S2N84SE«ilM3 63.4855 BALMER TWP · using a 4" diameter aluminum core barrel or casing and a steel...

S2N84SE«ilM3 6 3 . 4 8 5 5 BALMER TWP

FREEGOLD

FreeGold Recovery Inc.

Vancouver, B.C.

A Review and Verification by Wright Engineers Limited of the

Sampling Technique of Campbell Red Lake Mines Limited

Tailings Deposit

w XV WRIGHT ENGINEERS LIMITED VANCOUVER CANADA

PROJECT NO. 1^71-100 SEPT. 1986

o/n*6-/-c-9#'

w S3'4&3

Phor* «604)684-B3?1 • Cable-WRK3HTENG" • Tel«i 04-S436T WRK^T ENGINEERS LIMITED •/»!

1444 Alberni Street, Vancouver, British Columbia, Canada, V6G 2Z4

Project No.: 1*71

September 10, 1986

FreeGold Recovery Inc. 1333 West 8th Ave. Vancouver, B.C. V6H 3W4

Attention: Mr. Harry Barr, President

Dear Harry:

We are pleased to submit the accompanying report entit led:

"A Review and Verification by Wright Engineers Limited of the Sample Handling Technique of Campbell Red Lake Mine Tailings".

We wish you well in your endeavours to recover precious metals from this and other mine tailings you have sampled.

We look forward to assisting you in any way as your testwork and project progresses.

Yours very truly,

WRIGHT ENGINEERS LIMITED,

fi -AAJ-

S.3. Andrews, P. Eng. Manager Mineral Eng. Division

SJA/mph Encl.

(

1-1

Project No. 1471

August 11, 1986

A Review and Verification by Wright Engineers Limited

of the Sample Handling Technique of Campbell Red Lake Mine Tailings

INTRODUCTION;

Campbell Red Lake Mines Ltd. have been in production since 1949. During

this period the mine has produced continuously at rates up to the present 1100 tpd.

The resulting tailing product production is estimated at approximately 10 million tons.

After miscellaneous losses and mine backfill are decuted the tailing deposit is

estimated at about 8 million tons.

Large portions of this deposit are from earlier years of production when gold

recovery technology was not as efficient as it is today and the price of gold did not

warrant extra costs for higher recovery. The result is a tailing deposit that may be

retreated at a profit to the mine and the contractor.

FreeGold Recovery Inc. have entered into a contract with Campbell Red

Lake Mines (C.R.L.M.) to sample and measure the tailing deposit and, if economical,

to recover a portion of the gold values in the deposit. Wright Engineers Limited were

requested in this first stage to act as a third party to verify the sampling methods,

techniques in handling the samples through to assay determinations and method of

calculating the tonnage in the deposit.

1-2

SAMPLING;

The tailings deposit under consideration was divided into two areas; the first

being the decant or slime deposit area and the second is the 'normal' 'beach' type

tailings disposal area. The relative areas are approximately 38 acres and 80 acres

respectively. These areas have been divided for sampling purposes into 250 foot

sections with sample hole spacing every 250 feet making a 250 x 250 foot sampling

grid.

FreeGold Recovery Inc. personnel (with assistance from C.R.L.M.) are drill

sampling at the above grid locations. The drill being used is a Wink-Vibrahead Unit

using a 4" diameter aluminum core barrel or casing and a steel cutter head with core

retainer. The two piece drill units are skid mounted and can be easily moved with a

mini 4 x 4 rubber tired, Suzuki, A.T.V. (all terrain vehicle). Spare drill barrels, core

trays, sampling equipment, etc. are moved in a small two wheel trailer with the same

A.T.V. See also the accompanying photographs. The vibrating power unit easily drives

the core barrel through the tailing material. The rate of penetration of course slows

with increasing depth but is still an acceptable rate. Some difficulties have been

encountered when the drill hits an obstruction such as a previous submerged dyke, road

or logs (or near vertical tree).

When the hole is drilled completely the core barrels and core are pulled with

the winching device (see pictures). The core flows, or is vibrated or pushed out of the

core barrels into a plastic tray (one half of a 6" plastic pipe). The 4" x 5' core sections

are then resampled every 4" with a 5/8" dia. tube coring tool. This yields

approximately \5 cores out of the original 5' core length. This is a good representative

sample of each 5' section.

In the drilling process the core is compacted by the vibration resulting in a

core length less than the core barrel driven length. This is accurately measured and

compensation is made in the assay/depth and tonnage calculation.

1-3

The final secondary sample (series of 5/8" cores) was taken to the C.R.L.M.

sample preparation room. Here the sample is dried, rolled with a bottle and screened

to break up agglomerates. It is then thoroughly mixed on a 'rolling cloth' and 'cut' to

obtain the sample for actual assay. The samples are then taken to the new, very

modern, assay laboratory. The very competent C.R.L. staff of the assay department

then assay the samples by a combination of fire assay and atomic absorption

spectophotometer.

As a check on the above procedure during the writer's visit separate samples

were cut out of the 5' core sections for independent assay. One set of samples was

taken by the writer for assay at Chemex Laboratory in Vancouver. The other check

sample was sent to Lakefield Research for assay. This resulted in a three way assay

check, the results of which are as follows:

VEL/Chemex Sample No.

5751

5752

5753

5754

5755

Line

4450 E

4450 E

4450 E

4450 E

4450 E

Hole No.

4

5

5

5

5

Depth

0 - 5

0 - 5

5 - 10

10- 15

15-20

CRM

0.060

0.058

0.074

0.055

0.043

Chemex

0.064

0.054

0.072

0.050

0.042

Lakefield

0.058

0.053

0.074

-

_

,&66 .6*3 .0 7^

\

1-*

TONNAGE CALCULATION;

Tonnage calculations are based on a planimeter survey of the map, tailing

storage area, plus the dri l l hole depth along each of the 250 foot section lines (see

attached example). The unit weight of tailings in place was originally calculated (as a

first best estimate) using the value used for tailings backfill namely 19.0 f t^ / ton. This

value has been checked by taking several core samples from the various fineness

grades of tailings (from near discharge point to the slime pond area). The results of

these 'in place' density tests are as follows:

Line

0 + 310

*950E

5950E

Hole No.

5

5

5

Fines (slime area tailings)

Medium fines tailing

Coarse (sampled near

discharge plant)

ft3 per ton

25.2*

23.61

22.60

Using the above 'in place' tailings densities the tonnage of tailings in the two

areas, Area 1 (slime tailing) and Area 2 (medium and coarse tailing) is as follows:

Area 1 (slime tailing) @ 25.2* f t 3 /s t

Area 2 (medium and coarse tailing) (3 23.10 f tVs t average

Short Tons

89*,333

2,719,593

Total

Total rounded

3,613,926

3,600,000

1-5

CONCLUSION:

This tailings deposit, as is commonly the case, is a relatively homogenous

layering of finely ground tailing material. The sample grid pattern adopted will

produce a good representative primary series of samples of each five foot depth of the

deposit. The (»" diameter core provides a good weight and size for secondary sampling.

The secondary sampling at Un intervals also produces a good representative secondary

sample. The sample handling and assaying is carried out by technicians who are well

qualified and using equipment of the latest design for accurate assays. The cross

checking at Lakefield and Chemex indicate that there are no inherent errors in the

sampling system.

The tonnage, grade and contained value calculations have been checked and

found to be based on representative samples and good engineering practice.

S.3. Andrews, P. Eng.

Manager Mineral Eng. Div.

FREEGOLD - C.R.L.M. TAILINGS RECOVERY TEST

JULY 1986

Sampling of #2 test area of our abandoned tailings ponds was completed the first week of July. A total of 69 holes having an average depth of 17.9 ft. were drilled on test area #2 which used to be the primary pond. Using 19 ft.'/ton for tailings density a total of 3.3 million tons of 0.053 oz./ton was outlined in this pond.

SAMPLING PROCEDURE

Sampling procedure remained the same as in test area #1. The sonic drill vibrated the core tube to the bottom of the tailings. This length was called the total drill length. As the core barrels were removed the sample was placed in split 6" PVC pipes and sampled every 4" using a 5/8" circular tube. Because the core liquified and compacted in the core barrel the sample was never as long as the actual length of core drilled. To adjust for this ratio of actual length drilled to length of sample was calculated (ratio C/D on sheets).

EXAMPLE

Area #2, line 4200 Er hole #2

Total length drilled = 16.5 ft.

Length of Core extracted =14.0 ft.

Ratio = C/D = 14.0 =0.85 16.5 .

The first core box contained 4 ft.

4 ft. sample represents 4/0.85 = 4/̂ 71 ft.

This corrected length was used to calculate the weighted average.

TONNAGE CALCULATION

A volume was calculated using a planimeter area multiplied by the average depth. On test area #1 volume • was calculted in the same way and was checked using cross-sec.tional areas and also using a pit volume program. Results were all within 10% (See tonnage calculation for test area #1) ••

2 -

(

Total Tonnage, Grade, Contents

Total Tons - Area #1 1.2 Million Tons Area #2 3.3 Million Tons *Total 4.5 Million Tons

*Using 19.0 cu. ft./ton which is likely a maximum value.

x_

= 0.051 oz./ton

Average grade

52Ne4SE«>e43 63.4855 BALMER TWP O S O

An Investigation of

THE RECOVERY OF GOLD

from Campbell Red Lake tailing samples

submitted by

FREEGOLD RECOVERY INCORPORATED

Progress Report No. 1

N

Project No. L.R. 3181

NOTE: This report refers to the samples as received.

The practice of this Company in issuing reports of this nature is to require the recipient not to publish the report or any part thereof without the written consent of Lakefield Research.

LAKEFIELD RESEARCH A DIVISION OF FALCONBRIDGE LIMITED

September 10, 1986

I N T R O D U C T I O N

In a letter dated July 21, 1986, Mr. H. Barr of FreeGold Recovery

Incorporated authorized a test program on samples of Campbell Red Lake tailings

and a gravity concentrate recovered from them to investigate the recovery of

gold by flotation and cyanidation.

LAKEFIELD RESEARCH

R.S. Salter

General Manager

K.W. Sarbutt

Chief Project Engineer

Experimental Work by: G. Mcllmoyle

S U M M A R Y

1. Head Samples

Twenty-eight

5700E No. 7

5200E No. 4

4700E No. 6

4450E No. 6

tailing samples were received:

0-4' 4'-9' 9'-14' ,14'-19'*

0-5 1/2' 5 1/2-10 10 1/2-15 15 1/2-20

.0-5' 5-10' 10-15'

0-7' 7-12' 12-17' 17-22'

5450E No.

5200E No. 1/2' 1/2 ' 1/2 »

4450E No.

« •

4450E No.

4

6

5

4*

0-5'* 5-10' 10-15' 15-20'

0-6' 6-11* 11-16' 16-21'

0-5'* 5-10'* 10-15' 15-20'

Five of these samples(*) were selected for check analysis and these

samples were dried and pulverized. The remainder of the samples were combined

and pulped and samples removed for testwork and head analysis.

A gravity concentrate was also received. The sample was.pulped and

charges removed for testwork.

- 3 -

Summary - Continued

1. Head Samples - Cont'd

Head assays and a size fraction analysis were conducted:

Sample

4450E No. 4 5450E No. 4 0-5 4450E No. 5 0-5 4450E No. 5 5-10 5700E No. 7 14-19

Tailing Comp. Gravity Cone.

Assay, g/t Au

2.26 2.00 1.83 2.58 1.63

1.86 3.72

Assay %

As J Fe 1 S

0.18 0.29

8.49 11.3

0.46 . 1.07

Size Fraction Analysis

Tailing Comp.

Size Fraction

+65 Mesh -65 +100 Mesh -100 +150 Mesh -150 +200 Mesh -200 +270 Mesh -2 70 +400 Mesh -400 Mesh

Head (Calculated)

Weight %

0.5 3.6 9.3 11.8 10 9.7 55.1

100.0

Assay g/t Au

6.09 3.74 2.62 2.55 3.63 1.61 1.43

2.02

1 " •

% Distribution Au

1.51 6.68 12.06 14.92 18.00 7.74

39.07

100.00

Gravity Concentrate

+65 Mesh -65 +100 Mesh -100 +150 Mesh -150 +200 Mesh -200 +270 Mesh -2 70 +400 Mesh -400 Mesh

Head (Calculated)

4.5 13.8 24.0 19.4 13.7 10.3 14.3

100.0

12.4 4.67 4.13 3.92 • 3.66 2.94 3.69

4.29

13.02 15.04 23.13 17.74 11.70 7.07 12.31

100.0

1

Summary- Continued

2. Testwork - Tailings Composite

2.1. Direct Cyanidation

Direct cyanidation tests were conducted on the tailing composite to

investigate the effect of fineness of grind on the recovery of gold. Standard

conditions employed in these tests were:

1 g/L NaCN pH 10.5 33% solids 48 hours leach time

The conditions and results of the tests are summarized in Table No. 1.

Table No. 1

Direct Cyanidation

Test

No.

2 3 7

8

Grind kWh/t

(approx . )

0 7.25

14.50 21.75 1

% -400

.'Mesh

53.3 81.8 92.6 96.4

Reag., Cons

NaCN

0.42 2.17 -1.50 1.40

»., k g / t

CaO

1.34 0.98 0.71 0.83

% Au Ex t r ac t i on *

24 h

38 .1 45.0

-

48 h

41 .1 48 .9 51.5 53.8

Residue

g / t Au

1.14 0.97 0.91 0.89

Head

| g / t Au

1.94 1.90 1.88 1.93

Recovery increased as the fineness of grind was increased. Approximately

50 percent of the gold could be recovered after grinding to 80 percent passing

400 mesh. Increasing the fineness to 96 percent passing 400 mesh only increased

the recovery to 54 percent.

Summary - Continued

2.2. Flotation and Cyanidation of Flotation Products

Two tests were conducted to examine preconcentration of the tailings

by flotation with and without a grind. A concentrate was recovered with A-350 and

R-208 as collectors. The results of the flotation are summarized in Table No. 2.

Table No. 2

F l o t a t i o n Resul t s

Test

No.

1 9

Grind kHh/t

(approx.)

0 8.5

% -400 Mesh

53 83

F l o t a t Weight

%

7.1 10.1

ion Concentrate Au I %

g / t Rec 'y

7.36 8.61

27.4 J 48.9

T a i l ,

g / t Au

1.48 1.01

Head

g / t Au

1.90 1.78

Gold recovery increased after grinding but was still low at 49 %*

The concentrate from each test was cyanided. Standard conditions employed

in the cyanidation were: 1 g/L NaCN pH 10.5 33 % solids 48 hours

The conditions and results of the cyanidation are summarized in Table No. 3,

Table No. 3

Cyanid

Test No.

1

-9 1

a t ion of F l o t a t i o n Concentrate

Regrind kWh/t Cone.

56 36

Reag. Cons. , NaCN

6.89 4 .71

kg / t Cone. CaO

2.94 2.47

% Ex t rac t ion Au

58.5 49.8

Residue g / t Au

3.05 4.32

% Au Rec 'y C a l l

16.0 24.4

Even after fine regrinding only 50-60 percent of the gold in the flotation

concentrate was recovered by cyanidation giving low overall recoveries.

Summary - Continued

3. Testwork - Gravity Concentrate

Similar testwork to that conducted on the tailing composite was also conducted

on the gravity concentrate. The results of the direct cyanidation tests are

summarized in Table No. 4.

Table No. 4

Cyanidation of Gravity Concentrate

Test

No.

Grind kWh/t

(approx.) -400 Mesh

Reag. Cons, kg/t

NaCN CaO

Residue Assay Au g/t

% Au

Rec'y

Head Calc. Au g/t

4 5 10

0 7.25

21.75

14 49 92

1.40 1.30 1.45

_.

0.74 0.72 0.77

2.24 1.77 1.52

47.9 60.8 57.7

4.30 4.51 .3.59

After grinding recovery increased form 48 to approximately 60 percent

from the gravity concentrate.

One test was conducted in which the gravity concentrate was further upgraded

by flotation and the flotation concentrate was reground and treated by cyanidation.

The flotation concentrate amounted to 11 percent of the weight, assayed

15.8 g/t Au and the gold recovery was 47.5 percent.

In cyanidation 59.5 percent of the gold in the concentrate was recovered

for an overall recovery of 28.2 percent.

P - 7 -

Summary - Continued

4. Overall Results and Recommendations

Direct cyanidation of the tailings with no grinding gave 41 percent recovery..

The recovery could be increased to 50-55 percent with fine regrinding.

Flotation of ground tailings concentrated 50 percent of the Au in 10

percent of the weight. After further regrinding and cyanidation only 50 percent

of the Au in the concentrate was recovered by cyanidation for an overall recovery

of only 25 percent.

Recoveries were slightly higher from the gravity concentrate but it is not

known how much weight or recovery the gravity concentrate represents.

Size fraction analysis of the head sample showed that 65 percent of the Au

was in the minus 200 mesh fraction. Therefore flotation may have more potential for

pre-concentration than gravity. Further testwork should be directed towards improving

the flotation response. Evaluation of grind, collector type and activators is required.

If the flotation response can be substantially improved then further evaluation of

cyanidation of the flotation concentrate would be warranted.

HMTirrUB.

- 8 -

DETAILS OF TESTS

Test No. 1

Purpose:

Feed:

Grind:

Conditions:

To investigate the recovery of gold by flotation and cyanidation.

1 kg of tailings composite.

As received

Stage

Rougher 1 Rougher 2 Rougher 3 Rougher 4

Reagents Added, grams per tonne

CuSO- A350 j R208

250

20 10

10

20 10

[ DF250

Time, minutes

Cond.

8 I 2 2

8 5 4 2

Froth

5 3. 4 3

PH

-

8.2

Stage Flotation Cell Speed: r.p.m.

Rougher 500 g D-l 1100

J

- 9

Test No. 1 - Continued

Purpose:

Procedure:

Cyanidation of flotation concentrate.

The sample was pulped with water in a one litre bottle. NaCN and lime were added and the cyanidation was carried out on rolls in two 24 hour stages with the solution being changed after each stage. Between each stage the pulp was filtered and the residue washed three times with water. The residue was then repulped with fresh cyanide solution and the test continued.

Feed' 120 g flotation rougher concentrate

Solution Volume: 240 mL Pulp Density 33 % solids

Solution Composition: 1.0 g/L NaCN

pH Range: 10.5 with Ca(0H)2

Grind: 20 minutes in lab pebble mill.

Reagent Balance:

Time

Hours

Stage 1 0-5 5-24

Stage 2 24-29 ' 29-48

Total Stage 1 Total Stage 2

Total

Actus NaCN

0.242 0.088

0.242 0.189

0.33

0.431

0.761

Added, Grams

Ca(0H)2

0.10 0.050

0.10

0.15

0.10

0.25

Equiv NaCN

0.240 0.084

0.240 0.180

0.324

0.420

0.744

alent CaO

0.075 0.038

0.075

0.113

0.075

0.188

Residual

Grams NaCN I CaO

0.156 0.084

0.06 0.22

0.084

0.22

0.304

0.005

0

0

"

Consumed

Gra NaCN

0.084 0.156

0.180 0.02

0.24

0.20

0.44

ms CaO

0.075 0.038

0.07 0.005

0.113

0.075

0.188

PH

10.7- 9.6 10.6-10.4

10.8-10.5 10.4

-

-

gent Consumption (kg/t of cyanide feed) NaCN: 6.89 CaO: 2.94

- 10 -


Metallurgical Results

Product

Flotation Concentrate Flotation Tailing

Head (Calculated)

" • - —

Amount

7.05 92.95

100.00

Assays,mg/L,g/t

Au

7.36 1.48

1.90

% Distribution

Au

27.4 72.6

100.0

Cyanidation

24 h Pregnant+Wash 48 h Pregnant+Wash Cyanide Residue

Flot. Cone. (Calc.) ••

680 mL 2000 mL 63.9 g

63.9 g

0.36 0.015 3.05

7.36

14.3 1.7

11.4

27.4

52.1 6.4

41.5

100.0

Screen Analyses

Rougher Tailing

Mesh Size (Tyler)

+ 65 100 150 200 270 400

- 400

Total

% Retained Individual Cumulative

0.6 3.6 10.6 12.0 10.6 10.4 52.2

100.0

0.6 4.2 14.8 26.8 37.4 47.8 100.0

-

. % Passing Cumulative

99.4 95.8 85.2 73.2 62.6 52.2

-

11

Test No. 2

Purpose:

Procedure:

To investigate the recovery of gold by cyanidation of the unground tailing composite.

The sample was pulped with water in a two litre bottle. NaCN and lime were added and the cyanidation was carried out on rolls in two 24 hour stages with the solution being changed after each stage. Between each stage the pulp was filtered and the residue washed three times with water. The residue was then repulped with fresh cyanide solution and the'test continued.

Feed: 500 g tailings composite




Reagent Balance:

Time

Hours

Stage 0-5

5-24

Stage 24-29 29-48

Total Stage Total Stage

Total

1

2

1

2

Actua NaCN .•

1.05 0.105

1.05 0.105

1.155

1.155

Added, Grams

|Ca(0H)2

0.500

0.500

0.50

0 .5o

Equiv NaCN

1.0Q 0.10

1.00 0.10

1.10

1.10

ra lent CaO

0.375

0.375

0.375

0.375

Residual

Gr* NaCN

0.900 0.95

0.90 0.963

0.95

0.963

ims CaO

0.06 0.02

0.025 0.023

0.02

0.023

Consumed

Gr* NaCN

0.100 0.05

0.100 0.037

0.105

0.104

0.209

ims CaO

0.315 0.04

0.35 0.002

0.319

0.352

0.671

PH

11.2-10 10 .9- 9

11 .3 -11 . 11.3

-

-

9 8

2

Reagent Consumption (kg/t of cyanide feed) NaCN: 0.42 CaO: 1.34

- 12 -



Product

1. 24 h Pregnant+Wash 2. 48 h Pregnant+Wash 3. Cyanide Residue

Head (Calculated)

Amount

1260 mL -1925 mL 495.0 g

495.0 g

Assays,mg/L,g/t

Au

0.29 0.015 1.14

1.94

% Distribution

Au

38.10 3.03 58.87

100.0 j

Calculated Grades and Recoveries

Products 1 and 2 3185 1.935 41.13

Screen Analyses

Cyanide Residue

Mesh Size (Tyler)

+ 65 100 150 200 2 70 400

- 400

Total

% Retained Individual - J Cumulative

0.6 3.7 10.3 11.4 10.5 10.2 53.3

100.0

0.6 4.3 14.6 26.0 36.5 46.7 100.0

-

% Passing Cumulative

99.4 95.7 85.4 74.0 63.5 53.3

-

- 13

Test No. 3

Purpose:

Procedure:

Feed:

To repeat test No. 2 but with a grind.

The sample was pulped with water in a two litre bottle. NaCN and lime were added and the cyanidation was carried out on rolls in two 24 hour stages with the solution being changed after each stage. Between each stage the pulp was filtered and the residue washed three times with water. The residue was then repulped with fresh cyanide solution and the test continued.

500 g tailing composite




Grind: 20 minutes inlab pebble mill.

Reagent Balance:

Time

Hours

Stage 0-5 5-24

Stage 24-29 29-48


Total

1

Added, Grams

Actual NaCN |ca(0H)2

1.05 0.211

2 1.050 0.211

1

2

1.261

1.261

0.300 0.050

0.300

0.350

0.300

Equî NaCN

1.00 0.200

1.00 0.200

1.200

1.200

1

/a] ent CaO

0.225 0.038

0.225

0.263

0.22!

Residual

Grams NaCN J CaO

0.800 1 0.01 0.60 J -

0.80 0.714

0.60

0.714

0.03 0.03

0.0

0.03

1

Consumed

Grams NaCN ICaO

0.200 0.40

0.200 0.286

0.600

0.486

0.215 0.038

0.195

0.253

0.195

1.086 0.488

10 10

11

PH

.6-10.

.6-10.

0-11. 10.9

-

3 4

0


- 14 -



Product


Head (Calculated)

Amount

1590 mL 1700 mL 485.3 g

485.3 s.

Assays,mg/L,g/t

Au

0.26 0.021 0.97

1.90

% Distribution

Au

45.0 3.9 51.1

100.0


Products 1 and 2 3290 0.14 1 48.9

Screen Analyses

Cyanide Residue

Mesh Size j % Reta (Tyler) j Individual

+ 150 0.2 200 1.4 270 4.9 400 11.7

- 400 81.8

Total j 100.0

:.ned Cumulative

0.2 1.6 6.5 18.2 100.0

-


99.8 98.4 93.5 81.8

-

15

Test No. 4

Purpose:

Procedure:

Feed;

To investigate the recovery of gold by cyanidation of the unground gravity concentrate.


500 g gravity concentrate as received.




Reagent Balance:

Time

Hours

Stage 1 0-5 '; 5-24

Stage 2 24-29 29-48

Total Stage 1 Total State 2

Total

Added, Grams

Actua NaCN."

1.05 0.053

1.050 0.211

1.103

1.261

1 Ca(0H)z

0.300

0.300

0.300

0.300

Equiv NaCN

1.00 0.050

1.00 0.20

1.050

1.20

alent CaO

0.225

0.225

0.225

0.225

Residual

Gr* NaCN

0.950 1.0

0.80 0.552

1.0

0.552

tms CaO

0.07 0.02

0.035 0.035

0.02

0.035

Consumed

Grams NaCN CaO

0.050 0.00

0.20 0.448

0.050

0.648

0.698

0.155 0.05

0.165

0.205

0.165

0.370

PH

-

n.o-ii:o 11.0-10.8

11.0-11.2 11.1

-


- 16 -



Product


Head (Calculated)

Amount

1500 mL 1725 mL 496.9 g

496.9

Assays,mg/L,g/t

Au

0.64 0.04 2.24

4.30

% Distribution

Au

44.7 3.2 52.1

100.0


Products 1 and 2 3225 0.80 47.9

Screen Analyses

Cyanide Residue

Mesh Size (Tyler)

+ 65 100 150 200 270 400

- 400

Total

% Reta Individual

4.2 14.2 24.7 19.5 13.9 9.9 13.6

100.0

.ned Cumulative

4.2 18.4 43.1 62.6 76.5 86.4 100.0

-


95.8 81.6 56.9 37.4 23.5 13.6 -

-

- 17

Test No. 5

Purpose:

Procedure:

To repeat test No. 4 but with a grind.

The sample was -pulped with water in a two litre bottle. NaCN and lime were added and the cyanidation was carried out on rolls in two 24 hour stages with the solution being changed after each stage. Between each stage the pulp was filtered and the residue washed three times with water. The residue was then repulped with fresh cyanide solution and the test continued.

Feed: 500 g gravity concentrate



pH Range: 10.5•with Ca(0H)2

Grind: 20 minutes in 2 kg lab pebble mill.

Reagent Balance:

Time

Hours

Stage 0-5 5-24

Stage 24-29 29-48


Total

1

2

1

2

Ac tu'a NaCN

1.05 0.53

1.050 0.158

1.58

1.208

Added, Grams

1 Ca(0H)2

0.25 0.100

0.250

0.35

0.25

Equi\ NaCN

1.00 0.50

1.00 0.15

1.50

1.15

alent CaO

0.188 0.075

0.188

0.263

0.188

Residual

Gr* NaCN

0.500 1.00

0.85 1.00

1.00

1.00

tms CaO

0.02

0.04 0.04

0.02

0.04

Consumed

Gra NaCN

0.500 0.00

0.15 0.00

0.500

0.150

0.65

ms CaO

0.188 0.055

0.148

0.243

0.148

0.391

PH

10.8-10.3 10.9-10.6

11.0-11.1 10.9

-

Reagent Consumption ( k g / t of cyanide feed) NaCN: 1.30 CaO: 0.782

- 18 -



Product


Head (Calculated)

Amount

1490 mL 1525 mL 497.8 g

497.8

Assays,mg/L,g/t

Au

0.88 0.035 1.77

4.51

% Distribution

Au

58.40 2.36 39.24

100.0


Products 1 and 2 3015 0.45 60.76

Screen Analyses

Cyanide Residue

Mesh Size (Tyler)

+ 100 150 200 270 400

- 400

Total

% Retained Individual - Cumulative

0.2 2.3 9.2 19.1 19.9 49.3

100.0

0.2 2.5 11.7 30.8 50.7 100.0

-


99.8 97.5 88.3 69.2 49.3

-

- 19 -

Test No. 6

Purpose: To investigate the recovery of gold by flotation and cyanidation from the gravity concentrate.

Procedure: The 1 kg sample was pulped and floated in a 1000 g. cell.. All flotation was done in the one cell.

Feed: Gravity concentrate, as received

Conditions:

Stage

Rougher 1 Rougher 2 Rougher 3 Rougher 4 Rougher 5


A350 I R208

20 30

10 50

20 20

10 50

CuSOi,

300

DF250

'8 4 4 4 8

Time, minutes

Cond.

2 2 5 1 2

Froth

4 4 4 4 4

PH

7.0 7.0 7.1

Stage Rougher 1-5 Flotation Cell 1000 g D-2 Speed: r.p.m. 1400 % Solids 33

• - 20 -


Purpose:

Procedure: The sample was pulped with water in a two litre bottle. NaCN and lime were added and the cyanidation was carried out on rolls in two 24 hour stages with the solution being changed after each stage. Between each stage the pulp was filtered and the residue washed three times with water. The residue was then repulped with fresh cyanide solution and the test continued.

Feed: 200 g rougher concentrate flotation



pH Range: ' 10.5 with Ca(0H)2

Grind: 20 minutes in lab pebble mill.

Reagent Balance:

Time

Hours

0-17 17-24 24-29 29-48

Added, Grams

Actu'i NaCN

0.43 0.380 0.42 0.042

]1 Ca(0H)2

0.10 0.08 0.14

Total 1 1.262 J 0.32

Equiv* NaCN

0.40-0.36 0.00 0.04

0.800

l e n t CaO

0.075 0.060 0.105

0.24

Residual

Gra NaCN

0.04 0.40 0.36 0.292

0.292

ns CaO

-

0.00

Consumed

Gra NaCN

0.36 0.00 0.04 0.108

0.508

ns CaO

0.075

0.24

PH

10 .5 - 9.0 10 .5-10.4 10.4-10.4

9.9

-


- 21 -



Product

Flotation Cone. Flotation Tail.

Head (Calc.)

Cyanidation

24 h Pregnant+Wash 48 h Pregnant+Wash Cyanide Residue

Head (Calc.)

Amount

11.1 88.9

100.0

1100 mL 1620 mL 104.3 g

104.3 g |

Assays, %

Au

15.8 2.18

3.69

0.73 0.11 6.40

15.8

% Distribution Au

O'all

47.5 52.5

100.0 .

23.1 5.1 19.3

47.5

Ind.

48.7 10.8 40.5

100.0

• - 22 -

To repeat test No. 2 but with a finer grind.





pH Range: 10.5 with CaOOH)2

Grind: 40 minutes in lab pebble mill

Reagent Balance:

Time

Hours

0-17 17-23 23-40 40-48

Added, Grams

Ac tu'a NaCN

1.05 0.421 0.105 0.158

Total j 1.724

1 Ca(0H)2

0.325 0.150

0.475

Equivalent NaCN CaO

1.00-0.400 0.100 0.150

1.650

0.244 0.113

0.357

Residual

Gr* NaCN

0.600 0.900 0.850 0.900

0.900

uns CaO

"

-

Consumed

Gra NaCN

0.400 0.100 0.150 0.100

0.750

is CaO

0.244 0.113

0.357

PH

10 .5-10 .0 10 .6-10 .6 10 .6-10 .4 10 .4-10 .3

10.4


Test No. 7

Purpose:

Procedure:

Feed:

- 23 -



Product

L. Pregnant+Wash 2. Residue

Head (Calculated)

Amount

1400 tnL 507.4 g

507.4

Assays,mg/L,g/t

Au

0.35 0.91

1.88

% Distribution

Au

51.5 48.5

100.0

—

Screen Analyses

Cyanide Residue

Mesh Size (Tyler)

% Ret Individual

+ 400 1 7.4 - 400 92.6

Total 100.0

lined Cumulative

7.4 100.0

-


92.6

-

24 -

Test No. 8

Purpose:

Procedure:

Feed:

To repeat test No. 2 but with a still finer grind.

The sample was pulped with water in a two litre bottle. NaCN and lime were added and the cyanidation was carried out on rolls in one 48 hour stage. The pulp was filtered and the residue washed three times with water.





Grind: 60 minutes in lab pebble mill

Reagent Balance:

Time

Hours

0-17 17-23 23-40 40-48

Added, Grams

j Actua NaCN

1.05 0.474 0.105 0.105

Total 1 1.734

Ca(0H)2

0.325 0.150 0.050

0.525

Equiv NaCN

1.00 0.450 0.100 0.100

1.650

ilent CaO

0.244 0.133 0.038

0.415

Residual

Grams NaCN j CaO

0.550 0.900 0.900 0.950

0.950 -

Consumed

Grams NaCN CaO

0.450 0.100 0.100 0.050

0.700

0.244 0.133 0.038

0.415

PH

10.5-10.0 10.6-10.4 10.6-10.4 10.4-10.4

10.4


- 25 -


Meta l lu rg ica l Resu l t s

Product

1. Pregnant+Wash 2. Residue

Head (Calculated)

Amount

1460 mL 507.9 g

507.9g 1

Assays,mg/L,g/t

Au

0.36 0.89

1.93

% Distribution

Au

53.8 46.2

100.0

Screen Analyses

Cyanide Residue

Mesh Size (Tyler)

+ 400 - 400

% Ret Individual

3.6 96.4

Total 100.0

tined Cumulative

3.6 100.0

-


96.4

-

26

Test No. 9

Purpose:

Procedure:

Feed :

Grind:

Conditions:

To determine if grinding will enhance recovery of gold by flotation from the gravity tailings sample.

The tailings sample was ground in ball mill and floated in a series of roughers until tailings was visually barren.

1 kg of gravity .tailings as received.

15 minutes/kg in 2 kg lab ball mill.

Stage

Grind Rougher 1 Rougher 2 Rougher 3 Rougher 4


A350

_ 100 50 50 50

R208

-50 30 -

DF250

_ 4 -1

CuSC\

_ -50 50

20 1 50

Metso

H

1500 --

'

Time, minutes

Grind

15 ----.

1 Cond. Froth

mm

3 5 5 3

«. 5 3 3 3

pH

«, 9.8 9.5 8.8 a.3

Note:

Stage Flot. Cell Speed:rpm % Solids

Very heavy weight of silica in 1st rougher concentrate - stopped float and added metso to depress silica with some success.

Rougher 1-4 1000 g D-l 1600 33

m.

27


Purpose:

Procedure:

To investigate the solubility of gold.


Feed: 100 g rougher flotation concentrate




Grind: 1 kg sample ground 15 minutes in ball mill- float—rougher concentrate reground 20 minutes in pebble mill.

Reagent Balance:

Time

Hours

Added, Grams

Actual NaCN Ca(0H)2

Equivalent NaCN CaO

Residual

Grams NaCN CaO

Consumed

Grams NaCN CaO

PH

0-7 7-24 24-41 41-48

0.316 0.221 0.142 0.091

0.135 0.080 0.080 0.040

0.300 0.210 0.135 0.086

0.101 0.060 0.060 0.030

0.09 0.165 0.214 0.251

0.020

0.21 0.135 0.086 0.049

0.101 0.060 0.040 0.050

10.5- 9.9 10.9-10.0 10.7-10.3 10.7-10.3

Total 0.770 0.335 0.731 0.251 0.251 0.48 0.251 10.4


Product

flot. Concentrate Hot. Tailing . Head (Calculated)

Cyanidation Pregnant+Wash Residue Head (Calc.) i

Aimount

10.1 89.9

100.0

1040 mL 101.9 g 101.9 g j

Assays,mg/L,g/t

Au

8.61 1.01

1.78

0.42 4.32 8.61

% Distribution Au

O'all

48.9 51.1

100.0

24.4 24.5 48.9

Ind.

-

-

49.8 50.2 100.0

- 28 -

Test No. 10

Purpose: To investigate the effect of a still finer grind on the recovery of gold from the gravity concentrate by cyanidation.

Procedure: The sample was puLped with water in a two litre bottle. NaCN and lime were added and the cyanidation was carried out on rolls in two 24 hour stages with the solution being changed after each stage. Between each stage the pulp was filtered and the residue washed three times with water. The residue was then repulped with fresh cyanide solution and the test continued.

Feed: 500 g gravity concentrate



pH Range: 10.5 with Ca(OH)2

Grind: Ground 60 minutes in lab pebble mill.

Reagent Balance:

Time

Hours

0-17 17-23 23-40 40-48

Total

Added, Grams

Actua NaCN

1.05 0.447 0.105 0.105

1.652

1 Ca(0H)2

0.325 0.100 0.100

0.525

Equi> NaCN

1.00 0.425 0.100 0.100

1.625

a l e n t CaO

0.244 0.075 0.075

0.394

Residual

Gra NaCN

0.575 0.900 0.900 0.900

0.900

IIS

CaO

0.01

-

Consumed

Gra NaCN

0.425 0.100 0.100 0.100

0.725

ns CaO

0.244 0.075 0.065

0.384

pH

10.5-10 .0 10.4-10.2 10.5-10.4 10 .4-10 .4

10.4


29



Product

1. Pregnant+Wash 2. Residue

Head (Calculated)

Amount

1380 mL 512.4 g

512.4 g

Assays,mg/L,g/t

Au

0.77 1.52

3.59

% Distribution

Au

57.7 42.3

100.0

Screen Analyses

Cyanide Residue

Mesh Size (Tyler)

+ 400 - 400

% Ret Individual

8.4 91.6

Total 100.0

ained Cumulative

8.4 100.0

-


91.6

-

LAKEFIELD RESEARCH A Division of Falconbridge Limited Lakefield, Ontario September 10, 1986 / slk

52NMSEe«43 63.4855 BALMER TWP 0 3 0


Vancouver, B.C.

Gold Recovery Plant

Order of Magnitude Capital and Operating Cost

Campbell Red Lake Mines Limited's

Tailings Deposit

w w WRIGHT ENGINEERS LIMITED VANCOUVER CANADA

PROJECT NO. 1*71-100 OCT. 1986

\X-R10iT ENGINEERS LIMITED VflT Ptwnt 68«-M71 • C»W« "WRIGHTENG- • Tclei 0 4 - 5 0 6 7

1444 Alberni Street, Vancouver, British Columbia, Canada, V6G 2Z4

Project No.: 1471

November 3, 1986

FreeGold Recovery Inc. 1333 West 8th Ave. Vancouver, B.C. V6H 3W4

Attention: Mr. Harry Barr, President

RE: Gold Recovery from Campbell Red Lake Plant Tailing Pond

Dear Mr. Barr:

We are pleased to submit the accompanying report outlining a capital and operating cost estimate for a base case 2500 tons per day gold recovery plant. Other plant capacities have been considered by factoring the costs of the base case. Those costs are based on new equipment, are in order of magnitude and do not reflect any optimization.

We hope this information will help enable you to formulate your next step on this project.

Yours very truly,

WRIGHT ENGINEERS LIMITED,

1&J& Keith Remfert Senior Consultant

KR/mph Encl.

1.

/ 1

CAPITAL AND OPERATING COST

GOLD RECOVERY PLANT

INTRODUCTION

Wright Engineers Limited (WEL) were requested to formulate an order of

magnitude capital and operating cost estimate for the recovery of gold from tailings

using a standard sodium cyanide circuit. This study is a continuation of work relative

to reclaiming the tailings deposited at Campbell Red Lake Mines Ltd.

Information and reports upon which this study is based include the following:

Wright Engineers Ltd. - Review and Verification of Sampling Technique,

September 1986

Lakefield Research - Progress Report //I, September 10, 1986

FreeGoId Recovery Inc. - Report - Drilling and Sampling, Tonnage Estimate,

( August 20, 1986

FreeGoId Recovery Inc. and Lakefield Research - Testwork Results,

August 18,1986

FreeGoId Recovery Inc. - FreeGoId Campbell Tailings Recovery Test,

June, 1986

A preliminary flowsheet was prepared to establish equipment sizes and

numbers. Operating cost data was provided by Campbell Red Lake Ltd. August 19,

1986. Consumption of reagents established during testwork at Lakefield Research was

used.

FLOWSHEET DEVELOPMENT

Preliminary testwork indicates that direct cyanidation of the tailing material

may be the preferred flowsheet. Further grinding of this material to increase the gold

recovery is not economically favourable.

\ * y

2.

The basic flowsheet is given in the Study Basis section of this report.

Grinding circuits were not costed once it was obvious that the net revenue per ton

would be less than direct cyanidation before grinding.

STUDY BASIS

Ton: 2500 t/d Hours (net): 23 hr/d Nominal hourly rate: 110 t/hr Density solids (calc. purpose): 2.8 t/nr»3 Estimated tons available: 3.6 x 10^

Annual tonnage = 875000

Basic flowsheet:

Pond —• Traini Screen

' ' Leach

i i

•

• Sludge * '

f V r \jSl4K UUIlip

Filter

• •

Cla my

—• < Dake Tailing Pond

• Dea er« ixion 1 Filter|~» Refine

Leach:

pH: Time: Pulp density:

10.5 48 hr 60% solids

Base Cost Factors:

Power NaCN CaO Labour

Supervisor Ave. trade

$.09 kWh $1.70 kg $.18 kg

$25.50 loaded/hr $24.20 loaded/hr

CAPITAL COST ESTIMATE

The capital cost for the 2500 tpd plant outlined by the attached flowsheet is

10.0 x 106 dollars Canadian. A breakdown of the equipment is found overleaf.

file:///jSl4K

3.

MECHANICAL EQUIPMENT LIST

BASE CASE 2500 TPD FEED

Quantity Type

1 6' x 16' D.D. Screen

1 125'0 Thickener

6 11m 0 x 11m Cyanide Leach Tank

c/w Agitator

1 Leach Air Blower

2 6" x 6" Horizontal Centrifugal Pump (1 standby)

1 Primary Filter Feed Tank

5 12' 0 x 18' Primary Drum Filter c/w Repulper

1 Secondary Filter Feed Tank c/w Agitator


5 12' 0 x 16' Secondary Drum Filter c/w Repulper


2 42" 0 x 6 ' Primary Filtrate Receiver

1 4" x 3" Filtrate Pump

2 42" 0 Secondary Filtrate Receiver

4" x 3" Filtrate Pump

42" 0x6* Moisture Trap

Vacuum Pump (8000 cfm)

Sump Pump (2" vertical)


Sampler

10' 0 x 10' Pregnant Solution Tank

4" x 4" Horizontal Centrifugal Pump

600 ft2 Pressure Clarifier

3'-0" 0 x 3'-6" Precoat Tank c/w Agitator

2" Horizontal Centrifugal Pump

1)4" Horizontal Centrifugal Pump

Installed Order of Magnitude

Capital Cost (Cdn. $)

36,800

282,000

870,000

65,000

15,300

4,700

750,000

7,100

15,300

750,000

15,300

4,700

5,400

4,700

5,400

4,700

170,000

3,200

3,200

45,000

10,900

6,100

420,000

2,400

2,900

2,500

0.

Quantity

1

1

1

1

2

1

2

1

1

1

1

1

1

TOTAL

Building,

use 2.5 rr

Type

81 0 x 8' Clarified Solution Tank

0' 0 x 10" Deaeration Tower

Zinc Feeder

In-line Pump

2" x 2" x 20 frame Filter Press

(Clean bi-weekly)

12' 0 x 12' Barren Solution Tank

Horizontal Centrifugal Pump

Furnace - Gas Fired 1000// c/w Blower

Baghouse c/w Fan

Acid Leach Tank c/w Agitator

Exhaust Fan

In-line Filter


6" x 6" Thickener Pump

piping, etc. less tailing disposal and

multiplier

powe r supply,

say



5,300

6,100

12,000

5,000

25,000

12,900

12,200

75,000

20,000

10,000

11,000

0,500

6,100

9,500

3,725,200

x2.5

9,313,000

$10,000,000

5.

OPERATING COST ESTIMATE

The basis for operating cost and revenue were established by test datum

originating from Lakefield Resource testwork, report issued August 18, 1986. The four

tests referred to in this estimate are described below. Direct cyanidation of

recovered tailing pond material applies in all tests.

Test 2 No grinding of the plant feed.

Test 3 Grind plant feed to 81.8% -400 mesh (7.25 kWh/t)



Item

Power (Grinding) Power (Agitators) Power (Est. Balance) Sub-Total

Base Reagents NaCN CaO

Sub-Total

Grinding media Sub-Total

Maintenance Supplies; Take 5% of installed mech., divided by annual tonnage

Labour Supervisor Ave Trades

Sub-Total

Pond to Plant Cost (FGRI's suggested cost)

Unit Cost $

.09 kWh

.09 kWh

.09 kWh

1.70 kg • 18 kg

$25.50/hr $24.20/hr

Test 2 Units

0 3.1 3.7 6.8

.42 1.34

1 4 5

Cost

0 .28 .33 .61

.71

.24

.95

0

.21

.23

.88 1.11

1.50

Tes Units

7.25 3.1 3.7

14.05

2.17 .98

1 5 6

t 3 Cost

.65

.28

.33 1.26

3.69 .18

3.87

.35

.35

.21 +

.23 1.10 1.33

1.50

Tes Units

14.5 3.1 3.7

21.3

1.50 .71

1 5 6

t 7 Cost

1.31 .28 .33

1.92

2.55 .13

2.68

.75

.75

.21 +

.23 1.10 1.33

1.50

Test Units

21.7 3.1 3.7

28.5

1.40 .83

1 5 6

8 Cost

1.95 .28 .33

2.56

2.38 .15

2.53

1.10 1.10

.21 +

.23 1.10 1.33

1.50

TOTAL $/TON FEED 4.38 8.52+ 8.39+ 9.23H

REVENUE

6.

Test 2 Test 3 Test 7 Test 8

Recovery %

Head g/t

Recovered g/t

Value @ $18.Cnd/g ($400. U.S./oz.)

Less (Oper. Cost)

Net Sub-Total

41.1

1.94

.797

14.35

4.38

9.97

48.9

1.90

.929

16.72

8.52

8.20

51.9

1.88

.976

17.57

8.39

9.18

53.8

1.93

1.038

18.68

9.23

9.45

Less Cap. & Cost of Cap. (@ 12.5% interest)

Net Revenue (pre Taxes)

Project Life =

* Estimated Capital (2500 t/d)

Interest Cost (12.5%)

Total Capital

$ 3.72* #* «* #«

$ 6.25 /t i.e. $22.5 x 106 over life of the project

= 4.1 years

$10,000,000

3.6 x 10JT 875000 T/yr

3,380,000

$13,380,000 divided by 3.6 x 10^ tons = $3.72

•Grinding circuit concentrator is not considered further due to lower net revenue than established in Test 2 before cost of capital was considered.

CAPITAL COST FOR ALTERNATE FEED RATE

The base case capital cost has been factored using the following formula:

t0.6 C2 = Cj x

where:

( «

c2 Ci

T2

Tl

=

=

=

=

therefore:

Ci

Tl

For T2

For T2 For T2

z

=

=

=

cost of alternate plant

cost of base case plant

tonnage of alternate plant

tonnage of base case plant

$10,000,000

2500 TPD

1000 C 2 = $5.8 x 106

5000 C 2 = $15.2 x 106

10000 C2 = $23.0 x 106

Project Life

10.3 yr

2.1 yr

l y r

OPERATING COST FOR ALTERNATE FEED RATE

8.

TPD =

Power $/T

Reagents $/T

Supplies $/T

Labour $/T

Supervisor

Trades

Feed Costs $/T

Sub-Total Costs

Revenue $/T

Sub. Net $/T

Cost of Cap. $/T

Net $/T

2500

.61

.95

.21

-

.23

(4) .88

1.50

4.38

14.35

9.97

3.72

6.25

1000

.61

.95

.08

-

.5%

(4) 2.20

1.50

5.92

14.35

8.43

2.88

5.55

5000

.61

.95

.43

-

.12

(4) .44

1.50

4.05

14.35

10.30

4.85

5.45

10000

.61

.95

.86

-

.06

(6) .33

1.50

4.31

14.35

10.04

6.82

3.22

CONCLUSIONS

From the information available the following points apply:

1. Grinding of tailings for increased recovery is not economical.

2. The more favourable tonnage rates to be considered are 1000 tpd and

2500 tpd. The decision here is related to length of term and original capital

invested. The operating cost differential is principally the labour fraction

and the cost of capital.


Vancouver, B.C.

Gold Recovery Plant

Order of Magnitude Capital and Operating Cost

Campbell Red Lake Mines Limited's

Tailings Deposit

w w WRIGHT ENGINEERS LIMITED VANCOUVER CANADA

PROJECT NO. 1471-100 NOV. 1986

CAPITAL AND OPERATING COST

GOLD RECOVERY PLANT

INTRODUCTION

Wright Engineers Limited (WEL) were requested to formulate an order of

magnitude capital and operating cost estimate for the recovery of gold from tailings

using a standard sodium cyanide circuit. This study is a continuation of work relative

to reclaiming the tailings deposited at Campbell Red Lake Mines Ltd. and an

addendum to report issued in November 1986.

Information and reports upon which this study is based include the following:

Wright Engineers Ltd. - Review and Verification of Sampling Technique,

September 1986

Lakefield Research - Progress Report //I, September 10, 1986

FreeGold Recovery Inc. - Report - Drilling and Sampling, Tonnage Estimate,

August 20, 1986

FreeGold Recovery Inc. and Lakefield Research - Testwork Results,

August 18, 1986

FreeGold Recovery Inc. - FreeGold Campbell Tailings Recovery Test,

June, 1986

Preliminary test results supplied by FreeGold Recovery Inc. done at Hazen

Research, October 1986.

A preliminary flowsheet was prepared to establish equipment sizes and

numbers. Operating cost data was provided by Campbell Red Lake Ltd. August 19,

1986. Consumption of reagents established during testwork at Hazen Research was

used. The principal difference between this report and that of November 3, 1986 is

the reduction of leach retention time from 48 hours to 8 hours, and lower head grade.

FLOWSHEET DEVELOPMENT

Direct cyanidation processing of the tailing material was used as the base

case.

w

2.

The basic flowsheet is given in the Study Basis section of this report.

STUDY BASIS

Ton: 2500 t/d Hours (net): 23 hr/d Nominal hourly rate: 110 t/hr Density solids (calc. purpose): 2.8 t/m3 Estimated tons available: 3.6 x 10^

Basic flowsheet:

Annual tonnage = 875000

Pondj—* Trasn Screen

i •

Leach l

• Sludge 4

O size aump

Filter

•

Clarify

-Cake

—1

-—• Tailing Pond

Deaeration Filter —* Refine

Leach:

pH: Time:

10.5 8hrs

Pulp density: 60% solids

Base Cost Factors:

Power NaCN CaO Labour

Supervisor Ave. trade

$.09 kWh $1.70 kg $.18 kg

$25.50 loaded/hr $24.20 loaded/hr

CAPITAL COST ESTIMATE

The preliminary capital cost estimate for the 2500 tpd plant outlined by the

attached flowsheet is 7.0 x 10*> dollars Canadian. A breakdown of the equipment is

found overleaf.

3.

MECHANICAL EQUIPMENT LIST

BASE CASE 2500 TPD FEED

Quantity Type

1 6'x 16' D.D. Screen

1 125' 0 Thickener

6 6.5m 0 x 6m Cyanide Leach Tank

c/w Agitator

1 Leach Air Blower


1 Primary Filter Feed Tank

3 12' 0 x 18' Primary Drum Filter c/w Repulper

1 Secondary Filter Feed Tank c/w Agitator


3 12' 0 x IZ' Secondary Drum Filter c/w Repulper


2 42" 0 x 6 ' Primary Filtrate Receiver

1 4" x 3" Filtrate Pump

2 42" 0 Secondary Filtrate Receiver

4" x 3" Filtrate Pump

42" 0 x 6" Moisture Trap

Vacuum Pump (8000 cfm)



Sampler

10' 0 x 10' Pregnant Solution Tank

4" x 4" Horizontal Centrifugal Pump

600 ft2 Pressure Clarifier

3'-0" 0 x 3'-6" Precoat Tank c/w Agitator

2" Horizontal Centrifugal Pump

lYz" Horizontal Centrifugal Pump



36,800

282,000

372,000

50,000

15,300

4,700

450,000

7,100

15,300

450,000

15,300

4,700

5,400

4,700

5,400

4,700

170,000

3,200

3,200

45,000

10,900

6,100

420,000

2,400

2,900

2,500

li.

Quantity

1

1

1

1

2

1

2

1

1

1

1

1

1

TOTAL

Building,

use 2̂ 5 rr

Type

8' 0 x 8' Clarified Solution Tank

4' 0 x 10' Deaeration Tower

Zinc Feeder

In-line Pump

2" x 2" x 20 frame Filter Press

(Clean bi-weekly)

12' 0 x 12* Barren Solution Tank


Furnace - Gas Fired 1000// c/w Blower

Baghouse c/w Fan

Acid Leach Tank c/w Agitator

Exhaust Fan

In-line Filter


6" x 6" Thickener Pump

piping, etc. less tailing disposal and

multiplier

power supply,

say



5,300

6,100

12,000

5,000

25,000

12,900

12,200

75,000

24,000

14,000

11,000

4,500

6,100

9,500

2,612,200

x2.5

6,530,500

$ 7,000,000

5.

SAMPLE DATA

The basis for operating cost and revenue were established by test datum

originating from Hazen testwork, preliminary report issued November, 1986.

Bulk samples taken from the Campbell Red Lake mine tailing pond were

combined and mixed for cyanide leach test. The samples came from three locations of

//2 test area shown on the attached location map. FreeGold Recovery Inc. supervised

the sampling and sample preparation and describe the samples as follows:

above.

Sample

1

2

3

Location

Line 4450 Hole 8

Line 6200 Hole 7

Halfway between Hole 4 on Line 5200 and Hole 4 on Line 4950

Depth

18'

12'

12'

Amount

2 barrel volume

2 barrel volume

2 barrel volume

The composite sample head assay reported was .040 oz/t (1.24 g/t).

The following assays were provided for the individual drill holes addressed

Location Depth Assay oz/t July 1986

Line 4450 - Hole 8

Line 6200 - Hole 7

Line 5200 - Hole 4

Line 4950 - Hole 4

Cyanide leach tests were performed on the combined sample as described in

the report from Hazen Research Inc. overleaf.

26'

30'

22.5'

14.5'

.053

.043

.058

.084

Objective: Cyanide le

Saxtple: HRI 33738,

Time, hr pH, Initial

adjusted Temperature % Solids Ket Pulp Kt, s KaCN, g/l

Reagent Additions, grams CaO HaCN

Cumulative Consumptions: lb/ton of test feed

CaO HaCH

wt, g or

vol, ml

2 hour aliquot 50 4 hour aliquot SO R hour aliquot 60

2< hour aliquot 60

ech.

, 6 rash ore feed.

0 8.6

11.0 amb

30 1667

0.27 ?.33

Analyses, (1)

Au

0.1? O.IR 0.18^ 0,21

2 10.9

amb 30

1666 1.93

1.1 0.31,

Di

4 10,9

amb 31

1611 1.B7

l . l 0.62

ssolutton oz Au/t

0.012 0.012 0.013 0.016

e 10.0

a tib 31

\m 1.84

#

1.1 0.47

»

• 24 10.9

amb 32

1673 1,07

1.1 3.60

Test Ko. Project: Date Paget

•

•

Dissolution, t Au

30.0 30.0 32.5 37.6

2 008-112 Oct '86

1 of 1

24 hour F/W (3) 1*00 0.15 24 hour solids 499.0 0,026

0.014 35.0

Calculated feed Assayed feed

0.040

(1) Solids analyses for gold and/or silver arc reported as troy ounces per Short ton. and other metals as percentages. Liquor analyses are reported as mg/l.

(2) The percentage dissolutions ere based upon the calculated ffred(s). (3) Combined final pregnant and xash liquors.

Remarks:

6.

OPERATING COST ESTIMATE

Item

Power (Agitators)

Power (Est. Balance)

Sub-total

Base Reagents

NaCN

CaO

Sub-total

Maintenance Supplies; take 3% of

installed mech., divided by

tonnage

annual

Unit Cost $

.09/kWh

.09/kWh

1.70/kg

.18/kg

Units

kWh/t

1.2

3.7

U.9

kg/t

.55*

.50

Cost

.11

.33

Ak

.93

.09

1.02

.09

**Labour (Total averaged/hour) $77.52/hr .70

Pond to Plant Cost

(FGRI's suggested cost) 1.50

TOTAL $/TON FEED 3.75

* The NaCN consumption data is suspect. Hazen Research indicated

problems with the titration procedure during analysis.

** Reduced to a minimum crew of 13 total.

• 7.

REVENUE

Recovery %

Head g/t

Recovered g/t

Value @ $18.Cnd/g ($400. U.S./oz.)

Less (Oper. Cost)

Net Sub-Total

Less Cap. & Cost of Cap. {(§ 12.5% interest)

Net Revenue (pre Taxes)

Project Life =

* Estimated Capital = (2500 t/d)

Interest Cost (12.5%) +

Total Capital

COMMENTS

The following points apply:

a) Head assay results from the bulk composite range from .036 to .045 oz/t

compared to drill core assays from the same pits of .043 to .084 oz/t.

b) Hazen Research Laboratory recovery results for direct cyanidation for

24 hours is 37.5% compared to Lakefield's 38.1%.

c) Operating costs do not include costs for administration, maintenance shops,

mobile equipment, laboratory work or the contracting out of these services.

32.5

1.24

.403

7.25

3.75

3.50

$ 2.50*

$ 1.00 /t i.e. $3.6 x 106 over life of the project

3.6 x 10JT ., , 875000Wr = *#1 y e a r $

$7,000,000

2,006,000

$9,006,000 divided by 3.6 x 10^ tons = $2.50

l u 4

fc

I •

63.4855 BALMER TWP ,_, .. _

0-40

PHASE I REPORT

ON

FREEGOLD RECOVERY INC/CAMPBELL RED LAKE MINES

TAILING RECLAIMING PROGRAM

BY: Kelly Dolphin Vice-President Freegold Recovery

52N«4SEee43 63.4855 BALMER TWP 040C

CONTENTS

•.

Introduction

Drilling Procedure

Bulk Sampling Procedure

Tonnage Calculation

Summary

Recommendation

Test Area #1 Drill Logs

Test Area #2 Drill Logs

Screen Sizing Results >

Test Plant Results

References

PAGE :•

1

1 - 4

4 - 5

5 - 6

6

7

Appendix

Appendix

Appendix

Appendix

8

t

A

B

C

D

0

INTRODUCTION

The tailings area are divided into two areas. Test

area #1 which was the decant pond was drilled ,first due

to the fact that C. R. L. M. already have engineering plans

in progress to .extend the north dyke over the center of

test area #1 (See Fig. A).

Test area #2 which is considered to be the coarse

tailings contained in the dyking system. This tailings

area used to be the primary pond for the mine until

abandoned in 1984.

• DRILLING PROCEDURE

A vi-cor sonic drill was used to obtain the tailings

sample material. A grid spacing on test area #1 was used

consisting of 250 foot lines with hole spacing on each

line of 250 feet for a total of 28 holes drilled. The

same grid pattern was used for test area #2 for a total

of 69 holes in test area #2. All holes were drilled

within 1' of surveyed location unless otherwise specified

and were drilled completely through tailings.

A 4" casing was driven through the tailings with a

sonic head driven by a 7 h.p. Honda gas power motor. A

flex cable is used in connecting the head,to the power

pack. Drill casing was broken into 5. foot lengths, this

made the handling of the large volume of drill cores

easier on the drill operators.

- 3 -

The sonic.head vibrated the drill casing to the bottom of

the tailings. ; Bottom of tailings was usually noted as

blackish lake bottom type material with dead reeds visible

in it. A 12,000 lb. winch was then used to pull lip the

casing. The total length of drill casing was then measured

(D). When the holes is drilled the cased material is

compressed. This length is also measured (C). Because the

core liquified and compacted in the drill casing, the sample

was never as long as the actual drill length. To adjust for

this the following formula was used:

Example: Line #0+310 Hole #3 -

Total Drill Length (D)

Compacted Core Length (C) '

Ratio C/D = 15/12.5 =0.83

Drill cores were removed and placed in split 6" PVC

pipe used as core boxes. Drill cores were then sampled

every 4" using a 5/8" circular tube.

Drill samples were logged and assigned a number then

sent to the Campbell Red Lake Mines Assay Office for assay.

Remaining drill cores were then contained in 5 gal. buckets

for future analysis.

Freegold recovery retained Wright Engineering of

Vancouver to oversee the sampling procedures. Mr. Stu

Andrews was on sight for one day to visually inspect the

sampling procedures, he sampled independently and took the

sample with him for Assay (see Figure B). He agreed with all

sampling procedures and was satisfied with the drilling and

testing program. He was not satisfied with the material

- 4': -

He suggested the'following procedures for Freegold and

C.R.L.M. to use in1 determining true in place material.

density.

Step 1: Drill sample, measure drilled-core .length.

. and compacted core length.

Step 2: Pull drill core and weigh drill cores in casing.

Step.3: Transport drill core in casing to mill lab. Pull

drill core for a moisture content test and dry

weight.

He selected 3-locations within the two test areas to perform

the density test line 0+310 hole #5, Line-4 950 E. Hole #7

and Line 6200 E. Hole #6. Results will be forthcoming from

this test.

BULK SAMPLING

A number of bulk samples were run through Freegold

Recovery Test Plant. The plant consists of a 1' x'3* aries

screening unit with 10 mesh decking. Minus 10 mesh feed is *

then laundered into a salo SPV 181 slurry pump. Slurry feed

is then pumped into a Reichert Mark 7 L.G. spiral concentrator.

The spiral produces 3 splits, concentrates, middling and

tailing. The qurrent flow sheet is to have the tailing being

discarded out of the system. Middlings are being recycled

back into the head feed. Concentrates are then fed onto a

Gemen'i #250 finishing table. The table produces 4 splits,

Free Au, Table Con, Table Mids and Table Tails.

- 5 -

. There were a total of 8 samples splits collected per

bulk sample processed through the test plant. Composite

samples from each bulk sample were acquired by driving a

3/4" soil sampler into each 5 gal. bucket, as it's fed

to the test plant. The composite sample was then cross

checked by obtaining a composite slurry sample from the

screeners underfeed. Approximately every minute product

samples from the 3 spiral splits were obtained. All

table splits were bagged and tagged for transportation

to the mill lab. Test results from this preliminary

bulk sample were rather inconsistent. There was no •

product balancing or product performance adjusting

conducted in these test runs. This made it rather difficult

to account for and any recovery percentages or performance

ratios from the test plant. Part of Phase II program will

be to conduct ratios of Freegold test plant.

TONNAGE CALCULATION

Volume tonnage aalculations were performed by the

Engineering staff of Campbell Red Lake Mines. Two

different seniraos were used in the calculation. The first

one was calculated using a planimeter area multiplied by

the average depth of the test holes and divided by 19.0 cu.

ft./ton which is likely maximum volume. The second approach

was using the pit volume program. All results were within

10%.

- 6 -

1.2 million tons

3.3 million tons

4.5 million tons

Average grade = (1.2 x 10" x 0.046) (3.3 x 106 x 0.053)

4.5 x 106

= .051 oz./ton

0.051 oz./ton x 4.5 x 106 Tons = 230,100 oz./Au

SUMMARY

The drilling of the tailing areas went very well, once a

the drill crew modified the core catcher for the sonic drill.

Much thanks and appreciation must go Campbell Red Lake

Mines Surface Shops and people for their help and contributions

in developing the new drill tooling.

Freegold Recovery retained Lakefield Research to conduct

bench scale lap tests. Two 20 lb. samples were sent to

Lakefield by myself. One sample contained 28 different assay

samples from our drilj. cores. Lakefield was to select 5

samples for check analysis. The remainder of the samples were

to be combined for test work. The second sample was a gravity

con produced with one pass through the Reichert MK 7 spiral.

Lakefield is to examine two possible flowsheets for gold

recovery.

Flowsheet #1: Grinding and cyanidation

Flowsheet #2: Flotation and cyanidation of cons.

Total Tons - Area #1

Area #2

TOTAL

- 7 -

RECOMMENDATION '•

1. Bring Wright Eng. into the Phase II Program as soon as

Lakefield produces their test results. Wright can then

product a preliminary feasibility study on cap. cost

and overhead operating cost on a production plant.

2. Complete a screening analysis and size fraction test

on the tailing material. This will then determine the

approach that might be taken with regards to classification

and gravity recovery.

3. Backfill ammendability testing, part of this will be

completed in recommendation #2. An already slurryed

product could be classified and transported relatively

cheaply. This would help off-set some of the operating

cost on the production plant. This part of the project

would potentially open up new areas within the tailing

containment dyking area.

4. Talk with the regulatory boards and obtain an

understanding of their position: and concerns. This

will enable us to make the wisest decision with regards

to plant set up and operation.

5. Engineering and performance testing of the tailings

with the latest technology and equipment in grinding,

classification and recovery.

Kelly Dolphin, Vice-President, Freegold Recovery. .

KD*cy August 20, 1986.

8 -

R E F E R E N C E S

CAMPBELL RED LAKE MINES LIMITED

WRIGHT ENGINEERING

FREEGOLD RECOVERY INC.

t

• «

! •

l i r •

APPENDIX 'A • M

£

.b

APPENDIX 1 B m »

L

L'<

L:

APPENDIX .'C.'

L

L

L APPENDIX 'D m i

L

B

S2N84SE«ilM3 63.4855 BALMER TWP · using a 4" diameter aluminum core barrel or casing and a steel...

Documents

Transcript of S2N84SE«ilM3 63.4855 BALMER TWP · using a 4" diameter aluminum core barrel or casing and a steel...