121 121 166 166 Minimum 0.000 0.000 0.000 0.000 0.001 0.001 0.003 0.003 Maximum 11.018 8.065 3.979 3.622 8.431 8.065 11.018 7.813 Median 0.200 0.200 0.198 0.198 0.121 0.121 0.299 0.299 Arithmetic Mean 0.634 0.606 0.541 0.529 0.457 0.447 0.843 0.786 Standard Deviation 1.194 1.044 0.786 0.742 1.131 1.067 1.471 1.212 Coefficient of Variation 1.883 1.724 1.452 1.402 2.474 2.387 1.745 1.542



TABLE 17-8 DESCRIPTIVE STATISTICS OF U3O8 COMPOSITE VALUES BY LENS


MZ SZ NZ UZ Uncut Cut Uncut Cut Uncut Cut Uncut Cut


MT22a MT22b MT34a MT34b Uncut Cut Uncut Cut Uncut Cut Uncut Cut


DENSITY

Scott Wilson RPA received and reviewed rock density and bulk density data collected

by Strateco technicians and Geoanalytical Laboratories of the Saskatchewan Research

Council (SRC), Saskatoon.

In August 2006, Strateco commissioned SRC to make 22 rock density measurements

on crushed drill core using a pycnometer. Results ranged from 2.58 g/cc to 2.82 g/cc and

averaged 2.63 g/cc. These results appear reasonable, however, since rock density does

not include porosity, it cannot be used as a tonnage factor to convert the estimated

volume of the Mineral Resource to tonnes.

Scott Wilson RPA received an Excel spreadsheet of 28 samples labelled “bulk dry

density”. These data appear to be made by SRC, but no official certificates were

produced. Results range from 2.57 g/cc to 3.32 g/cc and average 2.68 g/cc. Removing



the one outlier value of 3.32 labelled “massive” lowers the average value to 2.65 g/cc.

Removing seven samples labelled “semi-massive” lowers the average to 2.61 g/cc.

In June 2008, Strateco technicians began making bulk density measurements using

the water immersion method. As noted in Section 13, core is sealed in plastic wrap to

preserve pore space. Many of the measurements were made on split drill core. The

rough split surface would not permit a tight plastic seal and therefore introduce volume

and ultimately lead to a significant downward bias in the measured bulk density. Scott

Wilson RPA chose to only consider the 60 measurements made on full core. Bulk density

results for those 60 measurements range from 2.27 g/cc to 2.83 g/cc and average 2.52

g/cc. Rock density values for the same 60 samples range from 2.28 g/cc to 2.99 g/cc and

average 2.62 g/cc. The average density of the entire Strateco dataset, including bulk

density measurements made on split core range from 2.03 to 2.84 and average 2.46. Scott

Wilson RPA made several statistical analyses that confirmed that the results from the

split core are biased low.

Considering the methods used and available data verification documents, Scott

Wilson RPA used a global bulk density of 2.52 t/m3 to convert volume to tonnes. Scott

Wilson RPA strongly recommends that the on-going bulk density measurement program

be reviewed and improved. All measurements should be taken on full core. The density

of a “standard”, with similar physical properties to the Matoush mineralization, should be

measured daily. The use of paraffin wax to seal core should also be considered. Scott

Wilson RPA notes that the application of paraffin precludes the option of chemical

analysis.

VARIOGRAPHY AND KRIGING PARAMETERS

Sage 2001 software was used to prepare and interpret variograms from two-metre

composite U3O8 values located within the mineralized wireframes (Figures 25-1 to 25-3

in Appendix 2). The downhole variogram is well developed and indicates a relative

nugget effect of 20%. Long range directional variograms were focused in the plane of

mineralization. In general, the longest ranges were interpreted in the direction of plunge



(southward from -10° to -45°). Equations 1 through 4 list the variogram models

interpreted and applied during the ordinary kriging process.

Equation 1: g i = Co + C1 Sph R1 + C2 Sph R2

Where

g I = Gamma (variogram function) for orientation i (major, semi-major, minor). It is equivalent to the variance of the deposit in that orientation

Co = Nugget effect, calculated from the downhole variogram. Cn = Sill of the nth nested spherical model Rn = Range in metres of the nth nested spherical model

Equation 2: semi-major = 0.2 + 0.25 Sph 25 + 0.35 Sph 50

Equation 3: major = 0.2 + 0.35 Sph 25 + 0.30 Sph 60

Equation 4: minor = 0.2 + 0.8 Sph 5

Scott Wilson RPA used a search ellipsoid measuring 80 m by 40 m by 4 m oriented in

the overall average plane of mineralization. A minimum of one to a maximum of eight

composites were used to estimate the grade of each block. Other search strategy criteria,

unique to each lens, are outlined in Table 17-9.



TABLE 17-9 MATOUSH - SEARCH STRATEGY AND KRIGING PARAMETERS


Main Lens South Lens Upper Lens North Lens Principal Azimuth 188 188 188 188 Principal Plunge -20 -20 -20 -20 Intermediate Azimuth 040 008 008 008 High-grade limit none none none none High-grade range 1 n/a n/a n/a n/a High-grade range 2 n/a n/a n/a n/a High-grade range 3 n/a n/a n/a n/a

MT-22a MT-22b MT-34a MT-34b Principal Azimuth 188 188 188 188 Principal Plunge -30 -20 -45 -20 Intermediate Azimuth 008 008 030 040 High-grade limit 5% none 5% none High-grade range 1 60 n/a 30 n/a High-grade range 2 30 n/a 15 n/a High-grade range 3 6 n/a 6 n/a

BLOCK MODEL SET-UP

The rotated (-8º) Gemcom block model is made up of 65 columns, 115 rows, and 65

levels for a total of 485,875 blocks. The model origin is at UTM coordinates 699,104 mE

5,760,154 mN and 653 m elevation. Each block is 10 m by 10 m by 3 m in size and

contains the following information:

• Estimated cut and uncut U3O8 grades related to mineralized blocks inside the mineralization wireframes (Figures 17-3 to 17-6).

• The percentage volume of each block within the mineralization wireframes.

• A global tonnage factor of 2.52 t/m3.

• Mineral Resource classification identifiers for resource blocks.

• The distance to the closest composite used to interpolate the block grade.

• The average distance to all composites used to interpolate the block grade.

• The number of composites used to estimate the block grade.

Strateco Resources Inc.Matoush Deposit

Longsection Section (Looking West)Block Grades, Lens Outlines, DDH Traces

SCO

TT WILSO

N R

PA

Figur e 17-3

MT-07-128

MT-07-127

MT-07-129MT-07-131

MT-07-125MT-07-15

MT-07-124

MT-07-132

MT-07-123

MT-07-121

MT-08-005

MT-08-009

MT-08-010

MT-08-016

MT-07-130

MT-07-104

MT-07-126

MT-08-018

MT-08-020MT-08-022

MT-07-13

MT-08-024

MT-07-12

MT-08-025

MT-08-027

MT-08-028

MT-08-030

MT-07-36

MT-07-62

MT-08-050

MT-08-051MT-08-037

MT-08-052

MT-08-038

MT-08-039

MT-08-054

MT-08-055

MT-07-22

MT-08-056

MT-07-20

MT-08-057

MT-07-19

MT-08-058

MT-07-17

MT-08-059

MT-08-042

MT-08-060

MT-07-16

MT-08-061

MT-07-09

MT-08-062

MT-08-043MT-08-063

AM20

MT-08-048

MT-08-003

AM23

MT-07-116

MT-07-110

MT-08-015

MT-07-101

MT-08-021

MT-07-30

MT-08-033

MT-08-035

MT-07-60

MT-07-21MT-07-18

MT-07-56

MT-06-17

MT-07-53

MT-07-50

MT-06-20

MT-07-120

MT-06-21

MT-07-114

MT-06-22

MT-08-017

MT-06-23

MT-07-90

MT-06-25

MT-08-034

MT-06-26

MT-06-27

MT-07-22A

MT-07-57

MT-07-108

MT-08-023

MT-07-01

MT-07-25

MT-08-019

MT-08-029

-50 0 50 100

Metres

-100

0 100

200

300

400

500

600

700

800

900

1000

0

100

200

300

400

500

600

700

0.00 0.050.05 0.200.20 0.400.40 0.600.60 0.800.80 1.00

> 1.00

U3O8%

ww

w.scottw

ilson.com

MT-22

a (low

grad

e)

MT-22

a (low

grad

e)

MT-22

a (low

grad

e)

MT-22a

MT-22a

North Lens

MT-34b

MT-34a

Upper LensMain Lens

SouthLens

December 2008

17-14

0

100

200

300

400

500

600

700

Strateco Resources Inc.Matoush ProjectLevel Plan 570

Block Grades, Lens Outlines, DDH Traces

SCOTT WILSON RPA

Figur e 17-4

MT-

07-8

7

MT-

07-8

8

MT-07-46

MT-

06-1

7

MT-

08-0

08

MT-

07-5

1

MT-07-44

MT-07-42

MT-

07-9

0

MT-

07-0

3

MT-

07-3

7

MT-

06-0

6

MT-

06-1

6

MT-06-05

MT-

06-1

5

MT-

07-0

2

MT-

07-3

5

MT-06-04

MT-06-14

MT-

06-3

8

MT-

07-3

2

MT-

07-2

9

MT-

06-3

7

MT-07-27

MT-

08-0

14

MT-

08-0

32

MT-06-12

MT-

08-0

31

MT-

07-1

1

MT-

07-3

9

MT-06-10

MT-

06-1

8

MT-06-09

MT-

06-0

8

MT-06-03

MT-

08-0

11

MT-

06-3

6

MT-

07-8

3

MT-

06-3

5

MT-

06-3

4

MT-07-05

MT-

06-3

3

MT-

07-1

04

MT-

06-3

2

MT-

06-3

1

MT-

07-3

4

MT-

06-0

2

MT-

07-3

1

MT-

06-0

1

MT-

07-2

6M

T-06

-30

AM22

MT-

08-0

07

MT-

07-0

8

MT-

06-2

9

MT-08-037

MT-

06-2

8

MT-

07-0

6

MT-07-56

MT-06-19

MT-

07-3

3

MT-07-54

MT-06-13

MT-

07-1

0

AM20MT-

07-0

7

MT-07-50

MT-08-016

MT-

06-1

1

MT-

06-0

7

MT-

07-0

4

MT-

07-2

8

MT-

07-4

9

MT-07-48

MT-

08-0

04

-25 0 25 50

Metres

6992

00

6993

00

6994

00

5760500

5760600

5760700

www.scottwilson.com

0.00 0.050.05 0.200.20 0.400.40 0.600.60 0.800.80 1.00

> 1.00

U3O8%

Sout

h Le

nsM

T-34

a Le

ns

Mai

n Le

nsN

orth

Len

sLens

December 2008

17-15

5760500

5760600

5760700

Strateco Resources Inc.Matoush Project

Vertical Section 1290NBlock Grades, Lens Outlines, DDH Traces

SCOTT WILSON RPA

Figur e 17-5

MT-07-49

MT-07-110

MT-06-18

MT-07-108

MT-06-29

-200

-100

0 100

200

300

100 Elev

200 Elev

300 Elev

400 Elev

500 Elev

600 Elev

www.scottwilson.com

0.00 0.050.05 0.200.20 0.400.40 0.600.60 0.800.80 1.00

> 1.00

U3O8%

MT-

22a

Lens

Nor

th L

ens

December 2008

17-16

200 Elev

300 Elev

400 Elev

500 Elev

600 Elev

Strateco Resources Inc.Matoush Project

Vertical Section 1120NBlock Grades, Lens Outlines, DDH Traces

SCOTT WILSON RPA

Figur e 17-6

MT-06-36

MT-07-07

40

60

80

100

120

500 Elev

520 Elev

540 Elev

560 Elev

580 Elev

600 Elev

620 Elev

www.scottwilson.com

0.00 0.050.05 0.200.20 0.400.40 0.600.60 0.800.80 1.00

> 1.00

U3O8%

Sou

th L

ens

MT-

34a

Lens

December 2008

17-17

520 Elev

540 Elev

560 Elev

580 Elev

600 Elev

620 Elev

500 Elev



CLASSIFICATION OF MINERAL RESOURCES

Definitions for resource categories used in this report are consistent with those set out

in CIM Definition Standards for Mineral Resources and Mineral Reserves (December

2005) and adopted by NI 43-101. In the CIM classification, a Mineral Resource is

defined as “a concentration or occurrence of natural, solid, inorganic or fossilized organic

material in or on the Earth’s crust in such form and quantity and of such grade or quality

that it has reasonable prospects for economic extraction”. Mineral Resources are

classified into Measured, Indicated and Inferred categories. A Mineral Reserve is defined

as the “economically mineable part of a Measured or Indicated Mineral Resource

demonstrated by at least a Preliminary Feasibility Study”. Mineral Reserves are

classified into Proven and Probable categories.

Scott Wilson RPA classified the Mineral Resources based on drill hole spacing, lens

thickness, continuity, and variogram ranges. Most of the AM-15 Zone was classified as

Indicated and the MT-22 and MT-34 zones as Inferred. There are no Mineral Reserves

reported at Matoush.

The drill hole spacing in the AM-15 Zone is generally less than 20 m. The Main and

South lenses were classified as Indicated. Most of the North Lens was also classified as

Indicated. The Upper Lens was classified as Inferred. Although there are some areas of

closely spaced drilling in the Main and South lenses, no blocks have been classified as

Measured because grade continuity has not been established to the confidence level

required for the Measured category, in Scott Wilson RPA’s opinion.

As discussed above, the MT-22a lens includes several low-grade subzones containing

weakly mineralized intersections that do not meet the cut-off grade. These subzones are

therefore not part of the Mineral Resource and were “carved out” using additional

wireframe models (Figure 17-7). Care was taken to maintain mineable continuity. The

average drill hole spacing in the MT-22a lens is 60 m to 70 m. Both MT-22a and MT-

22b were classified as Inferred. Several areas of MT-22a could be upgraded to Indicated

with a few additional holes.



The upper part of the MT-34a lens, which is considered the fault-offset extension of

the Main Lens, has an average drill hole spacing of less than 20 m and was classified as

Indicated. The MT-34b lens includes low-grade subzones that contain weakly

mineralized intersections but do not meet the cut-off grade. MT-34b was treated in a

similar manner as MT-22a and has been classified as Inferred.

FIGURE 17-7 LONG SECTION OF MT-22A SHOWING CLASSIFICATION

MINERAL RESOURCE REPORTING

Tables 17-10 and 17-11 list the Mineral Resources by lens, zone and cut-off grade.

Observations include:

• The average grade of the block model is consistent with the average grade of the composite values.

• Lenses MZ and MT-34a make up the bulk of the Indicated Mineral Resources. • Since the last estimate, there has been a substantial increase in Inferred

Mineral Resources and only a small increase in Indicated Mineral Resources. • Additional drilling will likely upgrade parts of MT-22a to the Indicated

category.

Low grade classified as waste.

Inferred

Inferred



• A large portion of the resources are greater than 0.25% U3O8. • A large portion of MT-34a classified as Indicated has an average grade greater

than 1% U3O8.

TABLE 17-10 MINERAL RESOURCE REPORT BY LENS, JULY 25, 2008


Tonnage Grade Pounds U3O8 (x 1,000) (%U3O8) (x1,000)

Indicated MZ 111 0.54 1,320 NZ 15 0.58 190 SZ 36 0.41 330 MT34A 88 0.97 1,890

Total Indicated 250 0.68 3,730 Inferred

NZ 7 0.18 30 UZ 9 0.11 20 MT22A 766 0.38 6,390 MT22B 35 0.38 290 MT34A 480 0.55 5,870 MT34B 47 0.46 480

Total Inferred 1,344 0.44 13,070

Notes: 1. CIM definitions were followed for Mineral Resources. 2. The cut-off grade of 0.05% U3O8 was estimated using a U3O8 price of US$55/lb

and assumed operating costs. 3. Wireframes at 0.05% U3O8 and a minimum true thickness of 1.5 metres were

used to constrain the grade interpolation. 4. High U3O8 grades were cut to 9% prior to compositing to two-metre lengths. 5. Several blocks less than 0.05% U3O8 were included for continuity or to expand

the lenses to the minimum thickness. 6. Totals may not sum correctly due to rounding



TABLE 17-11 MINERAL RESOURCE REPORT BY CUT-OFF GRADE, JULY 25, 2008


Cut-off Grade Tonnage Cut U3O8 Cut U3O8 (%) (x 1,000) (%) (lbs x 1,000)

Indicated Mineral Resources AM-15 1.25 6.4 1.434 202 1.00 20.2 1.202 535 0.75 40.8 1.033 930 0.50 68.1 0.864 1,297 0.25 113.3 0.664 1,659 0.05 161.7 0.515 1,838 MT-34 1.25 24.2 1.674 892 1.00 37.8 1.470 1,225 0.75 56.9 1.273 1,596 0.50 68.7 1.161 1,760 0.25 80.8 1.045 1,862 0.05 88.2 0.972 1,891 Total 1.25 30.6 1.624 1,095 Indicated 1.00 58.0 1.377 1,760 0.75 97.7 1.173 2,526 0.50 136.9 1.013 3,057 0.25 194.0 0.823 3,521 0.05 250.0 0.677 3,729 Inferred Mineral Resources AM-15 1.25 0.0 0.000 0 1.00 0.0 0.000 0 0.75 0.0 0.000 0 0.50 0.0 0.000 0 0.25 1.1 0.329 8 0.05 16.2 0.138 49 MT-22 1.25 33.3 2.099 1,542 1.00 54.1 1.709 2,039 0.75 78.2 1.445 2,490 0.50 132.0 1.102 3,207 0.25 436.8 0.578 5,569 0.05 800.5 0.379 6,680 MT-34 1.25 55.4 1.782 2,175 1.00 69.0 1.656 2,517 0.75 112.9 1.349 3,356 0.50 228.6 0.972 4,897 0.25 340.4 0.776 5,820 0.05 527.2 0.546 6,345 Total 1.25 88.7 1.901 3,716 Inferred 1.00 123.1 1.679 4,556 0.75 191.1 1.388 5,846 0.50 360.6 1.019 8,104 0.25 778.4 0.664 11,398 0.05 1,343.9 0.441 13,075

See Notes for Table 17-10 on page 17-20.



MINERAL RESOURCE VALIDATION

Scott Wilson RPA validated the block model by visual inspection, volumetric

comparison, swath plots, and a comparison with results using inverse distance squared

(ID2) grade interpolation. Visual comparison on vertical sections and plan views, and a

series of swath plots found good overall correlation between block grades and composite

grades.

The estimated total volume of the wireframe models is 649,200 m3, while the volume

of the block model at a zero grade cut-off is 649,600 m3. The small difference in volume

is due to Gemcom’s needling method that estimates the percentage of mineralization

within each block. Results are listed by lens in Table 17-12.

TABLE 17-12 VOLUME COMPARISON Strateco Resources Inc. - Matoush Project

Lens ID

Vol. Wireframe

(m3 x 1,000)

Vol. Block Model

(m3 x 1,000) Difference

(m3 x 1,000) Difference

(%) MT22a 311.2 312.3 1.1 0.3 MT22b 13.8 14.2 0.4 2.7 MT34a 234.1 233.7 -0.4 -0.2 MT34b 19.0 18.8 -0.2 -1.1 MZ 44.0 43.9 -0.1 -0.2 NZ 8.7 8.7 -0.1 -0.8 SZ 14.5 14.4 0.0 -0.3 UZ 3.8 3.6 -0.2 -6.6 Total 649.2 649.6 0.5 0.1

Scott Wilson RPA estimated the average grade of the Indicated Mineral Resources

using ID2 to be 0.70% U3O8. This compares to the average grade of 0.68% U3O8 as

estimated by ordinary kriging. Similarly for Inferred Mineral Resources, the estimated

average grade using ID2 is 0.45% U3O8 compared to kriging, which gave 0.44% U3O8.

The differences are within acceptable limits and may be due to the declustering effect of

the kriging method. Details at various cut-off values are given in Table 17-13.



TABLE 17-13 GRADE ESTIMATION COMPARISON Strateco Resources Inc. - Matoush Project

Zone Cut-off Tonnage Cut Kriged Cut ID2 Uncut Kriged

(x1,000) (%U3O8) (%U3O8) (%U3O8) Indicated AM-15 1.25 6.4 1.43 1.58 1.56 1.00 20.2 1.20 1.21 1.25 0.75 40.8 1.03 1.06 1.07 0.50 68.1 0.86 0.89 0.89 0.25 113.3 0.66 0.69 0.68 0.05 161.7 0.52 0.54 0.53 MT-34 1.25 24.2 1.67 1.67 1.70 1.00 37.8 1.47 1.48 1.49 0.75 56.9 1.27 1.28 1.29 0.50 68.7 1.16 1.17 1.17 0.25 80.8 1.05 1.05 1.05 0.05 88.2 0.97 0.98 0.98 ALL Indicated 1.25 30.6 1.62 1.65 1.67 1.00 58.0 1.38 1.39 1.41 0.75 97.7 1.17 1.19 1.19 0.50 136.9 1.01 1.03 1.03 0.25 194.0 0.82 0.84 0.84 0.05 250.0 0.68 0.70 0.69 Inferred AM-15 1.25 0.0 0.00 0.00 0.00 1.00 0.0 0.00 0.00 0.00 0.75 0.0 0.00 0.00 0.00 0.50 0.0 0.00 0.00 0.00 0.25 1.1 0.33 0.24 0.33 0.05 16.2 0.14 0.12 0.14 MT-22 1.25 33.3 2.10 2.20 2.20 1.00 54.1 1.71 1.82 1.78 0.75 78.2 1.44 1.56 1.49 0.50 132.0 1.10 1.19 1.13 0.25 436.8 0.58 0.62 0.59 0.05 800.5 0.38 0.40 0.38 MT-34 1.25 55.4 1.78 1.89 2.13 1.00 69.0 1.66 1.73 1.96 0.75 112.9 1.35 1.37 1.57 0.50 228.6 0.97 0.97 1.11 0.25 340.4 0.78 0.77 0.87 0.05 527.2 0.55 0.54 0.61 ALL Inferred 1.25 88.7 1.90 2.00 2.16 1.00 123.1 1.68 1.77 1.88 0.75 191.1 1.39 1.45 1.54 0.50 360.6 1.02 1.05 1.12 0.25 778.4 0.66 0.68 0.71 0.05 1,343.9 0.44 0.45 0.47



18 OTHER RELEVANT DATA AND INFORMATION INTRODUCTION

The Preliminary Assessment for the Project is based on the following assumptions:

• Mineral Resources as described previously, modified by the addition of dilution and extraction factors.

• Starting production rate of 500 tpd, ramping up to a maximum production rate

of 750 tpd (262,500 tpa) in Year 3. • Underground mining by blasthole stoping. • Backfill provided by cemented rock fill and waste rock fill from development. • On-site milling and processing, including crushing and grinding, leaching

using sulphuric acid, followed by Resin-in-Pulp, Resin Elution, Impurity Precipitation and calcining circuits to extract, purify, precipitate and thicken the final uranium product.

• Site layout as illustrated in Figure 18-1.

Scott Wilson RPA evaluated the portions of the study related to mining and site

operations. The processing portion of the study, including cost estimates, was assessed

by Melis within the battery limits of the process plant. The environmental aspects of the

Project, including permitting, licensing, monitoring, and closure, were evaluated by

Golder.

775.0

780.0

775.0

777.5

777.5

777.5

777.5

777.5

777.5

780.0

780.0

78

0.0

780.0

780.0

772.5

772.5

777.5

777.5

775.0

775.0

777.5

777.5

777.5

777.5

777.5

780.0

780.0

780.0

78

0.0

780.0

780.0

780.07

80.0

780.0

780.0

780.0

78

0.0

780.0

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

785.0

785.0

785.0

785.0

785.0

785.0

785.0

785.0

785.0

785.0

785.0

780.0

782.5

785.0

772.5

77

2.5

772.5

772.5

772.5

777.5

777.5

777.5

77

7.5

777.5

777.5

777.5

777.5

777.5

777.5

77

5.0

77

5.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

77

5.0

780.0

780.0

780.0

782.5

77

7.5

777.5

777.5

780.0

782.5

777.5

777.5

777.5

777.5

777.5

777.5

780.0

780.0

78

0.0

780.0

780.0

772.5

772.5

777.5

777.5

775.0

775.0

777.5

777.5

780.0

777.5

777.5

777.5

780.0

780.0

780.0

78

0.0

780.0

780.0

780.07

80.0

780.0

780.0

780.0

78

0.0

780.0

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

782.5

785.0

785.0

785.0

785.0

785.0

785.0

785.0

785.0

785.0

785.0

785.0

780.0

782.5

785.0

772.5

77

2.5

772.5

772.5

772.5

777.5

777.5

777.5

77

7.5

777.5

777.5

777.5

777.5

777.5

777.5

77

5.0

77

5.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

775.0

77

5.0

780.0

780.0

780.0

782.5

77

7.5

777.5

777.5

780.0

782.5

777.5777.5

CONTROL STATIONS

FINAL EFFLUENT PIPE

(APPROXIMATIVE LOCATION)

EXISTING ACCESS ROAD

BE KEPT INTACT

WETLAND TO

ANTICIPATE A

DIVERSION DITCH

PROJECTED ROAD

WETLAND TO

BE KEPT INTACT

ORE STORAGE AREA

WASTE STORAGE AREA

ACCESS ROAD AND

DRAINAGE DITCH

20

GARAGE

EXISTING CAMP

PROJECTED WIDENING

OF CAMP

COMPRESSOR

GENERATOR

FUEL DEPOT

44

35

FEEDING PIPE

(RAW WATER)

PROCESS PLANT

OFFICEMINE

RESCUE

PARKING AREA

OFFICE

WAREHOUSE

LEACHING BED

TREATMENT

POND

MWTP

DRY

10

5 TYP.

PROPANE

20VENTILATOR

PUMPING

STATION

PORTAL

COORDINATES

OF THE PORTAL

X=700 081

Y=5 760 792

WETLAND TO

BE KEPT INTACT

EXISTING ACCESS

ROAD

CONTRACTOR

AREA

5,6

70,2

50

mN

5,6

70,2

50

mN

5,6

70,5

00

mN

5,6

70,7

50

mN

5,6

71,0

00

mN

5,6

70,5

00

mN

5,6

70,7

50

mN

5,6

71,0

00

mN

700,000 mE 700,250 mE699,750 mE

Legend:

Existing Woodland Limit

Projected Woodland Limit

Projected Road and Wideningof the Existing Camp

Existing Access Road

Projected Infrastructure

0 25

Metres

50 75 100

N

December 2008 Source: Strateco Res., Inc., 2008.

Site Plan

Matoush Project

Strateco Resources Inc.

Québec, Canada

Figure 18-1

18-2




MINING OPERATIONS

Inferred and Indicated Resources used in the Preliminary Assessment are shown in

Table 18-1.

TABLE 18-1 MINERAL RESOURCES FOR PRELIMINARY ASSESSMENT

Strateco Resource Inc. - Matoush Project

Classification/Zone Tonnes (x1,000) Grade (%U3O8)

Indicated AM-15 162 0.51% MT-34 88 0.97% Total 250 0.68%

Inferred AM-15 16 0.14% MT-22 801 0.38% MT-34 527 0.54% Total 1,344 0.44%

Notes: 1. CIM definitions were followed for Mineral Resources. 2. The cut-off grade of 0.05% U3O8 was estimated using a U3O8 price of US$55/lb

and assumed operating costs. 3. Wireframes at 0.05% U3O8 and a minimum true thickness of 1.5 metres were

used to constrain the grade interpolation. 4. High U3O8 grades were cut to 9% prior to compositing to two-metre lengths. 5. Several blocks less than 0.05% U3O8 were included for continuity or to expand

the lenses to the minimum thickness. 6. Totals may not sum correctly due to rounding

MINE ACCESS Underground access to the Matoush Project will be via a five metre by five metre

ramp decline driven at a -15% gradient. The initial exploration ramp will be driven to a

vertical depth of 300 m, at which point a 1,000 m diamond drill drift will be established.

This will provide an effective drill platform for ongoing diamond drilling, particularly for

the deeper portions of the potential zones, as shown in Figure 18-2. To ensure adequate

ventilation, this phase will also include the development of a ventilation access drift on

the 180 m level and an exhaust raise to surface. The program will include a total of 2,400

m of main ramp access, approximately 1,700 m of access drifts, and about 320 m of



ventilation raise. The initial costs of the exploration ramp program, estimated to be

approximately $50 million, are not included in the current Preliminary Assessment.

The development required to progress to the initial production phase will continue

from these excavations, as shown in Figure 18-3. The exploration ramp will provide

access to the AM-15 Zone between the 180 m and 300 m levels. To access the MT-34

Zone, which is an extension of the AM-15 Zone to depth, the exploration ramp will be

extended to the 480 m level. Access to the MT-22 Zone, located north of the AM-15 and

MT-34 zones, will be achieved by a separate ramp driven off the 300 m level, providing

access between the 300 m and 615 m levels.

The potential for shaft access was explored, particularly for the MT-22 Zone as it

reaches depths greater than 600 m. However, it was determined that the benefits of this

access method would not offset the increased capital costs and necessary lead time. As a

result, the present study maintains ramp access for the Project. Optimization of the

access scenarios will be addressed at the pre-feasibility study level.

December 2008

Matoush Project

Exploration Ramp


Québec, Canada

Figure 18-2

SC

OT

TW

ILS

ON

RP

A

18

-5

ww

w.sco

ttwilso

n.co

m

December 2008

Matoush Project

Mine Development Layout


Québec, Canada

Figure 18-3

SC

OT

TW

ILS

ON

RP

A

18

-6

ww

w.sco

ttwilso

n.co

m



MINING METHOD The mining method selected for the Matoush property will be blasthole stoping. Due

to the variance in widths of the various zones, longhole stoping with both small and large

diameter holes is proposed. The mining methods are shown on the longitudinal and plan

sections in Figures 18-4 and 18-5.

LARGE DIAMETER LONGHOLE STOPING

Figure 18-4 shows the stoping arrangement that utilizes large diameter holes. This

method is applicable where the zone widths are approximately five metres or greater,

such as in the AM-15 and MT-34A zones. For this method, level spacing is assumed to

be 25 m (floor to back) and stope lengths are assumed to be 20 m.

Large diameter longhole stoping will incorporate a primary-secondary stoping

sequence. Once the primary stope has been mined, it will be backfilled with cemented

rock fill (CRF) prior to mining the secondary stopes. With the limited height of the stope

blocks in the AM-15 Zone, mining can commence in the bottom portion (205 m level to

230 m level), with subsequent mining in the upper portion (180 m to 205 m) once the

lower stopes have been backfilled. To limit the use of CRF, secondary stopes can also be

backfilled with waste rock from ongoing development.

SMALL DIAMETER LONGHOLE STOPING

Figure 18-5 shows the small diameter longhole stoping method. This method is

applicable to the zones with narrow widths, such as those found in the MT-22 and

MT34b zones. The widths for these zones average approximately 2.5 m and 3.0 m,

respectively. To facilitate grade control and precision drilling, these zones will need to

be developed with a maximum level interval of 15 m. In general, stope widths will be

approximately 100 m, and mining will proceed upwards from every second level.

Temporary pillars will be left to provide ground stability and avoid dilution from wall

rock by limiting the hydraulic radius of the exposed hanging wall. These pillars will also

be required to permit backfilling in the mining sequence. Once a 15 m level has been

mined, it is backfilled while the adjacent stope block is mined.



Mining on lower levels can be achieved with benching from the access drift, followed

by the development of an ore drift under the consolidated backfill. Alternately, an ore

drift could be developed under a sill pillar, with sill recovery in a subsequent pass. For

each set of 15 m levels that are mined concurrently, the bottom level will be backfilled

using CRF, while the top level will be backfilled using CRF or waste rock fill as

consolidation is not necessary. Depending on ground condition, the proposed stoping

method permits the reduction of stope lengths by leaving pillars (if required) and drilling

new slot raises to commence the stope advance.

The general stoping sequence will begin with central access to the zone, followed by

development to the extents of the zone. Mining will then proceed in a retreat fashion

toward this initial access. A footwall drift will be driven to allow access for mucking and

backfilling. In the MT-22 Zone, it will be necessary to develop three 15 m levels per year

to sustain the required production.

Mucking of the stopes will be carried out using either 5m³ or 3.2m³ scooptrams

equipped with remote controls. Ore will be loaded into 33 tonne capacity low profile

haulage trucks, which will then truck the ore to surface. Truck loading stations will be

established on each level for ease of loading. Tramming distances for the scooptrams

will also be kept to a minimum.

December 2008

Matoush Project

Large DiameterLonghole Stopes


Québec, Canada

Figure 18-4

18-9


December 2008

Matoush Project

Small DiameterLonghole Stopes


Québec, Canada

Figure 18-5

18-10




HIGH-GRADE ZONES While the majority of the deposit hosts grades typical for safe radon levels, there may

be areas where the grade is above acceptable levels to be mined by the proposed longhole

techniques. In such cases, a “Non Man Entry” mining method, such as the use of a raise

bore machine or similar technique, will need to be utilized. Although this was not

evaluated in the current study, such areas will be identified in later stages of the project

when the resource model has been refined, and an appropriate mining method designed to

ensure safe mining procedures will be incorporated.

DILUTION AND EXTRACTION

Stope dilution will be incurred from irregularities in the orebody, overbreak from

blasting, and deterioration of stope walls and back during mucking. For the current

evaluation, dilution has been estimated to be 15% for all stoping areas, which is typical

for the proposed mining method and design. This would represent 0.5 m to 1.0 m and 0.3

m to 0.5 m of dilution in the large and small diameter longhole stopes, respectively.

An extraction of 90% was considered appropriate for all stopes.

PRODUCTION RATE

The mine production rate over the Life of Mine (LOM) is based on the annual

production target of approximately 2.0 million pounds of U3O8. Considering the

estimated average grade of the deposit, the required production rate is approximately 600

tpd. The production rate using Taylor’s rule of thumb amounts to approximately 610 tpd,

based on a seven-day per week schedule, 350 days per year and the estimated resource of

1.6 million tonnes.

Scott Wilson RPA feels that based on the required amount of development to prepare

stopes each year, this production rate is achievable and can be sustained.

MINE VENTILATION Adequate ventilation will be critical to ensure effective and safe production. Initially,

the ramp will act as the fresh air source and the exhaust route will be provided with a

ventilation raise that will be driven from the 180 m level to surface. Once the ramp has



reached the 300 m level, an exhaust raise will be driven from that level to the 180 m level

to tie into the system.

An intake ventilation raise will be driven to surface prior to pre-production

development to prepare the mine for production. Fresh air will then downcast from this

raise to provide the required volume or air underground. Each of the levels will be tied

into the fresh air ventilation raise to provide rapid egress from any level. Once the

underground workers have reached the raise location, they will be in a fresh air “base”

and will have a secure access route to surface. Production levels will continue to tie into

this fresh air raise as development progresses.

Exhaust ventilation raises will be developed between levels as required. The AM-15

and MT-34 zones will require one main exhaust raise, while the MT-22 Zone will require

a system of raises at each end of the zone, which is over 500 m in strike length. As

mining progresses downwards, the exhaust raises will be arranged to tie in with the

exhaust system on the levels above. To ensure proper exhausting procedures, sufficient

auxiliary ventilation fans and ducting will also be required to properly ventilate active

development headings. The required safety monitoring system will also be installed to

provide adequate warning in case of high radon gas levels.

MINERAL PROCESSING

The mineral processing is covered in Section 16 of the report.

GEOMECHANICS AND GROUND SUPPORT

The Matoush uranium-bearing zones are hosted in sedimentary rock units. There are

two main units, Active Channel Facies (ACF) and Channel Bar Facies (CBF). The ACF

consists of massive to slightly cross-bedded, gritty, coarse-grained sandstone to

conglomerate. The conglomerates are dominantly matrix supported and the matrix is

composed of clay-silt and medium to coarse sands. The rocks are generally well

cemented with silica. The CBF consists of medium- to coarse-grained, well cross-



bedded, sorted, subarkosic sandstone. The quartz and feldspar grains are subangular to

subrounded. The sandstone is well cemented with silica. Figure 7-2 in Section 7 of this

report shows the cycles of ACF and BCF deposition with the various thicknesses.

Examination of the drill core indicates very continuous core and that core recovery has

been excellent. Presently, little is known about the detailed structural geology of the

Matoush property other than that the stratigraphy is flat-lying and cut by a number of

north and northeast trending faults. The current defined resources at Matoush are all

associated with a north trending fault structure occupied by a mafic dyke. The fault is

located on the east side or footwall of the mineralized zones. The fault appears to be 0.2

m to 0.5 m in thickness and will require attention when passing through it to access the

mineralized zones.

Data collected such as RQD values and discontinuity description including

orientation, spacing, persistence, roughness, weathering, width, infill and water and any

other factors affecting rockmass strength will be used to determine rockmass

characterization in order to develop recommendations for the excavation and support

design.

Samples collected from drill holes will permit laboratory testing including

Unconfined Compression tests, Brazilian Indirect Tensile Strength tests and Point Load

Index tests, which will be very useful in evaluating the rockmass strength and support

requirements.

With the proposed mining method, backfilling of the open stopes will be completed,

thereby reducing the standup time and any negative effects of leaving open stopes for

long periods of time. Backfilling will provide a means of support as well as a means of

disposing of waste rock produced from development.

For the present study, standard support methods are assumed for the ramp

development and in the mineralized zones. These will consist of 2.1 m rebar bolts on 1.2

m centres with #6 WWM (welded wire mesh) screen in the main ramp back, and 1.5 m

split set bolts in the walls. Support in the mineralized zones will include standard 16 mm



diameter rockbolts with #6 WWM screen and 1.5 m split sets in the walls. The need for

cable bolting of the hanging wall is not anticipated. To control and keep the radon gas

emanations to the required levels, it will be necessary to shotcrete the walls and the back

in addition to cementing the floor.

INFRASTRUCTURE

RAMP The main underground mode of access will be via a five metre by five metre ramp

decline driven at a gradient of -15%. This size of opening provides sufficient space for

the required equipment, and ensures that the production targets can be sustained.

Adequate ancillary openings for refuge stations, electrical substations, remuck bays,

safety bays, storage bays, and primary and secondary sumps will be driven off of the

main ramp. Ventilation raises will also connect with the ramp to provide sufficient

ventilation volume to meet regulatory requirements. The mineralized zones reach a depth

of over 600 m vertically. As a result, proper maintenance will be essential to maintain

equipment condition and meet production requirements. Haulage distances from the

deepest portions of the mine will be over four kilometres.

MATERIAL HANDLING Handling of mineralized and waste material will be by 33-tonne low profile trucks

from the underground to the surface stockpile or storage areas. The potential to

incorporate larger trucks will be evaluated at later stages in the Project. On surface, front

end loaders will be utilized to transfer mineralized material into the processing facility.

The tailings produced will be pumped to the tailings facility on surface. The possibility

of using the tailings as paste backfill at this time has not been considered in order to keep

the water usage to a minimum and thereby reduce potential radon emissions.

BACKFILL Backfill will be in the form of CRF or waste fill, sourced from underground waste

development. For the present study, it has been assumed that a paste backfill or CRF



plant will be present on surface. Capital and operating cost allowances have been made

accordingly.

VENTILATION During production, ventilation will be provided by a downcast raise from surface,

combined with several exhaust raises required to ensure a single-pass system of

ventilation air. The required volume of air will be in the range of 500,000 cfm to supply

all areas of the mine, which are spread over a kilometre laterally from north to south. To

maintain reasonable air speed at 2,500 fpm in the raise, the raise dimension would be in

the order of 4.8 m in diameter. The main downcast fresh air raise will be located

centrally to service the underground development headings and mine workings with

exhaust raises located to the south for the MT-34 Zone and to the north for the MT-22

Zone. Main fans complete with a 30 MBTU mine air heating system will provide the

required air volume to properly ventilate the underground. In addition, suitably sized

exhaust fans will be located on surface at the exhaust raises. A manway will be installed

in the fresh air downcast raise to permit egress from any area in the mine. Monitors with

adequate warning systems will be installed to record ventilation volumes and gas levels.

A schematic of the ventilation air flow is shown in Figure 18-6.

Exhaust Raise

Exhaust Raise

Fresh Air Route

Intake Raise

Exhaust Route

December 2008

Matoush Project

Ventilation Schematic


Québec, Canada

Figure 18-6

SC

OT

TW

ILS

ON

RP

A

18

-16

ww

w.sco

ttwilso

n.co

m



DEWATERING Main sumps will be established at the required levels in the mine and all water

pumped to surface will be directed to the Water Treatment Plant (WTP) for treatment

prior to being discharged. Some of this water may be used for the process plant operation

as well.

MAINTENANCE The maintenance of underground and surface mobile equipment will be carried out in

the surface garage. This facility will be approximately 40 ft. by 60 ft., and will host three

bays with an overhead crane. An underground garage will also be developed, and will

but used for minor maintenance repairs only.

A wash bay will be located near surface in the underground ramp. This will provide

for a washing facility for equipment prior to being sent to the surface garage for

maintenance. Water will again be collected and directed to the WTP for treatment.

POWER Both hydroelectric power and diesel-generated power were considered for the Project.

The estimated cost by Hydro Québec for a 161 kV power line extension from the Troilus

mine was estimated at $130 million. Other options for lines from the Eastmain or

Nikamo would cost approximately $180 million. These costs are for dedicated lines and

some reduction in cost would be possible by sharing with other users. Given the capital

requirements for a power line and the uncertainty of this arrangement, the present

evaluation was completed assuming the use of diesel power generated on site.

The total connected power will be 6 MW. The underground distribution would be via

a 4,160 V cable to the underground substation, stepped down by a transformer to 600 V.

The majority of underground power consumption will be from the electric jumbo drills,

main pump stations, and auxiliary ventilation fans.



COMPRESSED AIR AND SURFACE WATER Two main compressors, each providing 1,500 cfm, will be located on surface. Air

will be fed to two air receivers prior to supplying the underground. Due to the depth of

the underground zones, the need for booster compressors on the line may be required to

compensate for the length of the air piping. The main consumers of compressed air

underground will be the longhole drills, airleg drills and auxiliary equipment.

Compressed air is initially delivered via the air line in the ramp, and a permanent line will

be installed in the main ventilation raise. This will reduce the line length and the

resistance, thereby improving the system pressure drop.

Similar to the underground water, surface water drainage from the stockpile areas will

be collected and sent to the WTP for treatment. It will then be recycled or sent to final

effluent.

COMMUNICATIONS Communications underground will be provided via a leaky feeder communication

system. This will also provide monitoring functions for electric ground monitoring

systems, such as ground movement monitors (GMM). Surface communications will be

provided via satellite.

ROADS Road maintenance will be carried out by the surface crew using a grader, dozer and

truck. A portable crusher will be contracted to prepare sufficient crushed material from

the mine waste rock. This will serve as roadbed material and for surface pad preparation.

EQUIPMENT REQUIREMENTS

The production mining equipment required for the Project consists of the following:

• Four MT436B-33T low profile haul trucks • Two 35T Backfill low profile haul trucks c.w. pushplate • Two M2D (or equiv.) drill jumbo- 2 Boom • Two Axera (or equiv.) drill jumbo-1 Boom • One LH drill Elec.-Hydraulic (64-102 mm dia. holes)



• Two LH drills- Pneumatic (50-62.5 mm dia. holes) • One Grader unit for ramp • Two, Rock-breakers-hydraulic • Three ST1030- 5m³ scooptrams • Two ST710- 3.2m³ scooptrams • Two ML40 Scissor lifts • One RBM (or equiv.) roof bolter • One M40 Crane truck-with Hiab • Four Kubota M59 Service tractors • Two Anfo Loaders • One Shotcrete unit • Two F.E. Loader 5 yd.³. (mine and mill) • One F.E. Loader 2yd.³ • One Backhoe 2 yd.³ • Boom Truck (mill)

Additional equipment will include minor items such as ½ ton pick–up trucks. It

should be noted that a portion of the listed equipment fleet is assumed to be purchased

during the advanced exploration phase when the main ramp is developed to the 300 m

level. Scott Wilson RPA is of the opinion that the equipment list will be sufficient to

sustain the planned production rate of over the LOM.

The surface mobile equipment fleet will consist of front end loaders, a 10 wheel (11

m³) haul truck for miscellaneous haulage purposes (with snow plough quick coupler

system), a grader for road maintenance, a small bulldozer for site preparation and tailings

work, and other smaller equipment.

ADMINISTRATION

The office area includes an administration building, as well as warehousing facilities,

first aid and medical area, and a mine rescue training/emergency control station area.

Arrangements will be required to secure the proper mine rescue equipment from the

Government of Quebec, including BG-4 and ancillary apparatus to respond to an

underground emergency. The administration and mine staff office will be two-floor

building with dimension of 40 ft. by 60 ft. A 40 ft. by 60 ft. mine dry and a 40 ft. by 40

ft. area for a safety/training/first aid/mine rescue centre will also be built. A two-floor 30

ft. by 40 ft. area will be utilized for warehousing inventory.



TAILINGS AND WASTE ROCK DISPOSAL

Based on information provided to Golder, the annual tailings production for the

Project will be approximately 750 tpd, or in the range of 235,000 tpa to 285,000 tpa, with

a total tailings production of 1.8 million tonnes of fine tailings solids over the life of the

mine. There are two options for the mine tailings management: place a percentage of the

tailings solids production as paste backfill in the underground mine cavity, or have the

remaining solids deposited on surface in an engineered tailings management facility

(TMF). For the purpose of this report, it is assumed that all tailings solids production will

be deposited on the surface in an engineered TMF. The total mass of 1.8 million tonnes

of tailings solids at a dry density of 0.9 tonnes per cubic metre will occupy a volume of

approximately 2,000,000 cubic metres in the tailings management facility at the end of

the mine life.

An engineered TMF for the Project will consist of a compacted earth fill perimeter

dyke providing containment for the deposited tailings stream and the impacted direct

precipitation, and will also serve as a diversion of surface water runoff from entering the

tailings storage pond. Liner requirements for hydraulic containment of the tailings fluids

will be based on the hydraulic conductivity of the surficial soils and bedrock at the TMF.

These requirements will be determined in consultation with the regulatory authorities in

the future detailed design stage after the location and size for the required tailings

management facility has been confirmed.

To improve the storage volume available within the constructed perimeter dyke, the

tailings management facility will be located at the site of an existing depression at a

suitable location in proximity to the tailings production mill site. At this stage of the

Project, it is considered that two lakes located in the upper part of the project watershed

would be a suitable location for constructing the tailings management facility. To further

improve the available storage volume, it is proposed that the surficial soils around the

shoreline of the selected lake site and within the perimeter dyke alignment be excavated,

and selectively used for borrow material in the construction of the compacted earth fill

perimeter dyke.



The footprint area required that the project total tailings solids storage volume should

be about 700 m by 700 m (490,000 m2) assuming that the tailings deposit thickness was

in the order of four metres. This TMF may be created by constructing perimeter dykes in

the order of five metres high, assuming that the lake basin (drained) will have an average

depth of about two-metres, and that about three metres of tailings may be impounded

against the compacted earth fill dyke, leaving approximately two metres of freeboard for

water cover and wave run-up provision at the end of mine life.

In addition to the fine tailings solids, the TMF will receive process water and direct

precipitation, which will form a decant fluid at the surface of the tailings and will be

recycled back to the mill for use as process water. Some of the process water will remain

within the pore space of the tailings solids, and a minimum 0.5 m thick water cover will

be maintained over the deposited tailings during the active mine life. At the end of mine

life, a permanent cover will be provided for the TMF.

The approximate dimensions of the compacted earthfill perimeter dyke will be:

• height of 5 m

• crest width of 5 m

• side slopes of 3H:1V

• total perimeter length of 2,800 m

• compacted fill volume of approximately 280,000 cubic metres

Approximately 1.2 million tonnes of waste rock will be generated during the life of

the mine, 50% of which will be returned underground as backfill. There is no

environmental geochemistry data for the waste rock. The current assumptions are that all

waste rock will be stored within the mining infrastructure footprint and that careful

segregation of mineralized waste and “clean” waste rock would be carried out during

mining operations to optimize water treatment cost and facilitate closure.



ENVIRONMENT

PERMITS AND LICENSING The Project is subject to the federal and provincial environmental assessment

processes. The Project requires a licence from the Canadian Nuclear Safety Commission

(CNSC) and hence will be subject to the federal process, as the Canadian Environmental

Assessment Act (CEAA) states that a project requires an environmental assessment (EA)

when the project involves a federal licence. The Project will be subject to the provincial

process, as Quebec’s Environment Quality Act and the James Bay and Northern Quebec

Agreement (JBNQA) specify that mining projects are automatically subject to the

environmental and social impact assessment (ESIA) and review procedure.

Both levels of government may collaborate in a single cooperative environmental

assessment process, although this remains to be determined because projects on the

territory governed by the JBNQA are not tied to the Canada-Quebec Agreement on

Environmental Assessment Cooperation.

In accordance with the CEAA and its regulations, CNSC oversees EAs to make sure

nuclear projects are safe for the environment. CNSC works in concert with the Canadian

Environmental Assessment Agency, while its Commission Tribunal is responsible for

making many EA decisions.

The Project is located on the territory governed by the JBNQA south of the 55th

parallel. As such, the tripartite Quebec/Canada/Cree Evaluating Committee (COMEV)

would issue guidelines for the ESIA based on the Project’s preliminary information

statement. Once the ESIA is completed, it would be reviewed by the bipartite

Quebec/Cree Review Committee (COMEX). COMEX would communicate its

recommendations to the Administrator of the Ministry of Sustainable Development,

Environment and Parks (MDDEP), who would make a final decision based on the

COMEX recommendations.

Once the Project has EA/ESIA approval, it is in a position to advance into regulatory

permitting. There are a number of federal and provincial permits that may be required



depending on the specifics of the Project, including a Mine Facility Licensing Manual

which is specific to uranium mines.

BASELINE STUDY Environmental baseline data were gathered to obtain a general overview of the

physical, biological, and human environment contiguous to the Project. More field work

is required to complete the baseline data to the standard of an eventual environmental

impact assessment.

Field work was conducted in 2007 and 2008 for the Hydrology, Surface Water and

Sediment Quality, Hydrogeology and Groundwater Quality, Fish and Fish Habitat,

Wildlife and Birds, and Vegetation and Wetlands components.

A desktop review was done for all of the components mentioned above, as well as the

following: Air and Climate, Noise, Soil and Terrain, Archaeology, Visual, Socio-

Economics, Land and Resource Use, and First Nations.

In general, a Local Study Area (LSA) was outlined which extended approximately 3.5

km around the Project area and 750 m on either side of the winter access road; however,

the actual area studied can vary according to particular needs of specific environmental

components.

The desktop review for the Air and Climate component indicated that the four nearest

meteorological stations are located 85 km, 155 km, 225 km and 280 km away from the

Project site. The data from these stations give an indication of possible conditions at the

Project; however, site-specific data are needed. In order to meet this data requirement, a

meteorological station has been installed at the site in 2007. Meteorological data from

the meteorological station with complete data records located nearest to the Project site

indicates average yearly precipitation of 946 mm for the period of record (1971-2000).

The monthly average temperatures and monthly precipitation for this station are

respectively -21.0ºC and 57.2 mm for January, -2.2ºC and 58.8 mm for April, 14.6ºC and

131.0 mm for July, and 1.5ºC and 79.7 mm for October.



Available regional data on air quality are very limited, and since the nearest two

stations are located 350 km and 515 km away from the site, no data on current local air

quality conditions are available.

The desktop review for the Noise component revealed no major issue, as the site is in

a remote area, far from any permanent anthropogenic sources or receptors, and is not

classified in any provincial zoning category. Noise, however, may factor in when

considering spiritual or special areas used by First Nations, wildlife and birds, or health

and safety for on-site workers.

The desktop review and field work for the Hydrology component identified all main

water courses within the LSA, established the Project’s local watershed (estimated at 24.8

km2) and outlet point, and allowed to assess stream flows at three stations within the

Project local watershed in September and October 2007. The recorded flows during the

observation period ranged between 2.1 m3/s and 0.1 m3/s for a 6.7 km2 sub-watershed,

between 3.3 m3/s and 0.25 m3/s for an 11.1 km2 sub-watershed, and 0.3 m3/s and 0.07

m3/s for a 3.2 km2 sub-watershed. The data currently available are not sufficient to

establish flows in drought/flood conditions or average yearly flows; however, a field

program is currently under way and will provide additional data.

A field program for the Surface Water and Sediment Quality component including

collection and analysis of surface water and sediment samples was carried out in 2007.

Water samples were analyzed for metals, radionuclides, major ions, nutrients and

suspended solids. Sediment samples were analyzed for metals, radionuclides and organic

carbon. Analytical results indicated exceedances of provincial criteria for aluminum and

suspended solids for surface water samples and exceedances of federal criteria for

cadmium in sediment samples. There were no applicable criteria for radionuclides at the

time of writing the report. The exceedances found are believed to reflect the natural

background as there is no known industrial activity that may have affected the surface

water and sediments in the area.



The desktop review and field work for the Hydrogeology and Groundwater Quality

component allowed evaluating geological conditions and assessing in-situ hydraulic

conductivity of the main sandstone facies and the Matoush Fault. The median value of

estimated hydraulic conductivities for the various facies, contacts between sandstone

facies and the Matoush Fault ranged between 2 x 10-6 m/s and 5 x 10-6 m/s. These values

are typical of upper range conductivity values for sandstone and are consistent with

drilling fluids loss of return reported during the exploration drilling program. Shallow

groundwater samples were collected and analyzed for metals, major ions and

radionuclides. Analytical results showed no exceedances when compared with applicable

mine water discharge criteria (Québec Directive 019).

Bedrock geology is fairly consistent across the site, with one major sandstone unit

(Indicator Formation) present down to a depth of approximately 800 m and resting on

Archean granitic basement. At this point, only the six upper facies (ACF1 to 3 and CBF1

to 3) have been submitted to in situ permeability testing.

Overburden material cover is also very consistent across the site with a glacial till

blanketing the site and the nearest bedrock outcrop located kilometres away from the

Project site. There is no available field data on groundwater level in the overburden and

overburden permeability at the time of writing this report. Vertical gradients between the

overburden and the bedrock along with associated recharge and discharge areas can,

therefore, only be inferred from topography because there are no field data to substantiate

the interpretation.

The conceptual hydrogeological model is, therefore, fairly simple at this point and

can be summarized as a series of sandstone facies (ACF1 to 3 and CBF1 to 3) of similar

bulk hydraulic conductivity, which are likely to be anisotropic as a result of subhorizontal

facies contacts and the subvertical Matoush Fault.

The desktop review for the Soil and Terrain component initially provided regional

information on geology and drainage based on available government maps. Since then, a

detailed compilation of the Project exploration drilling data and detailed interpretation of



overburden geology (Poly-Geo, 2008) has been completed. The updated description of

regional and local bedrock and overburden geology has been presented in Section 7 of

this document.

As for drainage and topography, the terrain surrounding the Project site is generally

flat and is characterized by gentle rolling hills. Slopes are thus not very steep and vary

between 8% in the steepest hills north of the study area and 3% in other areas.

Topography in the Project area is strongly influenced by the shape of the glacial

deposits which show northeast-southwest alignments. The axis of most lakes is a

reflection of the topography of glacial till deposits and drumlins, which produce

topographic lows and highs aligned on a northeast-southwest axis. Maximum and

minimum elevations found on the study area are approximately 800 m in the northwest

and 715 m in the southeast of the site. The study watershed is approximately 24.8 km2

and flows to the southeast. On a regional scale, surface water from the Project site flows

to the south towards the projected biodiversity reserve that is some 19 km from the

Project.

None of the special status species identified during the Vegetation and Wetlands

desktop review was observed during the field work, although the habitat of the Project

site is suitable for one of these species, Hudsonia tomentosa, classified as a species

susceptible of being designated threatened or vulnerable.

The desktop review and field work for the Fish and Fish Habitat component revealed

that lakes in the area are acidic, oligotrophic, and the habitat features are highly uniform.

Brook trout (salvelinus fontinalis) was the dominant species in the 2007 and 2008

collections. Based on available historical information and the results of the 2007-08

sampling, none of the fish species known to be in the area are currently listed as

threatened or endangered under provincial or federal legislation. No spawning grounds

were confirmed during the 2007 field work; the results of the spring and fall 2008

spawning surveys have yet to be compiled. Fish flesh and bones were analyzed for

metals and radionuclides.



The desktop review and field work for the Wildlife and Birds component identified

species likely to occur in the area. The rusty blackbird, Euphagus carolinus, listed as

species of special concern by the Committee on the Status of Endangered Wildlife in

Canada (COSEWIC), was the only listed wildlife species observed in the LSA during

field work. A woodland caribou, Rangifer tarandus caribou, listed as threatened by

COSEWIC and as vulnerable under the provincial jurisdiction, was observed

approximately 19 km from the Project site. Other species at risk may be present given

the regional habitat.

The desktop review for the Archaeology component revealed that there are no known

archaeological sites; however, eight zones where work with respect to the Project is

foreseen show archaeological potential.

The desktop review for the Visual component included maps, photographs, imagery,

and a video of the visual landscape of the area consisting of forests, wetlands, old burned

areas, lakes, and cleared areas for the mine workers camp and the winter access road.

The desktop review for the Socio-Economics, Land and Resource Use, and First

Nations components compiled information on the demography, education, employment,

and health of the community, traditional knowledge, and available services and

infrastructures. The site is located on Category III Quebec public land that falls under the

JBNQA, where the Cree have exclusive rights to harvest certain aquatic species and fur-

bearing mammals and to participate in the administration and development of the land.

Information for these components will be completed during pre-consultation activities

that started in the fall of 2008.

SITE WATER BALANCE A preliminary site water balance has been made for the Project site based on available

regional meteorological data, estimated preliminary footprint of the surface infrastructure,

fresh water requirements for the ore processing plant, and order of magnitude estimations

of underground mine water seepage.



Since available site-specific data are currently limited and details of the surface

infrastructure including tailings and waste rock storage areas are not determined, the site

water balance was made on a yearly basis and was based on recorded average

precipitation data from a regional meteorological station located 150 km southeast of the

site.

Assumptions made for the water balance are as follows:

• Yearly precipitation within the mine infrastructure footprint including tailings and waste rock storage areas is collected, evaporation is assumed to be negligible and all collected water is included in the mine effluent.

• Yearly precipitation was assumed to be 950 mm. • Process water is lost (mainly as tailings pore water) at a rate of 40 m3/h1. • Precipitation on other mine-related infrastructures, such as the permanent

camp and roads, is not considered to be impacted by mining activities and is not considered as part of the mine effluent.

• All surface runoff outside the tailings storage facility and the mine

infrastructure footprint (including waste rock storage area) will be diverted and is not considered in the mine effluent.

• The area of the tailings storage facility is estimated at 490,000 m2. • The mine infrastructure footprint including waste rock storage area2 is

estimated to be 160,000 m2. • Underground mine seepage rate was estimated to be potentially in excess of

0,5 m3/s for a 1,000 m long, 600 m deep mine with subvertical tabular openings without seepage mitigation measures used (i.e., no grouting or cemented backfill). This is based on the assumption that the bulk hydraulic conductivity of the rock mass is in the order of 10-6 m/s and there is no hydraulic connection between the mine and the adjacent lakes and streams.

Based on these data and assumptions, the yearly water balance indicates positive

influx of water from the tailings area, waste rock area and mine infrastructure footprint

(0.6 million m3/yr) and from mine seepage water (15.7 million m3/yr) and losses (0.4

million m3/yr), resulting in a net effluent volume of approximately 15.9 million m3/yr.

1 Information provided to Golder by Melis Engineering Ltd. 2 Information provided to Golder by Scott Wilson Roscoe Postle Associates Inc.



WATER TREATMENT AND DISPOSAL SCHEME A water treatment plant will be constructed at the mine site. All mine water and

process water will be tested to ensure it meets the effluent quality objectives prior to

being discharged to the environment.

The Project site is located at the top of a watershed and the discharge rate of the mine

effluent is expected to be high as a result of potential high seepage rates from the

underground mine. Since average yearly stream flows near the site (within a radius of a

few kilometres around the site) are likely to be less than the mine effluent, and

considering the physical limitations of natural streams to take up additional flow without

generating uncontrolled erosion and the potential sensitivity of aquatic species, it is

unlikely that the mine effluent discharge point will be close to the mine site.

HISTORICAL ENVIRONMENTAL IMPACTS There are no known previous industrial activities within the Project site. The only

known activities other then fishing and trapping are related to mining exploration and

have included geophysical surveys, geological mapping, soil sampling and exploration

drilling. Since the Project drilling program has systematically reported that there has

been no return of drilling fluids, it appears that no drill cuttings have been brought to

surface.

Other impacts from fuel storage and waste water from the exploration camp would be

expected to be negligible, since the site is fairly new and has been submitted to strict

environmental controls.

Thus, impacts from previous activities at the site would be expected to be negligible.

MONITORING Monitoring for radiation in the underground mine and within the mining facilities will

be addressed by the Mine Facility Licensing Manual. Environmental monitoring will

include monitoring of air and surface water quality at several stations around the site,

including reference points outside the Project watershed. Aquatics species, including



tissue analysis, will also be monitored, as well as groundwater at critical locations around

the site.

LICENSING COSTS Initial licensing cost and Environmental Impact Study for the project would be

expected to be in the range of $8 million to $10 million. Yearly cost for maintaining the

licence would be in the range of $0.5 million to $0.9 million a year, which would equate

to $5 million to $9 million for a life of mine of 10 years. This does not include internal

costs related to health and safety and environmental monitoring required for operating the

mine.

MINE CLOSURE PLAN The mine closure plan will address the waste (i.e., tailings and waste rock) storage

areas both in terms of physical and chemical long-term stability. The closure plan will

also address decommissioning and safe disposal of the mining infrastructure; remediation

of impacts from mining, if any; disposal of remaining reagents/products or waste stored

on site; and water management. Finally, the closure plan will address long-term site

safety issues, in particular those related to the mine openings.

The mine closure costs will vary according to initial engineering and adherence to

good environmental practices at the site during the entire mining operation. Progressive

remediation and proactive management of potential environmental liabilities such as

waste rock and tailings storage areas can have a major effect on the closure costs.

Closure costs can range between $30 million for well-managed, low-impact sites and

$100 million for sites with more complex conditions such as existing impacts,

problematic waste rock and tailings, and complex water management issues. For the

purpose of this report, it is assumed that the site will be well managed, with low impacts,

and closure costs will be in the range of $30 million. Engineering and permitting costs

related to closure can be expected to range between $5 million and $10 million. It is

noted that the Quebec mining regulations and CNSC require that a mine closure plan be



prepared and the funds required for site closure set aside gradually in the early stage of

mining operations.

MARKETS

The principal commodity at Matoush is freely traded, at prices that are widely known,

so that prospects of sale of any production are virtually assured. Scott Wilson RPA

assumed a long-term uranium price of US$75/lb of U3O8 for the Base Case, as

recommended in the market analysis that follows. The following market study was

completed by TD Energy Associates Limited.

WORLD DEMAND AND SUPPLY The price of uranium more than quadrupled between January 2005 and July 2007,

primarily as a result of supply shortages and price speculation. A sustained lack of

investment during the 1980s and 1990s has left the uranium industry dependent on non-

mined production, and in 2007 less than 65% of annual uranium demand from utilities

was met from mine production.

Nuclear power currently accounts for approximately 16% of world production of

energy. World electricity demand is forecast to double by 2030 (World Nuclear



Association (WNA), 2007). If nuclear power’s share of future electricity production is to

remain at current levels, nuclear capacity will double by 2030. This is similar to the high

scenario in the recent WNA nuclear forecast (IEA, 2007) and results in a demand for

uranium in 2030 which is more than double current uranium demand.

Availability of uranium for new nuclear reactors is not considered an impediment to

growth of nuclear capacity, as the world known reserves of uranium are more than

adequate to satisfy requirements well past 2030.

The critical issues are how fast uranium production capability will be made available

and at what price. To meet forecast uranium demand, uranium production capacity will

have to more than double by 2020 and almost quadruple by 2030.

DEMAND

The uranium demand forecast used in this study is based on a nuclear build scenario

that maintains nuclear power’s share of electricity production close to the present day

level of 16%. The drivers for this scenario are concerns about climate change, the

increasing cost of fossil fuels and their security of supply, and the improved economics of

new nuclear plant. There also appears to be a growing recognition at the national level

that nuclear power is a key component in any carbon reduction energy strategy.

As uranium is purchased for the first cores of new reactors several years in advance of

start-up, the forecast for nuclear capacity is extended to 2030 while that for uranium only

goes out to 2020.

Nuclear Capacity Forecast to 2030 At the end of March 2008, there were 439 nuclear reactors worldwide with a capacity

of 372GWe. Nuclear capacity is forecast to increase to 520GWe by 2020, up 40% from

2006, and almost double by 2030. The main growth areas to 2020 are China, Korea and

India, where there has been rapid economic growth and strong electricity demand, and

also Russia, where the government is committed to nuclear expansion.



Post 2015, growth continues in these areas, but there is also a significant number of

new builds in North America and Central Europe as concerns about increasing fossil fuel

prices, security of supply and carbon emissions are translated into new orders for nuclear

reactors.

World Nuclear Capacity Forecast (GWe)

0

100

200

300

400

500

600

700

800

2006 2008 2010 2012 2014 2016 2018 2020 2022 2024 2026 2028 2030

GW

e

N America Europe & FSU East Asia Other

Uranium demand forecast to 2020

Reactor requirements are forecast to increase from 167 million pounds in 2006 by

almost 30% to 212 million pounds in 2015 and 40% to approximately 240 million pounds

by 2020.



World Uranium Requirements 2006 -2020

0

50

100

150

200

250

300

2006 2007 2008 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020

m lb

s U

3O8

N America Europe & FSU East Asia Other

Uranium demand is dependent on uranium requirement, which is based on installed

nuclear capacity, but is also influenced by several other factors. These include capacity

factors, the actual power generated by a reactor during year expressed as a percentage of

its generation at full power for the same period; technical factors, related to operation,

fuel cost optimization and fuel design; and purchases for first cores and to maintain

‘desired’ uranium inventory levels.

SUPPLY

Uranium production from existing mines is forecast to peak in 2012 as mines in

Kazakhstan reach full production and then gradually decline. Over 70% of this

production is forecast to be from mines in Africa, Australia, Canada and Kazakhstan.



Supply from Existing Mines 2006 to 2020

Australia

Africa

Canada

Kazakhstan

Russia

UkraineUSA UzbekistanOther

-

20 000

40 000

60 000

80 000

100 000

120 000

140 000

2006 2007 2008 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020

000

lbs

U3O

8

When forecast supply from existing mines is compared with world demand, it is clear

that there is a need for new mines to be developed. Some of the key factors that will

impact on the introduction of supply from new mines will be the political and regulatory

environment of the project and price.



World Supply from Existing Mines vs World Demand 2006 -2020

0

50000

100000

150000

200000

250000

300000

2006 2007 2008 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020

000

lbs

U3O

8

Demand Supply

Secondary Supply Sources

The gap between supply and demand historically has been filled by secondary

supplies and non-government inventory. As the diagram below shows, the end of the

HEU Agreement in 2013 will reduce secondary supplies available to the global markets

by over 40%.



Secondary Supply Sources

0

10 000

20 000

30 000

40 000

50 000

60 000

70 000

2003

2005

2007

2009

2011

2013

2015

2017

2019

000s

lbs

U3O

8

HEU I Re-Enriched Tails

Russian Government Stockpile Sales US/USEC Government Stockpile Sales

MOX & RepU

PRICE EXPECTATIONS AND CONTRACTING The uranium spot price peaked in July 2007 at $135/lb U3O8 and is currently in the

mid-$40s/lb U3O8 to mid-50s/lb U3O8 range. However, the spot price tends to reflect

conditions prevailing in the short-term market, which represents approximately 20% of

the total market for uranium, and is predominantly affected by perceptions of supply and

demand in the year as well as sellers’ external liquidity requirements. The market is very

thin and transactions as low as 100,000 lb of U3O8 can have a significant impact on price.



Source: The Ux Consulting Company, LLC http://www.uxc.com/

The fundamentals driving longer term supply and demand remain strong. Demand is

forecast to double by 2030 and production from existing mines currently accounts for less

than 65% of current demand. As a result of many years without investment in

exploration and development, the supply side of the industry has been slow to increase

production but swift to announce plans for future production.

Demand for the next 15 year period will be primarily from reactors that are in

operation, under construction, or at an advanced level in the planning process and should

be relatively predictable.

Production is significantly more uncertain. To meet demand, existing mines will

have to operate at full capacity and new mines will have to be brought on stream. The

situation is further complicated as two of the major new mines represent, between them,

close to 40 million pounds U3O8 per year (a third of anticipated production in 2008) and

any delays to timing will have a dramatic impact on availability of supply and pricing.

Until sufficient new production is brought on line to bring supply and demand into

balance, price is likely to be event-driven and for this reason two scenario-based forecasts

have been developed. The main differences between the scenarios are described below.

Neither scenario includes projects owned by junior mining companies which have not yet

reached a pre-feasibility stage.

http://www.uxc.com/



It is assumed that these projects will come into production to meet the demand for

new mines if the price is at an appropriate level. However, the current level of the spot

price coupled with recent events in the financial markets have severely reduced junior

mining companies access to funding, resulting in delays and potential cancellations of

some projects.

THE ANNOUNCED PLANS SCENARIO

The Announced Plans scenario assumes that all announced plans by existing mines

and planned mines (post pre-feasibility stage) will go ahead. The Olympic Dam mine

will expand production from 2013, ramping up to full capacity by 2019; Cigar Lake will

start production in 2011 and reach full capacity in 2014; Uranium One’s Dominion is

assumed to produce at a rate of 3.5 million pounds U3O8 per year; and Energy Resources

of Australia’s Jabiluka is assumed to commence production in 2016. This production

scenario has already been proved optimistic by Uranium One’s recent announcement that

the Dominion mine has been placed on care and maintenance.

REASONABLE DOUBT SCENARIO

Production from the Olympic Dam expansion is assumed to start in 2019 and peak at

10 million pounds rather than 22.5 million pounds. Cigar Lake production profile is

maintained, but start-up is delayed until 2014. Dominion is assumed to produce at 0.6

million pounds U3O8 per year and Jabiluka is assumed to start production in 2019.

This is still a conservative forecast, as production in all other areas is assumed to

develop in line with current announcements allowing approximately 20 million pounds of

new production from Kazakhstan by 2012, and approximately 15 million pounds of new

production from Africa by 2013 (6 million pounds from Trekkopje in Namibia). As the

recent shelving of the Dominion mine has proved, even assumptions about production

scaled down from existing and planned mines could be on the high side.

There is also some uncertainty regarding the availability of secondary supplies. The

HEU (highly enriched uranium) agreement ends in 2013, taking 20 million pounds of

equivalent U3O8 (e-U3O8) out of the market, but Russian re-enrichment of tails is

assumed to continue (9 million pounds). This has been assumed to continue under the



Reasonable Doubt scenario but could be considered optimistic, as there have been

announcements by the Russians that they will be ending re-enrichment of tails, and this

could impact on some or all of this quantity.

Production cost estimates for the projects outside Kazakhstan that will be required to

fill the need for new mines fall into a range between mid-$40 per lb and high $60 per lb.

It is assumed that in order to develop these mines and encourage exploration, the long-

term price will have to reflect this reality.

SUPPLY AND DEMAND – ANNOUNCED PLANS SCENARIO

Under the Announced Plans scenario, supply is in excess until 2014 and then stays in

balance until 2019, after which new supply is needed.

Price Implications:

• 2008 to 2012 - no upward pressure on price: o Long-term price range $60 to $90/lb o Spot price $45 to $65/lb

World Supply and Demand from Existing, Planned and Potential Mines and Secondary Sources - Announced Plans Scenario

0

50000

100000

150000

200000

250000

300000

350000

2008 2010 2012 2014 2016 2018 2020 2022 2024

000

lbs

U3O

8

Demand

Secondary Supply

Announced Plans Scenario for New mines

Existing Mines



• 2013 to 2018 - pressure on price as small changes to production could cause deficit

o Long-term price range $70 to $90/lb, market price: o Spot price range $65 to $120/lb

• 2019 to 2020 - new mines needed post 2020:

o If supply response inadequate: Long-term price range $80 to $100/lb, market Spot price range $80 to $140/lb

o If supply response adequate: Long-term/Spot price range $50 to $70/lb

• Post 2020: Long-term/Spot price range $50 to $70/lb

SUPPLY AND DEMAND - REASONABLE DOUBT SCENARIO

World Supply and Demand from Existing, Planned and Potential Mines and Secondary Sources – Reasonable Doubt Scenario for

Production

0

50000

100000

150000

200000

250000

300000

350000

2008 2010 2012 2014 2016 2018 2020 2022 2024

000 lbs U3O8

Demand

Existing Mines

Reasonable Doubt scenario for new mines

Secondary Supply



The Reasonable Doubt scenario differs from the previous primarily in the timing and

production levels of three mines.

Under this scenario, there is a need for additional new mines from 2013.

Price implications:

• 2008 to 2011 - no upward pressure on price: o Long-term price range $60 to $90/lb o Spot price $45 to $65/lb

• 2012 to 2016 - pressure on long-term price to bring on new mines:

o Long-term price range $85 to $110/lb, market price o Spot price range $65 to $140/lb

• 2017 to 2022:

o If supply response positive and new mines available in timely fashion: Long-term price range $75 to $90/lb, market Spot price range $70 to $90/lb

o If production response inadequate: Long-term price range $90 to $120/lb Spot price $80 to $160/lb

• Post 2020:

o Long-term/Spot price range $50 to $70/lb

PRICE FOR EVALUATION USE Under the Announced Plans scenario, it would be possible to sign contracts at prices

ranging from $60/lb to $90/lb over the LOM. An evaluation price of $75/lb over the

LOM would seem appropriate.

The price range for the Reasonable Doubt scenario is from $60/lb to $110/lb, and the

suggested evaluation price is $85/lb.



MARKETING STRATEGY CONTRACT STRATEGY

TABLE 18-2 STRATECO’S PLANNED PRODUCTION Strateco Resources Inc. – Matoush Project

2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022

M lbs U3O8 1 2 2 2 2 2 2 2 2 2 2

Matoush is planned to come into production at a time when the market will be

looking for new supply.

Three of several options that would be open to Strateco are direct marketing, sales via

an intermediary, or sales to a trader/primary producer for part of or total production.

Considering direct marketing by Strateco and assuming an environment of volatility

in the spot market and an upward trend in long-term market prices, it would be prudent to

adopt a contracting strategy which would allow at least 50% of production to be sold

under medium- to long-term contracts three to four years ahead of the delivery year. The

uncommitted production could then be sold either under long-term or spot market

contracts closer to the years of delivery, depending on market conditions.

Strateco would start marketing product three to four years before the first year of

production from Matoush.

The diagram below illustrates a marketing strategy under which Strateco sells

multiple term contracts of around 250,000 lb U3O8 per year for a term of four to six years

and leaves 200,000 lb to 500,000 lb per year for sale in the spot market in the year of

delivery. This strategy allows Strateco to sell into the market every year, thus keeping

abreast of market movements and benefiting from any improvements in market

conditions.



Volume Contracted and Year in which Contract Signed

0200400600800

100012001400160018002000

2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022

Year of delivery

000

lbs

U3O

8

2009 2010 2011 2012 2013 2014 2015 2016 2017 20182019 2020 2021 2022

Spot Sales

CONTRACT TERMS Two significant aspects of any long-term contract are:

1. Pricing

Some of the main pricing mechanisms are the following:

• Base Price Escalated A price is agreed at the time of contract signature and is fully or partially escalated by an agreed indicator for the duration of the contract.

• Market Price The price paid is the spot market price, as published by an agreed source, such as Ux Consulting Company or TradeTech, at the time of delivery (or average of an agreed period of time before the delivery), with or without a discount or premium (agreed at the time of signature). This price mechanism usually includes a floor and/or a ceiling price, which, if agreed, would escalate over the period of the contract. The price paid would be the spot price, if it is higher than the floor and lower than the ceiling. If the spot price is below the floor price, the floor price would be paid, and if it is above the ceiling price, the ceiling price would be paid.



• Formula Price The price would be any mix of the Base Price mechanism and the Market price mechanism or could include other published price indicators.

2. Delivery Location The price is the price paid at an agreed conversion facility, such as ConverDyn, Metropolis, Ill, USA; Cameco, Blind River, Canada; or Comurhex, Pierrelatte, France. Strateco would bear the cost of shipping to the conversion facility. Costs of shipping to France are almost 50% higher than shipping to North America.

TARGET MARKET The USA currently represents approximately one third of world uranium

requirements and by 2020 will still account for approximately 25%. Domestic production

in the USA is limited and the US nuclear program will be dependent on imports.

Strategically, the USA and Canada would appear to be Strateco’s primary target

markets:

• Large uranium requirement looking for secure imports • Lower costs of transport when compared with transport to Europe • Major nuclear utilities seeking to improve diversity of uranium supply

A secondary target would be Far Eastern or European utilities with conversion

contracts in North America.

CONTRACTS

Strateco does not have any existing contracts at present that would apply to the mine

operation. Future contracts will be necessary for the delivery of hydroelectric power, if

available, the delivery of both oxygen and sulphuric acid for the processing, the haulage

of yellowcake concentrate to the refinery, the supply of explosives and accessories for the

mine, and other minor items. Scott Wilson RPA is of the opinion that none of these

contracts will present any particular problems to hinder a smooth mining operation.



LIFE OF MINE PLAN

In order to advance the Project, the completion of a Feasibility Study will be required,

with additional work and refinement in several areas. The decision to proceed to the

construction phase will allow for financing and final permitting of the Project.

Construction will be dependent upon equipment lead times and the logistics of

transporting materials and equipment to site.

Scott Wilson RPA notes that the Life of Mine Plan (LOMP) does not consider

underground mine development which has been included in the initial exploration

program.

PRE-PRODUCTION SCHEDULE Construction at the mine is scheduled to start approximately two years prior to

production start-up. During this period, major equipment items for underground mining,

surface operations, ventilation, backfill and mine services will be purchased and

transported to site. Construction of the process plant, expected to last approximately 20

months, will also commence once supplies and materials have been shipped to site.

One year prior to production start-up, construction of the mine site buildings will

begin, including the expansion of the current camp facilities, and additional diesel

generator capacity will be installed. Pre-production development, expected to last

approximately one year, will also commence in the AM-15 Zone to provide access to the

required number of working areas for the start of production. Allowances are also made

to ensure sufficient access to fresh air and exhaust routes.

PRODUCTION SCHEDULE Production for the first two years of operations will come from the AM-15 Zone. In

the third year, the MT-34 and MT-22 zones will begin to supply the production.

A ramp-up period of two years has been assumed to reach the full production rate.

The production rates for Year 1 and Year 2 have been assumed to be 500 tpd and 675 tpd,



respectively. The full rate of 750 tpd will be reached in the third year of production, and

will be maintained through the remaining five years of mine life.

Although the schedule has not been optimized for grade, starting production in the

AM-15 Zone does provide access to a higher-grade portion of the deposits. In general,

the MT-22 Zone hosts lower grades. Therefore, during the years in which the MT-23 and

MT-22 zones are mined concurrently, the production from each zone was varied to

ensure a balanced profile and maintain at least 2.0 million pounds of U3O8 annually.

The LOMP is shown in Table 18-3.

Scott Wilson RPA notes that production grades for the LOMP have been assigned

according to the average grade per level mined. Applying more detail to the production

schedule is not warranted for this level of study.

-2 -1 1 2 3 4 5 6 7 8 TOTAL

PHYSICALS

Ore Production

AM15 & MT34 '000 t 175.0 236.3 131.3 65.6 65.6 131.3 16.1 821.1 Grade %U3O8 0.633% 0.454% 0.371% 0.736% 0.827% 0.520% 0.469% 0.542%

MT22 '000 t - - 131.3 196.9 196.9 131.3 172.3 828.5 Grade %U3O8 0.000% 0.000% 0.353% 0.492% 0.310% 0.223% 0.248% 0.333%

Total '000 t 175.0 236.3 262.5 262.5 262.5 262.5 188.4 1,649.7 Grade %U3O8 0.633% 0.454% 0.362% 0.553% 0.439% 0.372% 0.267% 0.437%Contained Metal '000 lbs 2,440.7 2,362.9 2,097.0 3,201.2 2,540.4 2,151.9 1,108.7 15,902.8

Production Rate tpd 500 675 750 750 750 750 538 - 689

METALLURGY

Mill Feed '000 t 175.0 236.3 262.5 262.5 262.5 262.5 188.4 1,649.7 Grade %U3O8 0.633% 0.454% 0.362% 0.553% 0.439% 0.372% 0.267% 0.437%Contained Metal '000 lbs 2,440.7 2,362.9 2,097.0 3,201.2 2,540.4 2,151.9 1,108.7 15,902.8

Recovered Metal 97.6% '000 lbs 2,382.1 2,306.2 2,046.7 3,124.4 2,479.4 2,100.2 1,082.1 15,521.2

YEAR

TABLE 18-3 LIFE OF MINE PLANStrateco Resources Inc. - Matoush Property

SCO

TT

WIL

SON

RPA

ww

w.scottw

ilson.com

Strateco R

esources Inc. – Matoush Project

Preliminary A

ssessment –D

ecember 17, 2008

Page 18-48



DEVELOPMENT SCHEDULE Development from the exploration program will give access to all the levels required

to mine the AM-15 Zone. Pre-production development includes lateral drifting on the

165, 180, 205, and 230 levels. Level access drifts will be extended from the ramp to

reach the mineralized zone, and a drift will be driven in ore to map the ore-waste contact.

Once this has been established, the footwall drift and crosscuts will be developed to

prepare for stope extraction.

A ventilation raise on the north end of the AM-15 Zone will be driven via alimak to

provide an intake route for fresh air, and drop raises will be developed between the pre-

production drifts to permit one-pass air movement through the openings. Additional

intake and exhaust raises will be developed from the 300 drill drift to provide a

ventilation route to the lower portion of the mine.

A pre-production development schedule is shown in Figure 18-7.

Once production begins, lateral and vertical development will advance in the AM-15

and MT-34 zones. A second independent ramp will be developed off the 300 drill level

to access the MT-22 Zone. As the MT-22 Zone is development-intensive and hosts a

relatively narrow mineralized zone, development is required in the first year of

production to prepare for extraction in the third year. All ramping is assumed to be at

minus 15% grade, and level spacing in the AM-15–MT-34 and MT-22 zones is assumed

to be 25 m and 15 m, respectively. All development is planned to be completed one year

prior to the end of mine life.

The single and multi-heading advance rates were assumed to be 6 m per day and 15 m

per day, respectively, amounting to an average of approximately 5,000 m per year. Re-

muck points and safety bays were assumed to be every 150 m and 30 m, respectively.

Additional allowances were made for the equipment setup time for alimak raising.

H1 H2

155 Level

Lateral Development

Ventilation Raises

180 Level

Lateral Development

Ventilation Raises

205 Level

Lateral Development

Ventilation Raises

230 Level

Lateral Development

Ventilation Raises

300 Level

Ventilation Raises

Ongoing (Lateral and Raising)

255 Level

280 Level

300 Level

325 Level

350 Level

375 Level

400 Level

425 Level

450 Level

475 Level

Ongoing (Lateral and Raising)

345 Level

363 Level

381 Level

399 Level

417 Level

435 Level

453 Level

471 Level

489 Level

507 Level

525 Level

543 Level

561 Level

579 Level

597 Level

615 Level

YEAR 7

MT-22 Zone

AM-15 & MT-34 Zones

YEAR 3 YEAR 4 YEAR 5 YEAR 6YEAR -1

ITEM YEAR 1 YEAR 2

December 2008

Matoush Project

Pre-productionDevelopment Schedule


Québec, Canada

Figure 18-7

18-50




Ventilation was a primary consideration in planning and scheduling the development.

To ensure that air was not being re-circulated, a one-pass ventilation system was ensured

by providing a fresh air source and exhaust route for each level.

CAPITAL COSTS

A pre-production capital cost of $297.3 million is estimated to bring the Project into

production. These costs include underground development, mine and process equipment

purchase and transportation to site, construction of surface facilities and infrastructure,

and installation of all services required to support the Project. The estimate includes

capital costs incurred over two years prior to the start of production, with no allowance

for escalation.

Direct costs for the mine and surface infrastructure, inclusive of equipment, materials

and labour, were estimated by Scott Wilson RPA. Mine and surface indirect costs,

inclusive of engineering, procurement, and construction management (EPCM), freight,

construction, owner’s costs, and contingencies were also estimated by Scott Wilson RPA.

Direct and indirect costs related to the process plant facilities were estimated by Melis.

Costs for advancing the Project to the construction stage, such as metallurgical testwork,

process development, permitting, environmental studies and project financing, are not

included in the capital costs.

Total Project capital costs are summarized in Table 18-4.



TABLE 18-4 CAPITAL COST SUMMARY Strateco Resources Inc. - Matoush Property

Pre-production Ongoing Total

Item ($ millions) ($ millions) ($ millions)

Mine 28.2 - 28.2 Process 149.9 - 149.9 Infrastructure 15.4 - 15.4 Indirects 49.9 - 49.9 Contingency 53.9 - 53.9 Subtotal 297.3 - 297.3

Sustaining Capital - 15.6 15.6 Reclamation - 30.0 30.0

Total 297.3 45.6 342.8

PRE-PRODUCTION CAPITAL The pre-production capital cost includes items from the time of the construction

decision, through two years of construction and into the start of production.

MINE CAPITAL

Underground mine capital for pre-production includes the purchase of mobile

equipment for the mine and surface facilities, lateral and ventilation raise development,

mine services, construction of the cement backfill plant, and equipment for pumping and

ventilation. A summary of mine capital costs during the pre-production period is

presented in Table 18-5.

TABLE 18-5 PRE-PRODUCTION MINE CAPITAL Strateco Resources Inc. - Matoush Property

Item Total ($ millions)

Underground Equipment 13.9 Surface Equipment 1.3 Development 6.7 Ventilation 1.3 Services 1.0 Backfill Plant (Cement) 4.0

Total 28.2



The underground and surface equipment costs are estimated from the fleet

requirements and recent quotations from manufacturers. Lateral development and

ventilation raise development is assumed to be carried out by owner crews and

contractors, respectively. Raise development by alimak, estimated to cost $3,600 per

meter, is based on quotations from the contractor. Capital costs for ventilation are

inclusive of intake, exhaust, and auxiliary fans, as well as the mine air heating and

propane system. Capital for mine services includes equipment and supplies for the shop,

electrical substations, communication systems, and explosives storage. The cement

backfill plant capital costs include allowances of $1.5 million for construction and $2.5

million for equipment and facilities.

PROCESS CAPITAL

Process capital costs are estimated by Melis for areas within the battery limits of the

process plant, and by Scott Wilson RPA for other areas. A tabulation of the process

capital is shown in Table 18-6.

The capital costs estimated by Melis were completed on an order-of-magnitude basis

to the level of detail typical for a Class IV estimate (-15% to -30% / +20% to +50%).

Equipment requirements and approximate sizing were defined by the design criteria and

process flow sheets. The installed cost of each piece of equipment or material included

the installation labour and material costs, based on budget quotes or Melis file data.

Labour costs were assumed to be $100 per hour, which includes regular and overtime

pay, travel, living allowance, room and board, and flights. The costs of process piping,

electrical, and instrumentation to site were estimated to be $10,000 per tonne, $20,000

per tonne, and $75,000 per tonne, respectively. Freight to site was assumed to be $4,000

per 38 tonne truckload, and an exchange rate of 0.95 was assumed in converting US to

Canadian currency.



TABLE 18-6 PRE-PRODUCTION PROCESS CAPITAL Strateco Resources Inc. - Matoush Property


Storage and Crushing 4.6 SAG Mill Grinding 13.0 Ball Mill 7.9 Neutral Thickening and Leaching 7.0 Resin Loading Circuit 13.2 Resin Elution 4.1 Impurity Precipitation 2.6 Gypsum Filter 3.1 Uranium Precipitation 1.5 Uranium Thickener 1.2 Uranium Calcining 4.5 Uranium Packaging and Scrubbing 2.2 Tailings Neutralization 5.4 Reverse Osmosis 6.6 Effluent Treatment 7.2 Ferric Sulphate and Hydrogen Peroxide 1.6 Oxygen and Magnesia 1.1 Barium Chloride 0.5 Lime 4.5 Flocculent Mixing and Acid Plant 1.6 Fresh and Process Water 3.7 Low and High Pressure, and Instrument Air 1.1 Process Plant/Crusher/Oxygen Plant Buildings 48.0 Reagents (First Fills) 2.7 Tailings Storage 0.5 Mobile Equipment 0.5

Total 149.9

INFRASTRUCTURE CAPITAL Infrastructure capital costs include buildings, power, and administrative items.

Estimated infrastructure capital costs are summarized in Table 18-7.



TABLE 18-7 PRE-PRODUCTION INFRASTRUCTURE CAPITAL Strateco Resources Inc. - Matoush Property


Buildings 10.0 Ore Pad 0.3 Power 4.3 Administration 0.8

Total 15.4

Building costs were estimated using a unit cost of $250 per square foot, which

includes an allowance for additional base support with pilings. Expansion of the camp,

totalling 230 rooms, was estimated at $30,000 per room. Capital costs for power include

the diesel generators, which were estimated at $1,100 per kWh installed. Administration

capital costs include vehicles, general office supplies, and communication systems.

INDIRECT CAPITAL COSTS

Indirect capital costs for the mine and infrastructure have been estimated by Scott

Wilson RPA, and costs for the process plant were estimated by Melis. A summary of the

indirect capital costs is shown in Table 18-8.

TABLE 18-8 INDIRECT CAPITAL Strateco Resources Inc. - Matoush Property


EPCM 23.0 Contractor Overhead and Profit 6.7 Freight 2.1 Construction Mobile Equipment 1.0 Owner’s Costs 16.0 Other 1.1

Total 49.9

With respect to the mine and infrastructure, capital costs for EPCM were calculated

as 10% of total direct capital costs. Freight was estimated to be 8% of direct capital costs

for relevant equipment items. Other indirect capital includes insurance and capital

spares, which were estimated to be 0.5% and 1.0% of total direct capital costs,

respectively.



Indirect capital costs for the process facilities accounted for EPCM and contractor

overhead and profit, which were estimated at 12.5% and 4.5% of direct capital costs,

respectively. Scott Wilson RPA notes that freight for process plant items has been

included as a direct cost.

Owner’s costs are inclusive of labour, room and board, and materials costs for staff

and hourly personnel working on site during the pre-production period. Based on

manpower levels and annual salaries, total owner’s costs are estimated at $16.0 million.

CONTINGENCY

In Scott Wilson RPA’s opinion, many of the direct capital costs were estimated at a

greater level of detail than anticipated for a Preliminary Assessment. Rather than

applying a global percentage to the capital cost estimate, contingencies were varied by

item depending on the level of detail pertaining to the cost estimate.

Mine equipment and building capital, being primarily based on vendor quotations,

were assigned contingencies ranging from 5% to 10%, averaging 9.5% for the total, and

process plant capital costs were assigned a contingency of 25%. The average

contingency for indirect capital items, amounting to approximately 21%, was based on

the average contingencies for direct items.

The total pre-production contingency amounted to $53.9 million, or 22% of direct and

indirect pre-production capital costs. An additional allowance was included for mill

capital spares, based on 1% of total direct process equipment costs.

ONGOING CAPITAL A summary of the ongoing capital costs is tabulated below (Table 18-9).



TABLE 18-9 ONGOING CAPITAL Strateco Resources Inc. - Matoush Property


Sustaining Capital 15.6 Closure and Reclamation 30.0

Total 45.6

Sustaining capital includes costs for waste development that have been capitalized

over the LOM. Closure costs are inclusive of a man-made lined cell, engineered cap for

the tailings, dismantling of the site infrastructure, long-term care and maintenance, and

contouring and revegetation of the site.

EXCLUSIONS The following items are also excluded from the capital cost estimate:

• Project financing • Land acquisition, leases rights of way and water rights • Escalation during construction • Permits and fees • Environmental impact studies • Any additional civil, concrete work due to the adverse soil condition and

location • Taxes • Import duties and custom fees • Cost of geotechnical investigation • Sunk costs • Pilot Plant and other testwork • Exploration drilling • Costs of fluctuations in currency exchanges • Project application and approval expenses • Future expansion • Relocation of any facilities, if required • All facilities outside Process Plant layout battery limit • Permanent communications • Townsite • Rail service • Construction camp

Scott Wilson RPA notes that the capital cost estimate described above is exclusive of

items that will be required during the exploration phase of the Project. Costs for a portion



of the equipment fleet, diesel generators, and service equipment are considered to be sunk

in relation to the scope and timeline of the Project. Direct and indirect costs for these

items amount to approximately $13 million.

OPERATING COSTS

The average unit operating costs over the LOM is $303 per tonne milled. The

operating costs include mining, power, maintenance, site services, and G&A estimated by

Scott Wilson RPA. Estimates for process operating costs were carried out by Melis, with

the exception of labour costs.

Operating costs for the Project are summarized in Table 18-10:

TABLE 18-10 UNIT OPERATING COST SUMMARY

Strateco Resources Inc. - Matoush Property

Item Total ($/t milled) Mining 82.80 Process 107.77 Power 35.75 Maintenance 24.84 Site Services 28.96 G&A 22.41

Total 302.53

MINING COSTS Mining costs, estimated to average of $82.80 per tonne milled, include labour for

owner and contract crews, explosives, consumables, and equipment operation.

Estimated underground mine operating costs are shown in Table 18-11.



TABLE 18-11 MINING OPERATING COST ESTIMATE Strateco Resources Inc. - Matoush Property

Item Total ($/t milled)

Labour 65.95 Definition Drilling 5.09 Development 6.39 Stoping 2.37 Backfill 3.00

Total 82.80

Labour costs have been developed using projected manpower levels, estimated

salaries, and loading factors representing potential fringe benefits and scheduled

overtime. Manpower, as detailed in a later section, includes staff labour for mine

supervision, engineering, geology, production, and maintenance.

Costs for definition drilling were based on an allowance of $1.2 million per year.

With the exception of alimak raising, it has been assumed that all development and

stoping will be conducted by owner crews. Developed from first principles, development

and stoping costs included explosives, consumables and supplies, such as water piping,

ventilation tubing and electrical cabling, and ground support. Alimak raising, estimated

to cost $3,600 per metre advance, was based on contract quotations.

Backfill costs, estimated to be $3.00 per tonne milled, account for non-labour costs

associated with sourcing and delivering waste rock fill and cemented rock fill to the

underground.

PROCESSING COSTS Processing costs, estimated to average $107.77 per tonne milled, include labour,

reagents, consumables, and maintenance.

Estimated process operating costs are shown in Table 18-12.



TABLE 18-12 PROCESS OPERATING COST ESTIMATE Strateco Resources Inc. - Matoush Property

Item Annual Total ($ ‘000)

Barium Chloride 15 Caustic Soda (NaOH) 15 Crusher Wear Parts 19 Diesel (Yellowcake Calcining) 360 Diesel (Steam Generation) 610 Ferric Sulphate (45%) 154 Flocculant, Anionic Polyacrylamide 7 Flocculant, Non-ionic Polyacrylamide 35 Grinding Mill Liners 500 Hydrogen Peroxide (70%) 271 Laboratory Supplies 250 Lime (98% CaO) 2,393 Magnesia (MgO) 48 Oxygen(1) 2,100 Product Drums 28 RIP Resin 378 Steel Grinding Balls (Grinding) 194 Steel Grinding Balls (Lime Slaking) 6 Sulphuric Acid (2) 9,196 Total Consumables 16,579 Maintenance Consumables 1,150 Contingency (15%) 2,659

Total 20,388

Process operating costs, as estimated by Melis, accounted for mill reagents, mill and

laboratory consumables, power, and maintenance consumables. Reagent and

consumables costs were estimated from unit costs and consumptions rates. Maintenance

consumables are estimated at 2% of the mechanical costs. The estimate includes

contingency, estimated at 15% of total reagent and consumables cost.

Scott Wilson RPA estimated the labour component of the processing costs. Costs

were estimated using annual loaded salaries and projected manpower levels based on the

organization chart provided by Melis.



POWER COSTS Average power costs over the LOM are estimated to be $35.75, and account for the

total power consumption on site.

Estimated power operating costs are shown in Table 18-13.

TABLE 18-13 POWER OPERATING COST ESTIMATE Strateco Resources Inc. - Matoush Property

Item Total ($/t milled) Mine 13.38 Process 19.32 Assay Lab 0.51 Tailings 0.75 Backfill 0.75 Surface 1.04 Total 35.75

Scott Wilson RPA has assumed that the mine, mill and surface will be powered

through the use of diesel generator sets. The potential to power the site using a power

line was considered. Although a reduction in operating cost would be realized with the

use of a power line, the significant capital requirements to install a line far exceed these

benefits.

Power costs for each item are calculated based on equipment unit horsepower (HP),

loading and utilization factors, and a diesel cost of $1.00 per litre.

MAINTENANCE COSTS Maintenance costs, estimated to be $24.84 per tonne milled over the LOM, account

for the mine, surface, and services. Process maintenance costs, as estimated by Melis, are

included with the process operating costs.

A summary of the estimated maintenance operating costs is shown in Table 18-14.



TABLE 18-14 MAINTENANCE OPERATING COST ESTIMATE Strateco Resources Inc. – Matoush Property

Item Total ($/t milled) Mine 21.64

Surface 2.24 Services 0.96

Total 24.84

The maintenance operating costs are estimated based on the equipment inventory,

projected annual operating hours, and hourly costs. The hourly costs include allowances

for parts, fuel, and consumables, such as lube, tires and wear parts.

SITE COSTS Site operating costs average $28.96 over the LOM, as summarized in Table 18-15.

TABLE 18-15 SITE OPERATING COST ESTIMATE Strateco Resources Inc. - Matoush Property


Labour 18.02 Camp Operations 8.78 Tailings Disposal 0.25 Building Installation and Maintenance 0.13 Airstrip and Road Maintenance 0.68 Equipment 0.59 Services 0.51

Total 28.96

Labour is estimated based on projected manpower levels and loaded annual salaries.

Camp operations are based on a contracted rate of $35 per person per day for food and

janitorial services. Equipment costs include allowances for the plant maintenance shop,

light and heavy equipment, compressors, and associated freight, fuel handling, and

distribution. Costs for services include operation and maintenance of the water supplies,

sewage, heating system, telecommunications, and services relevant to the plant.



GENERAL AND ADMINISTRATIVE COSTS G&A costs are estimated to average $22.41 over the LOM, as shown in Table 18-16.

TABLE 18-16 G&A OPERATING COST ESTIMATE Strateco Resources Inc. - Matoush Property


Labour 18.42 Air Travel 1.22 Training 0.31 Off-site Overhead 0.31 Freight 0.41 Other 1.74

Total 22.41

Labour costs are based on projected manpower levels and loaded annual salaries. Air

travel costs include charter and commercial flights. Off-site overhead is to account for

head office administrative costs. Other items included in the G&A operating cost include

insurance, property tax, consulting fees, licensing, and office supplies.

TAXES

Due to the uncertainty of the taxes to be applied to the Project at this early stage, the

scoping study was prepared on a before tax basis.

MANPOWER REQUIREMENTS

Crews for construction and operations are assumed to work two ten-hour shifts per

day for periods of two weeks followed by two weeks off site. Table 18-17 shows a

summary of the total manpower required for the operation. Non-integer quantities for

“On shift” and “On site” manpower arises from the counting of unique personnel who are

on site half of the time, and do not have a cross-shift.



TABLE 18-17 TOTAL MANPOWER REQUIREMENTS Strateco Resources Inc. - Matoush Property

Area On shift On site Total

Administration 16 16 24 Human Resources 13 17 34 Underground Mine 52 83.5 160.5 Process 20.5 28.5 55 Maintenance 17 24 46

Total 118.5 169 319.50

Tables 18-18 through 18-22 tabulate the estimated manpower requirements by area.

TABLE 18-18 ADMINISTRATION MANPOWER REQUIREMENTS Strateco Resources Inc. - Matoush Property


Management 1 1 1 Accounting 3 3 5 Superintendents 3 3 3 Community Relations 1 1 1 Safety 1 1 1 Warehouse and Purchasing 3 3 5 Travel 2 2 4 Environmental Monitors 1 1 2 Receptionist 1 1 2

Total 16 16 24

TABLE 18-19 HUMAN RESOURCES MANPOWER REQUIREMENTS Strateco Resources Inc. - Matoush Property


Administrative Assistant 1 1 2 Radiation 3 5 10 Training 2 2 4 Logistics and Catering 1 1 2 Security 3 5 10 Nurse 1 1 2 Ventilation/Radiation Tech. 2 2 4

Total 13 17 34



TABLE 18-20 MINE MANPOWER REQUIREMENTS Strateco Resources Inc. - Matoush Property


Supervision 5 9 15 Engineering 8.5 8.5 15.5 Geology 7 9 16 Production 31.5 57 114

Total 52 83.5 160.5

TABLE 18-21 PROCESS MANPOWER REQUIREMENTS Strateco Resources Inc. - Matoush Property


Superintendent 1 1 1 Foreman 1 1 1 Metallurgist 1 1 2 Technicians 2 2 4 Assay Lab 2 2 4 Trainer 0.5 0.5 1 Maintenance 3 3 6 Operators 10 18 36

Total 20.5 28.5 55

TABLE 18-22 MAINTENANCE MANPOWER REQUIREMENTS Strateco Resources Inc. - Matoush Property


Foreman 1 1 1 Electrician 5 7 13 Mechanics 3 6 12 Welders 2 4 8 Equipment Operator 1 1 2 Services 5 5 10

Total 17 24 46

Base salaries and wages have been based on comparable projects and operations. A

loading factor of 40% was applied to the base salaries to account for fringe benefits and

scheduled overtime.

ECONOMIC EVALUATION

A pre-tax Cash Flow Projection has been generated from the LOM production

schedule, capital and operating cost estimates, and is summarized in Table 18-23. A

summary of the key criteria is provided below. The cash flow considers the Project from



the time of the construction decision and includes costs for permitting, pre-feasibility, and

feasibility studies.

ECONOMIC CRITERIA PHYSICALS

• Mine life: 7 years

• Pre-production period: Two years

• Operations: 350 days per year

• Productions rate:

o 500 tpd in Year 1 o 650 tpd in Year 2 o 700 tpd in Year 3 to end of mine life

• Total mill feed: 1.6 million tonnes grading 0.437% U3O8

• Metallurgical recovery: 97.6%

• Average annual production: 2.2 million lbs U3O8

REVENUE

• Metal price: US$75.00 per lb U3O8

• Exchange rate: C$1.00 = US$0.85

• Payable metal: 100%

• Transport change: $0.10 per lb U3O8

• Royalty: 2% NSR

• Average NSR value: $813 per tonne milled

COSTS

• Operating costs (per tonne milled):

o Mining: $ 82.80 o Process: $ 107.77 o Power: $ 35.75 o Maintenance: $ 24.84 o Site Services: $ 28.96 o G&A: $ 22.41 o Total: $ 302.53

• Capital costs: Life of Mine

o Pre-production: $ 193 million



o Indirects $ 50 million o Contingency $ 54 million o Sustaining: $ 16 million o Closure: $ 30 million o Total: $ 343 million

• Unit cash costs (per pound of U3O8)

o Operating: $ 27.33 o Capital: $ 18.77 o Total: $ 46.11

-2 -1 1 2 3 4 5 6 7 8 TOTAL

PHYSICALS

Ore Production

AM15 & MT34 '000 t 175.0 236.3 131.3 65.6 65.6 131.3 16.1 821.1 Grade %U3O8 0.633% 0.454% 0.371% 0.736% 0.827% 0.520% 0.469% 0.542%

MT22 '000 t - - 131.3 196.9 196.9 131.3 172.3 828.5 Grade %U3O8 0.000% 0.000% 0.353% 0.492% 0.310% 0.223% 0.248% 0.333%

Total '000 t 175.0 236.3 262.5 262.5 262.5 262.5 188.4 1,649.7 Grade %U3O8 0.633% 0.454% 0.362% 0.553% 0.439% 0.372% 0.267% 0.437%Contained Metal '000 lbs 2,440.7 2,362.9 2,097.0 3,201.2 2,540.4 2,151.9 1,108.7 15,902.8

Production Rate tpd 500 675 750 750 750 750 538 - 689

METALLURGY

Mill Feed '000 t 175.0 236.3 262.5 262.5 262.5 262.5 188.4 1,649.7 Grade %U3O8 0.633% 0.454% 0.362% 0.553% 0.439% 0.372% 0.267% 0.437%Contained Metal '000 lbs 2,440.7 2,362.9 2,097.0 3,201.2 2,540.4 2,151.9 1,108.7 15,902.8

Recovered Metal 97.6% '000 lbs 2,382.1 2,306.2 2,046.7 3,124.4 2,479.4 2,100.2 1,082.1 15,521.2

REVENUE

Metal Price 75.00 US$/lb 75.00$ 75.00$ 75.00$ 75.00$ 75.00$ 75.00$ 75.00$ 75.00$ Exchange Rate 0.85 US$/C$ 0.85 0.85 0.85 0.85 0.85 0.85 0.85 0.85 Gross Revenue '000 C$ 210,186$ 203,488$ 180,590$ 275,684$ 218,775$ 185,315$ 95,477$ 1,369,515$

Transport 0.10 per lb '000 C$ 238$ 231$ 205$ 312$ 248$ 210$ 108$ 1,552$ Net Smelter Return (NSR) '000 C$ 209,948$ 203,257$ 180,386$ 275,372$ 218,527$ 185,105$ 95,369$ 1,367,963$

Royalty 2.0% '000 C$ 4,199$ 4,065$ 3,608$ 5,507$ 4,371$ 3,702$ 1,907$ 27,359$ NSR after Royalty '000 C$ 205,749$ 199,192$ 176,778$ 269,864$ 214,156$ 181,403$ 93,461$ 1,340,604$

OPERATING COSTS

Mining '000 C$ 20,070$ 19,230$ 19,848$ 19,626$ 20,320$ 19,626$ 17,867$ 136,587$ Process '000 C$ 25,397$ 25,397$ 25,397$ 25,397$ 25,397$ 25,397$ 25,397$ 177,781$ Power '000 C$ 8,425$ 8,425$ 8,425$ 8,425$ 8,425$ 8,425$ 8,425$ 58,974$ Maintenance '000 C$ 5,855$ 5,855$ 5,855$ 5,855$ 5,855$ 5,855$ 5,855$ 40,984$ Site Services '000 C$ 6,826$ 6,826$ 6,826$ 6,826$ 6,826$ 6,826$ 6,826$ 47,781$ General and Administrative (G&A) '000 C$ 5,282$ 5,282$ 5,282$ 5,282$ 5,282$ 5,282$ 5,282$ 36,974$ Total '000 C$ 71,855$ 71,015$ 71,633$ 71,411$ 72,105$ 71,411$ 69,652$ 499,082$

Unit CostsMining C$/t milled 114.68$ 81.40$ 75.61$ 74.77$ 77.41$ 74.77$ 94.82$ 82.80$ Process C$/t milled 145.13$ 107.50$ 96.75$ 96.75$ 96.75$ 96.75$ 134.78$ 107.77$ Power C$/t milled 48.14$ 35.66$ 32.09$ 32.09$ 32.09$ 32.09$ 44.71$ 35.75$ Maintenance C$/t milled 33.46$ 24.78$ 22.30$ 22.30$ 22.30$ 22.30$ 31.07$ 24.84$ Site Services C$/t milled 39.00$ 28.89$ 26.00$ 26.00$ 26.00$ 26.00$ 36.22$ 28.96$ G&A C$/t milled 30.18$ 22.36$ 20.12$ 20.12$ 20.12$ 20.12$ 28.03$ 22.41$ Total C$/t milled 410.60$ 300.59$ 272.89$ 272.04$ 274.69$ 272.04$ 369.63$ 302.53$

C$/lb U3O8 30.16$ 30.79$ 35.00$ 22.86$ 29.08$ 34.00$ 64.37$ 32.15$

OPERATING PROFIT '000 C$ 133,894$ 128,177$ 105,145$ 198,453$ 142,051$ 109,992$ 23,809$ 841,522$

CAPITAL COSTS

Direct Capital CostsMine

UG Equipment '000 C$ 6,920$ 6,920$ 13,840$ Surface Equipment '000 C$ 650$ 650$ 1,300$ Development '000 C$ -$ 6,726$ 6,726$ Ventilation '000 C$ 635$ 635$ 1,270$ Services '000 C$ 511$ 511$ 1,023$ Backfill Plant '000 C$ 2,000$ 2,000$ 4,000$

ProcessEquipment '000 C$ -$ 148,886$ 148,886$ Tailings '000 C$ -$ 500$ 500$ Mobile Equipment '000 C$ -$ 500$ 500$

InfrastructureMine Site Buiildings '000 C$ -$ 10,000$ 10,000$ Ore Pad '000 C$ -$ 250$ 250$ Power '000 C$ -$ 4,318$ 4,318$ Administration '000 C$ -$ 830$ 830$

Total Direct Capital '000 C$ 10,716$ 182,726$ -$ -$ -$ -$ -$ -$ -$ -$ 193,443$

Indirect Capital Costs '000 C$ 1,281$ 48,648$ -$ -$ -$ -$ -$ -$ -$ -$ 49,928$

Contingency '000 C$ 2,628$ 50,677$ -$ -$ -$ -$ -$ -$ -$ -$ 53,305$ Capital Spares '000 C$ -$ 575$ -$ -$ -$ -$ -$ -$ -$ -$ 575$ Sustaining Capital '000 C$ -$ -$ 2,556$ 2,985$ 2,590$ 3,182$ 2,620$ 1,631$ -$ -$ 15,564$ Closure '000 C$ -$ -$ -$ -$ -$ -$ -$ -$ -$ 30,000$ 30,000$ Total Capital Costs '000 C$ 14,625$ 282,627$ 2,556$ 2,985$ 2,590$ 3,182$ 2,620$ 1,631$ -$ 30,000$ 342,815$

C$/lb U3O8 22.09$

PRE-TAX CASH FLOWAnnual '000 C$ (14,625)$ (282,627)$ 131,338$ 125,192$ 102,556$ 195,271$ 139,431$ 108,361$ 23,809$ (30,000)$ 498,706$ Cumulative '000 C$ (14,625)$ (297,251)$ (165,913)$ (40,721)$ 61,834$ 257,105$ 396,536$ 504,897$ 528,706$ 498,706$

INTERNAL RATE OF RETURN % 37.1%

NET PRESENT VALUE (NPV) 5.0% C$ million 341.61$ 8.0% C$ million 271.20$ 10.0% C$ million 231.85$ 15.0% C$ million 154.11$

UNIT COST OF PRODUCTION

Operating US$/lb U3O8 25.64$ 26.17$ 29.75$ 19.43$ 24.72$ 28.90$ 54.71$ 27.33$ Capital US$/lb U3O8 18.77$ Total US$/lb U3O8 25.64$ 26.17$ 29.75$ 19.43$ 24.72$ 28.90$ 54.71$ 46.11$

YEAR

TABLE 18-23 PRE-TAX CASH FLOW SUMMARYStrateco Resources Inc. - Matoush Property





CASH FLOW ANALYSIS Considering the Project as a stand-alone operation, the undiscounted cash flow is

$499 million over the seven-year mine life. Operating cash flows average $120 million

per year, and simple payback of pre-production capital is achieved after 2.4 years of

operation.

The pre-tax Internal Rate of Return (IRR) is 37.1%. Pre-tax Net Present Value

(NPV) of the Project at various discount rates is as follows:

• Pre-tax NPV @ 5.0% $ 342 million




In Scott Wilson RPA’s opinion, a discount rate of 10% is appropriate for Base Case

evaluation at this stage of the Project. The economic analysis contained in this report is

based, in part, on Inferred Resources, and is preliminary in nature. Inferred Resources

are considered too geologically speculative to have mining and economic considerations

applied to them to be categorized as Mineral Reserves. There is no certainty that the

production and economic forecasts on which this report is based will be realized.

PROJECT SENSITIVITIES Project risks have been evaluated for economic and non-economic factors. Key

economics risks were examined by performing cash flow sensitivities to head grades,

metal price, operating costs and capital costs. The results are shown in Figure 18-8 and

Table 18-24. The cash flow was evaluated using a variation of ±20% for each variable.

The Project is most sensitive to head grade and metal price, followed by operating

cost and capital cost. The break-even uranium price for the Project is US$53 per lb.



FIGURE 18-8 SENSITIVITY ANALYSIS

$-

$50

$100

$150

$200

$250

$300

$350

$400

$450

0.80 0.90 1.00 1.10 1.20

Sensitivity Factor

Pre-

Tax

NPV

@ 1

0% (C

$ m

illio

ns)

Grade Metal Price Operating Cost Capital Cost

TABLE 18-24 SENSITIVITY ANALYSES Strateco Resources Inc. – Matoush Project

Parameter Variables Units -20% -10% Base +10% +20%

Metal Price US$/lb U3O8 60.00 67.50 75.00 82.50 90.00

Head Grade % U3O8 0.35 0.39 0.44 0.48 0.53

Capital Cost $ millions 274 309 343 377 411

Operating Cost C$/t 242 272 303 333 363

NPV @ 10% Units -20% -10% Base +10% +20%

Metal Price $ millions 74.3 153.1 231.9 310.6 389.4

Head Grade $ millions 74.5 153.2 231.9 310.5 389.2

Capital Cost $ millions 285.4 258.6 231.9 205.1 178.3

Operating Cost $ millions 289.3 260.6 231.9 203.1 174.4

Further to the present cash flow analysis, the effect of an increase in the cut-off grade

to 0.10% U3O8 was reviewed. The net effect would be a reduction of approximately

188,000 tonnes of resources at a grade of 0.069% U3O8. This would result in a loss of



approximately 288,000 lbs of uranium. The loss in revenue ($25 million) is evaluated at

about 1.8% of the total revenue and, when weighted against the reduction in operating

costs (10 months less production), the net result would be positive, in the order of $40

million to $50 million. This would improve the Project NPV by approximately $10

million to $15 million.



19 INTERPRETATION AND CONCLUSIONS Based on the observation made during the site visits and a review of the available

data, Scott Wilson RPA offers the following conclusions:

• The Mineral Resources, as presented, are estimated consistently with the CIM guidelines. At a cut-off grade of 0.05% U3O8, Indicated Mineral Resources are estimated to total 250,000 tonnes grading 0.68% U3O8 containing 3.73 million pounds U3O8. Inferred Mineral Resources are estimated to total 1.34 million tonnes grading 0.44% U3O8 containing 13.07 million pounds U3O8. The Mineral Resources are contained within three zones: AM-15, MT-22 and MT-34.

• There are no Mineral Reserves estimated at Matoush.

• Preliminary metallurgical testwork, initiated at SGS Lakefield in 2007 under

the direction of Melis as part of development work on the Matoush Project, confirm that high uranium extractions (approximately 98%) can, on average, be achieved from the Matoush mineralization under medium free acid concentrations using oxygen as oxidant. The leach test results show that a low pressure oxygen sulphuric acid leach is the leach option of choice for the Matoush mineralization, achieving very high uranium extractions of 98% or better.

• Considering the Project as a stand-alone operation, the undiscounted cash flow

is $499 million over the seven-year mine life. Operating cash flows average $120 million per year, and simple payback of pre-production capital is achieved after 2.4 years of operation.

The economic analysis contained in this report is based, in part, on Inferred Resources, and is preliminary in nature. Inferred Resources are considered too geologically speculative to have mining and economic considerations applied to them to be categorized as Mineral Reserves. There is no certainty that the production and economic forecasts on which this report is based will be realized.

• Based upon capital costs of $343 million, operating costs of $303 per tonne

and a metal price of 75$/lb of uranium, the Project has a pre-tax IRR of 37.1% and a pre-tax NPV at 10% of $231.9 million. The Project would generate some 15.5 million pounds of uranium (U3O8) over a seven-year period.

• While operating costs were evaluated with the most recent labour and material

cost information available, there are items that are in constant flux due to the present economic situation. A periodic review of certain key components would be required. For example, the most important item in the processing



cost is that for H2SO4, which represents about 55% of the consumables cost, and oxygen, which accounts for about 13% of the consumables cost.

• The mining method recommended is the longhole method, which can be

utilized for the bulk of the mineralization. A “Non Man Entry” method would need to be developed for the areas where the grade exceeds acceptable levels for normal mining operations. The typical method in this case would be the raise bore method. Detailed plans would need to be developed in this case.

• The Matoush Project site is located at the top of a watershed, and the

underground mine is entirely located within fairly permeable sandstone, which will likely be the largest contributor to the mine effluent, and average yearly stream flows within a radius of a few kilometres around the site are likely to be less than the mine effluent average discharge rate. Considering the physical limitations of natural streams to take up additional flow without generating uncontrolled erosion and the potential sensitivity of aquatic species, it is unlikely the mine effluent discharge point will be close to the mine site. There will likely be a trade-off between water treatment cost, water segregation cost, and distance of effluent relative to the mine facilities.

• The permitting process for the Matoush Project tailings facility may be

lengthy and construction costs of the tailings facility may be high if an artificial depression in the ground has to be built and/or if the tailings area has to be lined to very stringent criteria (such as a high security cell).

• Discussions with regulatory agencies may be required regarding fish habitat

loss, impact on wetlands, and the projected biodiversity reserve that is approximately 19 km away from the site.

• Considering the size of the Matoush property, with the numerous interpreted

fault zones and nine areas with anomalous radioactive boulders, there is a vast exploration potential to be investigated. Exploration drilling should continue along the MFZ and on any other major fault structures where geophysical interpretations and uranium-mineralized float boulders suggest potential centres of mineralization. There is also potential for unconformity-type uranium deposits on the Matoush property.



20 RECOMMENDATIONS Within the last year, two new zones of mineralization have been outlined on the

Matoush property where the MFZ is the primary host structure with kilometres of

untested potential. Without proven direct geophysical means of pinpointing mineralized

areas along the MFZ, diamond drilling will remain the principal exploration technique.

A recommended Phase 1 program to run from August 1, 2008, through June 30, 2009,

includes extensive diamond drilling to explore the limits and upgrade the resources of the

three zones already located, AM-15, MT-22 and MT-34. As time and resources permit,

drilling will be directed at targets selected along the north and south extensions of the

MFZ and also at targets associated with interpreted, basement-rooted fault zones.

Consideration should also be given to prioritize areas that might be explored for

unconformity-type uranium mineralization. In all cases, the objective will be the location

of other mineralized centres with potential to host substantial tonnages of economic

uranium mineralization.

The recommended Phase 1 budget includes funds to complete the various studies

such as the current Preliminary Assessment, an Environmental Impact Assessment and

Socio-Economic Study by Golder, and continued metallurgical test work by Melis. To

improve the logistics of transporting supplies, fuel and heavy equipment to the property, a

substantial upgrade to the winter road access is recommended.

Given the size of the Matoush Project, a Phase 2 program is proposed which would

include significant diamond drilling of untested existing targets, plus targets developed

over time as a synthesis of existing and newly acquired data pinpoints new target areas

with potential for uranium mineralization. Advancing to Phase 2 work will in part be

contingent on positive results from Phase 1 work. Nonetheless, it is felt that at the end of

Phase 1 there will still be a number of untested, top priority drill targets. Details of the

programs recommended by Strateco are shown in Table 20-1. Scott Wilson RPA has

reviewed the details and concurs with the recommended program and budget.



TABLE 20-1 PROPOSED BUDGET – STRATECO MATOUSH PROJECT Strateco Resources Inc. - Matoush Project

Phase 1 Program - August 1, 2008 to June 30, 2009 $ Head Office Services 1,000,000Project Management/Staff Costs 1,000,000Expense Accounts/Travel Costs 100,000Communications - telephone/fax/radio/hardware/software 120,000Camp Costs Supplies-food, lumber 700,000 Fuel-oil, propane 1,200,000Consultants Mining 200,000 Geophysics 325,000 Technical Support 150,000Diamond Drilling ($200/m) all included except fuel AM-15, MT-22 MT-34 (15,000 m) 3,000,000 North Extension (6,000 m) 1,200,000 South Extension, éclat (9,000 m) 1,800,000Downhole probing 150,000Transportation: Charter aircraft, trucks, skidoos 1,000,000Metallurgical Testing 300,000Environmental Impact Assessment 1,300,000Socio-Economic Study 550,000Market Study 70,000Winter Road Access 1,200,000Mapping/topography 30,000Tenure - Option Payments, Fees, Permits 250,000Shipping-couriers, freight 100,000

Subtotal 15,745,000

Contingencies - 10% 1,574,500

Total 17,719,500

Phase 2 Program – 2009-2010 Head Office and camp maintenance, continued drilling, geophysics, resource estimation, and exploration ramp development 50,000,000

Other recommendations include:

• The ongoing bulk density measurement program be reviewed and improved.

All measurements should be taken on full core. A “standard” with similar physical properties to the Matoush mineralization should be measured daily. The use of paraffin wax to seal core should also be considered. Scott Wilson RPA notes that the application of paraffin may preclude the option of chemical analysis.



• Several enhancements to the QA/QC program including the regular use of standards and coarse reject duplicates. Coarse rejects, blanks and CRMs should be submitted at a rate of 1 in 50 or at least one each per sample batch. SRC should be instructed to prepare a second pulp, also at a rate of 1 in 50, to be submitted to a second laboratory. The alternative lab should use similar digestion and analyses methods as used at SRC. Results from all QA samples should be compiled in a separate database with associated queries to identify QA failures.

• Based on results of the current Preliminary Assessment, an underground

exploration program is warranted, beginning as early as June or July 2009.

• Tailings and waste rock will have to be submitted to a geochemical characterization program according to provincial guidelines (Quebec Directive 019) and additional geochemical testing and modelling of the waste rock and tailings material over the long term will likely be required by the CNSC.

• A field investigation for the potential site(s) of the tailings facility and for the potential borrow sources will have to be carried out. The requirements of the applicable guidelines and the cost of a surface disposal facility should be assessed. Similarly, the potential costs/benefits for using tailings for cemented backfill in the underground mine should be assessed considering the potential benefit with respect to reduced mine seepage and the potential increase in control measures for radiation in the mine. Further field work may be required depending on the location of the tailings pond, namely for fish and fish habitat.

• Hydraulic conductivity of sandstone facies ACF4 (400 m thick), the basal conglomerate (28 m thick), the basement regolith (4 m thick) and the Archean granitic basement has not been assessed. Some of these geological features may provide conditions for permeable zones. Assessment of deep bedrock hydraulic conductivity and deep groundwater chemistry should be assessed in the future in order to evaluate potential mine water inflows and mine seepage water quality.

• Mine water seepage in deep uranium mines within or at the contact with sandstones are known to pose major challenges to mining operations. Detailed mine seepage estimations should be carried out and use of mitigation measures should be assessed in order to reduce the operational risks and reduce the volumes of water that could potentially require treatment.

• Consideration to discharging the mine effluent towards the northwest (Eastmain River watershed) rather than to the southeast (projected biodiversity reserve watershed) should be given during the baseline studies and the applicable baseline programs should be developed and implemented. Also, a model predicting water quality changes in the receiving water body should be



developed once predictions are available for treated effluent quality and quantity.

• Survey of wetlands that would be subject to impact or to partial or complete loss once the final infrastructure plan is defined should be carried out to help evaluate potential impacts or to determine wetlands value and assess the appropriate compensation effort that may be required by the MDDEP (Ministère du Développement durable, de l’Environnement et des Parcs).

• Assessment of fish habitat along the road leading to the borrow pits and along shoreline of water bodies within the borrow pit areas should be completed. Additional metal and radionuclide analysis on fish flesh and bones should be conducted from additional water bodies within and outside the Matoush Project site.

• Additional field surveys for wildlife should be conducted to increase confidence in the relative abundance, distribution and habitat use by wildlife within the study areas, and to better determine if listed wildlife is present or not.

• Where archaeological potential has been identified, an archaeological inventory should be conducted before construction starts to ensure that the Project does not impact on archived heritage.

• Cost for consumables should be reviewed periodically in an effort to improve the mine and processing operating costs, given the present day flux in certain consumables prices.

• Direct marketing of Matoush production is recommended. Assuming an

environment of volatility in the spot market and an upward trend in long-term market prices, it would be prudent to adopt a contracting strategy which would allow at least 50% of production to be sold under medium- to long-term contracts three to four years ahead of the delivery year. The uncommitted production could then be sold either under long-term or spot market contracts closer to the years of delivery, depending on market conditions.



21 REFERENCES Baker, D.J., 1980: The metamorphism and structural history of the Grenville Front near

Chibougamau; unpub. PhD Thesis, University of Georgia; 344 p. Bernier, L. and Moorehead, J., 2000: Controles structureaux, caractéristiques

pétrographiques et minéralogiques de la Kimberlite Otish; MER Québec, MB2000-14.

Bisson, E., Delorme, E., Rougerie, E and Solari, A., 1984: Rapport Final 1983, Monts

Otish – Cogema (Canada) Ltée; filed by MRN, Québec as GM 57709 and GM 57710. Caillat, C., and Raynal, M., 1984: Rapport de Fin de Campagne Eté 1984, Vol. 1 of 10,

Cogema (Canada) Ltée, filed as GM 42517 Chown, E.H., 1979: Structure and metamorphism of the Otish Mountains area of the

Grenvillian Foreland Zone, Quebec; Geol. Soc. America Bull., Vol. 90, Pt II, pp 178-196.

Chown, E.H., 1984: Mineralization controls in the Aphebian Formations, Chibougamau,

Mistassini and Otish Areas; Chibougamau-Stratigraphy and Mineralization; CIM Special Volume 34; pp 229-243.

Chown, E.H. and Caty, J.L., 1973: Stratigraphy, Petrography and Paleocurrent Analysis

of the Aphebian Clastic Formations of the Mistassini-Otish Basin; GAC Special Paper 12; pp 49-71.

Cook, R.B., and Ross, D.A., 2007: Technical Report on the Matoush Uranium Project,

Central Quebec, Canada, prepared by Scott Wilson RPA for Strateco Resources Inc., September 27, 2007.

Cook, R.B., and Ross, D.A., 2008: Technical Report on the Mineral Resource Update for

the Matoush Uranium Project, Central Quebec, Canada, prepared by Scott Wilson RPA for Strateco Resources Inc., September 16, 2008.

Di Prisco, G., 2007: Mineralogical Characterization of Drill Core Samples from the

Matoush Property, Northern Quebec; Report to Strateco by Terra Mineralogical Services, 36 p.

Elson, S. and Langridge, R., 2007: Interpretation and logistics Report, 2006 UTEM 3

Survey, Matoush property, Quebec; Lamontagne Geophysics Ltd; 14 p plus Appendices.

Gatzweiler, R., 1987: Uranium Mineralization in the Proterozoic Otish Basin, Central

Quebec, Canada. In Uranium mineralization, new aspects on geology, mineralogy, geochemistry and exploration methods. Monograph series on mineral deposits 27, Gerbruder Borntraeger, Berlin-Stuttgart, p. 27-48.



Genest, S., 1989: Analyse du Bassin d’Otish, Protérozoïque inférieur, Québec. Université de Montréal, Thèse de doctorat, 277 pages.

Girard, R., 2006: Matoush Project, Uranium Exploration in the Lac De L’Hippocampe

Area, Otish Mountains; 2006 Technical Report prepared for Strateco Resources Inc.; 63 p; filed on SEDAR, Oct.5, 2006.

Hardy, F., Paquet, G. and Hardy, C. 2008: Geological and Geomorphologic Study-

Matoush Property; Report prepared by Poly-Geo Inc; 34 pages. Hoeve, J. 1984: Host rock alteration and its application as an ore guide at the Midwest

Lake uranium deposit, northern Saskatchewan; CIM Bull. Vol.77. no 868, p 63-72. International Energy Agency, 2006: World Energy Outlook 2006. Jefferson, C.W., Thomas, D.J., Gandhi, S.S., Ramaekers, P., Delaney, G., Brisbin, G.,

Cutts, C., Portella, P. and Olson, R.A., 2007: Unconformity-associated uranium deposits of the Athabasca Basin, Saskatchewan and Alberta; in EXTECH IV: Geology and Uranium Exploration Technology of the Proterozoic Athabasca Basin, Saskatchewan and Alberta, (ed) C.W. Jefferson and G. Delaney; Geological Survey of Canada, Bull. 588, p 23-76.

Jenkins, C., 1984: Uranerz Exploration and Mining Limited, Assessment Report, Project

71-90, Lac Matoush; filed by MER as GM 41931. Lafontaine, J., 2007: Summary of the Sedimentary Rocks Observed in the Matoush area

of the Otish Mountain Sedimentary Basin; internal Strateco report. Fielder, B.C., Armstrong, K.C., and Melis, L.A., 2008: Summary of Metallurgical

Testwork, Process Plant Operations, Design, Capital and Operating Cost Estimate, prepared for Strateco Resources Inc. by Melis Engineering Ltd., October 29, 2008.

Pauwels, Andre M., 2005: Evaluation Report, Otish Mountain Property, Xemplar Energy

Corp; 60 p; filed on SEDAR, May 9, 2006. Rhys, David, 2007: Matoush Property – Field Visit and Core Evaluation; Strateco

internal report; 18 pp. SGS, 2003: Geology and Mineral and Petroleum Resources of Saskatchewan;

Saskatchewan Industry and Resources; Saskatchewan Geological Survey; Misc. Report 2003-7, 173 p.

Wilson, R.D., 2008: Evaluation of the Strateco gamma logging program; internal report

to Strateco Resources Inc., 6 pp. World Nuclear Association 2007: The Global Nuclear Fuel Market: Supply and Demand,

2007-2030.



22 SIGNATURE PAGE This report titled “Technical Report on the Preliminary Assessment of the Matoush

Project, Central Quebec, Canada” and dated December 17, 2008, was prepared and

signed by the following authors:

(Signed & Sealed) Dated at Toronto, Ontario December 17, 2008 Normand L. Lecuyer, P.Eng. Principal Mining Engineer (Signed & Sealed) Dated at Toronto, Ontario December 17, 2008 R. Barry Cook, M.Sc., P.Eng. Associate Geologist (Signed & Sealed) Dated at Toronto, Ontario December 17, 2008 David A. Ross, M.Sc., P.Geo. Senior Geologist (Signed & Sealed) Dated at Toronto, Ontario December 17, 2008 Bruce C. Fielder, P.Eng. Principal Process Engineer Melis Engineering Ltd. (Signed & Sealed) Dated at Toronto, Ontario December 17, 2008 Normand D’Anjou, ing. Golder Associés Ltée



23 CERTIFICATE OF QUALIFICATION NORMAND LECUYER

I, Normand Lecuyer, P.Eng., as an author of this report entitled “Technical Report on the Preliminary Assessment of the Matoush Project, Central Quebec, Canada” prepared for Strateco Resources Inc. and dated December 17, 2008, do hereby certify that:

1. I am Principal Mining Engineer with Scott Wilson Roscoe Postle Associates Inc. of

Suite 501, 55 University Ave Toronto, ON, M5J 2H7. 2. I am a graduate of Queen’s University, Kingston, Canada, in 1976 with a B.Sc.

(Hons.) degree in Mining Engineering. 3. I am registered as a Professional Engineer in the provinces of Ontario

(Reg.#26055251) and Québec (Reg.# 34914). I have worked as a mining engineer for a total of 30 years since my graduation. My relevant experience for the purpose of the Technical Report is:

• Review and report as a consultant on numerous exploration and mining projects around the world for due diligence and regulatory requirements.

• Vice-President Operations for a number of mining companies. • Mine Manager at an underground gold mine in Northern Ontario, Canada. • Manager of Mining/Technical Services at a number of base-metal mines in

Canada and North Africa. • Vice-President Engineering at two gold operations in the Abitibi area of

Quebec, Canada. 4. I have read the definition of "qualified person" set out in National Instrument 43-101

(NI43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI43-101.

5. I visited the Matoush Project on June 18, 2008. 6. I am responsible for overall preparation of the Technical Report. 7. I am independent of the Issuer applying the test set out in Section 1.4 of National

Instrument 43-101. 8. I have had no prior involvement with the property that is the subject of the Technical

Report. 9. I have read National Instrument 43-101, and the Technical Report has been prepared

in compliance with National Instrument 43-101 and Form 43-101F1.



10. To the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

Dated this 17th day of December, 2008

(Signed & Sealed)

Normand Lecuyer, P.Eng.



BARRY COOK I, R. Barry Cook, M.Sc., P.Eng., as an author of this report entitled “Technical Report

on the Preliminary Assessment of the Matoush Project, Central Quebec, Canada” prepared for Strateco Resources Inc. and dated December 17, 2008, do hereby certify that:

1. I am an Associate Consulting Geologist with Scott Wilson Roscoe Postle Associates

Inc. of Suite 501, 55 University Ave Toronto, ON, M5J 2H7.

2. I am a graduate of Queen’s University, Kingston, Ontario, Canada, in 1962 with a Bachelor in Science degree in Geological Engineering and in 1964 with a Master of Science degree in Geological Engineering.

3. I am registered as a Professional Engineer in the Province of Ontario (Reg. #

9202011) and as a Professional Engineer/Professional Geologist in the Northwest Territories (Reg. # L797). I have worked as a geologist for a total of 43 years since my graduation. My relevant experience for the purpose of the Technical Report is:

a. 40 years of active experience in mineral exploration including uranium

exploration in the Thelon Basin. b. Familiarity with the Athabasca unconformity-type uranium model and to a

lesser extent with the IOCG and intragranitic type uranium models. c. Attendance at a number of short courses and conferences and on field trips

concerning a variety of uranium deposits. 4. I have read the definition of "qualified person" set out in National Instrument 43-101

(NI43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI43-101.

5. I visited the Matoush Project on June 15-16, 2007.

6. I prepared sections 4 to 15 and collaborated on sections 1, 2, 19, and 20.

7. I am independent of the Issuer applying the test set out in Section 1.4 of National Instrument 43-101.

8. On September 27, 2007, I co-authored a NI43-101 Technical Report on the Matoush property. On October 16, 2008, I co-authored a NI 43-101 Technical Report on the Mineral Resource Update for the Matoush Project.

9. I have read National Instrument 43-101, and the Technical Report has been prepared in compliance with National Instrument 43-101 and Form 43-101F1.





(Signed & Sealed)

R. Barry Cook, M.Sc., P.Eng.



DAVID ROSS I, David A. Ross, P.Geo., as an author of this report entitled “Technical Report on the

Preliminary Assessment of the Matoush Project, Central Quebec, Canada” prepared for Strateco Resources Inc. and dated December 17, 2008, do hereby certify that: 1. I am a Senior Geologist with Scott Wilson Roscoe Postle Associates Inc. of Suite

501, 55 University Ave., Toronto, ON, M5J 2H7. 2. I am a graduate of Carleton University, Ottawa, Canada, in 1993 with a Bachelor of

Science degree in Geology and Queen’s University, Kingston, Ontario, Canada, in 1999 with a Master of Science degree in Mineral Exploration.

3. I am registered as a Professional Geoscientist in the Province of Ontario (Reg.#1192)

and for Incidental Practice registered with the Ordre des Géologues du Québec (#106). I have worked as a geologist for a total of 14 years since my graduation. My relevant experience for the purpose of the Technical Report is:

a. Mineral Resource estimation work and reporting on numerous mining and exploration projects around the world.

b. Exploration geologist on a variety of gold and base metal projects in Canada, Indonesia, Chile, and Mongolia.

4. I have read the definition of "qualified person" set out in National Instrument 43-101

("NI43-101") and certify that by reason of my education, affiliation with a professional association (as defined in NI43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI43-101.

5. I visited the Matoush Project most recently on July 15-16, 2008. 6. I am solely responsible for Section 17 of the Technical Report and collaborated on

sections 1, 2, 19, and 20. 7. I am independent of the Issuer applying the test set out in Section 1.4 of National

Instrument 43-101. 8. On September 27, 2007, I co-authored a NI 43-101 Technical Report on the Matoush

property. On October 16, 2008, I co-authored a NI 43-101 Technical Report on the Mineral Resource Update for the Matoush Project.

9. I have read National Instrument 43-101, and the Technical Report has been prepared

in compliance with National Instrument 43-101 and Form 43-101F1.




Dated this 17th day of December, 2008 (Signed & Sealed) David A. Ross, M.Sc., P.Geo



BRUCE C. FIELDER I, Bruce C. Fielder, P.Eng., as an author of this report entitled “Technical Report on

the Preliminary Assessment of the Matoush Project, Central Quebec, Canada” prepared for Strateco Resources Inc. and dated December 17, 2008, do hereby certify that:

1. I am Principal Process Engineer with Melis Engineering Ltd. of Suite 100, 2366 Avenue C North, Saskatoon, Saskatchewan S7L 5X5.

2. I am a graduate of the University of Alberta, Edmonton, Alberta, Canada, in 1981 with a B.Sc. degree in Metallurgical Engineering.

3. I am registered as a Professional Engineer in the Province of Saskatchewan (Certificate No. 10309). I have worked as a process engineer for a total of 27 years since my graduation. My relevant experience for the purpose of the Technical Report is:

• 27 years direct and indirect experience in the processing of uranium, and • Previous preparation of capital and operating cost estimates

4. I have read the definition of "qualified person" set out in National Instrument 43-101 (NI43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI43-101.

5. I visited the Matoush Project on September 3 and 4, 2008.

6. I am responsible for preparation of Item 16 and contributed to items 1, 19, and 20 of the Technical Report.


8. I have had no prior involvement with the property that is the subject of the Technical Report.






(Signed & Sealed)

Bruce C. Fielder



NORMAND D’ANJOU I, Normand D’Anjou, as an author of parts of sections 18, 19 and 20 of this report

entitled “Technical Report on the Preliminary Assessment of the Matoush Project, Central Quebec, Canada” prepared for Strateco Resources Inc. and dated December 17, 2008, do hereby certify that:

1. I am Engineer with Golder Associates Ltd. of 9200 boul. De l’Acadie, Montreal, Quebec.

2. I am a graduate of École Polytechnique de Montréal, Montréal, Qc and Université Laval, Québec City, Qc in 1986 and 1991 respectively with a Bachelor in Geological Engineering and a Master’s degree in Hydrogeology.

3. I am registered as a Professional Engineer in the Province of Québec (Reg.

#42764). I have worked as a geological engineer for a total of 20 years since my graduation. My relevant experience for the purpose of the Technical Report is: • Mining environment consulting for a total of 15 years • Water balance, water quality, waste rock and tailings management

4. I have read the definition of "qualified person" set out in National Instrument

43-101 (NI43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI43-101.

5. I visited the Matoush Project on June 14 to 15, 2007.

6. I am responsible for preparation of subsections pertaining to the environment

of sections 18, 19, 20 of the Technical Report.


8. I have had no prior involvement with the property that is the subject of the

Technical Report.






(Signed & Sealed)

Normand D’Anjou



24 APPENDIX 1 ANALYTICAL PROCEDURES

SASKATCHEWAN RESEARCH COUNCIL – REGULAR ICP PACKAGE METHOD SUMMARY FOR PROCEDURES ICP4-3 AND ICP4-3R

1. Report Scope This method summary report was requested by Strateco Resources. The following report contains a description of Processing Methods performed on all samples received for Strateco Resources as well as the quality control procedures employed to verify the results produced; including the interpretation of control data results and charts for standards used during the procedures.

2. Laboratory Background

The Geoanalytical Laboratories at the Saskatchewan Research Council are unique facilities offering high quality analytical services to the exploration industry. The facilities consist of modern equipment maintained and engineered towards the needs of our customers. The laboratory has a Quality Assurance program dedicated to actively seeking to evaluate and continually improve the internal quality management system. These measures include:

• Standards Council of Canada Accredited ISO/IEC 17025 Laboratory for Mineral Analysis Testing

• Participating in a CANMET base metal and precious metal proficiency testing program.

• Routine Quality control practices • Computerized sample management • Dedicated, experienced key personnel who are available to answer any

questions raised by our customers • Continual review and improvement of all operations at the facility

3. Quality Control

All Quality Control data generated at SRC Geoanalytical Laboratory is reviewed by the Quality Assurance Specialist. The Quality Control Techniques used for verifying all results generated include:

• Data verification:

At least 2 levels of data verification performed prior to reporting results.

• Instrument Calibration: All instrumentation is maintained and calibrated prior to sample analysis. This is to ensure the stability of the instrument during analysis.

• Analysis of Blanks: A digested blank sample is analysed by the same procedure as normal samples along with the sample group to ensure that there is no contamination from the sample preparation.

• Analysis of duplicates:



One sample preparation and analysis is duplicated with each group of samples to ensure that the repeatability of the results generated. All duplicate analysis must be within specified limits. Duplicate Error Determination

Percentage errors were determined by taking the difference between the two results and dividing it by average. Acceptance Criteria for duplicates All duplicates must be within the appropriate range of each other (within 10%). Any duplicate ranges greater than this are reported to the customer.

• Analysis of reference (QC) samples: Various QC samples are analysed with each group of samples and are used to monitor analytical performance. The results of QA analysis is charted using control charts/compare program on LIMS.

• Control Charts: Quality Control charts are produced for each QC standard for all elements analyzed. Upper and lower limits are set at 3 standard deviations. Appropriate corrective action is initiated and the effectiveness of that action is evaluated internally before reporting the final results to the customer. The corrective action taken that may follow a deviation could involve one or more of the following: 1) Reanalyze 2) Re-digest and reanalyze 3) Reprocess original data

• Interpretation: Control Charts

The mean value is represented by the dashed line; the 3 standard deviation line is represented by the black lines (both upper and lower limits). Analytical results for each standard are represented by the black (♦) points. The concentrations are displayed along the vertical axis and the number of analyses displayed along the horizontal axis

• Monitoring of control Charts: Control Charts are monitored by QA on a regular basis. Any deviations or bias observed for a QC in the method shall be documented and reported to the customer.



4. Sample Receipt and Preparation

All samples are received and entered into the Laboratory Information Management System (LIMS). All analytical data which is generated during the analysis of samples is the property of the customer and shall be controlled by QA measures at the laboratory. Depending on the sample type (non-mineralized, mineralized, etc.) each preparation step was performed in designated sample preparation areas. Rock samples were dried in the original plastic bags @80C overnight, then jaw crushed to 60% -2mm and 100-200g sub sample split out using a riffler. The sub sample was pulverized to 90% - 106 microns using a grinding mill (puck and ring or agate, depending on sample). The grinding mills were, at minimum, cleaned between samples, silica sand cleaning was employed in between groups. The pulp was transferred to a labeled plastic snap top vial.

5. Sample Analysis/Testing Overview The samples are tested using validated documented procedures by trained personnel. All samples are digested prior to analysis by ICP. Two separate digestions were preformed: Partial and Total.

6. Sample Preparation: Digestions

Total: Total Digestions were performed on an aliquot of sample for the analysis of the requested elements by ICP. A 0.25g aliquot of pulp was digested to dryness in a Teflon beaker using a hotplate in a mixture of concentrated HF:HNO3:HClO4. The residue was dissolved in 15 ml of dilute HN03. Partial: Partial Digestions were performed on an aliquot of sample for the analysis of the requested elements by ICP. A 2g aliquot of pulp was digested in a digestion tube in a mixture of HNO3:HCl, in a hot water bath for approximately 1 hour, then diluted to 15ml using deionized water.

7. Geochemical Analysis

QC measures and data verification procedures applied ICP-OES: Mutli element on partial and total digestion: Two separate analyses were done for the partial and total digestions. Instruments were calibrated using certified commercial solutions. The instruments used were PerkinElmer Optima 300DV, Optima 4300DV or Optima 5300DV.



Table 1: Detection Limits: Partial Digestion Total Digestion

Ag 0.1 ppm Al2O3 0.01% Ag 0.2 ppm Li 1 ppm W 1 ppm As 0.2 ppm CaO 0.01% Ba 1 ppm Mo 1 ppm Y 1 ppm Bi 0.2 ppm Fe2O3 0.01% Be 0.2 ppm Nb 1 ppm Yb 0.1 ppm Co 0.1 ppm K2O 0.002% Cd 0.2 ppm Nd 1 ppm Zr 1 ppm Cu 0.1 ppm MgO 0.001% Ce 1 ppm Ni 1 ppm Zn 1 ppm Ge 0.2 ppm MnO 0.001% Co 1 ppm Pb 1 ppm Hg 0.2 ppm Na2O 0.01% Cr 1 ppm Pr 1 ppm Mo 0.1 ppm P2O5 0.002% Cu 1 ppm Sc 1 ppm Ni 0.1 ppm TiO2 0.001% Dy 0.2 ppm Sm 0.5 ppm Pb 0.02 ppm Er 0.2 ppm Sn 1 ppm Sb 0.2 ppm Eu 0.2 ppm Sr 1 ppm Se 0.2 ppm Ga 1 ppm Ta 1 ppm Te 0.2 ppm Gd 0.5 ppm Tb 0.3 ppm U 0.5 ppm Hf 0.5 ppm Th 1 ppm V 0.1 ppm Ho 0.4 ppm U 2 ppm Zn 0.1 ppm La 1 ppm V 1 ppm

8. ICP Analysis

QC measures and data verification procedures applied includes the preparation and analysis of standards and blank. The standards used are BL-1, BL-4a, BL2a, BL-3, BL-5, RS211 and an in-house standard (UHU-1). The selection of standards is based on the radioactivity level of the samples to be analyzed. The ICP is calibrated before use using a certified U reference standard. An additional certified Fe2O3 standard is analysed to check for the interference of iron in the analysis. The calibration is checked after every 20 samples analysed in addition to the analysis of standards. Table 1: Detection Limit (%) for U3O8 by ICP

U3O8 (%) 0.001 %

9. Reporting Management has developed quality assurance procedures exist to ensure that all raw data generated in-house is properly documented, reported and stored to meet the confidentiality requirements of all our customers. All raw data is recorded on internally controlled data forms. Electronically generated data is calculated and stored on computers. All computer generated data is backed up on a daily basis. Access to samples and raw data is restricted to authorized SRC Geoanalytical personnel at all times. All data is verified by key personnel prior to reporting results. Laboratory reports are generated using SRC’s Laboratory Information Management System (LIMS). All QC standards and duplicate analysis is included in the final report which is sent to the customer. Hard copy and electronic copies of all reports are retained for a defined time period.



Samples with very low uranium content were analyzed by fluorimetry as follows, with steps 1 through 4 as outlined in the procedure described above.

1. Sample Analysis/Testing Overview The samples were tested using validated documented procedures by trained personnel. All procedural steps were followed in the analysis–no deviations from the method were requested. All samples were digested prior to analysis by Fluorimetry. The digestion preformed for the samples were Total, see section 6.

2. Sample Preparation: Digestions

Fluorimetric: Uranium was determined on the Total digestion. A 0.1ml aliquot of digestion solution was pipetted into a 90% Pt 10% Rh dish and the liquid evaporated @80C. A NaF/LiK pellet was placed on the dish and fused on a special propane rotary burner for 3 minutes then cooled to room temperature. Two calibration blanks and two calibration standards as well as one blank, two QC/QA standards and one replicate were fused with each group of samples.

3. Geochemical Analysis

QC measures and data verification procedures applied Uranium Fluorescence analysis by Jarrel Ash Fluorimeter: The fluorescence of the fused pellets was then measured on a modified Jarrel Ash Fluorimeter. All data are stored on computers. All calculations are done by computer. Calibration standards are made from 10,000ppm U commercial certified solution.

4. Quality Control: Duplicates Duplicate Error Determination

Percentage errors were determined by taking the difference between the two results and dividing it by average. Acceptance Criteria for duplicates All duplicates must be within the appropriate range of each other (within 10%). Any duplicate ranges greater than this are reported to the customer.

5. Standards

Geochemical Grade Standards:

The Quality Control standard utilized to monitor analytical performance of this method is CANMET standard BL1. Limits for this standard are documented below.

U 220 ppm ± 10 ppm



ANALYTICAL PROCEDURES EMPLOYED BY SGS

Method 9-6-1 Determination of Major Element Oxides and Rare Earth Oxides by Borate Fusion-XRF

1. Parameter(s) measured, unit(s):

SiO2, Al2O3, Fe2O3, MgO, CaO, Na2O, K2O, P2O5, MnO, TiO2, Cr2O3, Ni, Co, La2O3, Ce2O3, Nd2O3, Pr2O3, Sm2O3, BaO, SrO, ZrO2, HfO2, Y2O3, Nb2O5, ThO2, U2O8 , LOI; %

2. Typical sample size:

0.2 to 0.5 g

3. Type of sample applicable (media): Rocks, oxide ores and concentrates

4. Sample preparation technique used:

Samples are crushed and pulverized to -150 mesh. This method is used to report, in percentage, the whole rock suite (SiO2, Al2O3, Fe2O3, MgO, CaO, Na2O, K2O, P2O5, MnO, TiO2, Cr2O3) and Ni, Co as well as the rare earth oxides (La2O3, Ce2O3, Nd2O3, Pr2O3, Sm2O3), and other major element oxides (BaO, SrO, ZrO2, Hf O2, Y2O3, Nb2O5, ThO2, U2O8). Sample preparation entails the formation of a homogenous glass disk by the fusion of 0.2 to 0.5 g of rock pulp with 7g of lithium tetraborate/lithium metaborate (50/50). The LOI at 1000°C is determined separately gravimetrically. The LOI is included in the matrix-correction calculations, which are performed by the XRF instrument software.

5. Method of analysis used:

The disk specimen is analyzed by WDXRF spectrometry. 6. Data reduction by:

The results are exported via computer, on line, data fed to the Laboratory Information Management System with secure audit trail. Corrections for dilution and summation with the LOI are made prior to reporting.



7. Figures of Merit:

element Limit of Quantification (LOQ) % SiO2 0.01 Al2O3 0.01 MgO 0.01 Na2O 0.01 K2O 0.01 CaO 0.01 P2O5 0.01 TiO2 0.01 Cr2O3 0.01 V2O5 0.01 Fe2O3 0.01 MnO 0.01 Ni 0.01 Co 0.01 Ce2O3 0.02 Pr2O3 0.02 Sm2O3 0.03 BaO 0.02 La2O3 0.01 Nd2O3 0.02 ZrO2 0.01 Y2O3 0.02 SrO 0.02 Nb2O5 0.01

This method has been fully validated for the range of samples typically analyzed. Method validation includes the use of certified reference materials, replicates and blanks to calculate accuracy, precision, linearity, range, limit of detection, limit of quantification, specificity and measurement uncertainty.

8. Quality control:

One blank, one duplicate and a matrix-suitable certified or in-house reference material per batch of 20 samples.

9. Data approval steps: Step Approval Criteria 1. Sum of oxides Majors 98-101%;

Majors + NiO + CoO 98-102% 2. Batch reagent blank 2 x LOQ 3. Inserted weighed reference materials Statistical Control Limits 4. Weighed Lab Duplicates Statistical Control Limits by Range



10. Accreditation: This method is accredited by the Standards Council of Canada (SCC) and found to conform to the requirements of the ISO/IEC 17025 standard. See www.scc.ca for SGS Minerals Services Lakefield’s scope of accreditation.

METHOD 9-6-2 Determination of As, Sb, Th, U, Ba, Sn and Ta by XRF using Internal Standard Addition

11. Parameter(s) measured, unit(s):

arsenic, antimony, thorium, uranium, barium, tin, tantalum; (%) 12. Typical sample size:

6 g

13. Type of sample applicable (media): Rocks, soils, ores and concentrates

14. Sample preparation technique used:

This method is used as an alternative to fusion for analytical contexts where the analyte is volatile or if levels of detection below that attainable by fusion are required. The internal standard preparation technique forms a powder pellet specimen consisting of sample thoroughly mixed with a known amount of a reference compound and a binding agent. The compound is chosen to provide reference intensity as close to the analyte line wavelength as possible while avoiding interposing major element lines or absorption edges. Within an emission line series, this is usually the adjacent element in terms of atomic number.

15. Method of analysis used:

Xray fluorescence spectrometry 16. Data reduction by:

The results are exported via computer, on line, data fed to the Laboratory Information Management System with secure audit trail.

17. Figures of Merit:

element Limit of Quantification (LOQ) % As 0.003 Sn 0.002 Sb 0.003 Ta 0.003

ThO2 0.001 U3O8 0.002 Ba 0.003

18. Quality control:

One blank, one duplicate and a matrix-suitable certified or in-house reference material per batch of 20 samples.

19. Accreditation Status:

Standards Council of Canada to ISO/IEC 17025.

http://www.scc.ca/



25 APPENDIX 2 VARIOGRAPHY

FIGURE 25-1 DOWNHOLE VARIOGRAM

FIGURE 25-2 ALONG-STRIKE VARIOGRAM



FIGURE 25-3 DOWN-DIP VARIOGRAM



26 APPENDIX 3 SUMMARY OF METALLURGICAL TESTWORK, PROCESS PLANT OPERATIONS, DESIGN, CAPITAL AND OPERATING COST ESTIMATE, BY MELIS ENGINEERING LTD. DATED OCTOBER 29, 2008

SUITE 100, 2366 AVENUE C NORTH, SASKATOON, SK, CANADA S7L 5X5 PH: 306-652-4084 FAX: 306-653-3779 Email: [email protected] Web Site: www.meliseng.com

M E M O R A N D U M

October 29, 2008 Melis Project No. 490

To: Pierre H. Terreault, P.Eng., MPM, Vice President Operations & Engineering, Strateco Resources Inc.

Cc: Normand L Lecuyer, B.Sc., P.Eng., Principal Mining Engineer, Scott Wilson Mining

From: Bruce C. Fielder, P.Eng. Melis Engineering Ltd.

Re: Metallurgy and Process Report for Matoush Scoping Study

Attached, please find the Melis Engineering Ltd. report Strateco Resources Inc. Matoush Uranium Deposit, Quebec, Canada, Summary of Metallurgical Testwork Process Plant Operations, Design, Capital and Operating Cost Estimate –Preliminary.

This report is preliminary, as the capital cost estimate does not include the results of cost comparisons requested by Strateco Resources Inc. and still in progress. In particular, revision to the process plant crane and building cost may reduce the estimated cost. However, delivery of this report has been requested in the knowledge that it is preliminary.

This report may be used as an appendix to the scoping study document. An executive summary has been included in the report which can be extracted and included in the body of the scoping study report.

Yours truly, MELIS ENGINEERING LTD.

Bruce C. Fielder, P.Eng., K.C. Armstrong, P.Eng., Lawrence Melis, P.Eng., Principal Process Engineer Associate Metallurgist President

STRATECO RESOURCES INC.

MATOUSH URANIUM DEPOSIT, QUEBEC, CANADA

SUMMARY OF METALLURGICAL TESTWORK PROCESS PLANT OPERATIONS,

DESIGN, CAPITAL AND OPERATING COST ESTIMATE PRELIMINARY

MELIS Project No. 490

October 29, 2007

Prepared for

STRATECO RESOURCES INC.

by

Bruce C. Fielder, P.Eng. Principal Process Engineer

K.C. Armstrong, P.Eng., Associate Metallurgist

Lawrence A. Melis, P.Eng. President

MELIS ENGINEERING LTD. 2366 Avenue C North, Suite 100

Saskatoon, Saskatchewan S7L 5X5

Phone: (306) 652-4084 Fax: (306) 653-3779

EXECUTIVE SUMMARY

Melis Engineering Ltd Project No. 490 Strateco Resources Inc. - Matoush Uranium Deposit

October 29, 2008

EXECUTIVE SUMMARY

Metallurgical Testwork

Testwork on Matoush test composites made up from assay rejects was initiated at SGS Lakefield Research Limited (Lakefield) in Lakefield, Ontario in 2007 under the direction of Melis Engineering Ltd. (Melis) as part of development work on Strateco Resources Inc. (Strateco)’s Matoush Uranium Project located in the Otish Mountains of northern Quebec. This preliminary test program encompassed composite analyses, leach scoping tests, uranium recovery and upgrading, settling/filtration tests, effluent treatment and preparation of tailings to provide preliminary environmental data. In addition, quarter core samples were used for comminution (grinding) testwork.

Metallurgical test composites prepared for testing from assay rejects include sub-composites representing four diamond drill holes as well as a blend of all four sub-composites to provide an overall composite for initial testing. Key analyses of these composites are listed in the table below:

Strateco Resources Inc. – Matoush Uranium Project Matoush Test Composites – Summary of Analysis

Analyte Unit Composite 2006-11

Composite 2006-30

Composite 2007-03

Composite 2007-06

Overall Composite

U3O8 % 0.34 0.98 1.61 0.30 0.79 Fe2O3 % 2.01 0.72 1.17 0.64 1.13

As % <0.003 <0.003 <0.003 <0.003 <0.003 Mo % 0.027 <0.004 <0.004 <0.004 0.004 Se % 0.011 <0.005 0.011 <0.005 <0.005

The elemental analyses of the composites show that the Matoush composites prepared for testing were low in deleterious elements such as arsenic, molybdenum and base metals. Selenium was above the detection limit in two of the composites, which implies that the amount of selenium reporting to waste effluents will need to be taken into consideration in effluent treatment testwork.

Mineralogical examination of core samples revealed that the uranium mineralization in the Matoush deposit consists mainly of uraninite (UO2) and carnotite-weeksite K2(UO2)2V2O8·3(H2O) and K2(UO2)2(Si2O5)3·4(H2O). Gangue minerals include calcite, sericite-muscovite, chromite, apatite and tourmaline.

The uranium isotope ratio of the Matoush Overall Composite was measured to

EXECUTIVE SUMMARY


October 29, 2008

confirm that the uranium isotope ratio was close to the naturally occurring ratio of 0.71% U235/U (w/w). Within the limited accuracy of the gamma spectrometry analysis, the measured isotope ratio was approximately 0.52% U235/U (w/w), with a range of 0.37% U235/U (w/w) to 0.69% U235/U (w/w), close to the expected naturally occurring ratio.

Comminution (grinding) tests showed that the Matoush mineralization is of medium hardness. The grindability data were used to develop a preliminary grinding circuit design study using CEET2® technology. The goal of this preliminary study was to develop a suitable SAG and ball mill with pebble crusher (SABC) circuit capable of milling 28.8 t/h (636 t/d at 92% availability) for year 1 to 3, and would allow for expansion in year 4, to increase throughput capacity up to 48.0 t/h (1,061 t/d at 92% availability) to a final P80 of 100 µm.

For year 1-3, the selected SABC circuit design capable of treating 28.8 t/h (636 t/d at 92% availability) to an average P80 of 150 µm comprises:

• One SAG mill of 13’ diameter by 6’ EGL drawing 296 kW at the shell, followed by

• One ball mill of 9’ diameter by 16’ EGL drawing 310 kW at the shell.

• One 3’ diameter pebble crusher is required at all times.

This circuit uses 50 mm grate and 10 mm vibrating screen apertures making a transfer size T80 of 2.4 mm and average circulating load of 16% to the pebble crusher.

Subsequent to this study anticipated mill feed grades and production requirements changed, which reduced actual tonnage requirements to 24.2 t/h for the complete operating period with no further expansion required. The grinding circuit described above, with a capacity of 28.8 t/h, was kept as is for the purposes of this scoping study.

Settling tests completed on leach feed prepared from an overall Matoush composite yielded a thickener unit area of 0.08 to 0.10 m2/t/day.

Leach tests on an Overall Composite prepared from assay rejects confirmed that high uranium extractions (approximately 98%) can, on average, be achieved from the Matoush mineralization under medium free acid concentrations (30 g H2SO4/L) using oxygen as oxidant. The leach test results confirm that a low pressure oxygen sulphuric acid leach is the leach option of choice for the Matoush mineralization, achieving very high uranium extractions of 98 % or better. All four variability sub-composites yielded excellent extractions showing that uniform uranium extractions can be obtained from the Matoush deposit based on the

EXECUTIVE SUMMARY


October 29, 2008

mineralization tested.

Individual leach tests on four separate drill hole sub-composites yielded 98% uranium extractions on three of the four composites and a 95% uranium extraction on the higher grade composite; hence generally matching the uranium extraction achieved on the Overall Composite. The lower uranium extraction obtained on the higher grade composite can likely be increased by increasing the free acid in the leach to a value greater than 30 g H2SO4/L. Acid consumptions were variable ranging from 97.5 to 172.5 kg H2SO4/t and averaging 124 kg H2SO4/t; compared to the approximately 100 kg H2SO4/t consumption observed in the Overall Composite test.

The pregnant leach solution produced in some of the acid leach scoping tests was submitted to an ICP package analysis to give an indication of the elemental make-up of leach solution reporting to downstream uranium recovery. The resulting pregnant leach solution is quite low in deleterious elements, particularly arsenic, molybdenum, selenium and base metals, which will minimize the potential impact on waste treatment requirements. It is also very low in elements which would normally report to the final yellowcake uranium product, in particular molybdenum and vanadium. The relatively high phosphorus content originates from the apatite content in the mineralization.

Ion exchange tests showed that Purolite A660 resin outperformed Dowex 21K resin in terms of loading efficiency with complete uranium adsorption achieved in three stages, yielding a loaded resin containing 60 g U3O8/L. A 90% elution efficiency was achieved using 125 g H2SO4/L sulphuric acid solution at 45 ºC.

Yellowcake precipitation, tailings neutralization, effluent treatment testwork, tailings environmental tests and further resin absorption/elution tests are underway or planned at Lakefield.

EXECUTIVE SUMMARY


October 29, 2008

Process Flowsheet Description

This section has been prepared as a description of the uranium recovery circuits in the Matoush process plant. The operation and control of each unit operation associated with the Matoush process plant are detailed in this process description. These include crushing, grinding, leaching, resin in pulp and resin elution, product precipitation, calcining and packaging, tailings neutralization, effluent treatment, reagent preparation, water supply and utilities.

This process description has been prepared for inclusion in the scoping study. Changes to unit operations and control strategies will occur in conjunction with metallurgical design developments. Consequently, this process description will require revisions from time to time to reflect these changes.

A simplified flowsheet for the Matoush process plant is depicted in the flowsheet depicted in the figure below.

TITLE

AREA

DWG. No.

LOC.

DESIGN AUTH. APPROVAL:

DESIGN LEAD APPROVAL:

CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):

BRU CE C. FIELD ER, P.EN G. 07 /08/08

MATOUSH URANIUM PROJECT

ABCD

DATE DESCRIPTION26/05/0825/08/0815/09/0810/17/08 SCOPING STUDY RELEASE

DRAFT SCOPING STUDY RELEASESCOPING STUDY RELEASESCOPING STUDY SECOND RELEASE

BYBCF

BC F/KCA

BCFBCF

REVI

SION

S

NO.DATE DESCRIPTION BY DES. LEAD APP..REVIS

IONS

NO.DWG. NO. DESCRIPTIONREFE

RENC

E

FROM MINE

PROCESS MILL

SCOPING STUDY SUMMARY FLOWSHEET

490-F-100

MELIS E GI EERI G LTD.

PROCESS WATER

URANIUM

TREATED EFFLUENT DISCHARGESULPHURIC ACID,

OXYGEN

LIMEFERRIC SULPHATE

FLOCCULANT

HYDROGEN PEROXIDEMAGNESIUM HYDROXIDE

LIME LIME

GRINDING

LEACHING

IMPURITY PRECIPITATION

URANIUM PRECIPITATION

CALCINING

SAG MILL

BALL MILL

CYCLONES

LEACH TANKS

GYPSUM BELT FILTER

IMPURITY PRECIPITATION TANKS

URANIUM PRECIPITATION TANKS

URANIUM THICKENER

URANIUM CENTRIFUGE

URANIUM CALCINER

BARIUM CHLORIDEFERRIC SULPHATE

TAILINGS NEUTRALIZATION

TAILINGS MANAGEMENT FACILITY

REVERSE OSMOSIS

TO PROCESS

FRESH WATER

BARIUM CHLORIDEFERRIC SULPHATELIME, FLOCCULANT

BARREN SOLUTION

SAND FILTERS

SURFACE DRAINAGE

EFFLUENT TREATMENT

SULPHURIC ACID

URANIUM BIN

NEUTRAL THICKENER

FLOCCULANT

TAILINGS THICKENER

TAILINGS TANKS

MONITORING PONDS (3)

REJECTPERMEATE

MINE WATER

TO PROCESS

FRESH WATER TANK

PROCESS WATER TANK

MAIN PROCESS FLOWPROCESS FLOWINTERMITTENT / ALTERNATE FLOW

KEY

RESIN IN PULP CIRCUIT

DILUTION WATER

FRESH ELUATE TANK

RIP TANKS

ELUTION COLUMNS

RESIN SCREEN

LEAN ELUATE TANK

EFFLUENT SAND FILTERS

COARSE ORE BIN

c/w GRIZZLYAND APRON

FEEDER

FINE ORE BIN

FINE ORE BIN APRON FEEDERS

(2)

CRUSHER

CRUSHING

PEBBLE CRUSHER

Summarized below are the production basis for the Matoush process plant and a general description of the unit operations in the milling circuit.

EXECUTIVE SUMMARY


October 29, 2008

Production Basis

The process plant is designed to process 193,594 tonnes/a at a feed grade of 0.48% U3O8 and produce 907,185 kg (2,000,000 lbs) U3O8 per year of high quality uranium concentrate. The process plant is designed to operate at 95% availability for 350 days per year, or equivalently, 24 hours per scheduled day for 333 operating days per year.

The slurry circuits (grinding, leaching, resin in pulp and tailings neutralization) circuits are sized for a design feed rate of 24.2 metric tonnes per hour (MTPH) and the downstream solution circuits (resin elution, impurity precipitation, uranium precipitation and calcining and packaging) are sized to produce an average 114 kilograms per hour (kg/h) U3O8.

Overall uranium recoveries are expected to be approximately 97.6%.

Process Description

Crushing is used to reduce the size of run-of-mine ore prior to grinding. The major equipment in crushing consists of a jaw crusher. The crusher has been sized so as to operate for 12 hours/day.

The grinding circuit reduces the size of the process plant feed from the K80 of 85 mm exiting the crusher to a K80 of 150 µm, appropriate for leaching. The major equipment in grinding consists of a SAG mill and a ball mill.

The leaching circuit is used to extract uranium present in the ore (cyclone overflow). Leaching is accomplished through the addition of sulphuric acid and oxygen. The major equipment in leaching consists of the neutral thickener and six leach tanks.

The resin in pulp circuit is composed of the resin in pulp and the resin elution processes.

The resin in pulp preferentially extracts dissolved uranium from the pregnant leach solution by absorbing it into a designed resin. The major equipment in the resin in pulp is the eight tanks of the Kemix carousel, a vendor package.

The resin elution circuit removes the dissolved uranium from the resin, allowing the resin to be recirculated to the resin in pulp circuit. The uranium is stripped from the resin with a sulphuric acid solution. The major equipment in the resin elution circuit is the three resin elution columns.

The impurity precipitation circuit removes excess sulphuric acid and metals co-

EXECUTIVE SUMMARY


October 29, 2008

absorbed on the resin from solution. The precipitated impurities are separated from the uranium bearing solution. The major equipment in the impurity precipitation circuit are six impurity precipitation tanks and the gypsum belt filter.

The uranium precipitation circuit precipitates uranium from solution. The resulting product is thickened to a minimum density of 35% solids (w/w). Uranium is precipitated through the addition of hydrogen peroxide and the pH controlled through the addition of magnesium hydroxide. The major equipment in uranium precipitation are three uranium precipitation tanks and the uranium thickener.

The product calcining circuit cleans the uranium product, dries it and converts it to a high specific gravity product preferred by refineries. The major equipment in the product calcining circuit are the centrifuge, the calciner and three scrubbers. Final product is packaged in standard 205 L drums for shipment to the uranium refinery.

The tailings neutralization circuit treats uranium production process waste flows and precipitate from the water treatment plant. Lime addition increases the pH of the combined streams to pH 10 and the thickened tailings is pumped to the tailings management facility, or to the backfill plant. The major equipment in the tailings neutralization circuit are two tailings neutralization tanks and the tailings thickener.

The effluent treatment circuit removes dissolved impurities from effluent prior to its release or recycle. The major equipment in the effluent treatment circuit are reverse osmosis, the primary and secondary effluent treatment circuits, each consisting of two effluent treatment tanks and a clarifier, three effluent sand filters and three monitoring ponds.

Personnel

The process plant was assumed to operate 24 hours/day, 7 days/week with operators working 12 hour shifts. Two crews were assumed, one on site while the second was off site.

A total of 66 individuals are envisioned to operate (Mill Operations) and maintain (Mill Maintenance) the process plant.

An additional 40 individuals are envisioned to provide support services and administer the site.

Departments not included in the general administration personnel estimate were:

• Mining,

• Mine Maintenance,

EXECUTIVE SUMMARY


October 29, 2008

• Site Services, and

• Housekeeping and catering.

Capital Cost Estimate

The Matoush Scoping Study capital cost estimate is summarized in the table below.

Strateco Resources Inc. Matoush Uranium Project Scoping Study

Summary of Class IV Capital Cost Estimate Cost Area Labour (Hours) 496,200

Labour Cost 49,618,700Material Cost 57,445,370Buildings Cost 39,203,310Reagents, First Fills 2,618,940Total Direct Cost 148,886,320Contractor Overhead and Profit 6,700,000Engineering, Procurement, and Management 18,610,000Total Direct and Indirect Costs 174,196,320Contingency (25%) 43,550,000Capital Spares, 1% of Equipment Cost 1,150,000Total Estimated Capital Costs 218,896,320

Say, 220,000,000Estimated Weight, tonnes 25,910Estimated Operating and Building Power, kW 2,053

EXECUTIVE SUMMARY


October 29, 2008

Process Plant Operating Cost Estimate

The Matoush Scoping Study mill operating cost estimate, excluding operating personnel costs is summarized in the table below.

Strateco Resources Inc. Matoush Uranium Project Scoping Study

Summary of Mill Operating Cost Estimate Operating Cost Class $ (Cdn)/a $ (Cdn)/t $ (Cdn)/lb U3O8

Total Consumables 18,553 95.76 9.28

Electrical Power 3,780 19.51 1.89

Maintenance Consumables 1,150 5.94 0.58

Sub-Total 23,483 121.21 11.74

Contingency (15%) 3,522 18.18 1.76

Total 27,005 139.39 13.50

Simplified Plant Layout

The Matoush Plant is composed of three buildings:

• the process plant building,

• the oxygen plant, and

• the crusher building,

The process plant has been laid out with the following principles:

• that grinding be kept at the far end of the plant area from the laboratory and administration,

• that calcining and packaging be isolated at the far end of the plant area from the laboratory and administration, and

• that the crusher be isolated from the main process building to reduce dust.

The total area covered by the process plant building has been estimated at 7,156 m2, the crusher building area at 100 m2 and the Oxygen Plant area at 126 m2.

Including outside tanks and the tailings thickener, and assuming a 7 m wide paved area surrounding the buildings, tanks and thickener, the total area of the mill apron was estimated to be 13,100 m2.

EXECUTIVE SUMMARY


October 29, 2008

The simplified plant general arrangement is shown in the figure below.

84m.

\

GRINDING CIRCUIT

LEACHING CIRCUIT


CIRCUIT

IMPURITY PRECIPITATION AND GYPSUM FILTRATION CIRCUIT

SULPHURIC ACID STORAGE TANKS

RIP CIRCUIT

CALCINING CIRCUIT

URANIUM DRUM STORAGE

REAGENT PREPARATION AND STORAGE


CIRCUITFRESH AND PROCESS

WATER

REVERSE OSMOSIS

EFFLUENT TREATMENT

COMPUTER ROOM, MCC ROOM MILL DRY, MILL MECHANICAL SHOP

AND LABORATORY

OXYGEN PLANT

CRUSHING CIRCUIT

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):

BRU CE C. FIELD ER, P.EN G. 15 /0 8/09


AB

DATE DESCRIPTION15/09/0817/10/08

SCOPING STUDY RELEASESCOPING STUDY RELEASE

BYBCFBCF

NO.DATE DESCRIPTION BY DES. LEAD APP..NO.DWG. NO. DESCRIPTION


PROCESS PLANT

MATOUSH PLANT SIMPLIFIED LAYOUT

490-L-100

EDGE OF PAVED MILL TERRACE


CIRCUIT

Heat Recovery

The recovery of waste heat from process and other streams offers the potential to reduce operating cost. The streams which have been identified as offering potential for this are:

• The RIP discharge slurry may be used to pre-heat the leach feed slurry,

• Waste heat from power generation or the steam boilers may be used to pre-heat ventilation air, and

• The calciner discharge air may be used to pre-heat ventilation air.

A glycol heat exchanger system would be required to isolate the air streams. A direct contact heat exchanger would be sufficient for the slurry.

TABLE OF CONTENTS 1-1


October 29, 2008

TABLE OF CONTENTS

1.0 FLOWSHEET DESCRIPTION........................................................................1-1

1.1 Production Basis........................................................................................1-1

1.2 Overview....................................................................................................1-1

1.3 Personnel ...................................................................................................1-3

1.4 Consumables .............................................................................................1-4

1.5 Utilities.......................................................................................................1-4

2.0 SUMMARY OF METALLURGICAL TESTWORK......................................2-1

3.0 SUMMARY OF PLANT OPERATIONS.........................................................3-1

3.1 CRUSHING...............................................................................................3-13.1.1 Summary.....................................................................................3-13.1.2 Crushing .....................................................................................3-1

3.2 GRINDING ...............................................................................................3-13.2.1 Summary.....................................................................................3-13.2.2 Grinding......................................................................................3-1

3.3 LEACHING ..............................................................................................3-23.3.1 Summary.....................................................................................3-23.3.2 Leaching......................................................................................3-2

3.4 RESIN IN PULP .......................................................................................3-33.4.1 Summary.....................................................................................3-33.4.2 Resin In Pulp ..............................................................................3-3

3.5 RESIN ELUTION.....................................................................................3-43.5.1 Summary.....................................................................................3-43.5.2 Resin Elution...............................................................................3-4

3.6 IMPURITY PRECIPITATION................................................................3-43.6.1 Summary.....................................................................................3-43.6.2 Impurity Precipitation................................................................3-4

3.7 URANIUM PRECIPITATION ................................................................3-53.7.1 Summary.....................................................................................3-53.7.2 Uranium Precipitation................................................................3-5

3.8 CALCINING AND PACKAGING...........................................................3-6

3.9 Summary ...................................................................................................3-6

3.10 Calcining and Packaging ..........................................................................3-7

3.11 Dust Control ..............................................................................................3-8

3.12 TAILINGS NEUTRALIZATION ............................................................3-9



October 29, 2008

3.12.1 Summary.....................................................................................3-93.12.2 Tailings Neutralization...............................................................3-9

3.13 EFFLUENT TREATMENT .....................................................................3-93.13.1 Summary.....................................................................................3-93.13.2 Effluent Treatment .....................................................................3-9

3.14 REVERSE OSMOSIS.............................................................................3-103.14.1 Summary...................................................................................3-103.14.2 Reverse Osmosis .......................................................................3-10

3.15 Tailings Management Facility ................................................................3-10

3.16 REAGENTS AND UTILITIES ..............................................................3-113.16.1 Summary...................................................................................3-113.16.2 Ferric Sulphate .........................................................................3-113.16.3 Hydrogen Peroxide ...................................................................3-113.16.4 Oxygen Plant.............................................................................3-123.16.5 Caustic Soda .............................................................................3-133.16.6 Magnesia ...................................................................................3-133.16.7 Grinding Balls...........................................................................3-133.16.8 Barium Chloride.......................................................................3-133.16.9 Lime ..........................................................................................3-133.16.10 Flocculants ................................................................................3-143.16.11 Sulphuric Acid ..........................................................................3-143.16.12 Water Distribution ...................................................................3-153.16.13 Process and Instrument Air .....................................................3-153.16.14 Steam.........................................................................................3-153.16.15 Safety Shower System...............................................................3-16

4.0 SIMPLIFIED PLANT LAYOUT......................................................................4-1

5.0 HEAT RECOVERY ..........................................................................................5-1

6.0 SUMMARY OF METALLURGICAL ASSUMPTIONS.................................6-1

6.1 Production Rate ........................................................................................6-1

6.2 Operating Days .........................................................................................6-1

6.3 Head Grade ...............................................................................................6-1

6.4 Recovery ....................................................................................................6-1

6.5 Feed Rate...................................................................................................6-2

6.6 Concentrate Grade....................................................................................6-2

6.7 Concentrate Purity....................................................................................6-3

6.8 Discharge Water Quality ..........................................................................6-3

7.0 SUMMARY OF CONSUMABLES USAGE ASSUMPTIONS........................7-1



October 29, 2008

8.0 METALLURGICAL TESTWORK..................................................................8-1

8.1 Composite Preparation .............................................................................8-1

8.2 Composite Analyses ..................................................................................8-6

8.3 Uranium Speciation Analysis of Overall Composite................................8-8

8.4 Petrographic Analyses ..............................................................................8-9

8.5 Initial Leaching Tests..............................................................................8-11

8.6 Confirmation Leaching Tests .................................................................8-12

8.7 Pregnant Solution Analyses ....................................................................8-14

8.8 Ion Exchange Tests .................................................................................8-15

8.9 Settling Tests ...........................................................................................8-16

8.10 Filtration Tests ........................................................................................8-18

9.0 PRELIMINARY DESIGN CRITERIA ............................................................9-1

10.0 PROCESS PLANT PERSONNEL..................................................................10-1

10.1 Summary .................................................................................................10-1

10.2 Mill Operations Personnel ......................................................................10-1

10.3 Mill Maintenance Personnel ...................................................................10-2

11.0 SUPPORT AND GENERAL ADMINISTRATION PERSONNEL ..............11-1

11.1 Summary .................................................................................................11-1

11.2 Support Departments and Administration Personnel ...........................11-1

12.0 ESTIMATED PROCESS PLANT AND BUILDINGS CAPITAL COSTS - CLASS IV ESTIMATE ...................................................................................12-1

12.1 Summary .................................................................................................12-1

12.2 Basis of Estimate .....................................................................................12-1

13.0 CONSUMABLES, PROCESS AND BUILDING ELECTRICAL AND MAINTENANCE CONSUMABLES..............................................................13-1

13.1 Summary .................................................................................................13-1

13.2 Basis of Estimate .....................................................................................13-1

14.0 REFERENCES ................................................................................................14-1



October 29, 2008

APPENDICES

Appendix A: Matoush Process Flowsheets

Appendix B: Matoush Plant Simplified Layout

Appendix C: Matoush Mass Balance

Appendix D: Estimated Matoush Process Plant and Buildings Capital Cost Details - Class IV Estimate

Appendix E: Consumables, Process and Maintenance Consumables Operating Costs

Appendix F: Matoush Personnel Charts

FLOWHEET DESCRIPTION 1-1


October 29, 2008

1.0 FLOWSHEET DESCRIPTION

This section has been prepared as a description of the uranium recovery circuits in the Matoush process plant. The operation and control of each unit operation associated with the Matoush process plant are detailed in this process description. These include crushing, grinding, leaching, resin in pulp and resin elution, product precipitation, calcining and packaging, tailings neutralization, effluent treatment, reagent preparation, water supply and utilities.

This process description has been prepared for inclusion in the scoping study. Changes to unit operations and control strategies will occur in conjunction with metallurgical design developments. Consequently, this process description will require revisions from time to time to reflect these changes.

A simplified flowsheet for the Matoush process plant is depicted in Process Flowsheet No. 490-F-100. Summarized below are the production basis for the Matoush process plant and a general description of the unit operations in the milling circuit.

1.1 Production Basis

The process plant is designed to process 193,594 tonnes/a at a feed grade of 0.48% U3O8 and produce 907,185 kg (2,000,000 lbs) U3O8 per year of high quality uranium concentrate. The process plant is designed to operate at 95% availability for 350 days per year, or equivalently, 24 hours per scheduled day for 333 operating days per year.

The slurry circuits (grinding, leaching, resin in pulp and tailings neutralization) circuits are sized for a design feed rate of 24.2 metric tonnes per hour (MTPH) and the downstream solution circuits (resin elution, impurity precipitation, uranium precipitation and calcining and packaging) are sized to produce an average 114 kilograms per hour (kg/h) U3O8.

Overall uranium recoveries are expected to be approximately 97.6%.

1.2 Overview

Crushing is used to reduce the size of run-of-mine ore prior to grinding. The major equipment in crushing consists of a jaw crusher. The crusher has been sized so as to operate for 12 hours/day.

The grinding circuit reduces the size of the process plant feed from the K80

of 85 mm exiting the crusher to a K80 of 150 µm, appropriate for leaching. The major equipment in grinding consists of a SAG mill and a ball mill.



October 29, 2008

The leaching circuit is used to extract uranium present in the ore (cyclone overflow). Leaching is accomplished through the addition of sulphuric acid and oxygen. The major equipment in leaching consists of the neutral thickener and six leach tanks.

The resin in pulp circuit is composed of the resin in pulp and the resin elution processes.

The resin in pulp preferentially extracts dissolved uranium from the pregnant leach solution by absorbing it into a designed resin. The major equipment in the resin in pulp is the eight tanks of the Kemix carousel, a vendor package.

The resin elution circuit removes the dissolved uranium from the resin, allowing the resin to be recirculated to the resin in pulp circuit. The uranium is stripped from the resin with a sulphuric acid solution. The major equipment in the resin elution circuit is the three resin elution columns.

The impurity precipitation circuit removes excess sulphuric acid and metals co-absorbed on the resin from solution. The precipitated impurities are separated from the uranium bearing solution. The major equipment in the impurity precipitation circuit are six impurity precipitation tanks and the gypsum belt filter.

The uranium precipitation circuit precipitates uranium from solution. The resulting product is thickened to a minimum density of 35% solids (w/w). Uranium is precipitated through the addition of hydrogen peroxide and thepH controlled through the addition of magnesium hydroxide. The major equipment in uranium precipitation are three uranium precipitation tanks and the uranium thickener.

The product calcining circuit cleans the uranium product, dries it and converts it to a high specific gravity product preferred by refineries. The major equipment in the product calcining circuit are the centrifuge, the calciner and three scrubbers. Final product is packaged in standard 205 L drums for shipment to the uranium refinery.

The tailings neutralization circuit treats uranium production process waste flows and precipitate from the water treatment plant. Lime addition increases the pH of the combined streams to pH 10 and the thickened tailings is pumped to the tailings management facility, or to the backfill plant. The major equipment in the tailings neutralization circuit are two tailings neutralization tanks and the tailings thickener.



October 29, 2008

The effluent treatment circuit removes dissolved impurities from effluent prior to its release or recycle. The major equipment in the effluent treatment circuit are reverse osmosis, the primary and secondary effluent treatment circuits, each consisting of two effluent treatment tanks and a clarifier, three effluent sand filters and three monitoring ponds.

1.3 Personnel

The process plant was assumed to operate 24 hours/day, 7 days/week with operators working 12 hour shifts. Two crews were assumed, one on site while the second was off site.

A total of 66 individuals are envisioned to operate (Mill Operations) and maintain (Mill Maintenance) the process plant.

An additional 40 individuals are envisioned to provide support services and administer the site.

Departments not included in the general administration personnel estimate were:

• Mining,

• Mine Maintenance,

• Site Services, and




October 29, 2008

1.4 Consumables

Table 1, below, summarizes consumables used in the process plant circuits:

Table 1 Strateco Resources Inc. - Matoush Uranium Deposit

Summary Of Consumable Usage Consumables Used in Process Plant Circuit

Barium Chloride Impurity Precipitation Tailings Effluent Treatment

Ferric Sulphate Impurity Precipitation Tailings Effluent Treatment

Flocculant

Neutral Thickener Uranium thickener Tailings Thickener Effluent Treatment Clarifiers

Hydrogen Peroxide Uranium Precipitation Lime Neutralization Magnesium Hydroxide Uranium Precipitation Oxygen Leaching Resin Resin in pulp Steel Balls SAG Mill, Ball and Lime Mill

Sulphuric Acid Leaching Resin elution Effluent Treatment

Yellowcake Drums Packaging

1.5 Utilities

Water supply to the process plant consists of fresh water, process water and mine water. Reverse osmosis permeate is the main source of supply for fresh water, which is used wherever pure water is required. Process and mine water are used wherever pure water is not required.

Steam is used to heat the leach slurry to 50ºC.

Process and instrument air are supplied to the process plant. Two process air compressors supply process air; a bleed stream of process air is dried and de-oiled for use as instrument air.

A safety shower system is installed in the process plant. The system consists of a head tank containing warmed potable water and at least one safety shower/eye wash combination in each process circuit.

SUMMARY OF METALLURGICAL TESTWORK 2-1


October 29, 2008

2.0 SUMMARY OF METALLURGICAL TESTWORK

Testwork on Matoush test composites made up from assay rejects was initiated at SGS Lakefield Research Limited (Lakefield) in Lakefield, Ontario in 2007 under the direction of Melis Engineering Ltd. (Melis) as part of development work on Strateco Resources Inc. (Strateco)’s Matoush Uranium Project located in the Otish Mountains of northern Quebec. This preliminary test program encompassed composite analyses, leach scoping tests, uranium recovery and upgrading, settling/filtration tests, effluent treatment and preparation of tailings to provide preliminary environmental data. In addition, quarter core samples were used for comminution (grinding) testwork.

Metallurgical test composites prepared for testing from assay rejects include sub-composites representing four diamond drill holes (DDH 2006-11, DDH 2006-30, DDH 2007-03 and DDH 2007-06) as well as a blend of all four sub-composites to provide an overall composite for initial testing. Key analyses of these composites are listed in Table 2 below:

Table 2 Strateco Resources Inc. – Matoush Uranium Project

Matoush Test Composites – Summary of Analysis


Composite 2006-30

Composite 2007-03

Composite 2007-06

Overall Composite

U3O8 % 0.34 0.98 1.61 0.30 0.79 Al2O3 % 5.88 5.31 5.34 6.91 5.92 CaO % 0.45 0.62 1.79 0.67 0.85 Fe2O3 % 2.01 0.72 1.17 0.64 1.13 MgO % 2.22 0.52 0.56 0.47 0.86 P2O5 % 0.37 0.43 1.84 0.53 0.72 SiO2 % 82.3 87.0 80.4 86.3 84.5 V2O5 % 0.10 0.07 0.21 0.12 0.11 As % <0.003 <0.003 <0.003 <0.003 <0.003 Ba % 0.069 0.046 0.27 0.056 0.097 Co % 0.0047 0.0013 0.0061 0.0013 0.0029 Cu % 0.0031 0.0027 0.0110 0.0014 0.0039 Mo % 0.027 <0.004 <0.004 <0.004 0.004 Ni % 0.012 0.0026 0.0090 0.0034 0.0061 Pb % 0.097 0.180 0.270 0.0510 0.140 Se % 0.011 <0.005 0.011 <0.005 <0.005 Sr % 0.0074 0.0063 0.021 0.0056 0.0087 Y % 0.0009 0.0008 0.0022 0.0008 0.0011 Zn % 0.0025 0.0030 0.0023 <0.001 0.0018

The elemental analyses of the composites show that the Matoush composites



October 29, 2008

prepared for testing were low in deleterious elements such as arsenic, molybdenum and base metals. Selenium was above the detection limit in two of the composites, which implies that the amount of selenium reporting to waste effluents will need to be taken into consideration in effluent treatment testwork.

A comparison of the uranium assays, taken from a sample cut of the prepared sub-composites, to the drill indicated uranium grade (the grade of the sub-composites calculated on a weighted basis from the individual drill hole intersection analyses) is presented in Table 3 below for comparison.


Matoush Test Composites – Summary of Analysis


Composite 2006-30

Composite 2007-03

Composite 2007-06

Overall Composite

U3O8 - Assay % 0.34 0.98 1.61 0.30 0.79 U3O8 - Calculated % 0.41 1.08 1.77 0.33 0.87 % Difference 17 9 9 9 9

From a comparison of these results the assay head grade of the prepared sub-composites appears to be 9 % lower than the grade calculated from the drill core assays, ignoring Composite 2006-11 which appears anomalous relative to the other assay comparisons. Further metallurgical calculated head grades will become available for comparison as testwork is completed.

Mineralogical examination of core samples revealed that the uranium mineralization in the Matoush deposit consists mainly of uraninite (UO2) and carnotite-weeksite K2(UO2)2V2O8·3(H2O) and K2(UO2)2(Si2O5)3·4(H2O). Gangue minerals include calcite, sericite-muscovite, chromite, apatite and tourmaline.

The uranium isotope ratio of the Matoush Overall Composite was measured to confirm that the uranium isotope ratio was close to the naturally occurring ratio of 0.71% U235/U (w/w). Within the limited accuracy of the gamma spectrometry analysis, the measured isotope ratio was approximately 0.52% U235/U (w/w), with a range of 0.37% U235/U (w/w) to 0.69% U235/U (w/w), close to the expected naturally occurring ratio.

Comminution (grinding) tests showed that the Matoush mineralization is of medium hardness. The grindability data were used to develop a preliminary grinding circuit design study using CEET2® technology. The goal of this preliminary study was to develop a suitable SAG and ball mill with pebble crusher (SABC) circuit capable of milling 28.8 t/h (636 t/d at 92% availability) for year 1 to 3, and would allow for expansion in year 4, to increase throughput capacity up



October 29, 2008

to 48.0 t/h (1,061 t/d at 92% availability) to a final P80 of 100 µm.

For year 1-3, the selected SABC circuit design capable of treating 28.8 t/h (636 t/d at 92% availability) to an average P80 of 150 µm comprises:

• One SAG mill of 13’ diameter by 6’ EGL drawing 296 kW at the shell, followed by

• One ball mill of 9’ diameter by 16’ EGL drawing 310 kW at the shell.

• One 3’ diameter pebble crusher is required at all times.

This circuit uses50 mm grate and 10 mm vibrating screen apertures making a transfer size T80 of 2.4 mm and average circulating load of 16% to the pebble crusher.

Subsequent to this study anticipated mill feed grades and production requirements changed, which reduced actual tonnage requirements to 24.2 t/h for the complete operating period with no further expansion required. The grinding circuit described above, with a capacity of 28.8 t/h, was kept as is for the purposes of this scoping study.

Settling tests completed on leach feed prepared from an overall Matoush composite yielded a thickener unit area of 0.08 to 0.10 m2/t/day.

Leach tests on an Overall Composite prepared from assay rejects confirmed that high uranium extractions (approximately 98%) can, on average, be achieved from the Matoush mineralization under medium free acid concentrations (30 g H2SO4/L) using oxygen as oxidant. The leach test results confirm that a low pressure oxygen sulphuric acid leach is the leach option of choice for the Matoush mineralization, achieving very high uranium extractions of 98 % or better. All four variability sub-composites yielded excellent extractions showing that uniform uranium extractions can be obtained from the Matoush deposit based on the mineralization tested.

Individual leach tests on four separate drill hole sub-composites yielded 98% uranium extractions on three of the four composites and a 95% uranium extraction on the higher grade composite; hence generally matching the uranium extraction achieved on the Overall Composite. The lower uranium extraction obtained on the higher grade composite can likely be increased by increasing the free acid in the leach to a value greater than 30 g H2SO4/L. Acid consumptions were variable ranging from 97.5 to 172.5 kg H2SO4/t and averaging 124 kg H2SO4/t; compared to the approximately 100 kg H2SO4/t consumption observed in the Overall Composite test.



October 29, 2008

The pregnant leach solution produced in some of the acid leach scoping tests was submitted to an ICP package analysis to give an indication of the elemental make-up of leach solution reporting to downstream uranium recovery. The resulting pregnant leach solution is quite low in deleterious elements, particularly arsenic, molybdenum, selenium and base metals, which will minimize the potential impact on waste treatment requirements. It is also very low in elements which would normally report to the final yellowcake uranium product, in particular molybdenum and vanadium. The relatively high phosphorus content originates from the apatite content in the mineralization.

Ion exchange tests showed that Purolite A660 resin outperformed Dowex 21K resin in terms of loading efficiency with complete uranium adsorption achieved in three stages, yielding a loaded resin containing 60 g U3O8/L. A 90% elution efficiency was achieved using 125 g H2SO4/L sulphuric acid solution at 45 ºC.

Yellowcake precipitation, tailings neutralization, effluent treatment testwork, tailings environmental tests and further resin absorption/elution tests are underway or planned at Lakefield.

.

SUMMARY OF PLANT OPERATIONS Page 3-1


October 29, 2008

3.0 SUMMARY OF PLANT OPERATIONS

3.1 CRUSHING

3.1.1 Summary Crushing reduces the size of the run-of-mine ore to a size appropriate for feed to grinding. Crushing is sized to operate 12 hours per scheduled operating day, a feed rate of 48 tonnes/h.

3.1.2 Crushing A process flowsheet showing the Crushing circuit is attached in Appendix A. This flowsheet is:

• 490-F-101: Crushing Flowsheet. Run of mine ore, at approximately 5% moisture is stored in a 5,000 tonne live weight stockpile on the ore pad, which is covered with an A-Frame for protection from the elements. Three apron feeders under the ore pad discharge onto the mill feed conveyor, which feeds the jaw crusher. The jaw crusher discharges directly into the ore feed bin.

The ore feed bin, with a live volume of 280 m3, is designed to provide surge capacity between Crushing and Grinding.

The crushing dust scrubber collects ore dust as a slurry, which is then pumped to the cyclone feed pumpbox in Grinding.

3.2 GRINDING

3.2.1 Summary Grinding reduces the size of the feed to allow leaching of uranium to take place in a reasonable length of time. Grinding is sized for a feed rate of 24.2 tonnes/h, and to operate 24 hours per scheduled operating day.

3.2.2 Grinding A process flowsheet showing the Grinding circuit is attached in Appendix A. This flowsheet is:

• 490-F-102: Grinding Flowsheet. The Grinding circuit is sized for a design feed rate of 24.2 tonnes/h. Ore is metered from the ore feed bin via the ore feed bin apron feeders and discharged onto the grinding feed conveyor. The grinding feed conveyor discharges into the SAG (Semi-Autogenous Grinding) mill.



October 29, 2008

The SAG mill uses a large diameter and a small charge of steel balls to reduce the size of the ore feed. Ground ore with a density of 68% to 72% solids (w/w) is discharged from the SAG mill through grates onto the SAG mill discharge screen. Screen undersize is fed to the ball mill; screen oversize is fed to the pebble crusher.

The ball mill uses a higher ball charge of smaller balls to grind the ore finer than can be accomplished in the SAG Mill. The ball mill slurry, varying in density from 70% to 72% solids (w/w), discharges into the cyclone feedpumpbox and is pumped to the grinding cyclones.

The slurry in the ball mill discharge pumpbox is pumped to one of two cyclones for separating fine particles from coarse particles. The heavier coarse particles reporting to the cyclone underflow are returned to the ball mill for further grinding. The lighter fine particles rise up inside the cyclone vortex finder to be collected as cyclone overflow with a k80 of 150 µm.

The cyclone overflow is passed over a scalping screen to remove particles larger than the screen opening size of 425 micrometers (35 mesh). Along with cyclone underflow these oversize pieces are returned to the ball mill for further grinding. Cyclone overflow, at a density of 33% solids (w/w), is pumped to the neutral thickener.

The design of the Grinding circuit was based upon a preliminary assessment of grinding characteristics from tests on available Matoush composites. Testing of additional composites from other areas of the deposit will be required in future.

3.3 LEACHING

3.3.1 Summary The Leaching circuit is a slurry circuit. Leaching dissolves the uranium, and many other metals and ions, in a sulphuric acid solution.

3.3.2 Leaching A process flowsheet showing the Leaching circuit is attached in Appendix A. This flowsheet is:

• 490-F-103: Neutral Thickening and Leaching Flowsheet. The Leaching circuit is composed of the neutral thickener and six leach tanks. The neutral thickener increases the density of the leach feed slurry, increasing the residence time in leaching. Water separated from the leach feed slurry is recycled to ore sorting and grinding.



October 29, 2008

Leaching is the process of dissolving the uranium in the ore in a sulphuric acid solution. It is accomplished through the addition of sulphuric acid and oxygen to a final concentration of 30 g h2so4/l and an oxidation/reduction potential of 475 mV to 500 mV.

To quicken the leaching process, steam is added to heat the leach slurry to 50 ºC.

3.4 RESIN IN PULP

3.4.1 Summary The Resin In Pulp (RIP) circuit is a slurry circuit. The resin in pulp circuit adsorbs the uranium in beads of plastic resin specifically designed chemically to preferentially absorb uranium.

3.4.2 Resin In Pulp A process flowsheet showing the Resin In Pulp circuit is attached in Appendix A. This flowsheet is:

• 490-F-104: Resin Extraction Flowsheet. The Resin In Pulp circuit arrangement incorporates eight tanks, each incorporating a mixer and pumpcell. All tanks are placed at the same elevation which facilitates the carousel mode of operation.

Each tank contains a discrete batch of resin, each batch spending a specified time in the circuit before being drained and transferred to the resin elution circuit.

Loaded resin is transferred from Tank No. 1 to a resin elution column. After the transfer, eluted resin is transferred to the empty RIP tank, which is then placed on line as the final in the series, and all others advance in rank. The eluted resin is most efficient at absorbing the low concentration of uranium remaining in the leach discharge slurry, which has been contacted with resin in each of the other RIP tanks.

The resin screen separates loaded resin from leach discharge slurry, and the safety screen prevents resin from being pumped to tailings neutralization should a pumpcell screen develop a hole.



October 29, 2008

3.5 RESIN ELUTION

3.5.1 Summary Uranium liquor, consisting of tetravalent uranyl sulphate ([UO2(SO4)3]

4-) in a sulphuric acid solution, is created in the Resin Elution circuit.

3.5.2 Resin Elution A process flowsheet showing the Resin Elution circuit is attached in Appendix A. This flowsheet is:

• 490-F-105: Resin Elution Flowsheet. The Resin Elution circuit consists of three resin elution columns, three eluate tanks, two heat exchangers and a pneumatic conveyor system. Loaded resin is cleaned and transported from the Resin In Pulp circuit into one of three resin elution columns. During the resin elution cycle, two of the three resin elution columns will contain resin.

Lean eluate, a partially loaded eluate solution, is used for the first resin elution cycle on the loaded resin. This cycle partially elutes the resin and fully loads the eluate, now called concentrated eluate and stored in the concentrated eluate tank.

Fresh eluate, a 100 to 150 g H2SO4/L acid solution, is prepared and used to elute the partially eluted resin. The partially loaded eluate, now called lean eluate, is stored in the lean eluate tank and the resin in the second column, now fully eluted, will be transferred back the resin in pulp circuit when a tank becomes available at the start of the next resin elution cycle.

Heat exchangers are used to ensure that the temperature of the resin elution solutions remains between 45ºc and 50ºc to speed the resin elution process.

3.6 IMPURITY PRECIPITATION


4-) in a sulphuric acid solution, is purified in the Impurity Precipitation circuit.

3.6.2 Impurity Precipitation The Impurity Precipitation circuit is made up of six impurity precipitation tanks, the gypsum belt filter, the pregnant solution filters and their respective supporting tanks, pumpboxes and pumps. The process flowsheets showing



October 29, 2008

the impurity precipitation circuit are attached in Appendix A. These flowsheets are:

• 490-F-106: Impurity Precipitation Flowsheet, and

• 490-F-107: Gypsum Filtration Flowsheet. The six impurity precipitation tanks are used to precipitate impurities from the concentrated eluate. Although the Resin In Pulp circuit rejects most of the impurities dissolved during leaching, minor quantities of iron, arsenic, molybdenum and silica may report to the Concentrated Eluate with the uranium. If allowed to pass through to the Uranium Precipitation circuit they would precipitate to some degree with the uranium.

If necessary, ferric sulphate is added to the front end impurity precipitation tanks to precipitate arsenic and molybdenum as ferric arsenate and molybdate. Lime solution is added to the tanks to precipitate the excess sulphate in the Concentrated Eluate as gypsum.

Discharge from the impurity precipitation tanks contains a significant quantity of gypsum (CaSO4

.2H2O). This is separated from the aqueous stream by the Gypsum Belt Filter. The filtrate is clarified in the Pregnant Solution Sand Filters and is pumped to the Uranium Precipitation circuit. The washed filter cake is pumped to the Resin In Pulp circuit feed launder so that any entrained uranium can be recycled.

3.7 URANIUM PRECIPITATION


4-) in a sulphuric acid solution, is precipitated in the Uranium Precipitation circuit.

3.7.2 Uranium Precipitation The Uranium Precipitation circuit is made up of the uranium precipitation heat exchangers, three uranium precipitation tanks, the uranium thickener, the barren strip tank and their respective supporting pumpboxes and pumps. The process flowsheets showing the Uranium Precipitation circuit are attached in Appendix A. These flowsheets are:

• 490-F-108: Uranium Precipitation Flowsheet, and

• 490-F-109: Uranium thickener Flowsheet.



October 29, 2008

The uranium precipitation feed heat exchangers, one in operation and one spare, control the temperature of the feed solution to 25ºC which allows a dense precipitate to form in the uranium precipitation tanks.

The three uranium precipitation tanks are used to precipitate uranium as uranium peroxide (UO4•xH2O). Uranium is precipitated with hydrogen peroxide to ensure molybdenum and silica remain dissolved; magnesium hydroxide slurry is used to maintain the pH as magnesium sulphate, unlike calcium sulphate (gypsum), is soluble. The combination of hydrogen peroxide precipitation and magnesium hydroxide pH control results in a very pure precipitate.

The precipitated uranium slurry is thickened in the uranium thickener which also serves as surge storage between the Uranium Precipitation and Product Calcining circuits.

The barren solution tank stores barren solution and allows any entrained uranium precipitate in the overflow of the uranium thickener time to settle to the coned bottom of the tank. Occasionally, the barren solution tank transfer pump is used to transfer the settled precipitate to the uranium thickener.

3.8 CALCINING AND PACKAGING

3.9 Summary

This section deals with the calcining and packaging of the uranium precipitate in steel drums for shipment to a refinery where calcined yellowcake (U3O8) is converted to uranium trioxide (UO3).

The process of drying uranium at a temperature of approximately 870°C in a multiple hearth furnace is called calcining. The uranium calciner consists of a multiple hearth furnace heated with a series of burners using diesel as fuel. The uranium centrifuge discharge consisting of thickened uranium precipitate at 50% to 67% solids (w/w) is fed to the top hearth. It becomes dry as it travels from the top hearth to the bottom hearth. This calcined yellowcake containing approximately 99% U3O8 (w/w) and 0.2% moisture (w/w) is discharged from the bottom hearth to the lump disintegrator to break any lumps formed in its passage through the calciner.

The calcined lump-free yellowcake is stored in the uranium bin. As required, yellowcake is drawn from the yellowcake bin and packaged in 205 L steel drums. The filled drums are shipped by truck to a uranium refinery.



October 29, 2008

3.10 Calcining and Packaging

The Calcining and Packaging circuit is made up of the Uranium Wash Tank, the Uranium Centrifuge, the Uranium Calciner, the Uranium Lump Disintegrator, the Bucket Conveyor, the Uranium Bin, the three scrubbers and their respective supporting pumpboxes and pumps. The process flowsheets showing the Calcining and Packaging circuit are attached in Appendix A. These flowsheets are:

• 490-F-110: Uranium Centrifuge and Calciner Flowsheet, and

• 490-F-111: Uranium Packaging and Scrubbers Flowsheet. Thickened uranium slurry from the Uranium Thickener is pumped to the Uranium Wash Tank which evens out the feed density and flowrate, allowing more stable operation of the Uranium Centrifuge. The Uranium Centrifuge dewaters the feed to the Uranium Calciner which removes dissolved impurities such as magnesium sulphate, reduces the fuel consumption and helps protect the refractory in the Uranium Calciner.Uranium slurry discharges the Uranium Centrifuge directly into the Uranium Calciner.

The Uranium Calciner will be a vendor package. For the purposes of this study, a direct fired multiple hearth vertical calciner was assumed. The Uranium Calciner both dries and drives off the hydration water from the uranium peroxide (UO4

.xH2O), producing a mixed uranium containing mainly uranium tetraoxide (UO4) and yellowcake (U3O8). This uranium is denser than the uranium peroxide feeding the calciner, allowing uranium drums to be filled to the maximum weight.

The Uranium Lump Disintegrator is used to break up lumps exiting the Uranium Calciner by feeding them into a rotating trommel screen sealed at the end. Lumps larger than the screen holes remain in the screen and break each other up autogenously.

Uranium is transferred from the Uranium Calciner to the Uranium Lump Disintegrator Trommel by gravity and from the Uranium Lump Disintegrator Trommel to the Uranium Bin by a Bucket Conveyor. The Uranium Bin acts as surge capacity between the Uranium Calciner and the Packaging Circuit.

Uranium is packaged in 205 L drums. Empty drums are placed on a motorized roller conveyor which moves them into a partially enclosed room with a plexiglass window underneath the Uranium Bin which is kept under



October 29, 2008

slight negative pressure by the packaging scrubber. The rotary valve at the bottom of the Uranium Bin is operated by an operator standing outside the room and the drum filled to either the maximum weight of 400 kg or until the drum is full. The full drum is moved out of the room by the roller conveyor and an empty drum moved in for filling. Once out of the room, an operator places a lid on the full drum and tightens the sealing ring. The roller conveyor moves the lidded drum through a washing station, after which the drum is moved by a forklift to a scale where the weight is recorded and marked on the drum.

3.11 Dust Control

Uranium calcining and packaging have been designed to hold uranium dust production to a minimum. At times, due to equipment malfunction, fugitive uranium dust will escape into the area.

Three scrubbers are proposed for the Uranium Calcining circuit: the Uranium Calciner Scrubber, the Uranium Calciner Room Scrubber and the Uranium Packaging Scrubber. These units remove uranium dust from the air stream exiting the Uranium Calciner, the room containing the Uranium Calciner, and the Uranium drum packaging unit operations, respectively. The Uranium Calciner and Uranium Calciner Room scrubbers run at all times, the Uranium Packaging Scrubber runs when uranium packaging is taking place.

To conserve fresh water, barren solution will be used in the scrubbers where applicable.



October 29, 2008

3.12 TAILINGS NEUTRALIZATION

3.12.1 Summary Tailings Neutralization refers to those activities connected to the long-term chemical stabilization of tailings.

3.12.2 Tailings Neutralization The process flowsheet showing Tailings Neutralization is attached in Appendix A. This flowsheet is:

• 490-F-112: Tailings Neutralization Flowsheet. The major equipment in tailings neutralization consists of two tailings neutralization tanks and a tailings thickener. The process is to raise the pH of the streams which will be sent to the tailings management facility to pH 10.0, and, with the addition of barium chloride and ferric sulphate, to precipitate radium and arsenic.

Thickened in the tailings thickener to reduce the volume of slurry, the thickener underflow is pumped to the tailings management facility.

Tailings can also be pumped to the paste backfill plant.

3.13 EFFLUENT TREATMENT

3.13.1 Summary Effluent Treatment refers to those activities connected to the cleaning of effluent water prior to re-use or release.

3.13.2 Effluent Treatment The process flowsheet showing Effluent Treatment is attached in Appendix A. This flowsheet is:

• 490-F-113: Effluent Treatment Flowsheet. The Effluent Treatment circuit consists of two stages of treatment operated in series: primary effluent treatment and secondary effluent treatment. Each consists of two effluent treatment tanks and one clarifier. Following the secondary effluent treatment stage are three effluent sand filters, an effluent discharge tank and three monitoring ponds.

Primary and secondary effluent treatment remove metallic ions that are not precipitated by the increase in pH in tailings neutralization, and treat reverse osmosis plant reject. Primarily, ions precipitated consist of radium 226, molybdenum, arsenic and selenium.



October 29, 2008

With the assumptions of the existing water balance, on average, no water will exit the site. The effluent sand filters provide backup to the reverse osmosis plant for times of year when water flowrates may not be average.

The monitoring ponds are the last point of control for water discharged from site. A pond is filled, and a water sample composited from the filling water is assayed. Should the assays indicate that the water meets discharge limits the pond is discharged to the watershed. Should the water not meet discharge limits, the pond is pumped to the tailings management facility.

3.14 REVERSE OSMOSIS

3.14.1 Summary Reverse Osmosis refers to those activities connected to the cleaning of effluent water prior to re-use or release.

3.14.2 Reverse Osmosis The process flowsheet showing Reverse Osmosis is attached in Appendix A. This flowsheet is:

• 490-F-114: Reverse Osmosis Flowsheet

• The Reverse Osmosis Plant will be a package plant and so is not detailed.

The purpose of the Reverse Osmosis (RO) Plant is to produce clean water from tailings thickener overflow and excess tailings management facility supernatant. The only water exiting the site is RO permeate, or, if necessary, secondary clarifier overflow after a final polishing step in the effluent sand filters.

The reject from the Reverse Osmosis Plant, containing contaminants, is sent to the Effluent Treatment circuit for treatment.

3.15 Tailings Management Facility

The process flowsheet showing the Tailings Management Facility is attached in Appendix A. This flowsheet is:

• 490-F-114: Reverse Osmosis Flowsheet The Tailings Management Facility has two important functions: the storage of tailings in a geochemically inert manner, and as surge capacity for water.

The Tailings Management Facility is envisioned to be a sub-aqueous facility, in which the tailings are discharged underwater so that there is a



October 29, 2008

greatly reduced chance of oxidation of the tailings (enhancing geochemical stability) and of wind dispersal of tailings outside the Tailings Management Facility area.

3.16 REAGENTS AND UTILITIES

3.16.1 Summary Reagents are chemicals and materials which are purchased off site, or produced on site from materials purchased off site. Oxygen from the oxygen plant is included as a reagent, the off-site material used in its production being electricity. Steel balls for grinding ore and lime are reagents but do not have process flowsheets. Reagents include ferric sulphate, hydrogen peroxide, oxygen, magnesia, barium chloride, lime, flocculants, sulphuric acid and steel balls.

Utilities are supplies of water, compressed air or steam utilized in the milling process. Utilities are RO Permeate, Process Water, Process Air, Instrument Air and water for the Safety Shower system.

3.16.2 Ferric Sulphate The process flowsheet showing ferric sulphate (Fe2SO4)3)is attached in Appendix A. This flowsheet is:

• 490-F-115: Ferric Sulphate and Hydrogen Peroxide Flowsheet. The ferric sulphate unloading, mixing, and distribution system contains a Concentrated Ferric Sulphate Tank to store the ferric sulphate unloaded directly from a transport truck and Ferric Sulphate Distribution Pumps to pump the ferric sulphate to the Impurity Precipitation, Tailings Neutralization and Effluent Treatment circuits.

3.16.3 Hydrogen Peroxide The process flowsheet showing hydrogen peroxide (H2O2) is attached in Appendix A. This flowsheet is:

• 490-F-115: Ferric Sulphate and Hydrogen Peroxide Flowsheet. The hydrogen peroxide unloading, dilution, and addition system will be, for safety reasons, a vendor designed system. Spontaneous dissolution of hydrogen peroxide is a danger wherever it is stored; the hydrogen peroxide vendor is best qualified to design storage and addition systems and ensure they are properly passivated. Hydrogen peroxide vendors require that unloading and addition systems pass their safety standards prior to



October 29, 2008

delivering hydrogen peroxide, and require one of their personnel be on site for the first delivery.

The hydrogen peroxide system shown is based on an estimate prepared by Degussa Canada Inc. It contains a Hydrogen Peroxide Storage Tank, in which the 70% hydrogen peroxide delivered by truck is diluted to 50% for storage. Hydrogen peroxide is transferred from the Hydrogen Peroxide Storage Tank to the Hydrogen Peroxide Day Tank and from there to the Hydrogen Peroxide Dosing Tank which feeds two metering pumps delivering hydrogen peroxide to the two Uranium Precipitation Tanks.

3.16.4 Oxygen Plant The process flowsheet showing the Oxygen Plant is attached in Appendix A. This flowsheet is:

• 490-F-116: Oxygen Plant and Caustic Soda Flowsheet. Oxygen (O2) is used in leaching.

The oxygen plant will be owned and operated under contract, the oxygen produced in the plant purchased by the operator.

The Oxygen Plant will be a VPSA (Vapour Pressure Swing Adsorption) package plant and so is not detailed. Typical VPSA systems are either single or double bed, non-cryogenic systems. They include a rotary-lobe feed air blower, vacuum blower (for double bed systems), one or two adsorbent vessels, an oxygen surge tank, switching valves and computer controls.

In the single-bed system, the blower draws in air, compresses it and sends it to the adsorbent vessel to remove impurities, leaving 90 to 94 percent pure oxygen as the product. The adsorbent is then regenerated as the blower removes gas by reducing the pressure inside the vessel. The waste gas (nitrogen, water and carbon dioxide) is then discharged into the air. Since oxygen is not produced during regeneration, the system includes a low-pressure surge tank to allow for continuous oxygen supply. The double bed system uses a similar adsorption process cycle that relies on swings in pressure -- from above one atmosphere to below atmospheric pressure (vacuum) -- to cycle each bed sequentially from adsorption to desorption. One bed is always adsorbing impurities to separate oxygen from air, while the other bed regenerates. Thus, the beds alternately produce oxygen into a surge tank which ensures that product is available continuously at a consistent pressure and purity.



October 29, 2008

3.16.5 Caustic Soda The process flowsheet showing caustic soda (NaOH) is attached in Appendix A. This flowsheet is:

• 490-F-116: Oxygen Plant and Caustic Soda Flowsheet. Caustic soda is used to regenerate resin poisoned with silica.

3.16.6 Magnesia The process flowsheet showing magnesia is attached in Appendix A. This flowsheet is:

• 490-F-117: Magnesia and Grinding Balls Flowsheet. Magnesium oxide, MgO, is converted to magnesium hydroxide, Mg(OH)2 in the Magnesia mix tank. The term Magnesia is commonly used to refer to both magnesium oxide and magnesium hydroxide.

The magnesia system consists of a Magnesia Silo, feeding by screw to a Magnesia Mix Tank, which is pumped to a Magnesia Distribution Tank. Because the magnesia solution is used in the Uranium Precipitation circuit, permeate from the Reverse Osmosis Plant was specified for use as dilution water.

3.16.7 Grinding Balls The process flowsheet showing grinding balls is attached in Appendix A. This flowsheet is:

• 490-F-117: Magnesia and Grinding Balls Flowsheet. Grinding balls are used in the SAG mill, the ball mill and the lime mill.

3.16.8 Barium Chloride The process flowsheet showing barium chloride (BaCl2

.2H2O) is attached in Appendix A. This flowsheet is:

• 490-F-118: Barium Chloride and Lime Flowsheet. The barium chloride system consists of a Barium Chloride Mix Tank, a Barium Chloride Distribution Tank, one Barium Chloride Transfer Pump and two barium Chloride Distribution Pumps. Because the barium chloride solution is used in Effluent Treatment, permeate from the Reverse Osmosis Plant was specified for use as dilution water.

3.16.9 Lime The process flowsheet showing lime (CaO) is attached in Appendix A. This flowsheet is:



October 29, 2008

• 490-F-118: Barium Chloride and Lime Flowsheet. The lime slaking circuit is designed to supply 4 tonnes/h of equivalent CaO. It consists of a Lime Silo, Lime Mix Screw, a Lime Tower Mill, Lime Cyclone, Lime Storage Tank and associated pumpboxes and pumps.

The lime slaking circuit is run intermittently to maintain a sufficient supply of slaked lime in the Lime Storage Tank. The slaked lime is supplied to Impurity Precipitation, the Tailings Neutralization circuit and to Effluent Treatment.

Pebble lime is fed from a Lime Silo with a Lime Mix Screw to the Lime Tower Mill. The Lime Tower Mill is fed with pebble lime controlled by the speed of the screw conveyor.

Grinding (slaking) of lime is carried out in the Lime Mill. Ground lime, with a density 15% to 20% solids (w/w), is discharged from the Lime Mill and is stored in the Lime Storage Tank, equipped with an agitator

3.16.10 Flocculants The process flowsheet showing flocculants is attached in Appendix A. This flowsheet is:

• 490-F-127: Flocculants and Sulphuric Acid Flowsheet. Flocculants are used in the Neutral Thickener, Tailings Thickener, Uranium thickener and Effluent Clarifiers to speed the settling of solids and increase the underflow density. Up to three different types of flocculant may be used, depending upon the settling characteristics of the solids in each thickener or clarifier. Anionic polyacrylamide flocculant is used in the effluent treatment clarifiers, non-ionic polyacrylamide flocculant or flocculants are used in the Neutral Thickener, Tailings Thickener and Uranium thickener.

Each flocculant is prepared using a separate, skid-mounted mix unit.

3.16.11 Sulphuric Acid The process flowsheet showing sulphuric acid (H2SO4) is attached in Appendix A. This flowsheet is:

• 490-F-119: Flocculants and Sulphuric Acid Flowsheet. Sulphuric acid is used in Leaching, Resin Elution and Effluent Treatment

Sulphuric acid will be purchased as 94% H2SO4 w/w solution and stored in tanks on site. A maximum 14 day supply will be stored on site.



October 29, 2008

3.16.12 Water Distribution The process flowsheet showing water distribution is attached in Appendix A. This flowsheet is:

• 490-F-120: Fresh and Process Water Distribution Flowsheet. Two types of water are used in the process plant: Fresh and Process Water. Process Water is water containing various concentrations of impurities and is used where the quality of the required water is not important to the process. Fresh water is very pure water and is used where clean water is of great importance: in the impurity and uranium precipitation circuits, and for dilution of hydrogen peroxide.

Fresh Water is made up of permeate from the Reverse Osmosis plant, with make-up water being fresh water. Process Water is made up of Secondary Clarifier overflow and Mine Water, with make-up water being Fresh Water.

The fire water loop is supplied from the Fresh Water Tank, which is therefore larger than would be required were process flows alone taken from this tank.

3.16.13 Process and Instrument Air The process flowsheet showing process and instrument air is attached in Appendix A. This flowsheet is:

• 490-F-121: Air and Steam Distribution Flowsheet. Process air is the air used in the process or for maintenance. Process air, at 345 kPa pressure, is used as delivered from the Process Air Compressors. Process air contains entrained water and oil from the Process Air Compressors and is used where these are of no consequence.

Instrument air is process air that has been dried and de-oiled for use in powering process control instruments. As a rule, the presence of oil or water in instrument air shortens the life of such instruments.

3.16.14 Steam The process flowsheet showing steam is attached in Appendix A. This flowsheet is:

• 490-F-121: Air and Steam Distribution Flowsheet. Steam is used in leaching, and for heating various and miscellaneous process equipment.



October 29, 2008

3.16.15 Safety Shower System The process flowsheet showing the safety shower system is attached in Appendix A. This flowsheet is:

• 490-F-122: Safety Shower System Flowsheet. The safety shower system delivers clean, warmed water to the safety and eye showers in the process plant. The size and flowrate of components of this system are specified by code.

SIMPLIFIED PLANT LAYOUT Page 4-1


October 29, 2008

4.0 SIMPLIFIED PLANT LAYOUT

The simplified plant general arrangement is attached in Appendix B. This general arrangement is:

• 490-L-100: Matoush Plant Simplified Layout

The Matoush Plant is composed of three buildings:

• the process plant building,

• the oxygen plant, and


The process plant has been laid out with the following principles:

• that grinding be kept at the far end of the plant area from the laboratory and administration,

• that calcining and packaging be isolated at the far end of the plant area from the laboratory and administration, and

• that the crusher be isolated from the main process building to reduce dust.

The total area covered by the process plant building has been estimated at 7,156 m2, the crusher building area at 100 m2 and the Oxygen Plant area at 126 m2.

Including outside tanks and the tailings thickener, and assuming a 7 m wide paved area surrounding the buildings, tanks and thickener, the total area of the mill apron was estimated to be 13,100 m2.

HEAT RECOVERY Page 5-1


October 29, 2008

5.0 HEAT RECOVERY

The recovery of waste heat from process and other streams offers the potential to reduce operating cost. The streams which have been identified as offering potential for this are:

• The RIP discharge slurry may be used to pre-heat the leach feed slurry,

• Waste heat from power generation or the steam boilers may be used to pre-heat ventilation air, and

• The calciner discharge air may be used to pre-heat ventilation air.

A glycol heat exchanger system would be required to isolate the air streams. A direct contact heat exchanger would be sufficient for the slurry.

SUMMARY OF METALLURGICAL ASSUMPTIONS Page 6-1


October 29, 2008

6.0 SUMMARY OF METALLURGICAL ASSUMPTIONS

6.1 Production Rate

As defined by Strateco Resources Inc., the production rate assumed for the Matoush uranium project was 2,000,000 lbs U3O8/a.

6.2 Operating Days

The Matoush process plant was assumed to have a scheduled 350 days of operation/a, with an operating time of 95% of those days. The net operating days per annum is thus 333. This is detailed in Table 4, below.


Scheduled Operating Time Extraction or Loss Unit Value Comment

Scheduled Operating Days Days/a 350 Melis

Operating Time % 95 Melis Net Operating Days Days/a 333 Calculated

6.3 Head Grade

As defined by Strateco Resources Inc., the grade for the first three years of operation was 0.50% U3O8, following which the grade will decrease to 0.48% U3O8. The mass balance was calculated assuming the 0.48% U3O8 feed grade.

6.4 Recovery

The average net recovery, as defined by testwork and calculated in the mass balance, was 97.6%. The average net recovery is detailed in Table 5 below.


Uranium Recovery Extraction or Loss Unit Value Comment

Uranium Extraction % 98.0 Testwork

RIP Loss (Residue Loss) % 0.3 Calculated from Kemix Typical Data



October 29, 2008


Uranium Recovery Extraction or Loss Unit Value Comment

Loss in Gypsum Solids % 0.1 Calculated

Average Net Recovery % 97.6 Calculated

6.5 Feed Rate

The production rate, operating days, feed grade and recovery together define a feed rate of 193,594 tonnes/a, 581 tonnes/day or 24.2 tonnes/h.

The relationship between mill feed grade and mill feed tonnage, for an annual production rate of 2,000,000 lbs U3O8, is shown below in Figure 1.

Figure 1 Feed Grade vs. Tonnes/d for 2,000,000 lbs U3O8/a

6.6 Concentrate Grade

Based on the performance of the hydrogen peroxide precipitation process at other operating uranium mills, the concentrate grade is expected to be ≥99% U3O8.

y = 278.8x-1.0045

0

100

200

300

400

500

600

700

800

900

1,000

0.00 0.20 0.40 0.60 0.80 1.00 1.20

Mill Feed Grade, % U3O8

Mill

Fee

d, T

onn

es/d



October 29, 2008

6.7 Concentrate Purity

Based on the performance of the hydrogen peroxide precipitation process at other operating uranium mills, the concentrate purity is expected to meet the “Limit Without Penalty” guidelines listed in Table 6 below.


Uranium Ore Concentrate Impurities And Maximum Concentration Limits (From ASTM C967-02A)

Analyte Unit Limit Without Penalty

Limit Without Reject

As 100 x %/%U 0.05 0.10

B 100 x %/%U 0.005 0.10

Ca 100 x %/%U 0.05 1.0

CO3 100 x %/%U 0.20 0.50

F 100 x %/%U 0.01 0.10

Halogens (exclusive of F) 100 x %/%U 0.05 0.10 Fe 100 x %/%U 0.15 1.00

Mg 100 x %/%U 0.02 0.50

H2O 100 x %/%U 2.0 5.0

Mo 100 x %/%U 0.10 0.30

P 100 x %/%U 0.10 0.70 K 100 x %/%U 0.20 3.0

Si (as SiO3) 100 x %/%U 0.50 2.5

Na 100 x %/%U 0.50 7.5 S 100 x %/%U 1.0 4.0

Th 100 x %/%U 1.0 2.5

Ti 100 x %/%U 0.01 0.05 V 100 x %/%U 0.06 0.30

Zr 100 x %/%U 0.01 0.10 234U µg/g U 56 62

6.8 Discharge Water Quality

All water discharged from the Matoush site will be treated to meet the Maximum Monthly Arithmetic Mean Concentration limits typical for specified analytes as listed in Table 7 below. [(Government of Saskatchewan) Mineral Industry Environmental Protection regulations and the (Government of Canada) Metal Mining Effluent Regulations].



October 29, 2008


Monthly Arithmetic Mean Concentration Discharge Limits

Analyte Units Maximum Monthly Arithmetic Mean

Concentration Discharge Limits

pH units 6.0 – 9.5 TSS mg/L 15 As mg/L 0.5 Cu mg/L 0.3

Mo mg/L 0.5 Ni mg/L 0.5 Pb mg/L 0.2 Se mg/L (1)

U mg/L 2.5

Zn mg/L 0.5 Pb210 Bq/L 0.92 Ra226 Bq/L 0.37 Th230 Bq/L 1.85

Note: 1. No limit currently specified.

SUMMARY OF REAGENT USAGE ASSUMPTIONS Page 7-1


October 29, 2008

7.0 SUMMARY OF CONSUMABLES USAGE ASSUMPTIONS

The reagent usage assumptions used to calculate reagent and consumables consumptions in the mass balance, pending further testwork, are listed in Table 8, below. The mass balance is attached in Appendix C.


Summary of Reagent and Consumables Usage Assumptions Criteria Unit Value Notes

Barium Chloride

Tailings Neutralization g BaCl2.2H2O/m3 35 Melis

Secondary Effluent Treatment g BaCl2.2H2O/m3 35 Melis

Ferric Sulphate

Impurity Precipitation g Fe/m3 25 Melis

Tailings Neutralization g Fe/m3 40 Melis

Primary Effluent Treatment g Fe/m3 50 Melis

Secondary Effluent Treatment g Fe/m3 50 Melis

Flocculant, Non-ionic Polyacrylamide Neutral Thickener g/t ore 60 Melis Uranium Thickener g/t 150 Melis

Tailings Neutralization g/t 60 Typical

Flocculant, Anionic Polyacrylamide Primary Effluent Treatment mg/L 7 Melis Secondary Effluent Treatment mg/L 7 Melis

Hydrogen Peroxide

Dosage (kg 70% Soln/t U3O8) kg/t U3O8 280 Melis

Lime Impurity Precipitation Final Circuit pH Units 3.2 Melis

Tailings Neutralization kg CaO/t ore 40 Typical

Primary Effluent Treatment g/m3 0 Melis

Secondary Effluent Treatment g/m3 11 Melis

Magnesia

Magnesia Dosage kg MgO/t U3O8 150 Melis

Oxygen O2 Requirement tonnes O2/d 30 Estimated

Resin

SUMMARY OF REAGENT USAGE ASSUMPTIONS Page 7-2


October 29, 2008


Summary of Reagent and Consumables Usage Assumptions Criteria Unit Value Notes

Resin Loading g U3O8/L 55 Melis/Testwork

Proportion of Resin in Loading Tank % (v/v) 20.0 Kemix Typical

Resin Loss m3/ktonne ore 0.30 Purolite

Steel Grinding Balls SAG Mill Grinding Ball Consumption kg/t 0.75 Typical Ball Mill Grinding Ball Consumption kg/t 0.75 Typical

Sulphuric Acid Leaching kg 100% H2SO4/t ore 104 Testwork

Acid Elution Volume BV 3.5 Melis file data

Elution H2SO4 Concentration g/L 100 Melis file data

Secondary Effluent Treatment g H2SO4/L 0.07 Melis

METALLURGICAL TESTWORK Page 8-1


October 29, 2008

8.0 METALLURGICAL TESTWORK

Testwork on Matoush samples was initiated at Lakefield in Lakefield, Ontario under the direction of Melis as part of development work on Strateco’s Matoush Uranium Project. This preliminary test program encompassed composite analyses, leach scoping tests, uranium recovery and upgrading, settling/filtration tests, effluent treatment and preparation of tailings to provide preliminary environmental data. In addition, quarter core samples were used for comminution (grinding) testwork.

8.1 Composite Preparation

The selection of samples from the Matoush deposit was reviewed with Jonathan Lafontaine of Strateco during and subsequent to a site visit by Lawrence Melis of Melis on June 7 and 8, 2007. Four drill holes were selected by Strateco as representing the Matoush mineralization identified as of June 2007. Test composites were to be made up on a weighted basis according to drill hole intersections using coarse assay rejects stored at the Saskatchewan Research Council (SRC) in Saskatoon, Saskatchewan. The samples for the metallurgical composites were selected with a range of uranium grades which were thought typical of the mineralization, providing variability of high and low grades.

This provided the following composites for testing:

• Composite 2006-11 – representing a blend of the intersections 300.5 m - 317.0 m, 327.9 m - 328.4 m, 328.8 m - 329.5 m, 330.2 m- 330.5 m, and 332.5 m- 332.9 m.

• Composite 2006-30 - representing a blend of the intersections 250.0 m – 273.0 m and 278.0 m- 280.0 m.

• Composite 2007-03 - representing a blend of the intersections 291.0 m – 308.5 m.

• Composite 2007-06 - representing a blend of the intersections 315.0 m – 338.0 m.

• Overall Composite – a blend of the four drill hole sub-composites to be used in the initial testing.

Some of the assay rejects samples received at Lakefield were wet and had to be dried prior to compositing. This compromised the integrity of the samples being used in this preliminary test program. Hence, further confirmation leach tests on samples prepared from drill core for comminution testing (see below) were



October 29, 2008

provided.

The samples and drill hole intersections comprising Composite 2006-11, Composite 2006-30, Composite 2007-03 and Composite 2007-06 are shown in Table 9 to Table 12, below.


Make-up of Composite 2006-11Sample Number

From (m)

To (m)

Length (m)

Total Weight (kg)

Weight to Composite (kg)

U3O8

(%)

295206 300.5 301.4 0.9 1.76 0.734 0.014

295207 301.4 302.0 0.6 1.30 0.489 0.384

295208 302.0 302.6 0.6 1.38 0.489 0.822

295209 302.6 303.0 0.4 1.06 0.326 0.002

295210 303.0 304.0 1.0 2.00 0.815 0.020

295211 304.0 304.7 0.7 1.48 0.571 0.135

295212 304.7 305.2 0.5 1.16 0.408 0.018

295213 305.2 305.9 0.7 0.70 0.571 0.003

295214 305.9 306.2 0.3 0.70 0.245 0.146

295216 306.2 306.5 0.3 0.66 0.245 0.595

295217 306.5 306.8 0.3 0.78 0.245 1.62

295218 306.8 307.2 0.4 0.66 0.326 2.20

295219 307.2 307.6 0.4 1.34 0.326 1.35

295220 307.6 307.9 0.3 0.64 0.245 3.69

295221 307.9 308.2 0.3 0.50 0.245 0.865

295215 308.2 309.0 0.8 0.82 0.652 0.325

295224 309.0 310.0 1.0 0.92 0.815 0.006

295225 310.0 310.7 0.7 0.62 0.571 0.311

295226 310.7 311.1 0.4 0.62 0.326 1.54

295227 311.1 311.6 0.5 1.34 0.408 2.17

295228 311.6 312.0 0.4 0.52 0.326 1.36

295229 312.0 312.5 0.5 0.96 0.408 0.577 295230 312.5 313.0 0.5 1.34 0.408 0.050 295231 313.0 314.0 1.0 2.52 0.815 0.059 295232 314.0 315.0 1.0 2.16 0.815 0.002 295233 315.0 316.0 1.0 2.36 0.815 0.002 295234 316.0 317.0 1.0 2.34 0.815 0.007 295235 327.9 328.4 0.5 1.02 0.408 0.019 295236 328.8 329.5 0.7 1.16 0.571 0.025 295238 330.2 330.5 0.3 0.62 0.245 0.041 295239 332.5 332.9 0.4 0.82 0.326 0.028

Weighted Average – Composite 2006-11 15.0 0.41



October 29, 2008



From (m)

To (m)

Length (m)

Total Weight (kg)


U3O8

(%) 295467 250.0 251.0 1.0 1.06 0.800 0.0005 295468 251.0 252.0 1.0 0.94 0.800 0.0002 295469 252.0 253.3 1.3 1.94 1.040 0.0009 295470 253.3 253.7 0.4 1.66 0.320 0.0033 295471 253.7 254.7 1.0 1.82 0.800 0.149 295472 254.7 255.1 0.4 0.88 0.320 0.079 295473 255.1 255.4 0.3 0.60 0.240 1.00 295475 255.4 256.0 0.6 1.16 0.480 0.969 295476 256.0 256.7 0.7 1.58 0.560 2.97 295477 256.7 257.3 0.6 0.54 0.480 4.05 295478 257.3 257.8 0.5 1.28 0.400 0.431 295479 257.8 258.3 0.5 0.74 0.400 1.41 298005 258.3 258.6 0.3 0.88 0.240 0.400 295481 258.6 259.0 0.4 0.96 0.320 1.48 295482 259.0 259.3 0.3 0.68 0.240 0.542 295483 259.3 259.7 0.4 0.70 0.320 3.60 295484 259.7 260.1 0.4 0.98 0.320 1.43 295485 260.1 261.1 1.0 2.32 0.800 0.191 295486 261.1 261.7 0.6 0.14 0.140 0.586 295487 261.7 262.0 0.3 0.56 0.240 11.2 295488 262.0 262.5 0.5 0.94 0.400 0.918 295490 262.5 263.0 0.5 0.92 0.400 1.00 295491 263.0 263.4 0.4 0.80 0.320 8.90 295492 263.4 263.8 0.4 0.78 0.320 1.09 295493 263.8 264.2 0.4 0.48 0.320 2.83 295494 264.2 264.5 0.3 0.72 0.240 16.6 295495 264.5 265.0 0.5 1.30 0.400 2.14 295496 265.0 265.5 0.5 0.90 0.400 0.468 295497 265.5 266.0 0.5 0.98 0.400 0.174 295498 266.0 266.5 0.5 1.14 0.400 0.184 295499 266.5 267.0 0.5 1.12 0.400 0.264 298001 267.0 267.5 0.5 0.90 0.400 0.446 298002 267.5 268.0 0.5 1.02 0.400 0.084 298003 268.0 268.5 0.5 1.12 0.400 0.435 298004 268.5 269.0 0.5 0.96 0.400 0.214 298007 269.0 270.0 1.0 2.20 0.800 0.053 298008 270.0 271.0 1.0 2.36 0.800 0.045 298009 271.0 272.0 1.0 2.46 0.800 0.010 298010 272.0 273.0 1.0 2.44 0.800 0.003 298011 278.0 279.0 1.0 1.94 0.800 0.005 298012 279.0 280.0 1.0 2.22 0.800 0.010




October 29, 2008



From (m)

To (m)

Length (m)

Total Weight (kg)


U3O8

(%)

298325 291.0 291.5 0.5 1.20 0.286 0.015

298326 291.5 292.5 1.0 2.16 0.571 2.111 298327 292.5 293.0 0.5 1.24 0.286 1.958

298329 293.0 293.3 0.3 0.70 0.171 0.094 298330 293.3 294.0 0.7 1.02 0.400 0.955

298331 294.0 295.0 1.0 2.32 0.571 1.651 298332 295.0 296.0 1.0 2.30 0.571 2.3

298333 296.0 297.0 1.0 2.04 0.571 2.064 298335 297.0 298.0 1.0 1.88 0.571 1.392

298336 298.0 299.0 1.0 2.02 0.571 2.972 298337 299.0 299.4 0.4 0.82 0.229 3.833 298338 299.4 299.9 0.5 1.18 0.286 0.248

298339 299.9 300.1 0.2 0.44 0.097 1.946 298340 300.1 300.7 0.6 0.86 0.360 0.028

298341 300.7 301.0 0.3 0.14 0.140 0.519 298342 301.0 303.0 2.0 1.36 1.143 5.684

298343 303.0 303.9 0.9 0.46 0.460 2.476 298345 303.9 304.3 0.4 0.72 0.200 0.955

298346 304.3 305.0 0.8 1.26 0.429 0.088 298348 305.0 305.8 0.8 1.20 0.457 0.066

298349 305.8 306.5 0.7 2.04 0.400 0.073 298350 306.5 306.9 0.4 0.76 0.200 1.285 298351 306.9 307.6 0.7 1.62 0.400 0.159 298352 307.6 308.5 1.0 2.28 0.543 0.026




October 29, 2008



From (m)

To (m)

Length (m)

Total Weight (kg)


U3O8

(%)

298378 315.0 316.0 1.0 2.20 0.783 0.002 298379 316.0 317.0 1.0 1.96 0.783 0.007

298380 317.0 317.9 0.9 2.12 0.665 0.002 298381 317.9 318.4 0.5 1.08 0.391 0.110

298382 318.4 319.1 0.7 1.80 0.548 0.032 298383 319.1 319.9 0.8 1.58 0.626 0.155

298384 319.9 320.0 0.1 0.16 0.078 0.783 298385 320.0 320.2 0.2 0.26 0.157 0.149

298387 320.2 321.0 0.9 1.52 0.665 1.28 298388 321.0 321.6 0.6 1.38 0.446 0.916 298389 321.6 322.1 0.5 1.02 0.376 0.130

298390 322.1 322.8 0.8 1.62 0.587 1.38 298391 322.8 323.4 0.6 1.50 0.470 0.030

298392 323.4 324.0 0.6 1.50 0.470 0.091 298394 324.0 324.2 0.2 0.22 0.117 0.744

298395 324.2 324.8 0.7 1.68 0.540 0.162 298396 324.8 325.4 0.5 1.16 0.399 0.096

298397 325.4 325.8 0.4 0.40 0.313 2.47 298398 325.8 326.4 0.7 0.86 0.509 0.520

298399 326.4 326.7 0.3 0.30 0.235 6.24 298400 326.7 327.4 0.7 1.40 0.548 0.089 298401 327.4 328.1 0.7 1.14 0.548 0.070

298402 328.1 329.0 0.9 1.36 0.704 0.314 298403 329.0 329.8 0.8 1.76 0.650 0.063 298405 329.8 330.7 0.8 1.00 0.650 0.044 298406 330.7 331.5 0.8 0.96 0.657 0.057 298407 331.5 333.0 1.5 1.00 1.000 0.220 298408 333.0 334.0 1.0 1.94 0.783 0.098 298409 334.0 335.0 1.0 1.98 0.783 0.001 298410 335.0 336.0 1.0 2.30 0.783 0.001 298411 336.0 337.0 1.0 2.30 0.783 0.002 298412 337.0 338.0 1.0 2.20 0.783 0.002




October 29, 2008

The Overall Composite was subsequently prepared on a weighted basis (according to drill hole intersection) as a blend of the four individual drill hole sub-composites as listed in Table 13 below.


Make-up of Overall Composite

Composite No. Length

(m) Weight to

Composite (kg) U3O8

(%) 2006-11 18.4 7.9 0.41 2006-30 25.0 10.7 1.08 2007-03 17.5 7.5 1.77 2007-06 23.0 9.9 0.33

Weighted Average – Overall Composite 0.87

8.2 Composite Analyses

The four individual drill hole sub-composites and the Overall Composite were sampled as part of the riffling of test charges and submitted for analysis. The head assays are listed in Table 14 and Table 15 below, Table 14 providing the whole rock analysis and Table 15 providing the elemental analyses.

Table 14 Strateco Resources Inc. – Matoush Uranium Project Matoush Test Composites - Whole Rock Analysis, %

Analyte Composite 2006-11

Composite 2006-30

Composite 2007-03

Composite 2007-06

Overall Composite

U3O8 0.34 0.98 1.61 0.30 0.79 Al2O3 5.88 5.31 5.34 6.91 5.92 CaO 0.45 0.62 1.79 0.67 0.85 Cr2O3 0.30 0.55 0.96 0.21 0.47 Fe2O3 2.01 0.72 1.17 0.64 1.13 K2O 2.62 2.02 1.91 2.55 2.26 MgO 2.22 0.52 0.56 0.47 0.86 MnO <0.01 0.01 0.01 <0.01 0.02 Na2O 0.35 0.39 0.22 0.52 0.40 P2O5 0.37 0.43 1.84 0.53 0.72 SiO2 82.3 87.0 80.4 86.3 84.5 TiO2 0.97 0.06 1.00 0.25 0.51 V2O5 0.10 0.07 0.21 0.12 0.11 LOI 1.60 1.24 2.09 1.26 1.39



October 29, 2008

Table 15 Strateco Resources Inc. – Matoush Uranium Project Matoush Test Composites – Elemental Analysis, %

Analyte Composite 2006-11

Composite 2006-30

Composite 2007-03

Composite 2007-06

Overall Composite

Ag <0.003 <0.003 <0.003 <0.003 <0.003 As <0.003 <0.003 <0.003 <0.003 <0.003 Ba 0.069 0.046 0.27 0.056 0.097 Be 0.00009 0.00016 0.0003 0.00009 0.00015 Bi 0.0059 <0.003 0.0080 <0.003 <0.003 Cd <0.0002 <0.0002 <0.0002 <0.0002 <0.0002 Co 0.0047 0.0013 0.0061 0.0013 0.0029 Cu 0.0031 0.0027 0.0110 0.0014 0.0039 Li 0.0008 <0.0005 <0.004 <0.004 <0.0005

Mo 0.027 <0.004 <0.004 <0.004 0.004 Ni 0.012 0.0026 0.0090 0.0034 0.0061 Pb 0.097 0.180 0.270 0.0510 0.140 Sb <0.001 <0.001 <0.001 <0.001 <0.001 Se 0.011 <0.005 0.011 <0.005 <0.005 Sn <0.002 <0.002 <0.002 <0.002 <0.002 Sr 0.0074 0.0063 0.021 0.0056 0.0087 Tl <0.003 <0.003 <0.003 <0.003 <0.003 Y 0.0009 0.0008 0.0022 0.0008 0.0011 Zn 0.0025 0.0030 0.0023 <0.001 0.0018

The elemental analyses of the composites show that the Matoush composites prepared for testing are low in deleterious elements such as arsenic, molybdenum and base metals. Selenium is above the detection limit in two of the composites, which implies that the amount reporting to waste effluents will need to be taken into consideration in effluent treatment testwork.



October 29, 2008

8.3 Uranium Speciation Analysis of Overall Composite

The uranium isotope ratio of the Overall Composite was measured to confirm that the uranium isotope ratio was close to the naturally occurring ratio of 0.71% U235/U (w/w). The measured values, performed by Becquerel Laboratories Inc. of Mississauga Ontario using gamma-ray spectrometry are listed in Table 16 below:


Matoush Overall Composite – Uranium Speciation AnalysisIsotope ppm U-238 7,600 ± 400 U-235 40 ± 10

Within the limited accuracy of the gamma spectrometry analysis the measured isotope ratio is approximately 0.52% U235/U (w/w), with a range of 0.37% U235/U (w/w) to 0.69% U235/U (w/w), close to the expected ratio of 0.71% U235/U (w/w). If warranted a more precise measurement could be obtained by delayed neutron counting (DNC) for U235 and neutron activation analysis (NAA) for U238.



October 29, 2008

8.4 Petrographic Analyses

During the June 7 and 8, 2007 site visit by Lawrence Melis, a series of samples, 18 in total representing high and low grade mineralization as well as country rock, were selected for petrographic analyses. These samples, suitably prepared as polished thin sections (PTS) by the University of Saskatchewan Geology Department, were forwarded to Terra Mineralogical Services (Giovanni Di Prisco) in Peterborough, Ontario, along with three other samples subsequently selected by Jonathan Lafontaine of Strateco from DDH MT-07-54 to represent host rock lithology. The PTS sample list is presented in Table 17 below.


Selection of Samples for Polished Thin Section ExaminationDDH No. Intersection (m) Type

06-02 290.9 Mineralization06-04 302.3 Mineralization06-04 307.5 Mineralization06-10 315.7 Mineralization06-10 308.0 Mineralization06-10 327.0 Mineralization06-18 272.0 Mineralization06-30 257.0 Mineralization06-30 264.4 Mineralization07-09 938.0 Granitoid Rock07-09 941.0 Granitoid Rock07-09 945.0 Granitoid Rock07-09 977.7 Granitoid Rock07-09 987.0 Granitoid Rock07-22 769.7 Granitoid Rock07-22 771.5 Granitoid Rock07-35 295.5 Mineralization07-48 213.5 Mineralization

MT-07-54 223.5 Host Rock-CBF MT-07-54 230.0 Host Rock-ACF MT-07-54 196.7 Host Rock-Saccharoid

Mineralogical information derived from these samples, excerpted from Terra’s report, is summarized below.

The two main uranium-bearing minerals identified in the Matoush mineralization are uraninite (UO2) and carnotite-weeksite K2(UO2)2V2O8·3(H2O) and K2(UO2)2(Si2O5)3·4(H2O). Very minor amounts of uranophane



October 29, 2008

[CaH2(SiO4)2(UO2)·5(H2O)] were thought to be also present.

Uranium correlates very well with chromium, vanadium, phosphorus, lead, and boron. Chromium is predominantly hosted in chromite (Fe2+Cr2O4) and in the latest growth zones of tourmaline [NaAl6M3 (Si2O6)3(BO3)3 (OH,F)4 where M: Fe, Mn, Mg, Li, Al]. Tourmaline deposition correlates well with uranium mineral abundance, therefore chromium could represent the most useful pathfinder for primary uranium mineralization as uraninite. Vanadium remobilization during late-stage alteration of uraninite resulted in the growth of a carnotite-weeksite solid solution, which contains vanadium. This mineral association suggests that vanadium could be the most useful pathfinder for secondary uranium deposition as carnotite. Phosphorus is hosted in apatite [Ca5(PO4)3(OH,F,Cl)] and is generally intergrown with uraninite, therefore, the abundance of apatite should also correlate to uranium abundance. Finally, lead, contained as a by-product in uraninite, could indicate the close proximity of uraninite depositions.

Three types of uraninite deposition were identified. Uraninite types I and II, as well as carnotite, present as a broad range of grain sizes (2 mm to 60 mm), and often complex textural intergrowths with gangue minerals, or sandstone matrix. Therefore, these two uraninite types and carnotite would require relatively fine grinding (K80 of 30 µm to 40 µm) in order to achieve good mineral liberation, or at least become accessible to leaching solutions. In contrast, the third type of uraninite (type III) is commonly coarse grained, correlates well with high-grade samples, and mainly forms simple intergrowth textures that would liberate well, or would be exposed to leaching solution with a coarser grind target (K80 of 70 µm to 100 µm). The mineralized samples commonly contain calcite (CaCO3) (from a few percent up to 20%) and very fine-grained sericite-muscovite (a few percent up to 30%). The occurrence of calcite in a leach circuit could impact on reagent costs (calcite being an acid consumer), whereas fine-grained micas could effect solid/ liquid separation.



October 29, 2008

8.5 Initial Leaching Tests

Ten leach tests were completed on the overall Matoush composite, a blend of all four drill hole sub-composites assaying 0.79% U3O8, to provide initial leach conditions for the Matoush mineralization. The conditions used in these tests are summarized in Table 18 below.


Matoush Overall Composite - Summary of Leach Test ConditionsTest Conditions Reagent Additions

Test No. Temp ºC

Average FA g H2SO4/L

Avg. ORP mV

(Ag/AgCl)

Grind K80

µm H2SO4

kg/t NaClO3

kg/t O2

mL/min H2O2

kg/t Fe3+

g/L

MU-1 50.9 13.8 451 ~100 57.5 1.0 - - ~1 ** MU-1R 50.8 13.7 483 ~100 52.8 0.9 - 2.5 MU-2 50.6 45.8 478 ~100 138.1 0.9 - 3.0 MU-4(1) 54.9 16.0 (12 h) 470 ~100 53.7 - 2.6 (5-12 PSI) - 1.3 MU-5 50.8 13.3 484 ~75 57.9 1.2 - 3.1 MU-4R(2) 52 30 456 100 76.8 - 5 - 2 MU-6(3) 53 - 92 100 - - - 54.5 - MU-7 57 26 471 100 94.5 0.9 - - - MU-7R(4) 50 23 500 100 92.4 0.8 - - - MU-8 54 26 460 75 94.2 0.8 - - - MU-9 52 42 501 100 140.3 - 5 - 2 MU-10 50 15 446 100 64.4 50 -

Notes: 1. Free acid not maintained in last 12 hours of leach, difficulty with O2 sparging. 2. Repeat of Test MU-4 where difficulties were experienced with oxygen sparging in test. 3. Carbonate Leach at pH 8.9 using 101.5 kg Na2CO3/t, 40.6 kg NaHCO3/t and 109 kg H2O2 (50%)/t for emf control.

4. Repeat of Test No. MU-7



October 29, 2008

Test results are summarized in Table 19 below.


Matoush Overall Composite - Summary of Leach Test Results % U3O8 % U3O8 Extraction Test No.

Feed Residue Weight Loss, % Final Preg Sol’n

g U3O8/L 8 hours 12 hours 24 hours MU-1 0.74 0.10 2.0 3.95 80.2 84.2 86.8 MU-1R 0.73 0.096 1.0 3.81 81.2 83.9 87.1 MU-2 0.76 0.013 3.0 4.38 97.5 98.0 98.4 MU-4(1) 0.74 0.17 4.0 3.40 77.8 77.8 -MU-5 0.76 0.11 4.0 4.06 82.2 83.5 84.9 MU-4R(2) 1.53(5) 0.021 5.0 3.93 97.7 98.1 98.5 MU-6(3) 0.97 0.28 7.0 3.77 61.8 66.9 71.1 MU-7 1.01 0.035 1.0 4.84 94.1 95.8 96.5 MU-7R(4) 0.91 0.025 3.0 4.42 95.0 96.0 97.3 MU-8 1.03 0.038 - 4.95 90.3 94.5 96.3 MU-9 0.90 0.017 4.0 4.06 96.7 97.2 98.2 MU-10 0.78 0.078 5.0 3.47 86.4 90.4 90.4

Notes: 1. Free acid not maintained in last 12 hours of leach, difficulty with O2 sparging, only 12 hour leach result reported.

2. Repeat of Test MU-4 where difficulties were experienced with oxygen sparging in test. 3. Carbonate Leach at pH 8.9 using 101.5 kg Na2CO3/t, 40.6 kg NaHCO3/t and 109 kg H2O2 (50%)/t

for emf control. 4. Repeat of Test No. MU-7. 5. Anomalous value.

8.6 Confirmation Leaching Tests

Repeat testing of three of the four variability composites was completed to test the metallurgical leaching response using only low pressure oxygen in place of oxygen and sodium chlorate. In addition a 12 kg bulk leach of the overall composite was completed to prepare pregnant leach solution and leach residue for downstream uranium recovery tests and waste treatment tests. Hydrogen peroxide was used as the oxidant in this test in order to be able to complete it under atmospheric conditions (due to the amount of material leached) instead of low pressure oxygen.



October 29, 2008

Test conditions are summarized in Table 20 and results are tabulated in Table 21 below.


Matoush Individual Composites - Summary of Additional Leach Test ConditionsTest Conditions Reagent Additions

Composite Test No. Temp ºC

Avg. FA g H2SO4/L

Avg. ORP mV

Grind K80

µmH2SO4

kg/tH2O2

kg/t O2 PSI Fe3+

g/L2006-11 MU-11R 56 43 466 87 124.1 - 5 - 2006-30 MU-12R 55 33 436 99 102.3 - 5 - 2007-03 MU-13R 51 31 460 99 91.2 - 5 - 2007-06 MU-14 54 30 474 91 97.5 - 5 - Overall MU-15 59 28 481 100 103.5 3.39 - -


Matoush Individual Composites - Summary of Additional Leach Test Results % U3O8 Final Preg Sol’n % U3O8 Extraction

Composite Test No. Feed Residue

Weight Loss % g U3O8/ L g Fe3+/ L 6 hours 12 hours 24 hours

2006-11 MU-11R 0.39 0.007 4 1.78 0.86 97.9 98.2 97.9 2006-30 MU-12R 1.19 0.014 6 5.51 0.89 97.1 98.3 98.6 2007-03 MU-13R 0.34 0.004 5 1.65 1.26 97.2 97.9 98.6 2007-06 MU-14 0.33 0.005 2 1.53 1.16 97.5 97.9 98.2 Overall MU-15 0.85 0.015 0.5 3.60 1.12 94.8 97.9 97.9

.

The above results confirm that a low pressure (5 PSI) oxygen sulphuric acid leach is the leach option of choice for the Matoush mineralization, achieving a uranium extraction of 98 % or better. All four variability sub-composites yielded excellent extractions showing that uniform extractions can be obtained from the Matoush deposit based on the mineralization tested. The optimized Matoush leach conditions can be summarized as follows:

• A mesh-of-grind K80 of 150 µm,

• An ORP (oxidation-reduction potential) of approximately 475 mV controlled by the addition of oxygen under low (5 PSI) pressure,

• A target free acid concentration of 30 g H2SO4/L representing an acid consumption of 104 kg H2SO4/t, and

• A 12 hours leach residence time.



October 29, 2008

8.7 Pregnant Solution Analyses

The pregnant leach solution produced in some of the acid leach scoping tests was submitted to an ICP package analysis to give an indication of the elemental make-up of leach solution reporting to downstream uranium recovery. The results are summarized in Table 22 below. Two results are for the Overall Composite and two results are for Composites 2006-30 and 2007-3.


Matoush Leach Tests – Pregnant Leach Solution Analyses, mg/L

Analyte Overall

Composite Test MU-7

Overall Composite Test MU-8

Composite 2006-30

Test MU-12

Composite 2007-03

Test MU-13 Ag < 3 < 3 < 10 < 15 Al 1100 1100 740 1200 As 7 5 7 < 6 Ba < 0.2 < 0.2 < 0.08 0.32 Be 0.4 0.36 < 0.4 0.68 Bi < 4 < 4 < 4 < 2 Ca 790 800 920 1100 Cd < 0.09 < 0.09 < 0.09 < 0.09 Co 4.3 4.3 3.1 8 Cr 280 280 220 470 Cu < 15 < 15 5 16 Fe 2700 3000 2000 2500 K 400 390 270 280 Li < 2 < 2 < 2 < 2

Mg 310 290 120 130 Mn 24 23 22 28 Mo 1.6 1.3 2.1 1.6 Na 180 140 110 120 Ni 10 10 7 13 P 1600 1700 940 3300

Pb 11 11 12 12 Sb < 2 < 2 < 2 < 3 Se < 3 < 3 < 3 < 3 Sn < 2 < 2 < 2 < 2 Sr 3 2.6 1.4 5.1 Ti 21 17 15 40 Tl < 3 < 3 < 3 < 3 U 4100 4200 4550 7100 V 43 37 19 65 W < 2 < 2 < 2 < 2 Y 4.9 4.6 3.1 8.5 Zn 3 3 < 5 < 2



October 29, 2008

The resulting pregnant leach solution is quite low in deleterious elements, particularly arsenic, molybdenum, selenium and base metals, which will minimize the potential impact on waste treatment requirements. It is also very low in elements which would normally report to the final yellowcake uranium product, in particular molybdenum and vanadium. The relatively high phosphorus content originates from the apatite content in the mineralization. The somewhat high sodium content is due to the addition of sodium chlorate as oxidant for the leach; in practice the oxidant will be oxygen hence the salt content of the leach liquor will be low.

8.8 Ion Exchange Tests

Pregnant solution produced in the bulk leach test, Test MU-15 (which achieved a 98% uranium extraction and produced a pregnant leach solution assaying 3.60 g U3O8/L and 1.12 g Fe3+/L) was collected and use for ion exchange uranium recovery tests. Two resins were tested, Dowex 21K resin and Purolite A660 resin, both strong base ion exchange resins with a quaternary amine function group. The pregnant solution was contacted sequentially in five stages with resin, the barren solution from each stage sequentially contacted with fresh resin. Results achieved are summarized in Table 23 below.


Resin Adsorption Results Barren Solution Uranium Concentration, g U3O8/L Resin Loading, g/L % Loading Efficiency Resin Stage 1 Stage 2 Stage 3 Stage 4 Stage 5 U3O8 Fe SiO2 3 Stages 4 Stages 5 Stages

21K 2818 2170 1203 670 246 83.1 1.7 0.20 77.7 88.7 96.4 A660 2005 388 40 2 <1 60.1 0.38 0.79 99.3 100 100

The Purolite A660 resin outperformed the Dowex 21K resin in terms of loading efficiency with complete adsorption achieved in three stages, yielding a loaded resin containing 60 g U3O8/L.

A single test was completed where the A660 loaded resin was eluted with 125 g H2SO4/L sulphuric acid solution at 45 ºC. A 90% elution efficiency was achieved using 14 bed volumes. Further testwork, which will include continuous tests in addition to batch tests, will be required in future to optimize elution conditions.



October 29, 2008

8.9 Settling Tests

Settling tests were completed on leach feed and leach discharge from Test MU-15, a bulk test completed on an overall Matoush composite, to provide design data for reference purposes.

Leach Feed

Results on leach feed, neutral pulp after grinding to a K80 of 100 µm, are summarized in Table 24 below.


Summary of Static Settling Test Results on Test MU-15 Leach Feed Thickener Unit Area

(m2/t/day)(1) Feed

% Solids (w/w)

M10 Flocculant

(g/t)

Underflow Density %Solids

(w/w) Underflow Hydraulic 15 10 65 0.11 0.03 20 10 63 0.12 0.06 25 10 61 0.13 0.09 25 15 65 0.13 0.04 25 20 65 0.09 0.04 25 30 63 0.06 0.02 25 60 61 0.05 0.01

Note: 1. No safety factor applied

The above data show that neutral pulp settling is solids controlled [higher unit area versus the hydraulic (overflow) unit area], the higher flocculant dosage reduces thickener area requirements, and a lower feed pulp density (normally achieved with self-diluting feed wells) reduces unit area requirements and assists in achieving maximum underflow densities.



October 29, 2008

Leach Discharge

Thickener data for leach discharge, summarized in Table 25 below, was developed in the event a conventional counter-current-decantation/solvent extraction circuit is considered for the project in future (the present design concept uses resin-in-pulp to minimize capital costs and avoid the need to transport and store organic reagents).


Summary of Static Settling Test Results on Test MU-15 Leach Discharge Thickener Unit Area

(m2/t/day)(1) Feed

% Solids (w/w)

M10 Flocculant

(g/t)

Underflow Density %Solids

(w/w) Underflow Hydraulic 10 50 55 0.10 0.014 15 40 59 0.12 0.024 15 50 57 0.08 0.012 15 60 55 0.04 0.006 15 80 55 0.06 0.005 20 50 59 0.08 0.017


The above data show that leach acid pulp settling is solids controlled [higher unit area versus the hydraulic (overflow) unit area], the higher flocculant dosage reduces thickener area requirements with 60 g/t being the optimum amount, and a 15% solids (w/w) feed pulp density appearing to be the optimum.



October 29, 2008

8.10 Filtration Tests

For reference purposes in the event leach discharge filtration is considered for the project, vacuum filtration tests were completed on the leach slurry from Test MU-15 at 30 % solids (w/w) density. Results are summarized in Table 26 below.


Summary of Vacuum Filtration Test Results on Test MU-15 Leach Discharge Vacuum Filter Cycle Cake

Thickness (mm)

Cake Moisture (% H2O)

Form Time (min)

Dry Time (min)

Output (dry)(1)

(kg/m2/h)

Belt Filter 5 21.0 0.19 2.18 165 5 22.5 0.19 0.14 1191 5 24.0 0.19 0.01 1979

15 21.0 1.68 6.55 142 15 22.5 1.68 0.42 557 15 24.0 1.68 0.03 685

Disc or Drum Filter 5 22.1 0.19 0.31 628 9 21.7 0.60 1.01 349

13 21.5 1.26 2.10 242 17 21.4 2.16 3.59 185 21 21.3 3.29 5.48 150 25 21.2 4.66 7.77 126


PRELIMINARY DESIGN CRITERIA Page 9-1


October 29, 2008

9.0 PRELIMINARY DESIGN CRITERIA

A glossary of terms used in the preliminary design criteria is presented below in Table 27.


Glossary of Terms Used in Preliminary Design Criteria Term Description

% Percent

% (v/v) Volume percent % (w/w) Weight percent

% solids (w/w) Weight percent solids

% U3O8 Uranium, as Yellowcake /a per year (annum)

/day per day /h per hour

/L per litre /min per minute

µm microns, 10-6 metres

bar 100 kPa

BTU/lb British thermal unit, 251 Calories BV Bed Volume of resin

Dry t/m3 Unit of tailings density, excluding water ea. each

g grams

g/mole atomic weight h Hours

Kg kilograms kPa kilopascals

kW kilowatts

kWh kilowatt hours lbs pounds

m metres m2 square meters

m3 cubic metres mV (Ag/AgCl) millivolts, using the silver/silver chloride electrode as a zero point

Nm3 Normal cubic metres, ie, of air at standard temperature and pressure

ºC degree Celcius t tonnes



October 29, 2008

The preliminary design criteria used to calculate the mass balance, which will require updating as further testwork is completed, are listed in Table 28, below. The mass balance is attached in Appendix C.


Preliminary Design CriteriaCriteria Unit Value Comment General

Mill Feed Tonnage t/a 193,735 Strateco

Scheduled Operating Days days/a 350 Melis

Operating Time % 95 Melis

Net Operating Days days/a 333 Calculated

Mill Feed Tonnage t/h 24.2 Calculated

Mill Feed Tonnage t/d 582 Calculated

Average Feed Grade % U3O8 0.48 Strateco

Pounds Fed U3O8/a (millions) lbs U3O8/a 2,050,138 Calculated

Pounds Produced U3O8/a (millions) lbs U3O8/a 2,000,000 Strateco

Instantaneous Design lbs U3O8/a 2,192,192 Calculated

Instantaneous Calciner Discharge kg U3O8/h 114 Calculated

Tank Freeboard m 0.5 Melis

Mine Water Inflow m3/h 50 Estimated

Average Surface Drainage m3/h 0 Estimated

Ore Moisture % H2O (w/w) 5.0 Estimated

Specific Gravity of Ore t/m3 2.75 Matoush

Uranium Distribution


RIP Loss (Residue Loss) % 0.3 Calculated, Kemex Data

Loss in Gypsum Solids % 0.1 Calculated

Average Net Recovery % 97.6 Calculated

Crushing

Design Operating Time h/d 12.0 Typical

Primary (SAG Mill) Grinding

SAG Mill Power Index - Range min. 27.9 - 91.4 Testwork

SAG Mill Power Index - Average min. 55.1 Testwork

SAG Mill Power Index - 75th Percentile

min. 65.7 Testwork

SAG Mill Power Index - 90th Percentile

min. 83.3 Testwork

Percent Solids in Mill % (w/w) 65 Typical

Percent Screen Undersize % (w/w) 85 Estimated



October 29, 2008


Preliminary Design CriteriaCriteria Unit Value Comment Percent Solids in Screen Oversize % (w/w) 75 Estimated

SAG Mill Feed K80 µm 130,000 Typical

SAG Mill Transfer Size µm 2,400 Testwork

SAG Mill Size mm 3,962 mm ø x 2,286 Testwork

SAG Mill Power Draw (Shell) kW 296 Testwork

SAG Mill Ball Charge % (v/v) 10 Typical

Grinding Ball Size mm 100 Typical

Grinding Ball Consumption kg/t 0.50 Typical

Design Temperature ºC 5 - 30 Melis

Pebble Crusher

Load to Pebble Crusher % (w/w) 15 Estimated

Pebble Crusher Size mm 914 Melis

Close Side Setting mm 200 Melis

Pebble Crusher Power Draw kW 75 Melis

Secondary (Ball Mill) Grinding

Ball Mill Work Index - Range kWh/t 14.6 - 16.3 Testwork

Ball Mill Work Index - Average kWh/t 15 Testwork

Ball Mill Work Index - 75th Percentile kWh/t 16 Testwork

Ball Mill Work Index - 90th Percentile kWh/t 16.0 Testwork

Percent Solids in Mill % (w/w) 65 Typical

Percent Recirculating Load % (w/w) 300 Typical

Ball Mill Feed K80 µm 2,400 Typical

Ball Mill Discharge K80 µm 150 Typical

Ball Mill Size mm 2,743 mm ø x 4,879 Testwork

Ball Mill Power Draw (Shell) kW 310 Calculated

Grinding Ball Size mm 51 Typical

Grinding Ball Consumption kg/t 0.50 Typical

Ball Mill Discharge Pumpbox Retention Time

min 10 Typical

Ball Mill Discharge Pumpbox Volume m3 21.6 Calculated


Classification Cyclones

Number ea. 2 Melis

Cyclone size mm 660 Estimated

Cyclone Feed Density % (w/w) 50 Typical

Cyclone Overflow Density % (w/w) 33 Typical

Cyclone Underflow Density % (w/w) 57 Calculated




October 29, 2008


Preliminary Design CriteriaCriteria Unit Value Comment

Neutral Thickening

Number ea. 1 Melis

Thickener Type - High Rate Melis

Design Feed Density % solids (w/w) 15 Melis

Design Unit Area m2/t/d 0.08 Estimated

Thickener Diameter mm 7,925 Calculated

Overflow Clarity g/m3 200 Melis

Underflow Density % solids (w/w) 50 Melis

Flocculant Dosage g/t ore 60 Melis

Neutral Thickener Overflow Tank Retention Time

min 10 Typical

Neutral Thickener Overflow Tank Volume

m3 4 Calculated


Leaching

Nominal Feed Tonnage tonnes/h 24.2 Calculated

Temperature ºC 50 Melis


Leach Residue Grade % U3O8 0.010 Testwork

Weight Loss % 5.0 Melis

Oxidant type O2 Melis

O2 Requirement tonnes O2/d 30 Estimated

Discharge Oxidation/Reduction Potential (ORP)

mV (Ag/AgCl) 475 Estimated

Acid type H2SO4 Melis

Design Acid Consumption kg 100% H2SO4/t

ore 104 Testwork

Discharge Free Acid g H2SO4/L 30.0 Testwork

Retention Time (Total) h 12 Testwork

Number of Leach Tanks ea. 6 Melis

Nominal Leach Discharge Density % solids (w/w) 44.2 Calculated

Leach Tank Volume m3 70 Calculated

U3O8 in Leach Solution g/L 4.30 Calculated

Leach Discharge Pumpbox Retention Time

min 5 Typical

Leach Discharge Pumpbox Volume m3 2.82 Calculated

Design Temperature ºC 50 Melis

Resin In Pulp (RIP): Resin



October 29, 2008


Preliminary Design CriteriaCriteria Unit Value Comment Resin Type - Purolite A660 Melis

Resin Specific Gravity t/m3 1.2 Typical

Resin Loading g U3O8/L 55 Melis/Testwork

Eluted Resin g U3O8/L 2 Melis

Adsorption Efficiency % 99.7 Calculated

Proportion of Resin in Loading Tank % (v/v) 20.0 Kemix Typical

Resin Volume in Tank m3 20.0 Calculated

Resin Inventory m3 220 Calculated

Resin Loss m3/ktonne ore 0.30 Purolite

Resin in Pulp (RIP): Resin Loading

Feed Flow Rate m3/h 42 Calculated

Feed Concentration g/L 4.30 Calculated

RIP Tank Retention Time h 2 Calculated

RIP Tank Volume m3 100 Kemix

Number of Tanks ea. 8 Kemix

Pulp Discharge Concentration g U3O8/L 0.009 Kemix Typical

Residue Pumpbox Retention Time min 5.0 Typical

Residue Pumpbox Volume m3 3.5 Calculated

Loaded Resin Transfer Flowrate m3/h 146 Kemix Typical

Loaded Resin Pumpbox Retention Time

min 5.0 Typical

Loaded Resin Pumpbox Volume m3 12.2 Calculated

Wash Water to Screens m3/h 10.0 Typical

Resin Screen Oversize Density % (w/w) 80.0 Typical


Resin in Pulp (RIP): Resin Elution

Cycle Time H 8.8 Calculated

Elution Cycles per Day ea. 3 Calculated

Number of Columns - 3 Assumed

Resin Bed Volume m3 20.0 Calculated

Resin Return Density % (w/w) 80 Melis file data

Acid Elution Volume BV 5.0 Melis file data

Acid Elution Tank Size m3 100 Calculated

Elution Column Dead Space % BV 50 Typical

Elution Column Volume m3 30 Calculated

Elution H2SO4 Concentration g/L 150 Melis file data



October 29, 2008


Preliminary Design CriteriaCriteria Unit Value Comment Elution Flow Rate BV/h 1.5 Melis file data

Concentrated Eluate U3O8 Concentration

g/L 15.0 Melis file data

Concentrated Eluate H2SO4 Concentration

g/L 127 Calculated

Elution Temperature ºC 45 - 50 Melis

Resin Holding Tank Size BV 1.25 Typical

Resin Holding Tank Volume m3 25 Calculated

Resin Transfer Tank Retention Time min 5 Typical

Resin Transfer Tank Volume m3 2.4 Calculated

Impurity Precipitation

Feed Flowrate m3/h 8.1 Calculated

Ferric Sulphate Addition Dosage g/m3 25 Melis

Aeration m3/h/m2 20 Melis

Final Circuit pH Units 3.2 Melis

Final Circuit SO4 g/L 0.030 Calculated

SG of CaSO4.2H2O kg/L 2.32 Reference

Percent U3O8 in Gypsum Solids % 0.01 Melis

Percent Lime to Tank No. 1 % 50 Melis



Retention Time (Total) h 10.0 Melis

Number of Impurity Precipitation Tanks

ea. 6 Melis

Impurity Precipitation Tank Size m3 19.6 Calculated

Impurity Precipitation Discharge Pumpbox Retention Time

min 5 Melis

Impurity Precipitation Discharge Pumpbox Volume

m3 1.0 Melis


Gypsum Belt Filter

Number ea. 1 Melis

Gypsum Filter Feed Tank Retention Time

h 4 Melis

Gypsum Filter Feed Tank Volume m3 50 Melis

Barren Solution Wash Flow m3/h 4.87 Melis

Filtrate Clarity g/m3 1,000 Melis

Belt Filter Discharge Density % 40 Melis



October 29, 2008


Preliminary Design CriteriaCriteria Unit Value Comment Repulped Gypsum Density % 25 Melis

Percent U3O8 in Washed Filter Cake % 0.05 Melis

Filter Design Factor Units 1.1 Melis

Filter Design Feed Rate t/h 1.8 Calculated

Filter Unit Area kg/h/m2 200 Melis

Filter Belt Area m2 8.9 Calculated

Gypsum Filter Repulp Tank Retention Time

h 1 Melis

Gypsum Filter Repulp Tank Volume m3 5.6 Melis


Pregnant Solution Sand Filters

Sand Filter Feed Flow m3/h 14.2 Calculated

Number of Sand Filters ea. 3 Melis

Unit Area Feed Flow m/h 34.7 Calculated

Diameter of Sand Filters m 0.610 Calculated

Unit Area Backwash Flow m/h 36.7 Calculated

Total Backwash Flow m3/h 10.7 Calculated

Unit Area Air Scour Flow m/h 20.4 Calculated

Total Area Air Scour Flow Nm3/h 6.0 Calculated

Backwash Length/Cycle min 30 Melis

Backwash Cycles/Day (3 filters) ea. 3 once/day/filter

Discharge Clarity g/m3 20 Melis

Pregnant Solution Tank Flow m3/h 13.5 Melis

Pregnant Solution Tank Retention Time h 6 Melis

Pregnant Solution Tank Volume m3 81.1 Melis


Uranium Precipitation

Design Feed Flow Rate m3/h 13.5 Calculated

Design Feed Grade g U3O8/L 8.5 Calculated

Number of Uranium Precipitation Tanks

ea. 3 Estimated

Retention Time (Total) h 6.0 Melis

Cooling Water to heat Exchangers m3/h 2.5 Estimated

Uranium Precipitation Tank Volume m3 40.6 Calculated

Barren Solution Concentration g U3O8/L 0.01 Melis



October 29, 2008


Preliminary Design CriteriaCriteria Unit Value Comment Final Circuit pH Units 4.2 Melis

Final Circuit SO4 g/L 0.003 Calculated

Percent U3O8 in Uranium Precipitate % 79 Calculated

U Precipitation Discharge Pumpbox Retention Time

min 10 Melis

U Precipitation Discharge Pumpbox Volume

m3 2.3 Melis


Uranium Thickener

Number of Thickeners ea. 1 Melis

Uranium Thickener Flocculant Dosage g/t 150 Melis


Underflow Density % 35 Melis

Feed Rate t/h 0.15 Calculated

Unit Area m2/t/d 3.0 Melis

Thickener Diameter mm 3,962 Calculated

Uranium Thickener Overflow Tank Retention Time

min 5 Melis

Uranium Thickener Overflow Tank Volume

m3 3.5 Calculated


Barren Solution


Barren Solution to Scrubbers % 80 Melis

Barren Solution Tank Retention Time h 4 Melis

Barren Solution Tank Volume m3 166.4 Melis


Uranium Centrifuge

Operating/Standby ea. 1/1 Melis

Uranium Wash Tank Retention Time h 4 Melis

Uranium Wash Tank Volume m3 2.4 Melis

Number of Centrifuges ea. 1 Melis

Wash Water m3/h 0.3 Melis

Centrate Clarity g/m3 1,000 Melis

Discharge Density % 50 Melis

Centrifuge Design Factor Units 1.5 Melis

Centrifuge Design Feed t/h 0.21 Calculated

Uranium Centrifuge Wash Water min 15 Melis



October 29, 2008



Pumpbox Retention Time

Uranium Centrifuge Wash Water Pumpbox Volume

m3 0.1 Melis


Uranium Calciner and Scrubbers

Calciner Design Factor Units 1.5 See Centrifuge

Calciner Design Feed t/h 0.21 See Centrifuge

Percent U3O8 in Calciner Discharge % 100 Melis

Bulk Density of Calcined Uranium t/m3 2.2 Melis

Percent Uranium to Scrubbers % 0.1 Melis

Calciner Scrubber Water m3/h/scrubber 11.4 Melis

Calciner Room and Packaging Scrubbers Water m3/h/scrubber 8.1 Melis

Uranium Centrifuge Wash Water Pumpbox Retention Time

min 5 Melis

Uranium Centrifuge Wash Water Pumpbox Volume

m3 2.3 Melis


Uranium Bin

Uranium Bin Retention Time days 30 Melis

Uranium Bin Volume m3 37 Calculated


Calcined Uranium

No. of Drums/Shipping Container ea. 35 Melis

Average Calcined Uranium in Standard Drum

kg 400 Melis

No. of Standard Drums/d of Calcined Uranium

drums/d 6.9 Calculated

No. of Standard Drums/a of Calcined Uranium

drums/a 2,270 Calculated

No. of Times Each Drum is Recycled ea. 4 Typical

No. of New Drums/a drums/a 568 Calculated

No. of 20' Shipping Containers/a of Calcined Uranium

ea. 65 Calculated

No. of Standard Drums/a of Dried Uranium

ea. 3,950 Calculated

No. of 20' Shipping Containers/a of Dried Uranium ea. 113 Calculated


Tailings Neutralization

Nominal Feed Flow m3/h 108 Calculated

Number of Neutralization Tanks ea. 2 Melis



October 29, 2008


Preliminary Design CriteriaCriteria Unit Value Comment Tank No. 1 pH Units 4 Melis

Tank No. 2 pH Units 10 Melis

Aeration m3/h/m2 20 Melis

Lime Addition to Tailings kg CaO/t ore 40 Typical

Weight Gain in Tailings t tailings/t ore 1.2 Testwork

Specific Gravity of Tailings Units 2.4 Typical

Total Retention Time h 2 Melis

Neutralization Tank Volume m3 108 Calculated

Design BaCl2.2H2O Addition g/m3 35 Melis

Design Ferric Sulphate Addition g Fe/m3 40 Melis

Tailings Thickener Feed t/h 108 Calculated

Tailings Thickener Flocculant Addition g/t 60 Typical


Tailings Thickener Underflow Density % (w/w) 55 Target

Unit Area m2/t/d 0.30 Estimated

Thickener Diameter mm 16,459 Estimated

Tailings Thickener Overflow Tank Retention Time

min 10 Melis

Tailings Thickener Overflow Tank Volume

m3 12.0 Calculated

Consolidated Tailings Weight Dry t/m3 0.9 Typical

Tailings Consolidation Density % solids (w/w) 59 Calculated


Effluent Treatment

Feed to Effluent Treatment m3/h 11 Melis

Effluent Treatment Stages ea. 2 Melis

First Stage pH Units 3.5 Testwork

Second Stage pH Units 7.5 Testwork

Tank Retention Time (Total Stage) h 1.5 Melis

Reaction Tanks per Stage ea. 2 Melis

Reaction Tank Volume m3 8.5 Melis

Acid Dosage, Primary g H2SO4/L 0.07 Melis

Acid Dosage, Secondary g H2SO4/L 0.00 Melis

Ferric Sulphate Dosage, Primary g Fe/m3 50 Melis

Ferric Sulphate Dosage, Secondary g Fe/m3 50 Melis

Design BaCl2

.2H2O Addition, Secondary

g/m3 35 Melis



October 29, 2008



Lime Dosage, Primary g/m3 0 Melis

Lime Dosage, Secondary g/m3 11 Melis

Clarifiers per Stage ea. 1 Melis

Clarifier Overflow Tank Retention Time

min 10 Melis

Clarifier Overflow Tank Volume m3 1.9 Calculated

Flocculant Dosage mg/L 7 Melis

Clarifier Overflow Clarity g/m3 50 Typical

Clarifier Underflow Density % (w/w) 18 Typical

Effluent Sand Filters

Number of Sand Filters ea. 3 Melis

Unit Area Feed Flow m/h 8.0 Calculated

Diameter of Sand Filters m 0.000 Calculated

Unit Area Backwash Flow m/h 36.7 Calculated

Total Backwash Flow m3/h 0.0 Calculated

Unit Area Air Scour Flow m/h 20.4 Calculated

Total Area Air Scour Flow Nm3/h 0.0 Calculated

Backwash Length/Cycle min 30 Melis

Backwash Cycles/Day (3 filters) ea. 3 once/day/filter


Nominal Feed Clarity g/m3 18 Melis

Average Backwash Recycle % #DIV/0! Calculated

Effluent Discharge Tank Retention Time

cycles 6 Melis

Effluent Discharge Tank Volume m3 0.0 Calculated

Reverse Osmosis

Feed to Reverse Osmosis m3/h 75 Calculated

Reverse Osmosis Permeate % of Feed 85 Typical

Monitoring Ponds

Monitoring Pond Discharge Flowrate m3/h 110 Calculated

Monitoring Pond Retention Time, each h 24 Melis

Monitoring Pond Active Volume, each m3 620 Calculated

Design Temperature ºC -40 - 35 Melis

Ferric Sulphate

45% Concentrate Strength kg Fe/m3 120 Reference

SG at 45% kg/L 1.528 Reference



October 29, 2008


Preliminary Design CriteriaCriteria Unit Value Comment pH of 45% Solution unit 1.5 Reference

Ferric Sulphate Tank Retention Time Truckloads 1.5 Melis

Ferric Sulphate Truckload Volume L 13,000 20 tonnes

Ferric Sulphate Tank Volume m3 19.5 Calculated


Hydrogen Peroxide

Type of System Package Storage and Distribution System Melis

Dosage (kg 70% Soln/t U3O8) kg/t U3O8 280 Melis

Concentration at Addition Point % 50 Melis

SG of 70% kg/L 1.288 Reference

pH of 70% Units 0.50 Reference (DuPont)

SG of 50% kg/L 1.195 Reference

pH of 50% Units 1.80 Reference (DuPont)

Concentration at 70% kg H2O2/m3 902 Reference

Concentration at 50% kg H2O2/m3 597.5 Reference

Percent Addition to Tank No. 1 % 100 Melis


Materials of Construction type Passivated SS Melis

Magnesium Hydroxide

Dosage kg/t U3O8 150 Melis

Concentration % 10 Melis

pH of 10% solution Units 10.5 Reference

SG of Mg(OH)2 kg/L 2.36 Reference

Mg(OH)2/SO4 (w/w) - 0.61 Reference

Overconcentration required ratio 1 Melis

Mg(OH)2/H2O2 (w/w) - 3.43 Reference

H2O/H2O2 (w/w) - 1.59 Reference

SG of MgSO4 SAG 2.66 Reference

MgSO4/Mg(OH)2 ratio 1.24 Reference

MgO/Mg(OH)2 ratio 0.69 Reference

Loop Pump Discharge m3/h 15.0 Melis

Magnesia Distribution Tank Retention Time

days 7.0 Melis

Magnesia Distribution Tank Volume m3 25.6 Calculated

Magnesia Mix Tank Retention Time days 5.0 Melis

Magnesia Mix Tank Volume m3 18.3 Calculated



October 29, 2008


Preliminary Design CriteriaCriteria Unit Value Comment Magnesia Loose Bulk Density SAG 1.00 Estimated

Magnesia Silo Volume m3 40.0 Calculated


Barium Chloride

BaCl2.2H2O Mix Concentration kg/m3 120 Melis

SG at Mix Concentration kg/L 1.09 Calculated

Barium Chloride Mix Tank Volume m3 18.3 As Magnesia Tank

Barium Chloride Distribution Tank Volume

m3 25.6 As Magnesia Tank


Lime

Mix Concentration % CaO 15 Melis

SG of CaO kg/L 3.315 Reference

Mix Concentration kg/m3 168 Calculated

Number of Lime Silos ea 1 Melis

Lime Silo Retention Time days 14

Lime Silo Volume m3 299 Calculated

Lime Mill Discharge Pumpbox Retention Time

min 5 Melis

Lime Mill Discharge Pumpbox Volume

m3 3.7 Calculated

Lime Storage Tank Retention Time h 16.0 Melis

Lime Storage Tank Volume m3 93 Calculated

Lime Loop Feed m3/h 10 Estimated


Flocculant

Mix Concentration for Neutral and Tailings Thickeners

% 0.2 Melis

Mix Concentration for Uranium Thickener and ET

% 0.05 Melis

pH of Mixed Flocculant Units 5.0 Reference

Typical Specific Gravity kg/L 1.2 Typical


Sulphuric Acid

Concentrated Acid Strength % 94.0 Reference

Concentrated Acid Concentration g H2SO4/L 1,700 Reference

Water in 94% Acid g H2O/L 128 Reference

Density of 94% H2SO4 kg/L 1.8276 Reference



October 29, 2008



Density of 100% H2SO4 kg/L 1.841 Reference

Acid Consumption t 100%/d 108 Calculated

Sulphuric Acid Storage Capacity days 14 Strateco

No. of Storage Tanks ea. 2 Melis

Volume of each Storage Tank m3 440 Calculated

Water Distribution

Total RO Permeate Available m3/h 64 Calculated

Total Fresh Water Requirement m3/h 0.0 Calculated

Total Process Water Flowrate m3/h 24.6 Calculated

Total Seal Water Flowrate m3/h 8.0 Calculated

Fresh Water Tank Retention Time h 2 Typical

Fresh Water Tank Volume m3 77 Calculated

Seal Water Flowrate m3/h 8 Estimated

Seal Water Tank Retention Time h 2 Typical

Seal Water Tank Volume m3 16 Calculated

Process Water Flowrate m3/h 25 Calculated

Process Water Tank Retention Time h 2 Typical

Process Water Tank Volume m3 49 Calculated


Steam Distribution

Distributed Steam Pressure kPa 700 Typical

Leach Temperature in, Minimum ºC 25 Melis

Estimated Maximum Steam Requirement for Leaching

tonnes/h 0.8 Calculated

PROCESS PLANT PERSONNEL Page 10-1


October 29, 2008

10.0 PROCESS PLANT PERSONNEL

10.1 Summary

A total of 74 individuals are envisioned to operate the process plant and staff mill maintenance.

The basis of the estimate was:

• Excepting senior staff, there will be two crews, one on site and one off site at any given time,

• Standard schedule will be 12 hours/day including lunch and break time,

• Operating personnel will be divided into two equal sized shifts per crew to allow 24 hour operation,

• Mill Maintenance will be divided into two unequal sized shifts per crew to allow 24 hour operation,

10.2 Mill Operations Personnel

The number and positions of personnel required to operate the process plant (Mill Operations) are detailed in the chart attached in Appendix D as:

• OC-001: Mill Operations Organizational Chart (1 of 2)

• OC-002: Mill Operations Organizational Chart (2 of 2) A total of 45 individuals, organized into two crews and two shifts per crew, plus supervision, are envisioned. A breakdown of these personnel is given below in Table 29.


Mill Operations and Laboratory PersonnelDescription NumberMill Operators 4 x 8.5 = 34Mill Foremen 4 x 1 = 4Mill Trainer 2 x 1 = 2Mill General Foremen 2 x 1 = 2Metallurgists 2 x 1 = 2Laboratory Supervisor 2 x 1 = 2Laboratory Staff 2 x 3 = 6Mill Superintendent 1TOTAL 53

PROCESS PLANT PERSONNEL Page 10-2


October 29, 2008

10.3 Mill Maintenance Personnel

The number and positions of personnel required to maintain the process plant (Mill Maintenance) are detailed in the chart attached in Appendix B as:

• OC-003: Mill Maintenance Organizational Chart A total of 21 individuals, organized into two crews plus supervision, are envisioned. A breakdown of these personnel is given below in Table 30.


Mill Maintenance PersonnelDescription NumberElectricians 2 x 2 = 4Instrument Technologists 2 x 1 = 2Millwrights 2 x 4 = 8Mill Maintenance Foremen 2 x 1 = 2Data Entry Clerks 2 x 1 = 2Mill Maintenance General Foremen 1Mill Maintenance Planner 1Maintenance Superintendent 1TOTAL 21

CAPITAL COST ESTIMATE FOR PROCESS Page 11-1


October 29, 2008

11.0 SUPPORT AND GENERAL ADMINISTRATION PERSONNEL

11.1 Summary

A total of 32 individuals are envisioned to provide support services and administer the site.

11.2 Support Departments and Administration Personnel

The number and positions of personnel required to provide support services and administer the site are detailed in the chart attached in Appendix D as:

• OC-003: Support and Administration Organizational Chart A total of 32 individuals, organized into two crews plus supervision, are envisioned. Support Departments include the laboratory, warehouse, radiation, health and safety, environmental, human resources, process plant office and supervision. A breakdown of these personnel is given below in Table 31.


Support and Administration PersonnelDescription NumberSite Services Supervisor 2 x 1 = 2Site Services Staff 2 x 2 = 4Warehouse Supervisor 2 x 1 = 2Warehouse Staff 2 x 2 = 4Environmental Technicians 2 x 2 = 4Radiation, Health and Safety Staff 2 x 2 = 4Environment, Radiation, Health and Safety Supervisor 1Human Resources Staff 2 x 2 = 4Human Resources Supervisor 1Site Secretary and Flight Clerk 2 x 2 = 4Environment, Radiation, Health and Safety Superintendent 1Mine Manager 1TOTAL 32



October 29, 2008

The basis of the estimate was:

• Standard schedule will be 12 hours/day including lunch and break time,

• All remaining departments will normally work a single shift per day. Departments not included in the support and general administration personnel estimate were:

• Mining,

• Mine Maintenance, and




October 29, 2008

12.0 ESTIMATED PROCESS PLANT AND BUILDINGS CAPITAL COSTS - CLASS IV ESTIMATE

12.1 Summary

The Matoush Scoping Study capital cost estimate is summarized in Table 32 below.

Table 32 Strateco Resources Inc.

Matoush Uranium Project Scoping Study Summary of Class IV Capital Cost Estimate

Cost Area Labour (Hours) 496,200

Labour Cost 49,618,700Material Cost 57,445,370Buildings Cost 39,203,310Reagents, First Fills 2,618,940Total Direct Cost 148,886,320Contractor Overhead and Profit 6,700,000Engineering, Procurement, and Management 18,610,000Total Direct and Indirect Costs 174,196,320Contingency (25%) 43,550,000Capital Spares, 1% of Equipment Cost 1,150,000Total Estimated Capital Costs 218,896,320

Say, 220,000,000Estimated Weight, tonnes 25,910Estimated Operating and Building Power, kW 2,053

The capital cost estimate details are attached in Appendix E.

12.2 Basis of Estimate

Included in the capital cost estimate were:

• the process plant and laboratory building,


• process equipment located in the mill building,

• the laboratory, and

• reagent first fills.



October 29, 2008

The capital cost estimate does not include:

• infrastructure costs such as fuel for construction, road maintenance,

• decommissioning (deconstruction and remediation) costs.

The battery limits for the capital cost estimates were as follows:

• receipt of ore at the crusher,

• receipt of reagents at the minesite,

• discharge of tailings to the Tailings Management Facility,

• discharge of treated effluent, and

• uranium packaging.

The capital costs were completed on an order-of-magnitude basis to the level of detail typical for a Class IV estimate (-15% to -30%/+20% to +50%). The estimate was completed in second quarter 2008 Canadian dollars.

Equipment requirements and approximate sizing were defined by the design criteria and process flowsheets, based on metallurgical data available as of the date of this report.

The cost estimate was prepared with separate sheets for each unit operation associated with each process, plus separate sheets for utilities and the process buildings. Each sheet lists the major pieces of equipment or material in that area. The installed cost of each piece of equipment or material was then estimated with the installation labour, unit cost and cost of field materials being included in the total installed mechanical cost. Line items in the capital cost estimates were costed based on budget quotes or on Melis file data.

For each unit area, the costs of process piping, electrical, instrumentation and freight to site were estimated as a percentage of total installed mechanical cost. The sum of these costs and the total installed mechanical cost were the total direct costs for that unit operation.

Indirect costs included contractor overhead and profit and engineering, procurement and management. Each was estimated as a percentage of the total direct costs for that option. Contractor overhead and profit was estimated at 4.5% of total direct costs, engineering, procurement, construction and management (EPCM) was estimated at 12.5% of direct costs. EPCM includes allowance for



October 29, 2008

commissioning.

The labour rate of $100/h was provided by Strateco Resources Inc. The labour rate does not include indirect costs.

Process equipment was sized based on the design criteria attached in Appendix B and shown on the process flowsheets attached in Appendix A. A 0.5 m freeboard was included for all tanks and pumpboxes.

Freight to site was estimated at $4,000 per 35 tonne truckload. The number of truckloads required was also estimated.

The US dollar conversion rate used was $0.95 Cdn/$ US.

OPERATING COST ESTIMATE FOR PROCESS Page 13-1


October 29, 2008

13.0 CONSUMABLES, PROCESS AND BUILDING ELECTRICAL AND MAINTENANCE CONSUMABLES

13.1 Summary

The Matoush Scoping Study mill operating cost estimate, excluding operating personnel costs is summarized in Table 33 below.

Table 33 Strateco Resources Inc.

Matoush Uranium Project Scoping Study Summary of Mill Operating Cost Estimate - DRAFT

Operating Cost Class $ (Cdn)/a $ (Cdn)/t $ (Cdn)/lb U3O8

Total Consumables 18,553 95.76 9.28

Electrical Power 3,780 19.51 1.89

Maintenance Consumables 1,150 5.94 0.58

Sub-Total 23,483 121.21 11.74

Contingency (15%) 3,522 18.18 1.76

Total 27,005 139.39 13.50

13.2 Basis of Estimate

Included in this operating cost estimate were:

• Mill reagents,

• Mill and laboratory consumables,

• Mill building electrical power, and

• Maintenance consumables.

Personnel cost is not included in the operating cost.

REFERENCES Page 14-1


October 29, 2008

14.0 REFERENCES

1. Melis Letter Status Report No. 1 - For the Period to November 20, 2007.

2. Melis Status Report No. 2 - For the Period November 20, 2007 to January 10, 2008.

3. Melis Status Report No. 3 - For the Period January 11 to April 2, 2008.

4. Melis Status Report No. 4 - For the Period April 3 to April 15, 2008.

5. Melis Status Report No. 5 - For the Period April 16 to June 9, 2008.

6. Melis Status Report No. 6 - For the Period June 10 to August 29, 2008.

7. SGS Lakefield Research Limited - Ore Characterization And Preliminary Grinding Circuit Design Using CEET2® Technology Based On Samples From The Matoush Deposit, April 16, 2008.

8. Terra Mineralogical Services - Mineralogical Characterization of Drill Core Samples From The Matoush Property Northern Quebec, December 3rd, 2007.


October 29, 2008

APPENDIX A

MATOUSH PROCESS FLOWSHEETS

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):

BRUCE C. FIELDER, P.ENG. 07/08/08


ABCD



BYBCF

BCF/KCA

BCFBCF

REVI

SION

S

NO.DATE DESCRIPTION BY DES. LEAD APP..REVI

SION

S


RENC

E

FROM MINE

PROCESS MILL

SCOPING STUDY SUMMARY FLOWSHEET

490-F-100


PROCESS WATER

URANIUM

TREATED EFFLUENT DISCHARGESULPHURIC ACID,

OXYGEN

LIMEFERRIC SULPHATE

FLOCCULANT

HYDROGEN PEROXIDEMAGNESIUM HYDROXIDE

LIME LIME

GRINDING

LEACHING



CALCINING

SAG MILL

BALL MILL

CYCLONES

LEACH TANKS

GYPSUM BELT FILTER

IMPURITY PRECIPITATION TANKS

URANIUM PRECIPITATION TANKS

URANIUM THICKENER

URANIUM CENTRIFUGE

URANIUM CALCINER

BARIUM CHLORIDEFERRIC SULPHATE



REVERSE OSMOSIS

TO PROCESS

FRESH WATER

BARIUM CHLORIDEFERRIC SULPHATELIME, FLOCCULANT

BARREN SOLUTION

SAND FILTERS

SURFACE DRAINAGE

EFFLUENT TREATMENT

SULPHURIC ACID

URANIUM BIN

NEUTRAL THICKENER

FLOCCULANT

TAILINGS THICKENER

TAILINGS TANKS

MONITORING PONDS (3)

REJECTPERMEATE

MINE WATER

TO PROCESS

FRESH WATER TANK

PROCESS WATER TANK


KEY


DILUTION WATER

FRESH ELUATE TANK

RIP TANKS

ELUTION COLUMNS

RESIN SCREEN

LEAN ELUATE TANK

EFFLUENT SAND FILTERS

COARSE ORE BIN

c/w GRIZZLYAND APRON

FEEDER

FINE ORE BIN


(2)

CRUSHER

CRUSHING

PEBBLE CRUSHER

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


FEED FROM MINE

CRUSHER APRON FEEDER

CRUSHING

CRUSHING FLOWSHEET

490-F-101

490-F-120

PROCESS WATER


KEY

1

CRUSHING OVERHEAD CRANE

490-F-102

FINE ORE BIN


(2)

TO GRINDING FEED CONVEYOR

2

JAW CRUSHER

CRUSHING AREA SUMP PUMP

490-F-102

TO BALL MILL CIRCUIT NO. 1 CYCLONE FEED PUMPBOX

490-F-122

CRUSHER AREA SAFETY SHOWER

FROM SAFETY SHOWER HEAD TANK

CRUSHING DUST SCRUBBER

HOSE STATIONS

490-F-102

TO CYCLONE FEED PUMPBOX

TO ATMOSPHERE

COARSE ORE BIN

c/w GRIZZLY

CRUSHER DISCHARGE CONVEYOR

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


NEUTRAL THICKENER OVERFLOW

SAG MILL

BALL MILL

490-F-103

490-F-103

TO NEUTRAL THICKENER

HOSE STATIONS

GRINDING AREA SUMP PUMP

GRINDING

GRINDING FLOWSHEET

490-F-102

490-F-120

PROCESS WATER

4 5 12

8

11

FROM ORE FEED BIN

490-F-101

490-F-101

FROM CRUSHING AREA SUMP PUMP

GRINDING FEED CONVEYOR

490-F-109

FROM BARREN SOLUTION TANK

51

GRINDING AND RESIN TANK AREA

OVERHEAD CRANE

490-F-122

GRINDING AREA SAFETY SHOWERS (2)


SAG MILLFEED CHUTE

13

2

15

16

PEBBLE CRUSHER

PEBBLE CONVEYOR

SCALPING SCREEN

SAG MILL DISCHARGE

SCREEN

CYCLONE FEED PUMPBOX

CYCLONE FEED

PUMPS

CYCLONES

6

10

BALL BUCKET

WEIGHTOMETER

7

3

490-F-117

GRINDING BALLS

490-F-101

FROM CRUSHING DUST SCRUBBER

BELT MAGNET

PROCESS VENTILATION

(TYP.)


KEY

FROM SCALPING SCREEN

SULPHURIC ACID

LEACH DISCHARGE PUMPS

490-F-102

490-F-119

LEACH DISCHARGE PUMPBOX

490-F-104

TO RIP FEED DISTRIBUTION LAUNDER

LEACH TANKS (6)c/w AGITATORS

HOSE STATIONS490-F-120PROCESS WATER

LEACH AREA SUMP PUMP

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


LEACHING

NEUTRAL THICKENING AND LEACHING FLOWSHEET

490-F-103

OXYGEN490-F-116

NEUTRAL THICKENER

MECHANISM

NEUTRAL THICKENER UNDERFLOW PUMPS

NEUTRAL THICKENER OVERFLOW PUMPS

NEUTRAL THICKENER OVERFLOW TANK

FLOCCULANT

490-F-119

490-F-102

11

13

14

17

490-F-121

TO GRINDING490-F-122LEACHING AREA

SAFETY SHOWERS (2)


STEAM145

113

120

PROCESS VENTILATION

(TYP.)MAIN PROCESS FLOWPROCESS FLOWINTERMITTENT / ALTERNATE FLOW

KEY

TITLE

R

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E

MELIS E GI EERI G LTD. RESIN EXTRACTION FLOWSHEET

490-F-104


KEYFROM LEACH DISCHARGE PUMPBOX

490-F-103 17

490-F-120PROCESS WATER

490-F-112

TO TAILINGS NEUTRALIZATION

20

490-F-105

RESIN TO ELUTION

RESIDUE PUMPBOX

RESIDUE PUMPS

LOADED RESIN PUMPS

SAFETY SCREEN

RESIN SCREEN

RIP TANKS (8)c/w PUMPCELLS (8)

(VENDOR PACKAGE)

490-F-105RESIN FROM RESIN TRANSFER TANK

RIP FEED DISTRIBUTION LAUNDER


23

19

22

19

30

490-F-107

GYPSUM SOLIDS

RIP AREA SUMP PUMP

HOSE STATIONS

40

31

136

24

490-F-122

RIP AREA SAFETY SHOWERS (2)


(MANUAL RECYCLE)

PROCESS VENTILATION

(TYP.)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E



RESIN ELUTION FLOWSHEET

490-F-105


KEY

FRESH ELUATE TANKw AGITATOR

FRESH ELUATE PUMPS 490-F-104

490-F-119

CONCENTRATED SULPHURIC ACID

490-F-120

FRESH WATER

490-F-104

RESIN FROM RESIN SCREEN

ELUTED RESIN HOLDING TANK ROTARY VALVE

23

31

ELUTION AREA SUMP PUMP

ELUTION COLUMNS

(3)

HOSE STATIONS

LEAN ELUATE TANK

LEAN ELUATE PUMPS

490-F-106

TO IMPURITY PRECIPITATION TANK NO. 1

CONCENTRATED ELUATE TANK

CONCENTRATED ELUATE PUMPS

29

28

490-F-121

PROCESS AIR

TO RIP TANKS

ELUTED RESIN TRANSFER TANK

27

STEAM

CONDENSATE RETURN

FRESH ELUATE HEAT

EXCHANGER

LEAN ELUATE HEAT EXCHANGER

24

26

25

30

490-F-122

RESIN ELUTION AREA SAFETY

SHOWER


121

125CAUSTIC

TANKc/w

AGITATOR

CAUSTIC PUMP

490-F-116

CAUSTIC SODA

490-F-121

490-F-121

PROCESS VENTILATION

(TYP.)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


FROM CONCENTRATED ELUATE TANK

490-F-105

490-F-120

FRESH WATER

490-F-115

FERRIC SULPHATE FEED

490-F-118LIME LOOP FEED

490-F-118LIME LOOP RETURN

490-F-121

PROCESS AIR


KEY

IMPURITY PRECIPITATION AREA SUMP PUMP

IMPURITY PRECIPITATION TANKS (6)

c/w AGITATORS

IMPURITY PRECIPITATION DISCHARGE PUMPBOX

IMPURITY PRECIPITATION DISCHARGE PUMPS


IMPURITY PRECIPITATION FLOWSHEET

490-F-106

HOSE STATIONS

91

32 33 34 35

36

29

490-F-107FROM GYPSUM FILTER AREA SUMP PUMP

GYPSUM FILTER FEED TANK AGITATOR

GYPSUM FILTER FEED PUMPS

490-F-107

TO GYPSUM BELT FILTER

490-F-107

PREGNANT SOLUTION FILTER BACKWASH

4241 37

490-F-122

IMPURITY PRECIPITATION AREA SAFETY SHOWER


106

PROCESS VENTILATION

(TYP.)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


490-F-104

GYPSUM SOLIDS TO RIP FEED DISTRIBUTION LAUNDER490-F-120

PROCESS WATERHOSE STATIONS

490-F-108

PREGNANT SOLUTION TO URANIUM PRECIPITATION TANK NO. 1

490-F-106

TO GYPSUM FILTER FEED TANK


KEY

GYPSUM FILTER AREA SUMP PUMP

GYPSUM BELT FILTER VACUUM PUMP

490-F-107

GYPSUM PRECIPITATION

GYPSUM FILTRATIONFLOWSHEET

GYPSUM FILTER REPULP TANK

AGITATOR

GYPSUM FILTERREPULP PUMPS

GYPSUM BELT FILTER VACUUM RECEIVERS AND FILTRATE PUMPS

VENDOR PACKAGE: GYPSUM BELT FILTER

BARREN SOLUTION

490-F-109

GYPSUM BELT FILTER

50

39

490-F-106

FROM GYPSUM FILTER FEED TANK

PREGNANT SOLUTION TANK

PREGNANT SOLUTION SAND

FILTER BACKWASH PUMP

PREGNANT SOLUTION PUMPS

43

490-F-106

TO GYPSUM FILTER FEED TANK

PREGNANT SOLUTION SAND

FILTERS (3)

490-F-121PROCESS AIR

38

4241

37

490-F-122GYPSUM FILTER AREA

SAFETY SHOWER


137

40

PROCESS VENTILATION

(TYP.)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


490-F-107

PREGNANT SOLUTION

490-F-115

HYDROGEN PEROXIDE

490-F-117

MAGNESIA LOOP FEED

490-F-121

PROCESS AIR


KEY


DISCHARGE PUMPBOX


DISCHARGE PUMPS

URANIUM PRECIPITATION AREA SUMP PUMP

490-F-109TO URANIUM THICKENER

490-F-108


URANIUM PRECIPITATIONFLOWSHEET

490-F-117

MAGNESIA LOOP RETURN

490-F-109

TO URANIUM THICKENER

URANIUM PRECIPITATION TANKS (3)

c/w AGITATORS

490-F-115

HYDROGEN PEROXIDE

HOSE STATIONS

490-F-120

FRESH WATER

(VENT) (VENT)

490-F-120

WATER RETURN


FEED HEAT EXCHANGERS

43

URANIUM THICKENER UNDERFLOW

490-F-109

126

126

44

490-F-122URANIUM PRECIPITATION AREA SAFETY SHOWER


(VENT)

94

9598

10099

PROCESS VENTILATION

(TYP.)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


FROM URANIUM PACKAGING AREA SUMP PUMP

490-F-110

URANIUM CENTRIFUGE CENTRATE

URANIUM THICKENERAND MECHANISM

URANIUM THICKENER UNDERFLOW PUMPS

URANIUM THICKENER AREA SUMP PUMP

490-F-108

490-F-108

URANIUM PRECIPITATION DISCHARGE

FROM URANIUM PRECIPITATION AREA SUMP PUMP

490-F-111

490-F-120RO PERMEATE

HOSE STATIONS

490-F-110

TO URANIUM WASH TANK


KEY

490-F-109

URANIUM THICKENER FLOWSHEET


TO BALL MILL DISCHARGE PUMPBOX

BARREN SOLUTION TANK

490-F-102

TO SCRUBBERS

490-F-111

490-F-111

URANIUM SCRUBBER WATER

FROM CALCINING AREA SUMP PUMP

490-F-110

490-F-108

RECYCLE TO URANIUM PRECIPITATION

FLOCCULANT

490-F-119

44

53

56

45

46

47

48

51

49

BARREN SOLUTION PUMPS

TO GYPSUM BELT FILTER

490-F-10750

490-F-122URANIUM THICKENER

AREA SAFETY SHOWER


BARREN SOLUTION TANK TRANSFER

PUMP

52

115

PROCESS VENTILATION

(TYP.)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


490-F-110

URANIUM CALCINING

URANIUM CENTRIFUGE AND CALCINER FLOWSHEET

FRESH WATER

HOSE STATIONS

URANIUM THICKENER UNDERFLOW

VENDOR PACKAGE: URANIUM CALCINER

URANIUM WASH TANK

c/w AGITATOR

CENTRIFUGE FEED PUMPS

490-F-120

490-F-109


TO URANIUM CALCINER SCRUBBER

TO URANIUM CALCINING ROOM SCRUBBER

URANIUM CENTRIFUGE

URANIUM CENTRATE PUMPBOX

URANIUM CENTRATE PUMPS


490-F-109

490-F-111

490-F-111

490-F-109

NO. 6 BUNKER OIL

TO PACKAGING

490-F-111


KEY

47

54

53

55

CALCINING AREA SUMP PUMP

BUCKET CONVEYOR

URANIUM CALCINER

URANIUM LUMP DISINTEGRATOR

490-F-122CALCINING AREA SAFETY SHOWER


127

URANIUM BIN c/w ROTARY VALVEand BIN VIBRATOR

55

PROCESS VENTILATION

(TYP.)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


CALCINED URANIUM

490-F-110

FROM URANIUM CALCINER ROOM

490-F-110

FROM URANIUM CALCINER

490-F-120

490-F-109



KEY

URANIUM PACKAGING AREA SUMP PUMP

URANIUM CALCINER SCRUBBER

URANIUM CALCINER ROOM SCRUBBER

490-F-111

URANIUM CALCINING

URANIUM PACKAGING AND SCRUBBERS FLOWSHEET

URANIUM PACKAGING SCRUBBER

FRESH WATER

URANIUM SCRUBBER WATER PUMPS

490-F-109


URANIUM SCRUBBER WATER PUMPBOX

490-F-109

BARREN SOLUTION

TO ATMOSPHERE TO ATMOSPHERE TO ATMOSPHERE

49

55

56

128

490-F-110

TO DRUM STORAGE AND SHIPPING

MOTORIZED ROLLER CONVEYOR

DRUM FILLING STATION

DRUM LIDDING STATION

DRUM WASHING STATION

DRUM SCALE

PRODUCT PACKAGING SYSTEM

55

LIME LOOP490-F-118

490-F-114

TO TAILINGS MANAGEMENT FACILITY

490-F-104

FROM RESIDUE PUMPS

TAILINGS NEUTRALIZATION TANKS (2)

c/w AGITATORS

HOSE STATIONS490-F-120PROCESS WATER

TAILINGS NEUTRALIZATION AREA SUMP PUMP

TAILINGS FEED LAUNDER

TAILINGS THICKENER UNDERFLOW PUMPS

TAILINGS THICKENER OVERFLOW PUMPSTAILINGS

THICKENER c/w MECHANISM

TAILINGS THICKENER OVERFLOW TANK

490-F-114TO REVERSE OSMOSIS

FLOCCULANT

490-F-119

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E



TAILINGS NEUTRALIZATION FLOWSHEET

490-F-112


KEY

FERRIC SULPHATE490-F-115

BARIUM CHLORIDE490-F-118

490-F-120

TO PROCESS WATER TANK

20

91

114

57

58

59

490-F-115

FROM H202 SUMP PUMP

STEAM490-F-121

490-F-113

EFFLUENT TREATMENT CLARIFIER UNDERFLOW

MINE WATER

490-F-122

TAILINGS NEUTRALIZATION AREA

SAFETY SHOWERS


490-F-113

LIME LOOP TO EFFLUENT TREATMENT

81

108

AIR BLOWER

107

143

102

TO PASTE BACKFILL PLANT

TAILINGS LINE BLOWOUT TANK

PROCESS AIR490-F-121 490-F-114

TO TAILINGS MANAGEMENT FACILITY

490-F-114

PROCESS VENTILATION

(TYP.)

TREATED EFFLUENT DISCHARGE

FERRIC SULPHATE490-F-115

BARIUM CHLORIDE490-F-118

LIME LOOP FROM TAILINGS NEUTRALIZATION490-F-112

FLOCCULANT490-F-119

PERMEATE DISCHARGE

490-F-114

EFFLUENT SAND FILTERS (3)

EFFLUENT DISCHARGE

TANK

EFFLUENT SAND FILTER BACKWASH

PUMPMONITORING PONDS (3)c/w DISCHARGE PUMPS

EFFLUENT TREATMENT AREA

SUMP PUMP

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


EFFLUENT TREATMENT

EFFLUENT TREATMENT FLOWSHEET

490-F-113


KEY

REVERSE OSMOSIS REJECT490-F-114

PRIMARY EFFLUENT TREATMENT TANKS (2)

c/w AGITATORS

PRIMARY CLARIFIER

c/w MECHANISM

PRIMARY CLARIFIER UNDERFLOW PUMPS

PRIMARY CLARIFIER OVERFLOW TANK

PRIMARY CLARIFIER OVERFLOW PUMPS

SECONDARY EFFLUENT TREATMENT TANKS (2)

c/w AGITATORS

SECONDARY CLARIFIER

c/w MECHANISM

SECONDARY CLARIFIER UNDERFLOW PUMPS

SECONDARY CLARIFIER OVERFLOW

TANK

SECONDARY CLARIFIER OVERFLOW PUMPS

490-F-112TO TAILINGS THICKENER

71 78

70

6875

6976

72

73

490-F-119SULPHURIC ACID

67 74

64

490-F-121PROCESS AIR

122

116

490-F-122

EFFLUENT TREATMENT AREA SAFETY SHOWER

500-SE-003


490-F-126

LIME LOOP RETURN

77

8180

85 86

8889

85

490-F-114

RECYCLE TO TAILINGS MANAGEMENT FACILITY

POND FILL SAMPLE STATION POND

DISCHARGE SAMPLE STATION

92

490-F-11479

490-F-12082

83

TO PROCESS WATER TANK

TO REVERSE OSMOSIS

87

66

103

PROCESS VENTILATION

(TYP.)

FROM TAILINGS THICKENER UNDERFLOW PUMPS

490-F-112


REVERSE OSMOSIS PLANT,c/w PRETREATMENT

SURFACE DRAINAGE

FROM TAILINGS THICKENER OVERFLOW PUMPS490-F-112

REJECT

PERMEATE490-F-120

TO FRESH WATER TANKRECLAIM BARGE

w PUMPS (2)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


TAILINGS MANAGEMENT

REVERSE OSMOSIS FLOWSHEET

490-F-114


KEY

TO EFFLUENT TREATMENT

490-F-113

TO EFFLUENT DISCHARGE TANK

490-F-113

59

61

58

60

64

62

66

63

FROM SECONDARY CLARIFIER OVERFLOW PUMPS490-F-113 83

65

FROM TAILINGS THICKENER UNDERFLOW PUMPS

490-F-112

SITE RUNOFF

MINE WATER

TO TAILINGS FEED LAUNDER

490-F-112MINE WATER POND c/w OIL/WATER

SEPARATOR MINE WATER PUMP

PROCESS VENTILATION

(TYP.)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


TO IMPURITY PRECIPITATION


KEY

HYDROGEN PEROXIDE

490-F-115

REAGENTS

FERRIC SULPHATE AND HYDROGEN PEROXIDE

FLOWSHEET

490-F-10690

70% H2O2 FROM TRUCK

490-F-120490-F-108

TO URANIUM PRECIPITATION TANK NO. 1

490-F-112

TO TAILINGS NEUTRALIZATION

HYDROGEN PEROXIDE AREA SUMP PUMP

HYDROGEN PEROXIDE DOSING TANK

HYDROGEN PEROXIDE

TRANSFER PUMPS

50% HYDROGEN PEROXIDE STORAGE TANK

FRESH WATER

VENDOR PACKAGE: HYDROGEN PEROXIDE

UNLOADING AND STORAGE490-F-108

TO URANIUM PRECIPITATION TANK NO. 2

HYDROGEN PEROXIDE METERING PUMPS

/242

CONCENTRATED FERRIC SULPHATE FROM TRUCK

490-F-120

PROCESS WATER

490-F-112


FERRIC SULPHATE

FERRIC SULPHATE TANK

FERRIC SULPHATE DISTRIBUTION PUMPS

FERRIC SULPHATE AREA SUMP PUMP

490-F-121STEAM

HOSE STATIONS


490-F-11392

9193

96

97

129

490-F-122 FERRIC SULPHATE AREA SAFETY SHOWER


490-F-122

HYDROGEN PEROXIDE AREA SAFETY SHOWERFROM SAFETY SHOWER

HEAD TANK

94

95

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E



KEY

490-F-116

REAGENTS

OXYGEN PLANT AND CAUSTIC SODA FLOWSHEET

OXYGEN PLANT(2 X 20 TONNES/DAY)

PACKAGE PLANT OWNED AND OPERATED BY AIR LIQUIDE490-F-103

TO LEACHING

CAUSTIC DRUM

490-105TO CAUSTIC TANK

CAUSTIC SODA

OXYGEN PLANT

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E



KEY

490-F-117

REAGENTS

MAGNESIA AND GRINDING BALLS FLOWSHEET

MAGNESIA SILO

MAGNESIA MIX SCREW

MAGNESIA SILO BIN VENT

MAGNESIA SILO BIN ACTIVATOR

MAGNESIA MIX TANKc/w AGITATOR

MAGNESIA TRANSFER

PUMP

MAGNESIA DISTRIBUTION TANKc/w AGITATOR

MAGNESIA DISTRIBUTION PUMPS

490-F-108

490-F-120

FRESH WATER

MAGNESIA LOOP TO URANIUM PRECIPITATION

490-F-108

MAGNESIA LOOP FROM URANIUM PRECIPITATION

MAGNESIA AREA SUMP PUMP

130

VENDOR PACKAGE: MAGNESIA UNLOADING, STORAGE AND MIXING

GRINDING BALLS

100 mm GRINDING BALL

STORAGE

75 mm GRINDING BALL STORAGE

25 mm GRINDING BALL STORAGE

490-105

TO SAG AND BALL MILLS

490-126TO LIME MILL

MAGNESIA

101

100

HOSE STATIONS

490-F-120

PROCESS WATER

490-F-112

LIME

LIME STORAGE TANK AGITATOR

LIME LOOP FEED PUMPS

LIME AREA SUMP PUMPLIME SILO

LIME MIX SCREW

LIME SILO BAGHOUSE

LIME SILO BIN ACTIVATOR

LIME TOWER MILL

LIME MILL DISCHARGE PUMPBOX

LIME MILL DISCHARGE PUMPS

LIME BLOWER

490-F-106IMPURITY PRECIPITATION LIME LOOP RETURN

490-F-106

IMPURITY PRECIPITATION LIME LOOP FEED

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


490-F-118

REAGENTS

BARIUM CHLORIDE AND LIME FLOWSHEET


KEYHOSE STATIONS

SOLID BARIUM CHLORIDE

490-F-120

FRESH WATER

490-F-113


BARIUM CHLORIDE

BARIUM CHLORIDE MIX TANK AGITATOR

BARIUM CHLORIDE TRANSFER PUMP

BARIUM CHLORIDE DISTRIBUTION TANK BARIUM CHLORIDE

DISTRIBUTION PUMPS

BARIUM CHLORIDE AREA SUMP PUMP


490-F-112

131

138

112

490-F-122

BARIUM CHLORIDE AREA SAFETY SHOWER


490-F-122LIME AREA

SAFETY SHOWER


490-F-113

TAILINGS NEUTRALIZATION AND EFFLUENT TREATMENT LIME LOOP RETURN

TAILINGS NEUTRALIZATION AND EFFLUENT TREATMENT LIME LOOP FEED

111

VENDOR PACKAGE: LIME UNLOADING, STORAGE AND

MIXING

490-F-117

GRINDING BALLS

104

107

106

107

102

103

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


490-F-103

SULPHURIC ACID TO LEACHING

SULPHURIC ACID

490-F-119

REAGENTS

FLOCCULANTS AND SULPHURIC ACID FLOWSHEET

490-F-105

SULPHURIC ACID TO FRESH ELATE TANK


KEY

NEUTRAL/TAILINGS THICKENER FLOCCULANT

490-F-120

PROCESS WATER

490-F-113

TO EFFLUENT CLARIFIERS

FLOCCULANT

HOSE STATIONS

URANIUM THICKENER FLOCCULANT

EFFLUENT CLARIFIER FLOCCULANT

490-F-109TO URANIUM THICKENER

TAILINGS THICKENER FLOCCULANT MIX PACKAGE

(PUMPS NOT INCLUDED)

URANIUM THICKENER FLOCCULANT MIX PACKAGE

(PUMPS NOT INCLUDED) EFFLUENT CLARIFIER FLOCCULANT MIX PACKAGE

(PUMPS NOT INCLUDED)

NOTE: ONE OR MORE FLOCCULANT MIX

PACKAGES MAY BE COMBINED AFTER

FLOCCULANT SELECTION490-F-112

TO TAILINGS THICKENER

490-F-103TO NEUTRAL THICKENER

116

118

119

490-F-113

SULPHURIC ACID TO ET

122123

490-F-120FRESH WATER

132

13

SULPHURIC ACID DISTRIBUTION PUMPS

115

114

113

117

121

120SULPHURIC ACID STORAGE TANKS

(2)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


490-F-114RO PERMEATE

490-F-120

UTILITIES

FRESH AND PROCESS WATER DISTRIBUTION FLOWSHEET


KEY

FRESH WATER

FRESH WATER TANK

PROCESS WATER TANK

JOCKEY PUMP

DIESEL FIRE PUMP

ELECTRIC FIRE PUMPS

FRESH WATER DISTRIBUTION

PUMPS135

134

PROCESS WATER PUMPS

140

490-F-110TO URANIUM WASH TANK

490-F-111TO SCRUBBERS

490-F-115

TO MAGNESIA MIX TANK490-F-117

TO BARIUM CHLORIDE MIX TANK

490-F-102TO GRINDING

490-F-107

TO GYPSUM FILTER REPULP TANK

490-F-118TO LIME MILL

490-F-119

138

130

131

490-F-108COOLING WATER RETURN

490-F-108

TO URANIUM PRECIPITATION FEED HEAT EXCHANGERS

490-F-112TAILING THICKENER OVERFLOW

490-F-118

TO STEAM GENERATION

126

126

SEAL WATER TANK

SEAL WATER PUMPS

TO SEAL WATER DISTRIBUTION

133

TO FIRE LOOP

490-F-113

SECONDARY CLARIFIER OVERFLOW

FIRE LOOP RETURN

TO 50% HYDROGEN PEROXIDE STORAGE TANK

129

490-F-106TO IMPURITY PRECIPITATION

490-F-109TO URANIUM THICKENER UNDERFLOW PUMPS

490-F-104

TO RIP CIRCUIT490-F-112TO TAILINGS HOSE STATIONS

490-F-115TO FERRIC SULPHATE HOSE STATIONS

490-F-119TO FLOCCULANT MIXING

490-F-121

TO FLOCCULANT MIXING

132

139

490-F-105TO FRESH ELUATE TANK124

136

125

127

128

137

141MINE WATER

490-F-112TO CRUSHING HOSE STATIONS

TO SPRINKLER SYSTEM

65

142

FROM LAKE

PROCESS VENTILATION

(TYP.)

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


PROCESS AND INSTRUMENT AIR

STEAM BOILER

TO STEAM DISTRIBUTION

490-F-120FRESH WATER

STEAM BOILER

490-F-121

UTILITIES

AIR AND STEAM DISTRIBUTION FLOWSHEET


KEY

490-F-103

TO LEACHING

490-F-106TO GYPSUM FILTER FEED TANK

490-F-112TO FERRIC SULPHATE UNLOADING

DIESEL OR FUEL OIL

TO HEAT RECOVERYOR ATMOSPHERE

145

STEAM BOILER PACKAGE

ALL

TO PROCESS AIR DISTRIBUTION

PROCESS AIR COMPRESSORS (2) PROCESS AIR

RECEIVER

ALL

TO INSTRUMENT AIR DISTRIBUTION

INSTRUMENT AIR RECEIVER

AIR DRYER/DE-OILER

490-F-105TO RESIN ELUTION

490-F-105CONDENSATE RETURN

CONDENSATE RETURN

ALL

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):

BRUCE C. FIELDER, P.ENG.


ABCD



BY

BCF/KCA

BCFBCFBCF

REVI

SION

S


SION

S


RENC

E


490-F-122

UTILITIES

SAFETY SHOWER SYSTEM FLOWSHEET


KEY

SAFETY SHOWER PUMPS

SAFETY SHOWER SURGE TANK

POTABLE WATER SYSTEM

FRESH WATERPOTABLE WATER DISTRIBUTION

SAFETY SHOWER HEAD TANK

WATER HEATER

SAFETY SHOWER LOOP RETURN

SAFETY SHOWER LOOP FEED

SEWAGE TREATMENT SYSTEM

POTABLE WATER DISTRIBUTION

GREY AND BLACK WATER


October 29, 2008

APPENDIX B

MATOUSH SIMPLIFIED PLANT LAYOUT

84m.

84m

.

\

GRINDING CIRCUIT

LEACHING CIRCUIT


CIRCUIT

IMPURITY PRECIPITATION AND GYPSUM FILTRATION CIRCUIT

SULPHURIC ACID STORAGE TANKS

RIP CIRCUIT

CALCINING CIRCUIT

URANIUM DRUM STORAGE

REAGENT PREPARATION AND STORAGE


CIRCUITFRESH AND PROCESS

WATER

REVERSE OSMOSIS

EFFLUENT TREATMENT

COMPUTER ROOM, MCC ROOM MILL DRY, MILL MECHANICAL SHOP

AND LABORATORY

OXYGEN PLANT

CRUSHING CIRCUIT

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



AB


SCOPING STUDY RELEASESCOPING STUDY RELEASE

BYBCFBCF

REVI

SION

S


SION

S


RENC

E


PROCESS PLANT

MATOUSH PLANT SIMPLIFIED LAYOUT

490-L-100

EDGE OF PAVED MILL TERRACE


CIRCUIT


October 29, 2008

APPENDIX C

MATOUSH MASS BALANCE

Melis Engineering Ltd.Project No. 490October 29, 2008

STRATECO RESOURCES INC. - MATOUSH URANIUM PROJECTMASS BALANCE

Stream Stream Slurry Solids Solution Solvent Sulphuric Acid Total pHNo. % (w/w) SG t/h m3/h t/h % U3O8 kg U3O8/h SG t/h m3/h g U3O8/L kg U3O8/h m3/h g H2SO4/L kg H2SO4/h kg U3O8/h

Crushing1 Feed to Crusher 95.0 2.75 51.0 18.6 48.5 0.480 233 1.00 2.55 2.55 0.000 0.000 2.55 0.000 0.000 233 7.0

SAG Mill Grinding2 From Ore Concentrate Bin 95.0 2.75 25.5 9.3 24.2 0.480 116 1.00 1.28 1.28 0.000 0.000 1.28 0.000 0.000 116 7.07 Pebble Crusher Discharge 75.0 1.91 4.85 2.5 3.64 0.480 17.4 1.00 1.21 1.21 0.000 0.000 1.21 0.000 0.000 17.4 7.03 SAG Mill Feed Conveyor Discharge 91.8 2.41 30.4 11.8 27.9 0.480 134 1.00 2.49 2.49 0.000 0.000 2.49 0.000 0.000 134 7.0

15 Thickener Overflow to SAG Mill 0.020 1.00 12.5 13.3 0.003 0.480 0.012 1.00 12.5 12.5 0.000 0.000 12.5 0.000 0.000 0.012 7.04 SAG Mill Discharge 65.0 1.71 42.9 25.2 27.9 0.480 134 1.00 15.0 15.0 0.000 0.000 15.0 0.000 0.000 134 7.05 SAG Mill Discharge Screen Undersize 63.7 1.68 38.0 22.6 24.2 0.480 116 1.00 13.8 13.8 0.000 0.000 13.8 0.000 0.000 116 7.06 SAG Mill Discharge Screen Oversize 75.0 1.91 4.85 2.5 3.64 0.480 17.4 1.00 1.21 1.21 0.000 0.000 1.21 0.000 0.000 17.4 7.0

Pebble Crushing7 Pebble Crusher Discharge 75.0 1.91 4.85 2.5 3.64 0.480 17.4 1.00 1.21 1.21 0.000 0.000 1.21 0.000 0.000 17.4 7.0

Ball Mil Grinding5 SAG Mill Discharge Screen Undersize 63.7 1.68 38.0 22.6 24.2 0.480 116 1.00 13.8 13.8 0.000 0.000 13.8 0.000 0.000 116 7.0

12 Cyclone Underflow 57.4 1.58 169 107 97.0 0.480 465 1.00 72.0 72.0 0.000 0.000 72.0 0.000 0.000 465 7.08 Ball Mill Discharge 58.6 1.59 207 130 121 0.480 582 1.00 85.8 85.8 0.000 0.000 85.8 0.000 0.000 582 7.0

Ball Mill Discharge Pumpbox and Grinding Cyclones8 Ball Mill Discharge 58.6 1.59 207 130 121 0.480 582 1.00 85.8 85.8 0.000 0.000 85.8 0.000 0.000 582 7.0

16 Thickener Overflow to Ball Mill Discharge Pumpbox 0.020 1.00 13.2 13 0.003 0.480 0.013 1.00 13.2 13.2 0.000 0.000 13.2 0.000 0.000 0.013 7.051 Barren Solution Recycle to Ball Mill Discharge Pumpbox 0.002 1.00 15.9 16.0 0.002 79.000 0.25 1.00 15.9 15.9 0.007 0.103 15.8 0.002 0.082 0.36 4.49 Process Water to Ball Mil Discharge Box 1.00 6.3 6.3 1.00 6.3 6.3

10 Grinding Cyclone Feed 50.0 1.47 242 165 121 0.480 582 1.00 121 121 0.000 0.000 121 0.000 0.000 582 7.011 Grinding Cyclone Overflow 33.0 1.27 73.5 58 24.2 0.480 116 1.00 49.2 49.2 0.000 0.000 49.2 0.000 0.000 116 7.012 Grinding Cyclone Underflow 57.4 1.58 169 107 97.0 0.480 465 1.00 72.0 72.0 0.000 0.000 72.0 0.000 0.000 465 7.0

Neutral Thickener11 Grinding Cyclone Overflow 33.0 1.27 73.5 58 24.2 0.480 116 1.00 49.2 49.2 0.000 0.000 49.2 0.000 0.000 116 7.0113 Flocculant Addition to Neutral Thickener 1.00 0.73 0.73 0.000 0.000 0.73 0.000 0.000 0.000 5.013 Thickener Overflow 0.020 1.00 25.7 25.7 0.005 0.480 0.025 1.00 25.7 25.7 0.000 0.000 25.7 0.000 0.000 0.025 7.014 Thickener Underflow 50.0 1.47 48.5 33.1 24.2 0.480 116 1.00 24.2 24.2 0.000 0.000 24.2 0.000 0.000 116 7.015 Thickener Overflow to SAG Mill 0.020 1.00 12.5 12.5 0.003 0.480 0.012 1.00 12.5 12.5 0.000 0.000 12.5 0.000 0.000 0.012 7.016 Thickener Overflow to Ball Mill Discharge Pumpbox 0.020 1.00 13.2 13.2 0.003 0.480 0.013 1.00 13.2 13.2 0.000 0.000 13.2 0.000 0.000 0.013 7.0

Leaching14 Thickener Underflow 50.0 1.47 48.5 33.1 24.2 0.481 117 1.00 24.2 24.2 0.000 0.000 24.2 0.000 0.000 117 7.0120 93\% Sulphuric Acid Addition to Leaching 1.83 2.92 1.60 0.000 0.000 0.20 1,700 2,715 0.000 -145 Steam Addition to Leaching 0.00 0.75 204.4 0.0037 0.75 204 0.7517 Leach Discharge 44.2 1.49 52.2 33.8 23.0 0.010 2.31 1.10 29.1 26.6 4.30 114 25.2 30.0 798 117 0.2

Resin Process: Resin Loading17 Leach Discharge 44.2 1.49 52.2 33.8 23.0 0.010 2.31 1.10 29.1 26.6 4.30 114 25.2 30.0 798 117 0.240 Gypsum Solids to RIP 25.0 1.17 6.50 5.58 1.63 0.050 0.81 1.00 4.88 4.87 1.34 6.53 4.87 0.002 0.010 7.35 4.430 Average Resin Exiting Elution 80.0 1.15 3.42 2.96 2.74 0.167 4.56 1.00 0.68 0.68 0.68 4.5618 Feed to RIP 44.1 1.47 62.1 42 27.4 0.028 7.69 1.07 38.3 35.8 3.37 121 34.4 25.4 798 129 0.319 Average Wash Water to Safety Screen 1.00 1.00 3.65 3.65 3.6520 Residue to Tailings Neutralization 44.1 1.48 62.1 42 27.4 0.010 2.75 1.08 42.0 38.8 0.009 0.35 38.1 20.2 798 3.10 0.421 Transfer to Resin Screen 44.1 1.48 214 146 94.4 1.08 145 134 131 20.2 2,702 0.000 0.422 Average Wash Water to Resin Screen 1.00 1.00 3.65 3.65 3.6523 Average Resin To Elution 80.0 2.59 3.42 3.00 2.74 4.58 125 1.00 0.68 0.68 0.68 125 7.024 Instantaneous Resin To Elution 80.0 2.59 33.3 29.2 26.6 4.58 1,221 1.00 6.66 6.66 6.66 1,221 7.0

Resin Process: Resin Elution121 Average 93% Sulphuric Acid to Fresh Eluate Tank 1.83 1.84 1.01 0.000 0.000 0.128 1,700 171025 Instantaneous 93% Sulphuric Acid to Fresh Eluate Tank 1.83 4.84 2.65 0.000 0.000 0.34 1,700 4,500125 Average Fresh Water to Fresh Eluate Tank 1.00 6.98 6.98 6.9826 Instantaneous Fresh Water to Fresh Eluate Tank 1.00 28.1 28.1 28.127 Fresh Eluate Solution (Batch) 1.10 32.9 30.0 1.10 32.9 30.0 0.000 0.000 28.4 150 4,500 0.000 -0.5

Page 1 of 5




28 Lean Eluate Solution (Batch) 1.10 32.9 30.0 1.10 32.9 30.0 5.00 150 28.4 142.2 4,267 150 -0.529 Concentrated Eluate Solution 1.09 8.82 8.1 1.09 8.82 8.06 15.0 121 7.68 126.7 1,021 121 -0.430 Average Resin Exiting Elution 80.0 1.15 3.42 2.96 2.74 0.17 4.56 1.00 0.68 0.68 0.68 4.56 7.031 Instantaneous Resin Exiting Elution 80.0 2.59 33.7 29.2 27.0 0.17 44.9 1.00 6.74 6.74 6.74 44.9 7.0

Impurity Precipitation29 Concentrated Eluate Solution 1.09 8.82 8.06 1.09 8.82 8.06 15.1 121 7.68 126.7 1,021 121 -0.490 Ferric Sulphate Addition 0.003 0.002 1.53 0.003 0.002 0.000 0.000 0.002 29.7 0.050 0.000 1.532 Lime Addition to Tank No. 1 15.0 1.12 1.94 1.74 0.29 1.12 1.94 1.74 0.000 0.000 1.65 0.000 0.000 0.000 13.233 Lime Addition to Tank No. 2 15.0 1.12 0.97 0.87 0.146 1.12 0.97 0.87 0.000 0.000 0.83 0.000 0.000 0.000 13.234 Lime Addition to Tank No. 3 15.0 1.12 0.58 0.52 0.087 1.12 0.58 0.52 0.000 0.000 0.50 0.000 0.000 0.000 13.235 Lime Addition to Tank No. 4 15.0 1.12 0.39 0.35 0.058 1.12 0.39 0.35 0.000 0.000 0.33 0.000 0.000 0.000 13.236 Impurity Precipitation Discharge 12.79 1.08 12.7 11.8 1.63 0.010 0.163 1.00 11.09 11.05 11.0 121 10.97 0.030 0.33 121 3.2

Gypsum Filter Feed Tank36 Impurity Precipitation Discharge 12.79 1.08 12.7 11.8 1.63 0.010 0.134 1.00 11.09 11.05 11.0 121 10.97 0.030 0.33 121 3.242 Average Pregnant Solution Filter Backwash 2.08 1.02 0.67 0.67 0.014 0.010 0.001 1.00 0.66 0.67 8.5 5.67 0.67 0.025 0.017 5.67 3.337 Gypsum Belt Filter Feed 12.25 1.07 13.4 12.5 1.64 0.008 0.14 1.00 11.7 11.7 10.8 127 11.6 0.030 0.35 127 3.2

Gypsum Belt Filter37 Gypsum Belt Filter Feed 12.25 1.07 13.4 12.5 1.64 0.008 0.136 1.00 11.7 11.7 10.8 127 11.6 0.030 0.35 127 3.250 Barren Solution to Gypsum Belt Filter 0.002 1.00 4.88 4.87 0.000 79.0 0.077 1.00 4.88 4.87 0.007 0.032 4.87 0.002 0.010 0.109 4.438 Belt Filter Filtrate 0.10 1.00 14.2 14.2 0.014 0.010 0.001 1.00 14.2 14.2 8.5 120 14.1 0.025 0.35 120 3.339 Belt Filter Cake 40.0 1.29 4.06 3.14 1.63 0.050 0.81 1.00 2.44 2.44 2.68 6.53 2.43 0.004 0.010 7.35 4.1137 Process Water for Gypsum Cake Repulp 1.00 1.00 2.44 2.44 0.000 0.000 2.44 0.000 0.000 0.000 7.040 Gypsum Solids to RIP 25.0 1.17 6.50 5.58 1.63 0.050 0.81 1.00 4.88 4.87 1.34 6.53 4.87 0.002 0.010 7.35 4.4

Pregnant Solution Sand Filters38 Belt Filter Filtrate 0.10 1.00 14.2 14.2 0.014 0.010 0.001 1.00 14.2 14.2 8.5 120 14.1 0.025 0.35 120 3.341 Instantaneous Pregnant Solution Filter Backwash 0.002 1.00 10.7 10.7 0.000 0.010 0.000 1.00 10.7 10.7 8.5 91 10.6 0.025 0.27 91 3.342 Average Pregnant Solution Filter Backwash 2.08 1.02 0.67 0.67 0.014 0.010 0.001 1.00 0.66 0.67 8.5 5.7 0.67 0.025 0.02 5.67 3.343 Pregnant Solution Tank Discharge 0.002 1.00 13.5 13.5 0.000 0.010 0.000 1.00 13.5 13.5 8.5 114 13.4 0.025 0.33 114 3.3

Uranium Precipitation43 Pregnant Solution Tank Discharge 0.002 1.00 13.5 13.5 0.000 0.010 0.000 1.00 13.5 13.5 8.5 114 13.4 0.025 0.33 114 3.394 50% H2O2 Addition To Tank No. 1 1.20 1.20 0.045 0.037 0.000 0.000 0.022 0.000 0.000 0.000 1.898 Mg(OH)2 Addition To Tank No. 1 1.06 1.06 0.16 0.15 0.000 0.000 0.15 0.000 0.000 0.000 10.595 50% H2O2 Addition To Tank No. 2 1.20 1.20 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 1.899 Mg(OH)2 Addition To Tank No. 2 1.06 1.06 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 10.544 Uranium Precipitation Discharge 1.05 1.01 13.7 13.6 0.14 78.9 114 1.00 13.6 13.6 0.010 0.14 13.6 0.003 0.041 114 4.2

Uranium Thickener44 Uranium Precipitation Discharge 1.05 1.01 13.7 13.6 0.14 79.0 114 1.00 13.6 13.6 0.010 0.136 13.6 0.003 0.041 114 4.253 Uranium Centrifuge Centrate 0.10 1.00 0.46 0.46 0.000 79.0 0.36 1.00 0.46 0.46 0.003 0.001 0.46 0.001 0.000 0.36 4.756 Scrubber Water To Uranium Thickener 0.002 1.00 27.8 27.8 0.000 94.8 0.44 1.00 27.8 27.7 0.005 0.14 27.6 0.001 0.041 0.58 4.552 Average Transfer to Uranium Thickener 35.0 1.44 0.017 0.01 0.006 79.0 5.92 1.00 0.011 0.011 0.007 0.000 0.011 0.002 0.000 5.92 4.4115 Flocculant Addition to Uranium Thickener 1.00 1.00 0.044 0.044 0.000 0.000 0.044 0.000 0.000 0.000 5.045 Uranium Thickener Feed 0.36 1.00 42.0 41.9 0.15 79.0 121 1.00 41.9 41.8 0.007 0.27 41.7 0.002 0.082 121 4.446 Uranium Thickener Overflow 0.020 1.00 41.6 41.6 0.008 79.0 6.58 1.00 41.6 41.5 0.007 0.27 41.4 0.002 0.082 6.85 4.447 Uranium Thickener Underflow 35.0 1.44 0.41 0.25 0.14 79.0 114 1.00 0.27 0.27 0.007 0.002 0.27 0.002 0.001 114 4.4

Barren Solution48 Barren Solution Tank Discharge 0.002 1.00 41.6 41.6 0.008 79.0 0.66 1.00 41.6 41.5 0.007 0.27 41.4 0.002 0.082 0.93 4.449 Barren Solution to Scrubbers 0.002 1.00 20.8 20.8 0.000 79.0 0.33 1.00 20.8 20.8 0.007 0.14 20.8 0.002 0.041 0.46 4.450 Barren Solution to Gypsum Belt Filter 0.002 1.00 4.88 4.87 0.000 79.0 0.077 1.00 4.88 4.87 0.007 0.032 4.87 0.002 0.010 0.109 4.4

51 Barren Solution Recycle to Ball Mill Discharge Pumpbox 0.002 1.00 15.9 16.0 0.002 79.0 0.25 1.00 15.9 15.9 0.007 0.103 15.8 0.002 0.082 0.36 4.4

52 Average Transfer to Uranium Thickener 35.0 1.44 0.017 0.012 0.006 79.0 5.92 1.00 0.011 0.011 0.007 0.000 0.011 0.002 0.000 5.92 4.4

Uranium Wash Tank and Centrifuge

Page 2 of 5




47 Uranium Thickener Underflow 35.0 1.44 0.41 0.25 0.14 79.0 114 1.00 0.27 0.27 0.007 0.002 0.27 0.002 0.001 114 4.4127 Fresh Water to Uranium Wash Tank 1.00 0.33 0.33 1.00 0.33 0.33 0.000 0.000 0.33 0.000 0.000 0.000 7.053 Uranium Centrifuge Centrate 0.10 1.00 0.46 0.46 0.000 79.0 0.36 1.00 0.46 0.46 0.003 0.001 0.46 0.001 0.000 0.36 4.754 Uranium Centrifuge Discharge 50.0 1.78 0.29 0.16 0.14 79.0 114 1.00 0.14 0.14 0.003 0.000 0.14 0.001 0.000 114 4.7

Uranium Calciner and Scrubbers55 Uranium Calciner Discharge 100 0.11 0.11 100 114 11449 Barren Solution to Scrubbers 0.002 1.00 20.8 20.76 0.000 79.0 0.33 1.00 20.8 20.8 0.007 0.14 20.8 0.002 0.041 0.46 4.4128 Fresh Water to Scrubbers 1.00 6.84 6.84 1.00 6.84 6.84 0.000 0.000 6.84 0.000 0.000 0.000 7.056 Scrubber Water To Uranium Thickener 0.002 1.00 27.8 27.77 0.000 94.8 0.44 1.00 27.8 27.7 0.005 0.14 27.6 0.001 0.041 0.58 4.5

Tailings Neutralization20 Residue to Tailings Neutralization 44.1 1.48 62.1 42 27.4 0.010 2.75 1.08 42.0 38.8 0.009 0.35 38.1 20.2 798 3.10 0.4143 Mine Water to Tailings Neutralization 50.0 50.0 50.0 50.081 Combined ET Clarifier Underflows 18.0 1.12 0.31 0.3 0.056 0.011 0.006 1.00 0.26 0.26 0.000 0.000 0.26 0.013 0.003 0.006 3.6108 Lime to Tailings 15.0 1.12 2.48 2.22 0.97 1.12 2.48 2.22 0.000 0.000 2.11 0.000 0.000 0.000 13.291 Ferric Sulphate to Tailings 0.020 0.013 1.53 0.020 0.013 0.000 0.000 0.018 29.7 0.38 0.000 1.5102 Barium Chloride to Tailings 0.012 0.011 1.09 0.012 0.011 0.000 0.000 0.011 0.000 0.000 0.000 7.057 Tailings Thickener Feed 23.5 1.31 124 95 29.1 0.011 3.11 1.00 94.8 94.8 0.000 0.000 94.8 0.000 0.000 3.11 10.0

Tailings Thickener57 Tailings Thickener Feed 23.5 1.31 124 95 29.1 0.011 3.11 1.00 94.8 94.8 0.000 0.000 94.8 0.000 0.000 3.11 10.0114 Flocculant to Tailings Thickener 1.00 1.00 0.87 0.87 0.000 0.000 0.87 0.000 0.000 0.000 5.058 Tailings Thickener Overflow 0.050 1.00 71.9 71.9 0.036 0.011 0.004 1.00 71.9 71.9 0.000 0.000 71.9 0.000 0.000 0.004 10.059 Tailings Thickener Underflow 55.0 1.47 52.8 35.9 29.1 0.011 3.10 1.00 23.8 23.8 0.000 0.000 23.8 0.000 0.000 3.10 10.0

Tailings Management Facility59 Tailings Thickener Underflow 55.0 1.47 52.8 36 29.1 0.011 3.10 1.00 23.8 23.8 0.000 0.000 23.8 0.000 0.000 3.10 10.060 Surface Drainage to TMF 0.0 0.0 1.00 0.0 0.0 0.000 0.000 0.0 0.000 0.000 0.00 7.061 Supernatant to Reverse Osmosis 0.010 1.00 3.6 3.6 0.000 0.011 0.0000 1.00 3.6 3.6 0.000 0.000 3.6 0.000 0.000 0.00 10.062 Consolidated Tailings 59.0 1.53 49.2 32 29.1 0.011 3.10 1.00 20.2 20.2 0.000 0.000 20.2 0.000 0.000 3.10 10.0

Reverse Osmosis61 Supernatant to Reverse Osmosis 0.010 1.00 3.6 3.6 0.000 0.011 0.0000 1.00 3.6 3.6 0.000 0.000 3.6 0.000 0.000 0.000 10.058 Tailings Thickener Overflow 0.050 1.00 71.9 71.9 0.036 0.011 0.004 1.00 71.9 71.9 0.000 0.000 71.9 0.000 0.000 0.004 10.083 Secondary Clarifier Overflow To Reverse Osmosis 0.005 1.00 0.000 0.0 0.000 0.011 0.000 1.00 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 7.563 RO Permeate 64.1 64.1 1.00 64.1 64.1 0.000 0.000 64 0.000 0.000 0.000 7.064 RO Reject 0.32 1.00 11.4 11.3 0.036 0.011 0.004 1.00 11.3 11.3 0.000 0.000 11.3 0.000 0.000 0.004 10.065 RO Permeate To Fresh Water Tank 1.00 38.3 38.3 1.00 38.3 38.3 0.000 0.000 38.34 0.000 0.000 0.004 7.066 Excess RO Permeate Discharge 1.00 25.8 25.8 1.00 25.8 25.8 0.000 0.000 25.8 0.000 0.000 0.004 7.0

Effluent Treatment, Primary Treatment64 RO Reject 0.32 1.00 11.4 11.3 0.036 0.011 0.004 1.00 11.3 11.3 0.000 0.000 11.3 0.000 0.000 0.004 10.067 93% Sulphuric Acid Addition to Primary ET 0.001 0.000 1.83 0.001 0.000 0.000 0.000 0.000 1,700 0.79 0.000 -68 Ferric Sulphate Addition to Primary ET 0.002 0.001 1.53 0.002 0.001 0.000 0.000 0.002 29.7 0.039 0.000 1.569 Barium Chloride Addition to Primary ET 0.004 0.003 1.09 0.004 0.003 0.000 0.000 0.003 0.000 0.000 0.000 7.070 Lime Addition to Primary ET 15.0 1.12 0.000 0.000 0.000 1.12 0.000 0.000 0.000 0.000 0.000 0.000 0.000 0.000 13.271 Flocculant to Primary ET 1.00 0.16 0.16 0.16 0.16 0.000 0.000 0.16 0.000 0.000 0.000 5.072 Primary Clarifier Overflow 0.005 1.00 11.2 11.2 0.001 0.011 0.000 1.00 11.2 11.2 0.000 0.000 11.2 0.014 0.16 0.000 3.573 Primary Clarifier Underflow 18.0 1.12 0.29 0.26 0.052 0.011 0.006 1.00 0.24 0.24 0.000 0.000 0.24 0.014 0.003 0.006 3.5

Effluent Treatment, Secondary Treatment72 Primary Clarifier Overflow 0.005 1.00 11.2 11.2 0.0006 0.011 0.000 1.00 11.2 11.2 0.0000 0.0000 11.2 0.014 0.157 0.000 3.586 Average Backwash Discharge 0.147 1.00 0.00 0.00 0.0000 0.000 0.000 1.00 0.00 0.00 0.0000 0.0000 0.00 0.000 0.000 0.000 7.574 93% Sulphuric Acid Addition to Secondary ET 0.000 0.000 1.83 0.000 0.000 0.0000 0.0000 0.000 1,700 0.000 0.000 -75 Ferric Sulphate Addition to Secondary ET 0.002 0.000 1.53 0.002 0.000 0.0000 0.0000 0.002 29.7 0.000 0.000 1.576 Barium Chloride Addition to Secondary ET 0.004 0.003 1.09 0.004 0.003 0.0000 0.0000 0.003 0.000 0.000 0.000 7.077 Lime Addition to Secondary ET 15.0 1.12 0.12 0.11 0.005 1.12 0.12 0.11 0.0000 0.0000 0.10 0.000 0.000 0.000 13.278 Flocculant to Secondary ET 1.00 0.16 0.16 0.0000 0.0000 0.16 0.000 0.000 0.000 5.079 Secondary Clarifier Overflow 0.005 1.00 11.5 11.5 0.001 0.011 0.0001 1.00 11.5 11.5 0.0000 0.0000 11.5 0.000 0.000 0.000 7.580 Secondary Clarifier Underflow 18.0 1.12 0.021 0.019 0.004 0.011 0.000 1.00 0.017 0.017 0.0000 0.0000 0.017 0.000 0.000 0.000 7.5

Page 3 of 5




81 Combined ET Clarifier Underflows 18.0 1.12 0.31 0.28 0.056 0.011 0.006 1.00 0.26 0.26 0.0000 0.0000 0.26 0.013 0.003 0.006 3.682 Secondary Clarifier Overflow To Process Water Tank 0.005 1.00 11.5 11.5 0.001 0.011 0.000 1.00 11.5 11.5 0.0000 0.0000 11.5 0.000 0.000 0.000 7.583 Secondary Clarifier Overflow To Reverse Osmosis 0.005 1.00 0.000 0.000 0.000 0.011 0.000 1.00 0.000 0.000 0.000 0.0000 0.000 0.000 0.000 0.000 7.584 Secondary Clarifier Overflow Discharge 0.005 1.00 0.00 0.00 0.000 0.011 0.000 1.00 0.00 0.00 0.000 0.0000 0.00 0.000 0.000 0.000 7.5

Effluent Sand Filters84 Secondary Clarifier Overflow Discharge 0.005 1.00 0.00 0.00 0.000 0.011 0.000 1.00 0.00 0.00 0.000 0.000 0.00 0.000 0.000 0.000 7.585 Instantaneous Backwash Flow 1.00 0.0 0.0 0.000 0.000 0.0 0.000 0.000 0.000 7.586 Average Backwash Discharge 0.147 1.00 0.00 0.00 0.000 0.000 0.000 1.00 0.00 0.00 0.000 0.000 0.00 0.000 0.000 0.000 7.587 Average Sand Filter Discharge 0.001 1.00 0.00 0.00 0.000 0.000 0.000 1.00 0.00 0.00 0.000 0.000 0.00 0.000 0.000 0.000 7.5

Monitoring Ponds87 Average Sand Filter Discharge 0.001 1.00 0.00 0.00 0.000 0.000 0.000 1.00 0.00 0.00 0.000 0.000 0.00 0.000 0.000 0.000 7.566 Excess RO Permeate Discharge 0.000 1.00 25.8 25.8 1.00 25.8 25.8 0.000 0.000 25.8 0.000 0.000 0.004 7.088 Feed to Monitoring Ponds 0.000 1.00 25.8 25.8 0.000 0.000 0.000 1.00 25.8 25.8 0.000 0.000 25.8 0.000 0.000 0.004 7.589 Discharge From Monitoring Ponds 0.000 1.00 110 110 0.000 0.000 0.000 1.00 110 110 0.000 0.000 110 0.000 0.000 0.004 7.5

Ferric Sulphate90 Ferric Sulphate to Impurity Precipitation 1.53 0.003 0.002 0.002 1.591 Ferric Sulphate to Tailings 1.53 0.020 0.0129 0.018 1.592 Ferric Sulphate Addition to ET 1.53 0.004 0.001 0.004 1.593 Total Ferric Sulphate Addition 1.53 0.026 0.016 0.024 1.5

Hydrogen Peroxide94 50% H2O2 Addition To Tank No. 1 1.20 1.20 0.045 0.037 0.022 1.895 50% H2O2 Addition To Tank No. 2 1.20 1.20 0.000 0.000 0.000 1.896 Total 50% Hydrogen Peroxide Consumption 1.20 1.20 0.045 0.037 0.022 1.897 Equivalent 70% H2O2 1.29 1.29 0.032 0.025 0.010 0.5129 Fresh Water to H2O2 Dilution 1.00 0.013 0.013 0.013 7.0

Magnesium Hydroxide98 Mg(OH)2 Addition To Tank No. 1 1.06 1.06 0.16 0.15 0.15 10.599 Mg(OH)2 Addition To Tank No. 2 1.06 1.06 0.000 0.000 0.000 10.5100 Mg(OH)2 Loop Feed 1.06 1.06 15.9 15.0 0.15130 Fresh Water to MgO Mixing 1.00 1.00 0.15 0.15101 Equivalent Solid MgO 2.36 0.012

Barium Chloride102 Barium Chloride to Tailings 0.012 0.011 1.09 0.012 0.011 0.011 7.0103 Barium Chloride Addition to ET 0.004 0.003 1.09 0.004 0.003 0.003 7.0104 Total Barium Chloride Addition 0.016 0.015 1.09 0.016 0.015 0.014131 Fresh Water to Barium Chloride Mixing 1.00 1.00 0.015 0.015 0.015105 Equivalent Solid Barium Chloride Addition 0.002

Lime Consumption106 Impurity Precipitation Lime Loop Feed 15.0 1.12 11.2 10.0 1.68 1.12 11.2 10.0 9.49 13.2107 Tailings and Effluent Treatment Lime Loop Feed 15.0 1.12 11.2 10.0 1.68 1.12 11.2 10.0 9.49 13.2108 Lime to Tailings 15.0 1.12 2.48 2.22 0.97 1.12 2.48 2.22 2.11 13.2109 Lime Addition to Primary ET 15.0 1.12 0.00 0.00 0.00 1.12 0.000 0.000 0.00 13.2110 Lime Addition to Secondary ET 15.0 1.12 0.12 0.11 0.005 1.12 0.12 0.11 0.10111 Total Lime Consumption 15.0 1.12 6.49 5.81 1.56 1.12 6.49 5.81 5.52 13.2138 Process Water to Lime Slaking 1.00 1.00 6.97 6.97 6.97 13.2112 Solid Lime Addition 0.89 1.56 0.00

Flocculant113 Flocculant Addition to Neutral Thickener 1.00 1.00 0.73 0.73 0.73 5.0114 Flocculant to Tailings Thickener 1.00 1.60 1.00 0.87 0.87 0.87 5.0115 Flocculant Addition to Uranium Thickener 1.00 1.00 0.044 0.044 0.044 5.0116 Flocculant to ET 1.00 1.00 0.32 0.32 0.32 5.0117 Solid Flocculant to Neutral/Tailings Thickener Floc Mixer 1.20 0.0008

Page 4 of 5




118 Solid Flocculant to Uranium Thickener Floc Mixer 1.20 0.00002119 Solid Flocculant to Effluent Treatment Floc Mixer 1.20 0.00016139 Process Water to Flocculant Mixing 1.00 1.60 1.60 1.60132 Fresh Water to Flocculant Mixing 1.00 0.36 0.36 0.36

93% Sulphuric Acid120 93\% Sulphuric Acid Addition to Leaching 1.83 2.92 1.60 0.20 1,700 2,715 -121 Average 93% Sulphuric Acid to Fresh Eluate Tank 1.83 1.84 1.01 0.128 1,700 1,710 -122 93% Sulphuric Acid Addition to ET 1.83 0.001 0.000 0.000 1,700 0.79 -123 Total Acid Consumption 1.83 4.76 2.60 0.33 1,700 4,426 -

Fresh Water Tank124 Fresh Water to Fresh Water Tank 1.00 0.00 0.0065 RO Permeate To Fresh Water Tank 1.00 38.3 38.3 7.0125 Average Fresh Water to Fresh Eluate Tank 1.00 6.98 6.98126 Fresh Water to Precip Heat Exchangers 1.00 2.50 2.50 7.0127 Fresh Water to Uranium Wash Tank 1.00 0.33 0.33 7.0128 Fresh Water to Scrubbers 1.00 6.84 6.84 7.0129 Fresh Water to H2O2 Dilution 1.00 0.013 0.013 7.0130 Fresh Water to MgO Mixing 1.00 0.15 0.15 7.0131 Fresh Water to Barium Chloride Mixing 1.00 0.015 0.015 7.0132 Fresh Water to Flocculant Mixing 1.00 0.36 0.36 7.0133 Fresh Water to Seal Water Tank 1.00 8.00 8.00 7.0134 Fresh Water to Process Water Tank 1.00 13.14 13.14 7.0135 Total Fresh Water Flowrate 1.00 38.3 38.3 7.0

Process Water9 Process Water to Ball Mil Discharge Box 1.00 6.31 6.31

136 Average Process Water to RIP 1.00 7.31 7.31137 Process Water for Gypsum Cake Repulp 1.00 2.44 2.44 7.0138 Process Water to Lime Slaking 1.00 6.97 6.97 7.0139 Process Water to Flocculant Mixing 1.00 1.60 1.60 7.0140 Total Process Water Flowrate 1.00 24.6 24.6 7.0134 Fresh Water to Process Water Tank 1.00 13.14 13.14 7.0141 Mine Water to Process Water Tank 1.00 0.00 0.00 7.0142 Secondary Clarifier Overflow To Process Water Tank 0.005 1.00 11.5 11.5 0.001 0.011 0.000 1.00 11.5 11.5 0.000 0.000 11.5 0.000 0.000 0.000 7.5

Mine Water141 Mine Water to Process Water Tank 1.00 0.00 0.00143 Mine Water to Tailings Neutralization 1.00 50.0 50.0144 Total Mine Water 1.00 50.0 50.0

Steam Generation145 Estimated Max. Steam Requirement for Leaching 0.0037 0.75 204

Page 5 of 5


October 29, 2008

APPENDIX D ESTIMATED MATOUSH PROCESS PLANT AND BUILDINGS

CAPITAL COST DETAILS - CLASS IV ESTIMATE


STRATECO RESOURCES INC. MATOUSH URANIUM PROJECT SCOPING STUDY

ESTIMATED PROCESS PLANT AND BUILDINGS CAPITAL COST DETAILS - PRELIMINARY CLASS IV ESTIMATE

PAGE AREA LABOUR LABOUR MATERIAL/BUILDING TOTAL COST(HOURS) COST ($CDN) COST ($CDN) ($CDN)

DIRECT COSTS

1 Ore Storage and Crushing 20,630 2,063,000 2,532,850 4,595,850 2 SAG Mill Grinding 46,830 4,683,000 8,328,580 13,011,580 3 Ball Mill Grinding 45,250 4,525,000 3,347,790 7,872,790 4 Neutral Thickening and Leaching 28,330 2,833,000 4,177,090 7,010,090 5 Resin Loading Circuit 36,000 3,601,000 9,610,700 13,211,700 6 Resin Elution 15,690 1,569,000 2,554,120 4,123,120 7 Impurity Precipitation 13,050 1,305,000 1,285,970 2,590,970 8 Gypsum Filter 13,070 1,307,000 1,749,960 3,056,960 9 Uranium Precipitation 7,510 751,000 770,660 1,521,660

10 Uranium Thickener 6,420 642,000 566,840 1,208,840 11 Uranium Calcining 17,210 1,721,000 2,773,520 4,494,520 12 Uranium Packaging and Scrubbing 11,650 1,165,000 1,041,670 2,206,670 13 Tailings Neutralization 25,106 2,510,600 2,908,340 5,418,940 14 Reverse Osmosis 26,500 2,650,000 3,891,420 6,541,420 15 Effluent Treatment 34,940 3,494,000 3,714,100 7,208,100 16 Reagents: Ferric Sulphate and Hydrogen Peroxide 7,690 769,000 834,080 1,603,080 17 Reagents: Oxygen and Magnesia 4,390 439,000 612,380 1,051,380 18 Reagents: Barium Chloride 2,770 277,000 221,570 498,570 19 Reagents: Lime 15,170 1,517,000 2,956,370 4,473,370 20 Reagents: Flocculant Mixing and Sulphuric Acid 5,500 550,000 1,068,400 1,618,400 21 Fresh and Process Water 17,940 1,794,000 1,853,300 3,647,300 22 Low and High Pressure, And Instrument Air 5,190 519,000 645,660 1,164,660 23 Process Plant, Crusher and Oxygen Plant Buildings 88,110 8,810,900 39,203,310 48,014,210 24 Reagents, First Fills 1,232 123,200 2,618,940 2,742,140

SUB-TOTAL DIRECT COSTS 496,200 49,618,700 99,267,620 148,886,320

INDIRECT COSTS

CONTRACTOR OVERHEAD AND PROFIT 6,700,000 ENGINEERING, PROCUREMENT AND CONSTRUCTION MANAGEMENT 18,610,000

TOTAL DIRECT AND INDIRECT COSTS 174,196,320

CONTINGENCY (25%) 43,550,000

CAPITAL SPARES, 2% OF PROCESS EQUIPMENT COST 1,150,000

TOTAL ESTIMATED CAPITAL COSTS 218,896,320

Say, 220,000,000


STRATECO RESOURCES INC. MATOUSH URANIUM PROJECT SCOPING STUDY

ESTIMATED PROCESS PLANT AND BUILDINGS CAPITAL COST DETAILS - PRELIMINARY CLASS IV ESTIMATEBY COST CENTRE

PAGE AREA INSTALLEDMECHANICAL PROCESS PIPING ELECTRICAL INSTRUMENTATION FREIGHT TOTAL COSTCOST ($CDN) COST ($CDN) COST ($CDN) COST ($CDN) COST ($CDN) ($CDN)

DIRECT COSTS

1 Ore Storage and Crushing 2,342,280 467,000 875,000 879,000 32,570 4,595,850 2 SAG Mill Grinding 6,321,300 1,899,000 2,367,000 2,375,000 49,280 13,011,580 3 Ball Mill Grinding 3,830,300 1,150,000 1,436,000 1,438,000 18,490 7,872,790 4 Neutral Thickening and Leaching 3,678,300 1,101,000 832,000 1,380,000 18,790 7,010,090 5 Resin Loading Circuit 8,501,500 1,280,000 1,705,000 1,700,000 25,200 13,211,700 6 Resin Elution 2,240,100 669,000 507,000 700,000 7,020 4,123,120 7 Impurity Precipitation 1,254,800 381,000 472,000 475,000 8,170 2,590,970 8 Gypsum Filter 1,483,200 448,000 560,000 558,000 7,760 3,056,960 9 Uranium Precipitation 767,500 232,000 234,000 284,000 4,160 1,521,660 10 Uranium Thickener 636,100 186,000 144,000 239,000 3,740 1,208,840 11 Uranium Calcining 2,270,300 851,000 510,000 856,000 7,220 4,494,520 12 Uranium Packaging and Scrubbing 1,157,900 349,000 257,000 437,000 5,770 2,206,670 13 Tailings Neutralization 2,915,660 732,000 655,000 1,097,000 19,280 5,418,940 14 Reverse Osmosis 4,185,000 631,000 837,000 879,000 9,420 6,541,420 15 Effluent Treatment 3,505,000 1,316,000 1,053,000 1,317,000 17,100 7,208,100 16 Reagents: Ferric Sulphate and Hydrogen Peroxide 844,590 251,000 186,000 318,000 3,490 1,603,080 17 Reagents: Oxygen and Magnesia 762,100 80,000 74,000 117,000 18,280 1,051,380 18 Reagents: Barium Chloride 288,900 35,000 69,000 104,000 1,670 498,570 19 Reagents: Lime 2,348,900 702,000 528,000 880,000 14,470 4,473,370 20 Reagents: Flocculant Mixing and Sulphuric Acid 1,328,400 67,000 70,000 136,000 17,000 1,618,400 21 Fresh and Process Water 1,545,100 1,164,000 346,000 580,000 12,200 3,647,300 22 Low and High Pressure, And Instrument Air 496,600 369,000 110,000 183,000 6,060 1,164,660 23 Process Plant, Crusher and Oxygen Plant Buildings 45,679,350 - - - 2,334,860 48,014,210 24 Reagents, First Fills 2,422,140 - - - 320,000 2,742,140

SUB-TOTAL 100,805,320 14,360,000 13,827,000 16,932,000 2,962,000 148,886,320

INDIRECT COSTS

CONTRACTOR OVERHEAD AND PROFIT 6,700,000 ENGINEERING, PROCUREMENT AND CONSTRUCTION MANAGEMENT 18,610,000

TOTAL DIRECT AND INDIRECT COSTS 174,196,320

CONTINGENCY (25%) 43,550,000

CAPITAL SPARES, 2% OF PROCESS EQUIPMENT COST 1,150,000

TOTAL ESTIMATED CAPITAL COSTS 218,896,320

Say, 220,000,000


STRATECO RESOURCES INC. STRATECO RESOURCES INC. MATOUSH URANIUM PROJECT SCOPING STUDY MATOUSH URANIUM PROJECT SCOPING STUDY

SUMMARY OF FREIGHT ESTIMATE: WEIGHT AND LOADS SUMMARY OF DIRECT ELECTRICAL POWER ESTIMATE

INSTALLED PEAK AVERAGEWEIGHT LOADS POWER LOAD OPERATING

AREA TONNES EA. PAGE AREA kW kW kW

1 Ore Storage and Crushing 285 8.1 1 Ore Storage and Crushing 127 76 762 SAG Mill Grinding 431 12.3 2 SAG Mill Grinding 566 339 3343 Ball Mill Grinding 162 4.6 3 Ball Mill Grinding 496 298 2614 Neutral Thickening and Leaching 164 4.7 4 Neutral Thickening and Leaching 342 205 1825 Resin Loading Circuit 220 6.3 5 Resin Loading Circuit 28 17 136 Resin Elution 61 1.8 6 Resin Elution 26 16 197 Impurity Precipitation 71 2.0 7 Impurity Precipitation 21 13 98 Gypsum Filter 68 1.9 8 Gypsum Filter 126 76 729 Uranium Precipitation 36 1.0 9 Uranium Precipitation 11 7 410 Uranium Thickener 33 0.9 10 Uranium Thickener 13 8 411 Uranium Calcining 63 1.8 11 Uranium Calcining 11 6 512 Uranium Packaging and Scrubbing 51 1.4 12 Uranium Packaging and Scrubbing 11 7 413 Tailings Neutralization 169 4.8 13 Tailings Neutralization 92 55 3514 Reverse Osmosis 82 2.4 14 Reverse Osmosis 197 118 11615 Effluent Treatment 150 4.3 15 Effluent Treatment 43 26 2816 Reagents: Ferric Sulphate and Hydrogen Peroxide 31 0.9 16 Reagents: Ferric Sulphate and Hydrogen Peroxide 9 5 517 Reagents: Oxygen and Magnesia 160 4.6 17 Reagents: Oxygen and Magnesia 904 542 54118 Reagents: Barium Chloride 15 0.4 18 Reagents: Barium Chloride 12 7 519 Reagents: Lime 127 3.6 19 Reagents: Lime 85 51 2520 Reagents: Flocculant Mixing and Sulphuric Acid 149 4.3 20 Reagents: Flocculant Mixing and Sulphuric Acid 11 7 321 Fresh and Process Water 107 3.0 21 Fresh and Process Water 493 296 2922 Low and High Pressure, And Instrument Air 53 1.5 22 Low and High Pressure, And Instrument Air 104 62 3223 Process Plant, Crusher and Oxygen Plant Buildings 20,430 584 23 Process Plant, Crusher and Oxygen Plant Buildings 250 250 25024 Reagents, First Fills 2,792 80 24 Reagents, First Fills - - -

Total 25,910 750 TOTAL DIRECT POWER 3,978 2,487 2,053

kWh/t 84.8kWh/lb U3O8 8.2

Melis Engineering Ltd. Page 1Project No. 490 STRATECO RESOURCES INC. October 29, 2008 MATOUSH URANIUM PROJECT SCOPING STUDY


AREA: ORE STORAGE AND CRUSHING

REFERENCE: FLOWSHEET NO. 490-F-101LABOUR EQUIPMENT/MATERIAL FIELD MATERIALS

SIZE POWER MATERIAL WEIGHT HOURS RATE TOTAL UNIT COST TOTAL UNIT COST TOTAL TOTAL ITEM DESCRIPTION /CAPACITY kW TONNE UNIT QTY PER UNIT ($CDN/HR) $ CDN $ CDN $ CDN $ CDN $ CDN $ CDN

1 Coarse Ore Bin 20 t - Concrete 84 m3 32 20 100 64,000 1,300 41,600 70 2,240 107,840

2 Crusher Apron Feeder 914 x 4,877 5.6 MS 4.8 ea 1 680 100 68,000 110,000 110,000 5,500 5,500 183,500

3 Jaw Crusher, FOB Montreal 508 x 914 75.0 MS 12.7 ea 1 680 100 68,000 210,000 210,000 10,500 10,500 288,500

4 Crusher Discharge Conveyor 762 x 30,000 25.0 MS 35 ea 1 410 100 41,000 490,000 490,000 24,500 24,500 555,500

5 Fine Ore Bin 620 t - Concrete 363 m3 142 20 100 284,000 1,300 184,600 70 9,940 478,540

6 Fine Ore Bin Apron Feeders 914 x 4,877 5.6 MS 4.8 ea 2 680 100 136,000 110,000 220,000 5,500 11,000 367,000

7 Crusher Dust Scrubber 6,000 m3/h 20.0 MS 4.0 ea 1 1,350 100 135,000 90,000 90,000 4,500 4,500 229,500

8 Crusher Area Overhead Crane 10T 10.0 MS 25.0 ea 1 410 100 41,000 50,000 50,000 2,500 2,500 93,500

9 Crusher Area Sump Pump 85 x 65 ( 25 m3/h) 5.6 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

10 Crusher Area Safety Shower 114 L/min - MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANICAL 152 240 - - 8,630 100 863,000 - 1,407,900 - 71,380 2,342,280

PROCESS PIPING 19 lot 1 2,800 100 280,000 187,000 187,000 - INCL 467,000

ELECTRICAL 20 lot 1 4,800 100 480,000 395,000 395,000 - INCL 875,000

INSTRUMENTATION 6 lot 1 4,400 100 440,000 439,000 439,000 - INCL 879,000

FREIGHT TO SITE 285 Loads 8.1 - 0 NIL 4,000 32,570 - NIL 32,570

TOTAL DIRECT COSTS 285 20,630 2,063,000 2,461,470 71,380 4,595,850



AREA: SAG MILL GRINDING



1 Grinding Feed Conveyor 762 x 15,000 15.0 MS 30 ea 1 270 100 27,000 140,000 140,000 7,000 7,000 174,000

2 Weightometer - 0.2 MS 0.3 ea 1 410 100 41,000 250,000 250,000 12,500 12,500 303,500

3 SAG Mill, c/w Liners, inching drive 3,962 ø x 2,286 373.0 MS 195.1 ea 1 8,100 100 810,000 4,020,000 4,020,000 201,000 201,000 5,031,000

4 Pebble Screen 1,219 x 3,048 5.6 MS 2.0 ea 1 270 100 27,000 24,000 24,000 1,200 1,200 52,200

5 Pebble Conveyor 762 x 30,000 5.6 MS 20.0 ea 1 300 100 30,000 170,000 170,000 8,500 8,500 208,500

6 Pebble Conveyor Belt Magnet 762 6.6 MS 2.7 ea 1 50 100 5,000 37,000 37,000 1,900 1,900 43,900

7 Pebble Cone Crusher 965 ø 150.0 MS 11.3 ea 1 680 100 68,000 325,000 325,000 16,200 16,200 409,200

8 Grinding Area Overhead Crane 10 T 10.0 MS 25.0 ea 1 410 100 41,000 50,000 50,000 2,500 2,500 93,500

9 Grinding Area Safety Shower 114 L/min - MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANICAL 566.0 286 - - 10,530 100 1,053,000 - 5,017,400 - 250,900 6,321,300

PROCESS PIPING 76 lot 1 11,400 100 1,140,000 759,000 759,000 - INCL 1,899,000

ELECTRICAL 53 lot 1 13,000 100 1,300,000 1,067,000 1,067,000 - INCL 2,367,000

INSTRUMENTATION 16 lot 1 11,900 100 1,190,000 1,185,000 1,185,000 - INCL 2,375,000





AREA: BALL MILL GRINDING


SIZE POWER MATERIAL MASS HOURS RATE TOTAL UNIT COST TOTAL UNIT COST TOTAL TOTAL ITEM DESCRIPTION /CAPACITY kW TONNE UNIT QTY PER UNIT ($CDN/HR) $ CDN $ CDN $ CDN $ CDN $ CDN $ CDN

1Ball Mill, c/w Liners and Inching Drive (Used)

2,743 ø x 4,879 373 MS 53.4 ea 1 6,750 100 2,196,000 1,100,000 1,100,000 55,000 55,000 3,351,000

2 Cyclone Feed Pump Box 2,700 x 2.700 x 3,500 H (23 m3) - MS/RL 10.2 ea. 1 220 100 22,000 102,000 102,000 5,100 5,100 129,100

3 Cyclone Feed Pump 156 x 104 56.0 CD4M 1.4 ea 2 220 100 44,000 70,500 141,000 3,600 7,200 192,200

4 Cyclones 660 ø - SRL 2.0 ea 2 50 100 10,000 25,000 50,000 1,300 2,600 62,600

5 Scalping Screen 1,219 x 3.048 5.6 MS 2.0 ea 1 270 100 27,000 24,000 24,000 1,200 1,200 52,200

6 Ball Bucket 0.5 m3 MS 1.0 ea 1 40 100 4,000 6,000 6,000 300 300 10,300

7 Grinding Area Sump Pump 85 x 65 ( 25 m3/h) 5.6 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

SUB-TOTAL INSTALLED MECHANICAL 496.195 73.8 - - 23,250 100 2,325,000 - 1,433,300 - 72,000 3,830,300

PROCESS PIPING - 46.0 lot 1 6,900 100 690,000 460,000 460,000 - INCL 1,150,000

ELECTRICAL - 32.0 lot 1 7,900 100 790,000 646,000 646,000 - INCL 1,436,000

INSTRUMENTATION - 10.0 lot 1 7,200 100 720,000 718,000 718,000 - INCL 1,438,000

FREIGHT TO SITE - 162 Loads 4.6 - 0 NIL 4,000 18,490 - NIL 18,490

TOTAL DIRECT COSTS 161.8 45,250 4,525,000 3,275,790 72,000 7,872,790



AREA: NEUTRAL THICKENING AND LEACHING



1 Neutral Thickener Tank 8,000 ø - LMS 9.3 ea 1 540 100 54,000 103,000 103,000 5,200 5,200 162,200

2 Neutral Thickener Mechanism 8,000 ø 2.2 LMS 4.8 ea 1 680 100 68,000 123,000 123,000 6,200 6,200 197,200

3 Neutral Thickener U/F Pumps 104 x 76 (73 m3/h) 11.2 CI 0.3 ea 2 220 100 44,000 26,000 52,000 1,300 2,600 98,600

4 Neutral Thickener O/F Tank 1,400 ø x 6,000 (7 m3) - LMS 2.9 ea 1 270 100 27,000 39,000 39,000 2,000 2,000 68,000

5 Neutral Thickener O/F Pumps 104 x 76 (73 m3/h) 6.6 CI 0.3 ea 2 220 100 44,000 24,000 48,000 1,200 2,400 94,400

6 Leach Tanks 8,000 ø x 9,600 (80 m3) - LMS 10.5 ea 6 810 100 486,000 265,000 1,590,000 13,300 79,800 2,155,800

7 Leach Tank Agitators 77Q60 44.8 LMS 1.0 ea 6 340 100 204,000 78,000 468,000 3,900 23,400 695,400

8 Leach Discharge Pumpbox 1,6100 x1,600 x 2,400 (5 m3) - LMS 4.0 ea 1 220 100 22,000 40,000 40,000 2,000 2,000 64,000

9 Leach Discharge Pump 85 x 65 ( 34 m3/h) 14.9 LMS 0.3 ea 2 220 100 44,000 26,000 52,000 1,300 2,600 98,600

10 Leach Area Sump Pump 85 x 65 ( 25 m3/h) 5.6 TPN 0.4 ea 1 220 100 22,000 10,500 10,500 600 600 33,100

11 Leach Area Safety Shower 114 L/min MS 0.0 ea 2 40 100 8,000 1,400 2,800 100 200 11,000

SUB-TOTAL INSTALLED MECHANICAL - 341.8 92 - - 10,230 100 1,023,000 - 2,528,300 - 127,000 3,678,300

PROCESS PIPING - 44 lot 1 6,600 100 660,000 441,000 441,000 - INCL 1,101,000

ELECTRICAL - 19 lot 1 4,600 100 460,000 372,000 372,000 - INCL 832,000

INSTRUMENTATION - 9 lot 1 6,900 100 690,000 690,000 690,000 - INCL 1,380,000

FREIGHT TO SITE - 164 Loads 4.7 - NIL NIL 4,000 18,790 - NIL 18,790




AREA: RESIN LOADING CIRCUIT



1 Carousel Package 40 m3/h 22 SS/RL 120 ea 1 9,450 100 945,000 7,000,000 7,000,000 350,000 350,000 8,295,000

2 Residue Pumpbox 1,6100 x1,600 x 2,400 (5 m3) - LMS 2.0 ea 1 220 100 22,000 40,000 40,000 2,000 2,000 64,000

3 Residue Pumps 85 x 65 ( 34 m3/h) 14.9 LMS 0.3 ea 2 220 100 44,000 26,000 52,000 1,300 2,600 98,600

4 RIP Area Sump Pump 85 x 65 ( 25 m3/h) 5.6 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

5 RIP Area Safety Shower 114 L/min MS 0.0 ea 2 40 100 8,000 1,400 2,800 100 200 11,000



ELECTRICAL - 38 lot 1 9,400 100 940,000 765,000 765,000 - INCL 1,705,000






AREA: RESIN ELUTION



1 Elution Columns Skid Mounted 3 columns, 3,050 ø x 4,120 (27 m3) 1.0 LMS 1.0 ea 1 1,490 100 149,000 1,300,000 1,300,000 65,000 65,000 1,514,000

2 Fresh Eluate Tank 3,580 x 6,200 (62 m3) - Plastic 1.6 ea 1 340 100 34,000 11,000 11,000 600 600 45,600

3 Fresh Eluate Tank Agitator 17Q10 7.5 SS 0.5 ea 1 270 100 27,000 12,300 12,300 700 700 40,000

4 Fresh Eluate Pumps 76 x 38, 16 m3/h 1.1 SS 0.1 ea 2 220 100 44,000 12,300 24,600 700 1,400 70,000

5 Fresh Eluate Heat Exchanger 16 m3/h - SS 1.0 ea 1 270 100 27,000 20,000 20,000 1,000 1,000 48,000

6 Lean EluateTank 3,580 x 6,200 (62 m3) - Plastic 1.6 ea 1 340 100 34,000 11,000 11,000 600 600 45,600

7 Lean Eluate Pumps 76 x 38, 16 m3/h 1.1 SS 0.1 ea 2 220 100 44,000 12,300 24,600 700 1,400 70,000

8 Lean Eluate Heat Exchanger 16 m3/h - SS 4.1 ea 1 270 100 27,000 20,000 20,000 1,000 1,000 48,000

9 Concentrated EluateTank 3,580 x 6,200 (62 m3) - Plastic 1.6 ea 1 340 100 34,000 11,000 11,000 600 600 45,600

10 Concentrated Eluate Pumps 76 x 38, 16 m3/h 1.1 SS 0.1 ea 2 220 100 44,000 12,300 24,600 700 1,400 70,000

11 Eluted Resing Holding Tank 2,438 ø x 5,791 (24.7 m3) - Plastic 1.6 ea 1 410 100 41,000 40,000 40,000 2,000 2,000 83,000

12 Eluted Resin Transfer Tank 2.4 m3/h - LMS 1.0 ea 1 220 100 22,000 6,000 6,000 300 300 28,300

13 Caustic Tank 3,580 x 6,200 (62 m3) - Plastic 1.6 ea 1 340 100 34,000 11,000 11,000 600 600 45,600

14 Caustic Tank Agitator 17Q10 7.5 SS 0.5 ea 1 - 100 - 12,300 12,300 700 700 13,000

15 Caustic Pump 76 x 38, 16 m3/h 1.1 SS 0.1 ea 1 220 100 22,000 12,300 12,300 700 700 35,000

16 Elution Area Sump Pump 65 x 40 ( 12 m3/h) 2.2 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

17 Elution Area Safety Shower 114 L/min - MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANIC - 25.9 17.4 - - 6,090 100 609,000 - 1,552,400 - 78,700 2,240,100

PROCESS PIPING - 27 lot 1 4,000 100 400,000 269,000 269,000 - INCL 669,000


INSTRUMENTATION - 6 lot 1 2,800 100 280,000 420,000 420,000 - INCL 700,000

FREIGHT TO SITE - 61 Loads 1.8 2,800 0 NIL 4,000 7,020 - NIL 7,020




AREA: IMPURITY PRECIPITATION



1 Impurity Precipitation Tank 3,500 ø x 4,200 (16.6m3) - LMS 3.6 ea 6 340 100 204,000 55,100 330,600 2,800 16,800 551,400

2 Impurity Precipitation Tank Agitator 14Q2 1.5 SS 0.1 ea 6 270 100 162,000 9,000 54,000 500 3,000 219,000

3 Impurity Precipitation Discharge Pumpbox 1,000 ø x 2,000 (1.1 m3) - LMS 2.1 ea 1 340 100 34,000 22,200 22,200 1,200 1,200 57,400

4 Impurity Precipitation Discharge Pumps 63 x 63, 9.9 m3/h 0.8 DI 0.3 ea 2 220 100 44,000 10,500 21,000 600 1,200 66,200

5 Gypsum Filter Feed Tank 5,740 ø x 6,890 (42m3) LMS 16.6 ea 1 340 100 34,000 171,000 171,000 8,600 8,600 213,600

6 Gypsum Filter Feed Tank Agitator 16Q5 3.7 LMS 0.2 ea 1 270 100 27,000 12,000 12,000 600 600 39,600

7 Gypsum Filter Feed Pumps 63 x 63, 10.6 m3/h 2.2 DI 0.3 ea 2 220 100 44,000 12,000 24,000 600 1,200 69,200

8 Impurity Precipitation Area Sump Pump 65 x 40 ( 12 m3/h) 2.2 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

9 Impurity Precipitation Area Safety Shower 114 L/min MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANICAL - 19.42 42 - - 5,750 100 575,000 - 646,500 - 33,300 1,254,800








AREA: GYPSUM FILTER


SIZE HP POWER MATERIAL WEIGHT HOURS RATE TOTAL UNIT COST TOTAL UNIT COST TOTAL TOTAL ITEM DESCRIPTION /CAPACITY kW TONNE UNIT QTY PER UNIT ($CDN/HR) $ CDN $ CDN $ CDN $ CDN $ CDN $ CDN

1Gypsum Belt Filters cw Vacuum Pump and Filtrate Pump 5 m2 57.8 SS 5.7 ea 1 1,350 100 135,000 475,000 475,000 23,800 23,800 633,800

2 Gypsum Filter Repulp Tank 2,438 ø x 2,743 (3.3 m3) - LMS 1.4 ea 1 340 100 34,000 21,100 21,100 1,100 1,100 56,200

3 Gypsum Filter Repulp Tank Agitator 14Q2 1.5 LMS 0.2 ea 1 270 100 27,000 7,000 7,000 400 400 34,400

4 Gypsum Filter Repulp Pumps 63 x 63, 10.6 m3/h 2.2 DI 4.1 ea 2 220 100 44,000 11,700 23,400 600 1,200 68,600

5 Pregnant Solution Sand Filters, 3 Skid Mounted 610 mm ø 57.8 LMS 5.7 ea 1 810 100 81,000 250,000 250,000 12,500 12,500 343,500

6 Pregnant Solution Sand Filter Backwash Pump 63 x 63, 13.1 m3/h 1.1 DI 0.3 ea 1 220 100 22,000 11,700 11,700 600 600 34,300

7 Pregnant Solution Tank 4,300 ø x 5,340 (68 m3) 10 10.4 ea 1 340 100 34,000 140,000 140,000 7,000 7,000 181,000

8 Pregnant Solution Pumps 63 x 63, 13.1 m3/h 0.8 DI 0.3 ea 2 220 100 44,000 23,300 46,600 1,200 2,400 93,000

9 Gypsum Filter Area Sump Pump 65 x 40 ( 12 m3/h) 2.2 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

10 Gypsum Filter Area Safety Shower 114 L/min MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANICAL - 126 33 - - 4,470 100 447,000 - 986,500 - 49,700 1,483,200

PROCESS PIPING - 18 lot 1 2,700 100 270,000 178,000 178,000 - - 448,000

ELECTRICAL - 13 lot 1 3,100 100 310,000 250,000 250,000 - - 560,000

INSTRUMENTATION - 4 lot 1 2,800 100 280,000 278,000 278,000 - - 558,000

FREIGHT TO SITE - 68 Loads 1.9 0 - 4,000 7,760 - - 7,760




AREA: URANIUM PRECIPITATION



1Uranium Precipitation Feed Heat Exchanger

Plate and Frame - SS 0.3 ea 2 270 100 54,000 20,000 40,000 1,000 2,000 96,000

2 Uranium Precipitation Tanks 3,500 ø x 4,200 (32.8 m3) - LMS 5.3 ea 3 340 100 102,000 82,000 246,000 4,100 12,300 360,300

3 Uranium Precipitation Tank Agitators 17Q30 1.5 LMS 0.5 ea 3 270 100 81,000 15,000 45,000 800 2,400 128,400

4Uranium Precipitation Discharge Pumpbox 1,219 x 1,219 x 1,524 high (0.8 m3) - LMS 1.5 ea 1 340 100 34,000 16,500 16,500 900 900 51,400

5Uranium Precipitation Discharge Pumps 63 x 63, 11.0 m3/h 2.2 DI 0.3 ea 2 220 100 44,000 23,300 46,600 1,200 2,400 93,000

6Uranium Precipitation Area Sump Pump 65 x 40 ( 12 m3/h) 2.2 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

7Uranium Precipitation Area Safety Shower

114 L/min MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANICAL - 11.2 20.4 - - 3,410 100 341,000 - 405,800 - 20,700 767,500




FREIGHT TO SITE - 36 Loads 1.0 0 NIL 4,000 4,160 - NIL 4,160

TOTAL DIRECT COSTS 36 7,510 751,000 749,960 20,700 1,521,660



AREA: URANIUM THICKENER



1 Uranium Thickener Tank 4,000 ø - LMS 3.0 ea 1 540 100 54,000 88,000 88,000 4,400 4,400 146,400

2 Uranium Thickener Mechanism 4,000 ø 3.7 LMS 1.7 ea 1 1,080 100 108,000 77,000 77,000 3,900 3,900 188,900

3 Uranium Thickener Underflow Pump 25 x 18, 0.5 m3/h 0.7 CD4M 0.1 ea 2 220 100 44,000 6,000 12,000 300 600 56,600

4 Barren Solution Tank 5,700 ø x 6,840 (156 m3) - LMS 13.5 ea 1 340 100 34,000 42,000 42,000 2,100 2,100 78,100

5 Barren Solution Tank Transfer Pump 63 x 63, 15.0 m3/h 2.2 DI 0.3 ea 1 220 100 22,000 11,700 11,700 600 600 34,300

6 Barren Solution Pumps 76 x 76, 40 m3/h 1.5 DI 0.3 ea 2 220 100 44,000 23,500 47,000 1,200 2,400 93,400

7 Uranium Thickener Area Sump Pump 65 x 40 ( 12 m3/h) 2.2 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

8 Uranium Thickener Area Safety Shower 114 L/min MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500



ELECTRICAL - 3 lot 1 800 100 80,000 64,000 64,000 - INCL 144,000






AREA:URANIUM CALCINING



1 Uranium Wash Tank 1,600 ø x 1,900 (2.6 m3) - LMS 1.2 ea 1 140 100 14,000 18,800 18,800 1,000 1,000 33,800

2 Uranium Wash Tank Agitator X6Q500 0.5 ea 1 140 100 14,000 2,000 2,000 100 100 16,100

3 Uranium Centrifuge Feed Pumps 25 x 18, 0.5 m3/h 0.7 CD4M 0.1 ea 2 220 100 44,000 6,000 12,000 300 600 56,600

4Uranium Centrifuge c/w Chutes, Diverter Flush Valves

170 kg U3O8/h 3.7 SS316 0.5 ea 1 270 100 27,000 108,000 108,000 5,200 5,200 140,200

5 Uranium Centrate Pumpbox 900 x 900 x 1,900 high (0.8 m3) - MS/RL 0.4 ea 1 340 100 34,000 5,500 5,500 300 300 39,800

6 Uranium Centrate Pumps 25 x 18, 0.5 m3/h 0.7 CD4M 0.1 ea 2 220 100 44,000 6,000 12,000 300 600 56,600

7 Uranium Calciner 170 kg/h 0.7 - 5.4 ea 1 2,030 100 203,000 1,400,000 1,400,000 70,000 70,000 1,673,000

8 Uranium Lump Disintegrator Tromm 170 kg U3O8/h 0.7 9.5 MS 1.7 ea 1 410 100 41,000 90,000 90,000 4,500 4,500 135,500

9 Uranium Transfer Bucket Elevator 3.7 MS 1.0 ea 1 540 100 54,000 25,000 25,000 1,300 1,300 80,300

10 Calcining Area Sump Pump 65 x 40 ( 12 m3/h) 2.2 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

11 Calcining Area Safety Shower 114 L/min MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANICAL - 13 11.1 - - 5,010 100 501,000 - 1,685,000 - 84,300 2,270,300




FREIGHT TO SITE - 63 Loads 1.8 - 0 NIL 4,000 7,220 - INCL 7,220




AREA: URANIUM PACKAGING AND SCRUBBING



1 Uranium Bin c/w Rotary Valve and Bin Vi 2,500 ø x 9,000 (40 m3 ) - MS 9.2 ea 1 1,080 100 108,000 67,200 67,200 3,400 3,400 178,600

2 Uranium Packaging Line 20 Drums per day 1.9 MS 4.0 ea 1 1,620 100 162,000 200,000 200,000 10,000 10,000 372,000

3 Uranium Calciner Scrubber Package 0.8 SS 4.0 ea 1 810 100 81,000 100,000 100,000 5,000 5,000 186,000

4 Uranium Calciner Room Scrubber - 0.8 MS 4.0 ea 1 810 100 81,000 75,000 75,000 3,800 3,800 159,800

5 Uranium Packaging Scrubber Package - 0.8 MS 4.0 ea 1 810 100 81,000 50,000 50,000 2,500 2,500 133,500

6 Uranium Scrubber Water Pumpbox 1.3 m x 1.3 m x 1.9 m high (2.3 m3) - LMS 1.7 ea 1 340 100 34,000 12,300 12,300 700 700 47,000

7 Uranium Scrubber Water Pumps 76 x 51 (24 m3) 2.2 SS 0.1 ea 2 220 100 44,000 10,426 20,900 600 1,200 66,100

8 Uranium Packaging Area Sump Pump 65 x 40 ( 12 m3/h) 2.2 TPN 0.4 ea 1 40 100 4,000 10,284 10,300 600 600 14,900

SUB-TOTAL INSTALLED MECHANICAL - 27.5 28 - - 5950 100 595,000 - 535,700 - 27,200 1,157,900

PROCESS PIPING - 14 lot 1 2100 100 210,000 139,000 139,000 - INCL 349,000


INSTRUMENTATION - 3 lot 1 2200 100 220,000 217,000 217,000 - INCL 437,000

FREIGHT TO SITE 51 Loads 1.4 0 0 NIL 4,000 5,770 - NIL 5,770




AREA: TAILINGS NEUTRALIZATION



1 Tailings Feed Launder 6,000 x 1,000 x 400 - LMS 1.1 ea 1 270 100 27,000 14,200 14,200 800 800 42,000

2 Tailings Neutralization Tank 4,900 ø x 5,880 - LMS 14.0 ea 2 540 100 108,000 190,000 380,000 9,500 19,000 507,000

3 Tailings Neutralization Tank Agitator 74Q15 11.2 LMS 0.5 ea 2 270 100 54,000 33,500 67,000 1,700 3,400 124,400

4 Tailings Neutralization Tank Blower - 2.2 NR 0.3 ea 1 220 100 22,000 10,000 10,000 500 500 32,500

5 Tailings Thickener Tank 16,459 ø MS 54 ea 1 810 100 81,000 510,000 510,000 25,500 25,500 616,500

6 Tailings Thickener Mechanism 16,459 ø 3.7 MS 30 ea 1 1,350 100 135,000 280,000 280,000 14,000 14,000 429,000

7 Tailings Thickener Overflow Tank 1,000 ø x 6,000 (5 m3/h) MS 2.2 ea 1 340 100 34,000 33,800 33,800 1,700 1,700 69,500

8 Tailings Thickener Overflow Pumps 93 m3/h 2.2 NR 0.3 ea 2 220 100 44,000 28,800 57,600 1,500 3,000 104,600

9 Tailings Thickener Underflow Pumps 61 m3/h 29.8 NR 0.3 ea 2 220 100 44,000 28,800 57,600 1,500 3,000 104,600

10 Tailings Line Blowout Tank 2,000 mm ø 57.8 LMS 5.7 ea 1 680 100 68,000 80,000 80,000 4,000 4,000 152,000

11Tailings Line Containment Excavation

4 km - - - m3 1,500 0.40 100 60,000 4.0 6,000 0.20 300 66,300

12 Tailings Lines 75 mm ø x 4 Km 57.8 LMS 5.7 m 4,000 1.0 100 400,000 25 100,000 2.0 8,000 508,000

13Tailings Line to Backfill Plant Containment Excavation

500 m - - - m3 190 0.40 100 7,600 4.0 760 0.20 - 8,360

14 Tailings Line to Backfill Plant 75 mm ø x 500 m 57.8 LMS 5.7 m 500 1.0 100 50,000 25 12,500 100 50,000 112,500

15 Neutralization Area Sump Pump 65 x 40 ( 12 m3/h) 2.2 TPN 0.4 ea 1 220 100 22,000 10,300 10,300 600 600 32,900

16 Neutralization Area Safety Shower 114 L/min - MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500




INSTRUMENTATION - 7 lot 1 5500 100 550,000 547,000 547,000 - INCL 1,097,000





AREA: EFFLUENT TREATMENT

REFERENCE: FLOWSHEET NO.490-F-113LABOUR EQUIPMENT/MATERIAL FIELD MATERIALS


1 Primary Effluent Treatment Tanks 2,100 ø x 2,520 mm - LMS 2.0 ea 2 340 100 68,000 31,000 62,000 1,600 3,200 133,200

2 Primary Effluent Treatment Tank Agitator 14Q3 2.2 LMS 0.5 ea 2 270 100 54,000 5,000 10,000 300 600 64,600

3 Primary Clarifier, Plate Type - 1.9 MS 15.0 ea 1 680 100 68,000 150,000 150,000 7,500 7,500 225,500

4 Primary Clarifier Overflow Tank 1,700 ø x 2,040 (4.3 m3) - MS 1.4 ea 1 270 100 27,000 9,900 9,900 500 500 37,400

5 Primary Clarifier Overflow Pump 76 x 51 (25 m3/h) 1.1 SS 0.1 ea 2 220 100 44,000 10,300 20,600 600 1,200 65,800

6 Primary Clarifier Underflow Pump 38 x 25 (2.1 m3/h) 1.1 SS 0.2 ea 2 220 100 44,000 10,800 21,600 600 1,200 66,800

7 Secondary Effluent Treatment Tanks 2,100 ø x 2,520 LMS 2.0 ea 2 340 100 68,000 31,200 62,400 1,600 3,200 133,600

8 Secondary Effluent Treatment Tank Agitator 14Q3 2.2 LMS 0.1 ea 2 270 100 54,000 5,000 10,000 300 600 64,600

9 Secondary Clarifier, Plate Type - 1.9 LMS 15.0 ea 1 680 100 68,000 150,000 150,000 7,500 7,500 225,500

10 Secondary Clarifier Overflow Tank 1,700 ø x 2,040 (4.3 m3) - MS 1.4 ea 1 270 100 27,000 9,900 9,900 500 500 37,400

11 Secondary Clarifier Overflow Pump 76 x 38 (16 m3/h) 1.5 SS 0.1 ea 2 220 100 44,000 10,000 20,000 500 1,000 65,000

12 Secondary Clarifier Underflow Pump 38 x 25 (4.3 m3/h) 1.5 SS 0.1 ea 2 220 100 44,000 10,800 21,600 600 1,200 66,800

13 Effluent Sand Filters, 3, Skid Mounted - - MS 3.1 ea 1 680 100 68,000 175,000 175,000 8,800 8,800 251,800

14 Effluent Discharge Tank 6,401 ø x 6,706 (200 m3) MS 11.7 ea 1 410 100 41,000 85,300 85,300 4,300 4,300 130,600

15 Effluent Sand Filter Backwash Pump 76 x 38 (16 m3/h) 0.8 SS 0.1 ea 2 220 100 44,000 4,700 9,400 300 600 54,000

16 Monitoring Ponds 1,030 m3 Plastic 1.2 ea 3 2,030 100 609,000 350,000 1,050,000 17,500 52,500 1,711,500

17 Monitoring Pond Discharge Pumps 180 m3/h 5.6 LMS 0.5 ea 3 220 100 66,000 21,000 63,000 1,100 3,300 132,300

18 Effluent Treatment Area Sump Pump 52 ( 12 m3/h) 1.5 A744 0.2 ea 1 220 100 22,000 10,500 10,500 600 600 33,100

19 Effluent Treatment Area Safety Shower 114 L/min MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANICAL - 43.0 63.6 - - 14,640 100 1,464,000 - 1,942,600 - 98,400 3,505,000


ELECTRICAL - 24 lot 1 5,800 100 580,000 473,000 473,000 - INCL 1,053,000






AREA: REVERSE OSMOSIS



1 Reclaim Barge pumps 152 ( 72 m3/h) 3.7 LMS 0.7 ea 2 220 100 44,000 19,565 39,100 1,000 2,000 85,100

2 Reclaim Barge 1.0 ea 1 540 100 54,000 10,000 10,000 500 500 64,500

3 Reverse Osmosis Package Plant 160 m3/h 190.0 MS/RL 30 ea 1 13,500 100 1,350,000 2,700,000 2,700,000 135,000 135,000 4,185,000

4 Mine Water Pond (Existing)

5 Mine Water Pump (Existing)









AREA: REAGENTS: FERRIC SULPHATE AND HYDROGEN PEROXIDE



1 Concentrated Ferric Sulphate Tank 3,900 ø x 4,680 (47.7 m3) - LMS 6.5 ea 1 340 100 34,000 100,000 100,000 5,000 5,000 139,000

2 Ferric Sulphate Distribution Pumps Metering 0.1 Metering 0.2 ea 2 140 100 28,000 3,000 6,000 200 400 34,400

3 Ferric Sulphate Area Sump Pump 52 ( 12 m3/h) 1.5 A744 0.2 ea 1 220 100 22,000 22,100 22,100 1,200 1,200 45,300

4 Ferric Sulphate Area Safety Shower 114 L/min MS 0.0 ea 1 40 100 4,000 1,390 1,390 100 100 5,490

5 Hydrogen Peroxide Package 36.5 m3 (Vendor Supplied) 0.1 Passivated SS 6.5 ea 1 2,030 100 203,000 300,000 300,000 15,000 15,000 518,000

6 Hydrogen Peroxide Metering Pump Metering 1.9 SS 0.2 ea 3 140 100 42,000 3,000 9,000 200 600 51,600

7 Hydrogen Peroxide Area Sump Pump 52 ( 12 m3/h) 1.5 A744 0.2 ea 1 220 100 22,000 22,100 22,100 1,200 1,200 45,300

8 Hydrogen Peroxide Area Safety Showe 114 L/min MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANICAL - 5.1 15 - - 3,590 100 359,000 - 461,990 - 23,600 844,590



INSTRUMENTATION 2 lot 1 1,600 100 160,000 158,000 158,000 - INCL 318,000





AREA: REAGENTS: OXYGEN AND MAGNESIA

REFERENCE: FLOWSHEET NO's 490-F-116/117LABOUR EQUIPMENT/MATERIAL FIELD MATERIALS


1Oxygen Plant, Owned and Operated by Air Liquide

20 tpd (14,000 x 9,000) 440 Misc. 60.0 ea 2 - - - - - - - -

1 Magnesia Silo and Mixing Package 14 MS 29.0 ea 1 1,620 100 162,000 385,000 385,000 19,300 19,300 566,300

2 Magnesia Distribution Tank 3,200 ø x 3,840 (25.5 m3) - LMS 4.2 ea 1 340 100 34,000 30,700 30,700 1,600 1,600 66,300

3 Magnesia Distribution Agitator 16Q10 7.5 LMS 0.3 ea 1 270 100 27,000 8,000 8,000 400 400 35,400

4 Magnesia Distribution Pump 76 x 38 (17.3 m3/h) 0.8 SS 0.1 ea 2 220 100 44,000 10,500 21,000 600 1,200 66,200

5 Magnesia Area Sump Pump 52 ( 12 m3/h) 1.1 A395 0.2 ea 1 220 100 22,000 5,600 5,600 300 300 27,900









AREA: REAGENTS: BARIUM CHLORIDE



1 Barium Chloride Mix Tank 2,900 ø x 3,480 (18.2 m3) - MS 3.7 ea 1 340 100 34,000 27,000 27,000 1,400 1,400 62,400

2 Barium Chloride Mix Tank Agitator 15Q7.5 5.6 MS 0.2 ea 1 270 100 27,000 6,000 6,000 300 300 33,300

3 Barium Chloride Transfer Pump 76 x 38, 18 m3/h 1.1 SS 0.1 ea 1 220 100 22,000 4,700 4,700 300 300 27,000

4 Barium Chloride Distribution Tank 3,200 ø x 3,840 (25.5 m3) - MS 4.0 ea 1 340 100 34,000 29,100 29,100 1,500 1,500 64,600

5 Barium Chloride Distribution Pump 76 x 38, 20 m3/h 2.2 SS 0.1 ea 2 220 100 44,000 11,500 23,000 600 1,200 68,200

6 Barium Chloride Area Sump Pump 52 ( 12 m3/h) 1.1 A395 0.2 ea 1 220 100 22,000 5,600 5,600 300 300 27,900

7 Barium Chloride Area Safety Shower 114 L/min MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500


PROCESS PIPING - 4 lot 1 500 100 - 35,000 35,000 - INCL 35,000




- TOTAL DIRECT COSTS 14.6 2,770 277,000 216,470 5,100 498,570



AREA: REAGENTS: LIME



1 Lime Package 5 tph 76.0 MS 80 ea 1 2,970 100 297,000 1,850,000 1,850,000 92,500 92,500 2,239,500

2 Lime Loop Pumps 76 x 38, (20 m3/h) 3.7 MS 0.2 ea 2 220 100 44,000 13,000 26,000 700 1,400 71,400

3 Lime Area Sump Pump 52 ( 12 m3/h) 1.5 A395 0.2 ea 1 220 100 22,000 10,000 10,000 500 500 32,500

4 Lime Area Safety Shower 114 L/min MS 0.0 ea 1 40 100 4,000 1,400 1,400 100 100 5,500

SUB-TOTAL INSTALLED MECHANICAL - 81.2 81 - - 3,670 100 367,000 - 1,887,400 - 94,500 2,348,900








AREA: REAGENTS: FLOCCULANT MIXING AND SULPHURIC ACID

REFERENCE: FLOWSHEET NO 490-F-119LABOUR EQUIPMENT/MATERIAL FIELD MATERIALS


1 Flocculant Mix System c/w Pumps - 1.5 (Package) 0.5 ea 3 270 100 81,000 76,000 228,000 3,800 11,400 320,400

2 Flocculant Metering Pumps 9,754 ø x 10,363 (737 m3) 0.3 plastic 0.0 ea 7 140 100 98,000 3,000 21,000 200 1,400 120,400

3 Acid Storage Tanks14 Days, 8,230 ø x 8,839

(444 m3)- MS 33 ea 2 600 100 120,000 240,000 480,000 12,000 24,000 624,000

4 Acid Storage Containment 110% of Storage Tank - Concrete 225 m3 90 5 100 45,000 1,300 117,000 70 6,300 168,300

4 Acid Distribution Pumps 38 x 25 (1.0 m3/h) 2.2 SS 0.1 ea 2 220 100 44,000 16,500 33,000 900 1,800 78,800

5 Sulphuric Acid Plant Safety Shower - MS 0.0 ea 3 40 100 12,000 1,400 4,200 100 300 16,500

SUB-TOTAL INSTALLED MECHANICAL - 11 143 - - 4,000 100 400,000 - 883,200 - 45,200 1,328,400 (Excludes Aggregate)





TOTAL DIRECT COSTS 149 5,500 550,000 1,023,200 45,200 1,618,400



AREA: FRESH AND PROCESS WATER

REFERENCE: FLOWSHEET NO. 490-F-120/122LABOUR EQUIPMENT/MATERIAL FIELD MATERIALS


1 Fresh Water Tank 4,572 ø x 5,182 (76.9 m3) - MS 10.6 ea 1 410 100 41,000 76,000 76,000 3,800 3,800 120,800

2 Fresh Water Distribution Pumps 27 m3/h 0.8 DI 0.2 ea 2 220 100 44,000 20,000 40,000 1,000 2,000 86,000

3 Jockey Pump 76 x 38 3.7 MS 0.1 ea 1 - 100 - - - - - -

4 Fire Pump 127 x 127 187.5 MS 0.0 ea 2 - 100 - - - - - -

5 Diesel Fire Pump 127 x 127 1.5 MS 0.0 ea 1 - 100 - - - - - -

6 Fire Loop Circulation Pump - 1.0 - 0.0 ea 1 - 100 - - - - - -

7 Items 3-5 on 9,144 x 10,728 Skid - - - 15.9 ea 1 410 100 41,000 315,000 315,000 15,800 15,800 371,800

8 Seal Water Tank 2,700 ø x 3,240 (16 m3) - MS 2.9 ea 1 340 100 34,000 20,700 20,700 1,100 1,100 55,800

9 Seal Water Pumps 76 x 38, 9.6 m3/h 3.7 DI 0.1 ea 2 220 100 44,000 9,400 18,800 500 1,000 63,800

10 Process Water Tank 3,962 ø x 4,572 (50.2 m3) - MS 8.0 ea 1 410 100 41,000 58,000 58,000 2,900 2,900 101,900

11 Process Water Pumps 16 m3/h 5.6 DI 0.2 ea 2 220 100 44,000 10,300 20,600 600 1,200 65,800

12 Safety Shower Surge Tank 2,438 x 2,437 x 3,048 (15 m3) 5.6 MS 3.5 ea 2 340 100 68,000 25,300 50,600 1,300 2,600 121,200

13 Safety Shower Head Pumps 76 x 38, 9.6 m3/h 3.7 DI 0.1 ea 2 220 100 44,000 9,400 18,800 500 1,000 63,800

14 Saftey Shower Water Heaters 30.0 MS 0.5 ea 2 220 100 44,000 1,000 2,000 100 200 46,200

15 Safety Shower Head Tank 3,000 ø x 3,60 (6 m3) 0.0 N/A 2.2 ea 1 340 100 34,000 15,700 15,700 800 800 50,500

16 Potable Water System 15 m3/d 16.8 DI 0.1 ea 1 1,350 100 135,000 250,000 250,000 12,500 12,500 397,500

SUB-TOTAL INSTALLED MECHANICA - 493 48.7 - - 6,140 100 614,000 - 886,200 - 44,900 1,545,100








AREA: LOW AND HIGH PRESSURE, AND INSTRUMENT AIR



1 High Pressure Air Compressor 706 Nm3/h, 760 kPa 50.0 N/A - ea 2 - 100 - - - - - -

2 High Pressure Air Receiver 915 ø x 2,440 - MS - ea 1 - 100 - - - - - -

3 Instrument Air Dryer and Oil Filter Cartridge type 2.0 N/A - ea 1 - 100 - - - - - -

4 Instrument Air Receiver 915 ø x 2,440 - MS - ea 1 - 100 - - - - - -

5 Items 1-4 on 3,000 x 9,000 mm skid 14 ea 1 810 100 81,000 261,000 261,000 13,100 13,100 355,100

6 Boiler 1.7 tph (860 HP) 1.5 MS 20 ea 1 680 100 68,000 70,000 70,000 3,500 3,500 141,500

SUB-TOTAL INSTALLED MECHANICAL - 104 34 - - 1,490 100 149,000 - 331,000 - 16,600 496,600








AREA: PROCESS PLANT, CRUSHER AND OXYGEN PLANT BUILDINGS

REFERENCE: 490-L-100LABOUR EQUIPMENT/MATERIAL FIELD MATERIALS


1 Excavation - - - - m3 27,800 0.40 100 1,112,000 4.0 111,200 0.20 5,600 1,228,800

2 Clay Cover - - Clay 8,050 m3 4,600 0.40 100 184,000 30 138,000 10 46,000 368,000

3 Two Seepage Barriers - - Plastic 100 m2 17,200 0.50 100 860,000 20 344,000 10 172,000 1,376,000

4 Concrete Floors and Walls - - Concrete 6,000 m3 2,400 5 100 1,200,000 1,300 3,120,000 70 168,000 4,488,000

5 Concrete Piers and Bases - - Concrete 5,250 m3 2,100 5 100 1,050,000 1,300 2,730,000 70 147,000 3,927,000

6 Roads - - Asphalt 6,105 m2 5,550 0.5 100 277,500 25 138,750 10 55,500 471,750

7Building Steel, Shell, Walls and Cranes 7,156 m2 - Steel 7,160 m2 7,160 0.40 100 286,400 2,500 17,900,000 200 1,432,000 19,618,400

8 Platforms, Stairs etc. - - Steel 750 t 750 30 100 2,250,000 7,700 5,775,000 400 300,000 8,325,000

9Process Plant Non-Process Electrical

- 240 - 80 m2 9,520 1.0 100 952,000 160 1,523,200 10 95,200 2,570,400

10 Crusher Building Concrete 10 m x 10 m - Concrete 125 m3 50 5 100 25,000 1,300 65,000 100 5,000 95,000

11 Crusher Building 100 m2 - Steel 100 m2 100 0.40 100 4,000 2,500 250,000 200 20,000 274,000

12Crusher Building Non-Process Electrical

- 10 - 0.8 m2 100 1.0 100 10,000 160 16,000 10 1,000 27,000

13 Distributed Control System - - - 10 ea. 1 5,000 100 500,000 700,000 700,000 35,000 35,000 1,235,000

14 Laboratory - - - 20 ea. 1 1,000 100 100,000 1,500,000 1,500,000 75,000 75,000 1,675,000

SUB-TOTAL INSTALLED MECHANICAL 250 20,430 100 8,810,900 - 34,311,150 - 2,557,300 45,679,350 (Excludes Aggregate)

PROCESS PIPING - - INCL - INCL - INCL -

ELECTRICAL - - INCL - INCL - INCL -

INSTRUMENTATION - - INCL - INCL - INCL -

FREIGHT TO SITE Loads (excludes aggregate) 583.7 - NIL NIL 4,000 2,334,860 - NIL 2,334,860

- TOTAL DIRECT COSTS 20,430 88,110 8,810,900 36,646,010 2,557,300 48,014,210



AREA: REAGENTS, FIRST FILLS

LABOUR EQUIPMENT/MATERIAL FIELD MATERIALSSIZE POWER MATERIAL WEIGHT HOURS RATE TOTAL UNIT COST TOTAL UNIT COST TOTAL TOTAL

ITEM DESCRIPTION /CAPACITY kW TONNE UNIT QTY PER UNIT ($CDN/HR) $ CDN $ CDN $ CDN $ CDN $ CDN $ CDN

1 Barium Chloride - - 20 t 20 1.0 100 2,000 1,000 20,000 - - 22,000

2 Caustic Soda (NaOH) - - 20 t 20 1.0 100 2,000 600 12,000 - - 14,000

3 Flocculant, Anionic Polyacrylamide - - 2 t 2 1.0 100 200 5,250 10,500 - - 10,700

4 Flocculant, Non-ionic Polyacrylamide - - 2 t 2 1.0 100 200 5,250 10,500 - - 10,700

5 Hydrogen Peroxide - - 24 t 24 1.0 100 2,400 1,060 25,440 - - 27,840

6 Lime (98% CaO) - - 400 t 400 0.2 100 8,000 193 77,200 - - 85,200

7 Magnesia (MgO) - - 20 t 20 0.2 100 400 510 10,200 - - 10,600

8 Product Drums - - 350 ea. 350 0.2 100 7,000 50 17,500 - - 24,500

9 Resin for RIP - - 264 m3 240 1.0 100 24,000 6,500 1,560,000 - - 1,584,000

10 Steel Grinding Balls (Grinding) - - 70 t 70 5.0 100 35,000 1,000 70,000 - - 105,000

9 Steel Grinding Balls (Lime Slaking) - - 20 t 20 5.0 100 10,000 1,000 20,000 - - 30,000

7 Sulphuric Acid - - 1,600 t 1,600 0.2 100 32,000 291 465,600 - - 497,600

SUB-TOTAL INSTALLED MECHANICAL - - 2,792 - - 1,232 100 123,200 - 2,298,940 - - 2,422,140

PROCESS PIPING - - - - - - - - - - - -

ELECTRICAL - - - - - - - - - - - -

INSTRUMENTATION - - - - - - - - - - - -

FREIGHT TO SITE - 2,792 Loads 80 - - - 4,000 320,000 - - 320,000

TOTAL DIRECT COSTS 2,792 1,232 123,200 2,618,940 - 2,742,140


October 29, 2008

APPENDIX E CONSUMABLES, PROCESS AND MAINTENANCE

CONSUMABLES OPERATING COSTS


TABLE 1STRATECO RESOURCES INC. - MATOUSH URANIUM PROJECT

CONSUMABLES LIST

Consumable Unit Consumption CommentsBarium Chloride tonnes/a 15 CalculatedCaustic Soda (NaOH) tonnes/a 25 EstimatedCrusher Wear Parts tonnes/a 19 0.1 kg/tonneDiesel (Yellowcake Calcining) L/a 407,000 0.4 kg/kg U3O8

Diesel (Steam Generation) L/a 675,000 Heating Leaching to 50ºCDiesel (Power Generation) L/a 3,780,000 $0.31/kWhFerric Sulphate (45%) tonnes/a 220 CalculatedFlocculant, Anionic Polyacrylamide tonnes/a 1 CalculatedFlocculant, Non-ionic Polyacrylamide tonnes/a 7 CalculatedGrinding Mill Liners ea. 2 Annual AllowanceHydrogen Peroxide (70%) tonnes/a 256 CalculatedLaboratory Supplies ea. - EstimatedLime (98% CaO) tonnes/a 12,400 CalculatedMagnesia (MgO) tonnes/a 94 CalculatedOxygen tonnes/a 10,000 EstimatedProduct Drums ea./a 568 Assumes Each is Recycled 4 TimesRIP Resin m3/a 58 0.30 m3/ktonne oreSteel Grinding Balls (Grinding) tonnes/a 194 1.0 kg/tonneSteel Grinding Balls (Lime Slaking) tonnes/a 6 0.5 kg/tonneSulphuric Acid (94% H2SO4 w/w) tonnes/a 38,000 114 tonnes of 94% H2SO4/dayNote: 1. Electrical power consumption does not include the oxygen plant; this power is included in the oxygen cost.

TABLE 2STRATECO RESOURCES INC. - MATOUSH URANIUM PROJECT

CONSUMABLES, PROCESS AND BUILDING ELECTRICAL AND MAINTENANCE CONSUMABLESOPERATING COST COST SUMMARY

Consumable Unit $/Unit $Cdn/a (1,000's) $ Cdn/t $Cdn/lb U3O8

Barium Chloride tonnes 1,000 15 0.08 0.01Caustic Soda (NaOH) tonnes 600 15 0.08 0.01Crusher Wear Parts tonnes 1,000 19 0.10 0.01Diesel (Yellowcake Calcining) Litres 1.00 407 2.10 0.20Diesel (Steam Generation) Litres 1.00 675 3.48 0.34Ferric Sulphate (45%) tonnes 700 154 0.79 0.08Flocculant, Anionic Polyacrylamide tonnes 5,250 6.6 0.03 0.003Flocculant, Non-ionic Polyacrylamide tonnes 5,250 34 0.18 0.02Grinding Mill Liners ea. 250,000 500 2.58 0.25Hydrogen Peroxide (70%) tonnes 1,060 271 1.40 0.14Laboratory Supplies ea. - 250 1.29 0.13Lime (98% CaO) tonnes 193 2,393 12.35 1.20Magnesia (MgO) tonnes 510 48 0.25 0.02

Oxygen(1) tonnes 210 2,100 10.84 1.05

Product Drums ea. 50 28 0.15 0.01RIP Resin m3 6,500 378 1.95 0.19

Steel Grinding Balls (Grinding) tonnes 1,000 194 1.00 0.10Steel Grinding Balls (Lime Slaking) tonnes 1,000 6 0.03 0.003

Sulphuric Acid (2) tonnes 291 11,058 57.08 5.53

Total Consumables - - 18,553 95.76 9.28

Diesel (Power Generation)(3) Litres 1.00 3,780 19.51 1.89

Maintenance Consumables unit (4) 1,150 5.94 0.58Sub-Total - - 23,483 121.21 11.74Contingency (15%) - - 3,522 18.18 1.76Total - - 27,005 139.39 13.50

Note: 1. Oxygen cost from Air Liquide, and includes the cost of electrical power. 2. Estimated 2012 sulphuric acid cost includes $250/tonne base cost, $15/tonne transfer cost and $3.25/km (return trip included) x 300 km, assuming 38 tonnes/load. 3. Electrical power cost does not include the oxygen plant, power to run the oxygen plant is included in the oxygen cost. 4. Annual maintenance consumables cost estimated at 2% of process equipment cost.


October 29, 2008

APPENDIX F MATOUSH PERSONNEL CHARTS

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



AB


SCOPING STUDY RELEASESCOPING STUDY SECOND RELEASE

BYBCFBCF

REVI

SION

S


SION

S


RENC

E

MATOUSH MILL

MILL OPERATIONS PERSONNEL CHART (1/2)

PC-001


MINE MANAGER

MILL SUPERINTENDENT

MILL GENERAL FOREMAN

MILL FOREMAN MILL FOREMAN

MILL OPERATORSLEAD HANDCONTROL ROOM OPERATORGRINDING OPERATORLEACHING AND RIP OPERATORPRECIPITATION AND CALCINER OPERATORTAILINGS NEUTRALIZATION ANDEFFLUENT TREATMENT OPERATORBOILER ROOM OPERATOR/SPARESPARE/TRAINEE OPERATORSPARE/TRAINEE OPERATOR

MILL OPERATORSLEAD HANDCONTROL ROOM OPERATORGRINDING OPERATORLEACHING AND RIP OPERATORPRECIPITATION AND CALCINER OPERATORTAILINGS NEUTRALIZATION ANDEFFLUENT TREATMENT OPERATORBOILER ROOM OPERATOR/SPARESPARE/TRAINEE OPERATOR

MILL GENERAL FOREMAN

MILL FOREMAN MILL FOREMAN

MILL OPERATORSLEAD HANDCONTROL ROOM OPERATORGRINDING OPERATORLEACHING AND RIP OPERATORPRECIPITATION AND CALCINER OPERATORTAILINGS NEUTRALIZATION ANDEFFLUENT TREATMENT OPERATORBOILER ROOM OPERATOR/SPARESPARE/TRAINEE OPERATORSPARE/TRAINEE OPERATOR

MILL OPERATORSLEAD HANDCONTROL ROOM OPERATORGRINDING OPERATORLEACHING AND RIP OPERATORPRECIPITATION AND CALCINER OPERATORTAILINGS NEUTRALIZATION ANDEFFLUENT TREATMENT OPERATORBOILER ROOM OPERATOR/SPARESPARE/TRAINEE OPERATOR

TOTAL MILL OPERATIONS (NOT INCLUDING LABORATORY)

MILL OPERATORS: 4 x 8.5 = 34MILL FOREMEN: 4 x 1 = 4MILL TRAINER: 2 x 1 = 2MILL GENERAL FOREMEN: 2 x 1 = 2METALLURGISTS: 2MILL SUPERINTENDENT: 1TOTAL 45

METALLURGIST

CREW 2CREW 1

METALLURGIST

SHIFT 1A SHIFT 2ASHIFT 1B SHIFT 2B

MILL TRAINER MILL TRAINER

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



AB



BYBCFBCF

REVI

SION

S


SION

S


RENC

E

MATOUSH MILL

MILL OPERATIONS PERSONNEL CHART (2/2)

PC-002


MINE MANAGER

MILL SUPERINTENDENT

TOTAL LABORATORY (INCLUDED IN MILL OPERATIONS)

LABORATORY SUPERVISOR: 2 x 1 = 2LABORATORY STAFF: 2 x 3 = 6TOTAL 8

LABORATORY SUPERVISOR

LABORATORY TECHNICIANS (3)

LABORATORY SUPERVISOR

LABORATORY TECHNICIANS (3)

CREW 1 CREW 2

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



AB



BYBCFBCF

REVI

SION

S


SION

S


RENC

E

CREW 1

MATOUSH MILL

MILL MAINTENANCE PERSONNEL CHART

PC-003


MINE MANAGER

MAINTENANCE SUPERINTENDENT

MILL MAINTENANCE GENERAL FOREMAN

MILL MAINTENANCE FOREMAN

MAINTENANCE CREWELECTRICIANELECTRICIANINSTRUMENTATION TECHNOLOGISTINSTRUMENTATION TECHNOLOGISTMILLWRIGHT/WELDERMILLWRIGHTMILLWRIGHT

TOTAL MILL MAINTENANCE

ELECTRICIANS: 2 x 2 = 4INSTRUMENT TECHNOLOGISTS: 2 x 2 = 4MILLWRIGHTS: 2 x 3 = 6MILL MAINTENANCE FOREMEN: 2 x 1 = 2DATA ENTRY CLERKS: 2 x 1 = 2MILL MAINTENANCE GENERAL FOREMAN: 1MILL MAINTENANCE PLANNER 1MAINTENANCE SUPERINTENDENT: 1TOTAL 21

MILL MAINTENANCE FOREMAN

MAINTENANCE CREWELECTRICIANELECTRICIANINSTRUMENTATION TECHNOLOGISTINSTRUMENTATION TECHNOLOGISTMILLWRIGHT/WELDERMILLWRIGHTMILLWRIGHT

MILL MAINTENANCE PLANNER

MINE MAINTENANCE

CREW 2

DATA ENTRY CLERK

DATA ENTRY CLERK

TITLE

AREA

DWG. No.

LOC.



CHECKED:

DRAWN:

DESIGNED:

DATESCALE (A1):



AB



BYBCFBCF

REVI

SION

S


SION

S


RENC

E

MATOUSH MILL

GENERAL AND ADMINISTRATION PERSONNEL

CHART

PC-004


MINE MANAGER

TOTAL GENERAL AND ADMINISTRATION

SITE SERVICES SUPERVISOR: 2 x 1 = 2SITE SERVICES STAFF: 2 x 2 = 4WAREHOUSE SUPERVISOR: 2 x 1 = 2WAREHOUSE STAFF: 2 x 2 = 4ENVIRONMENTAL TECHNICIANS: 2 x 2 = 4RADIATION, HEALTH & SAFETY STAFF: 2 X 2 = 4HES&R SUPERVISOR: 1 x 1 = 1HUMAN RESOURCES STAFF: 2 x 2 = 4SITE SECRETARY AND FLIGHT CLERK: 2 x 2 = 4HUMAN RESOURCES SUPERVISOR: 1 x 1 = 1HES&R SUPERINTENDENT: 1MINE MANAGER 1TOTAL 32

SITE SERVICES SUPERVISOR

SITE SERVICES STAFF (2)

WAREHOUSE SUPERVISOR

WAREHOUSE AND TOOL CRIB STAFF

(2)

SITE ADMINISTRATION

SUPERVISOR

HEALTH, ENVIRONMENT AND SAFETYSUPERVISOR

ENVIRONMENTAL TECHNICIANS (2)

RADIATION, HEALTH AND

SAFETY STAFF (2)

HUMAN RESOURCES STAFF

(2)

SITE SECRETARY AND FLIGHT CLERK

(2)

SITE SERVICES SUPERVISOR

SITE SERVICES STAFF (2)

WAREHOUSE SUPERVISOR

WAREHOUSE AND TOOL CRIB STAFF

(2)

ENVIRONMENTAL TECHNICIANS (2)

RADIATION, HEALTH AND

SAFETY STAFF (2)

HUMAN RESOURCES SUPERVISOR

HUMAN RESOURCES STAFF

(2)

SITE SECRETARY AND FLIGHT CLERK

(2)

CREW 1 CREW 2

17 mineral resource and mineral reserve estimates · 17 mineral resource and mineral reserve...

Documents