Leaching of a silver bearing sulphide by-product with cyanide, thiourea and chloride solutions

Minerals Engineering, Vol. 8, No. 3, pp. 257-271, 1995

Pergamon 0092-6875(94)00124-3

Copyright © 1995 Elsevier Science Ltd Printed in Great Britain. All rights reserved 0892-6875/95 $9.51)+0.00

LEACHING OF A SILVER BEARING SULPHIDE BY-PRODUCT WITH CYANIDE, THIOUREA AND CHLORIDE SOLUTIONS

M.F. ALMEIDA§ and M.A. AMARANTEt § Faculdade de Engenharia da Universidade do Porto, Rua dos Bragas, 4099 Porto Codex, Portugal t Instituto Geol6gico-Mineiro, Rua da Amieira, 4465 S. Mamede de Infesta, Portugal (Received 17 October 1994; accepted 11 November 1994)

ABSTRACT The aim of this research was to increase the value of sulphide tailings assaying 1000-1500 g/ton Ag per tonne and high arsenic (about 13%) content. The sulphides were characterized and it was verified that silver occurs both as mathildite, AgBiS2, and in solid solution in galena. Three alternative routes were assessed: cyanidation, thiourea leaching and chloride leaching. The cyanidation route showed that there was a high consumption of cyanide and led to the formation of an insoluble compound of silver thiocyanate. The leaching with thiourea was achieved without the introduction of Fe 3+ but it is not an economically viable route. Chloride leaching of the product in its original particle grain size showed very good extraction even without addition of Fe 3+. Operating under [Cl-]>_.3 M at 80-90°C for at least 3 h it was possible to dissolve 60-80% of the total silver, and higher recoveries of lead and bismuth were also achieved.

Keywords Hydrometallurgy, leaching, silver, sulphide, cyanide, thiourea, chloride

INTRODUC~ON Vale das Gatas mine is situated in the North of Portugal, its main activity being the production of W and Sn concentrates by means of gravity concentration, from which result a sulphide tailing containing between 0,10 and 0.15% Ag. There are approximately 4000 tonnes of this product on a dump. As about 1% sulphide is contained in the ore, two more tonnes of sulphides per day could be produced if the available treatment capacity was used. In an attempt to sell this tailing a flowsheet was established based on sulphide flotation, producing a concentrate with about 6 kg Ag/ton, yielding 10-12% and a silver recovery of about 80% [1]. However,

257

258

M.F. ALMEIDAand M. A. AMARANTE

at this time the silver price decreased and this treatment process was ruled out. The objective of thig research was to study the possibility of using a hydrometallurgical flowsheet at the mine to treat the two types of sulphides, i.e., the 4000 ton on the dump and the sulphides from normal activity of the mine. Considering the existing quantities and the plant capacity, man-power and equipment, this flowsheet should be simple, therefore cyanidation followed by cementation or electrolysis could be feasible. Although not leading to a very high Ag recovery, cyanidation can be applied at ambient temperature, it does not require any special equipment and can therefore be quickly implemented. As this method did not prove successful, thiourea and chloride leaching were tried as alternatives to cyanidation.

EXPERIMENTAL Following a mineralogical characterization of the two types of sulphides, the first phase of the testwork was aimed at the selection of the most promising methods. The second phase objective was to optimize the operating conditions of the methods selected, that is, cyanide, thiourea and chloride leaching, for economic feasibility. Commercial reagents and tap water were used for all the testwork at a 1:3 solids:liquid ratio and a 100 or 200 g feed mass. The pulp was agitated in beakers and all the tests were controlled by sampling at predetermined time and the residue was washed with constant volumes of a fixed pH solution. The recovery was based on the metals content of the sulphides and on the metals found in the final solutions. The residue content was checked and little discrepancy was found. The metallic elements were analyses by atomic absorption and the free cyanide by titration with silver nitrate using potassium iodide as indicator. Lime was used in cyanidation tests to raise the pH to an alkaline level. Sulphuric acid was added to adjust pH in the experiments with thiourea and chloride, these being carried out usually at 60-90°C, in which case beakers were covered with cellophane in order to prevent evaporation. Dry milling to less than 106 lam was used to test the influence of grain size on leaching, except in pilot scale tests for which the ore was wet milled in a rod mill in view of its quantity. Roasting was done in an electric furnace by carefully spreading the material on the hearth.

RESULTS AND DISCUSSION Characterization There are two types of sulphide tailings, Type A - - deposited in an old dump and Type B - - deposited in a more recent one, with chemical composition shown in Table 1. The particle grain sizes were of 70% and 50% greater than 212 btrn, respectively. An increase in silver content was observed as the particle size decreased. For example, the fraction below 150 lava assayed 2.2 kg Ag/ton. The following minerals were identified by both optical and SEM microscopy: arsenopyrite, pyrite, marcasite, sphalerite, galena, mathildite, other Ag+Bi sulphosalts (pavonite, neyite), native bismuth, bismuthinite, chalcopyrite and stannite. The dominant sulphides were arsenopyrite and the iron, zinc and lead sulphides [21.

Leaching of a silver bearing sulphideby-product

259

Silver usually associates with Ag+Bi sulphosalts of which mathildite (AgBiS2) is the most abundant and it also appears with galena (PbS). The mineralogical examination indicated that galena and mathildite occur in approximate quantities of 2 and 0.3%, respectively. As microanalysis indicates galena to contain only 2% Ag, and mathildite assays 30% Ag, about 70% of the total Ag in the ore appears to be in mathildite, therefore it is essential to promote its chemical degradation in order to obtain a high silver recovery. Moreover, the Type A sulphides were much more weathered than Type B sulphides by reason of their more ancient deposition leading to inter- and intragranular cracks, mainly in the neighbourhood of distinct mineralogical species. This was not evident in Type B sulphides [3]. These differences had already been shown by different flotation conditions and results: Type A sulphide requires sulphidization before the addition of flotation reagents Na2CO 3, NaCN, PAX and flotol, it requires much more flotation time than Type B sulphide and for the same final recovery of about 80% a concentrate with only 3.2 kg Ag/ton is obtained Ill.

TABLE 1 Chemical composition of sulphide samples (%) Type

Ag

J Bi

Pb

] Cu

J Zn

Fe

A

0.10

0.3

2.4

1.5

3.4

20.9

19.4

9.4

B

0.16

0.5

4.7

1.5

3.4

20.2

26.3

12.0

As

Screening Tests These included cyanidation of unmilled feed; cyanidation after milling; cyanidation after roasting; cyanidation after alkaline and acid leaching; thiourea leaching; thiosulphate leaching; chloride leaching; and others [3]. The main indications from these tests were as follows: - - cyanidation of the unmilled feed for 24 hours, with KCN concentration (2.5%) higher than usual, leads to less than 20% of the silver being dissolved for the Type B sulphide and slightly over 20% for Type A;

- - by contrast, after milling (to 90% minus 48 pm, 60% minus 24 pm, 40% minus 16 tam and 22% minus 8 pm) and leaching with a 0.8% KCN solution, recoveries of both sulphides were improved (flotation of the sulphide A cyanidation residue, adding only ZnSO 4 and KAX, gave a concentrate assaying about 4.5 kg Ag/ton, thus corresponding to a global recovery of 75% Ag contained in the feed);

- - conditioning the pulp with lime, or treating with hot caustic solution (10 M) or ammonia did not improve the recovery by cyanide leaching; - - roasting for 30 minutes at temperatures in the range of 450 to 750°C slightly improves the Ag recovery for Type A sulphides but can drastically decrease the Ag recovery from Type B, specially for temperatures over 700°C; - - some other prior treatments seem to improve silver dissolution by cyanidation significantly: nitric acid oxidation, oxidation in 1 M sulphuric acid solution containing ferric sulphate (silver recovery on Type B sulphide increased to >40%), melting the feed with sodium carbonate (1:2) (80% of the silver was dissolved by cyanidation of the milled melted product); - - prior treatment of sulphides with H2SO 4 or HC1 hot solutions does not lead to any improvement on silver dissolution by cyanidation;

260

M.F. ALMEIDAand M. A. AMARANTE

- - concentrated (2-10% of the feed added) thiourea solutions at 60-80°C, can solubilise about 70% of the silver contained in Type B sulphides; - - a solution containing 3 M NaCI + 0.5 M FeCI 3, working at pH
D E V E L O P M E N T OF L E A C H I N G R O U T E S From results obtained in the preliminary tests it was decided to further develop the following routes: cyanidation, thiourea and chloride leaching.

Cyanidation In the following a general flowsheet is evaluated involving cyanidation of the milled Type A sulphides followed by flotation of the residue and electrolysis and recirculation of the solution. Initially some laboratory tests were performed to assess the influence of reagent concentration on silver recovery and the free cyanide evolution with time. Figure 1 shows the results obtained from which the following, can be concluded: - - the highest Ag recovery is directly dependent on the initial cyanide concentration of the solutions; - - after the first 4 hours of agitation there is no significant improvement in silver recovery;

6o~ -

%

50

1.2

Initial KCN concentration 0.1% KCN

--L=- 0.6% KCN

A g 4ot~k~_~ r ~ 1 eC 8o

~<

1% KeN

¢

0.3% KCN

1

~ ~

% o.8 f 0.6 er e

O

~

v2°I~rer ~

/°'4 ~N

p,

[

0

0 0

5

10

15

Time, h

20

Fig. 1 Silver recovery and free cyanide versus time of leaching

25

Leaching of a silver bearing sulphide by-product

261

- - for cyanide concentration greater than 0.3%, about 20% of the silver is instantaneously dissolved, regardless of the initial concentration of the solution used; - - after an initial period of about I hour, the free cyanide in solution could be described by a linear law of the type [KCN] t = a + bt where t is the time measured from the beginning of the test and a and b are related to the solution initial cyanide concentration [4]; - - for an initial 0.3% KCN concentration, the dissolution kinetics seem to decrease at about 0.15% KCN; this value should therefore be considered when controlling cyanide addition at 0.3% initial concentration. Based on the above information it was decided to carry out a test with 100 kg of Type A sulphide which had been wet milled in a batch rod mill. The solids were washed twice with tap water, then a 50% solids pulp was prepared to which 1.5 kg of lime and 300 g of cyanide were added. Figure 2 shows silver and free cyanide concentration in the solution against time of leaching (curves 1). Silver concentration as a function of the number of times that the final solution passed through an electrolytic cell is also shown in this figure (curve 2). 350

0.35 •

+

=1

0.3

%

250

0.25

f

g

2O0

0.2

e

P

150

0.15

m

100

0.1

300~

A

r

e N 1

50

0

0

~_______~

I

I

I

I

1

I

1

2

3

4

5

6

0.05

N

0

7

Time, h / n. o f f l o w s through 1Dissolution

b 2 Electrolysis

Fig.2 Silver and free cyanide in leaching and electrolysis The results show that a maximum silver recovery of about 27% was achieved after 2 hours (>300 ppm Ag in the solution), but, while decanting the solution, about 15% of the dissolved silver precipitated as silver thiocyanate (in this case the final Ag concentration was slightly over 250 ppm). The leach residue was washed using the solution from the electrolysis step. Initial Ag concentration of this solution decreased from 135 ppm to 60 ppm after 4 hours and to 30 ppm after 7 hours of agitation, again due to silver thiocyanate precipitation. The 100 kg sample of Type A sulphide was subjected to 6 leaching cycles followed by electrolysis, after which 50 g of silver were deposited and 900 g NaCN and 4 kg of lime were consumed. However, despite the amount of lime added, the pulp pH was only 9.4 and was still decreasing. ME 8:3-B

262

M.F. ALMEIDAand M. A. AMARANTE

It was therefore decided to allow for pH stabilization, requiring another experiment in which the sulphides were washed with fresh water several times followed by the addition of 6 kg of lime and 2 days of conditioning, 300 g of cyanide then being added to the 100 ~ of solution. However, after 24 h of leaching, and despite the fact that the titrated free cyanide concentration was above 0.3% NaCN, the solution only contained traces of silver. In an attempt to explain this observation some of this solution was collected for leaching 100 g of fresh dry milled sulphides of Type A under the conditions of an usual laboratory test. However, because only silver traces were detected in the final solution, a 0.1% NaCN addition was made. After 24 hours of leaching 36% of the silver was dissolved. A test was performed by mixing this solution with an approximate volume of that resulting from leaching the last 100 kg of sulphides. An abundant white precipitate was obtained which was identified by X-ray diffraction as essentially AgSCN with some Ca(SCN)2.3H20. In conclusion: fine wet milling the feed increases significantly sulphide reactivity by helping the formation of sulphates, which react with cyanide producing thiocyanate whose stability is increased in the presence of ferric ions [5]. As a consequence, lime consumption is multiplied by four, and high insolubility of silver thiocyanate reveals cyanidation as an inappropriate method of extraction. It should also be noted that the free cyanide analysis via silver nitrate titration is no longer applicable in such circumstances, because SCNwill also react with silver. Thiourea

Leaching

These tests were carried out with Type B sulphides usually dry milled to minus 64 pro. It was also decided to test A sulphides only after assessing the economic feasibility of leaching B sulphides with thiourea. Table 2 shows the effect of temperature and ferric sulphate on thiourea leaching when 100 g of sulphides were tested with 300 m~ of solution at pH50°C in order to dissolve more than 40% of silver in the sulphides;

--

- - the detrimental influence of ferric sulphate in the recovery. Figures 3 and 4 show the thiourea and time effects at temperatures of 80 and 70°C. The beneficial influence of increasing thiourea addition on silver dissolution, can be seen at least up to 4 g of thiourea, and it is also seen that silver losses occur in the solution after a certain leaching time. Figure 5 shows the results obtained in 8 tests carried out between 60 and 90°C, using 3 g thiourea solutions in the presence of no ferric sulphate or 2 g. In order to evaluate the dissolution capacity of the above solutions, 100 g of fresh B sulphides were added to the pulp after 4 hours of agitation (note that the 6.5 hours value corresponds to the final volume including wash solution). The conclusions from these tests are similar to those obtained before. However, the effect of ferric sulphate at 90°C seems to be less unfavourable than at other temperatures. It can be observed that the operating conditions influence the beginning of precipitation that usually occurs between 1 and 2.5 hours after starting agitation. Also, after 4 hours, the dissolution decreases significantly in all cases.

Leaching of a silver bearing sulphide by-product

263

TABLE 2 Factors influencing recovery of silver via thiourea leaching

Thiourea, g

Sulphate, g

Temperature, °C

Ag recovered, % i

50

2

3

32.1

50

2

3

32.6

50

2

3

32.1

50

2

3

32.1

30

2

5

22.0

30

3

5

18.9

30

2

3

6.1

30

3

3

5.0

50

3

5

38.6

50

2

5

24.6

5O

I

5

37.5

50

3

1

1.9

50

5

2

7.9

70

1

3

49.0

70

2

3

40.1

70

3

5

58.0

5.2

70

250 200

A g P P m

+

150

4-

+

2 g ferric sulphate 100

70

C

50 0 0.5

I

J

1

1.5

J

I

2 2.5 Time, h

4 g thiourea

i

_.

3

1

3.5

3 g thiourea

Fig.3 Influence of thiourea on silver recovery

4

264

M. F, ALMEIDA and M. A. AMARANTE

60

%

_

1 g ferric sulphate

_

2 hours

50

A g

40

80

r e c o v e r e

C

30 20 lO

d o o

J

]

'

I

I

L

0.5

1

1.5

2

2.5

3

~

3.5

J

]

4

4.5

5

Thiourea, g Fig.4 Silver concentration versus time of leaching for 2 levels of thiourea addition

%

50 ~ ~ _ A

A g

40

r

30

e c o v e

20

r

3 g thiourea

A

~ X

10

e

d

0 ~

0

0.5

1

1.5

2

2.5

3

3.5

4

4.5

5

5.5

Time, h

×

60, 0

!

60, 2

>~

70, 0

[]

70, 2

80, 0

o

80, 2

A

90, 0

X

90, 2

Leaching temperature,

C,

Ferric sulphate, g

Fig.5 Silver recovery dependence on time, temperature and ferric sulphate addition

6

6.5

Leaching of a silver bearing sulphide by-product

265

The instability of thiourea which is oxidized to formamidine disulphide, then cyanamide, sulphidric acid and sulphur [6,7,8] is increased by ferric ion, which is usually necessary to balance the reaction redox potential as in the case of Au dissolution shown in equation 1 where E°=+0.04 V, 2Au + 2CS(NH2) 2 + NH2(NH)CSSC(NH)NH 2 + 2H + = 2Au(CS(NH2)2)2 +

(1)

Another contribution is the degeneration of thiourea in a sulphate medium that is followed by the transformation of the resulting product in FeSO4(CS(NH2))2+ [6] as shown by reaction (2): Fe 3+ + SO42- + CS(NH2) 2 = FeSO4(CS(NH2))2 +

(2)

Under the same leaching conditions the results also show that the recovery achieved with Type B sulphides after milling to minus 64 prn is greater than with the unmilled product, but less than that obtained on Type A under the same conditions. However, for the recoveries obtained, it is easy to conclude that the reagent costs exceed the actual silver market value, and thus this process would not be economically viable in any case.

Chloride Leaching The mineralogical information indicates that most of the zinc is in sphalerite, lead is in galena, copper is in chalcopyrite, bismuth is in mathildite and native bismuth, and the silver is associated with mathildite and galena. Therefore, the approximate percentage of each mineral in the Type A sulphides is 5.06% ZnS, 2.77% PbS, 4.33% CuFeS 2 0.27% AgBiS 2 and 0.16% Bi. The probable reaction of chalcopyrite with ferric chloride proceeds according to equation (3) [9], CuFeS 2 + 3FeC14- = CUC132- + 4FeC12 + 2S ° + CI-

(3)

but the following reactions can also occur, CuFeS 2 + 4FeC13

> CuC12 + 5FeCI 2 + 2S °

(4)

PbS + 2FeC13

> PbC12 + 2FeCI 2 + S °

(5)

ZnS + 2FeC13

> ZnC12 + 2FeCI 2 + S °

(6)

AgBiS 2 + 4FeC13 Bi + 3FeCI 3

> AgC1 + BiC13 + 4FeCI 2 + 2S ° > 3FeC12 + BiCI 3

(7) (8)

The stoichiometric amount of ferric chloride necessary to oxidize the minerals in sulphides is 612.3 g of FeCI3.6H20 per kg of Type A sulphides. It was decided to test the influence of the oxidant, the temperature and the time on the dissolution of the main elements in the sulphides by an experimental factorial design at T = 80, 90 and 100°C and t = 1/2, 1 and 1.5 hours, Because it is well known that chloride leaching implies the use of solutions where chloride concentration is >3 M (most of the metals are present there as anions [9]), the 3 M [C1-] solutions used were obtained by adding 5 ml of H2SO 4 and the following amounts of FeC13.6H20 and NaCI to 300 mQ of the final solution: 19.84 g and 39.70 g; or 61.23 g and 12.88 g; or 81.07 g and 0 g. An estimation of the experimental variance of the metal recoveries based on the three first centred replicates of the factorial design was obtained from Table 3. The results suggest the following:

266

M.F. ALMEIDAand M. A. AMARANTE

- - the highest recoveries of silver, bismuth and lead seem to be associated with tests where 61.23 g of FeCI3.6H20 was added; - - in all experiments the recoveries of copper and zinc are smaller than those for the other metals; - - however, due to the very high estimated experimental variance, almost all comparisons between the recoveries of any 2 experiments are risky (note that for a 95% confidence interval the calculated values of t.s/~Jn are 20.1, 10.6, 1.8, 7.9 and 3.3% respectively for Ag, Bi, Cu, Pb and Zn recoveries, all very large compared to the mean average values after 1 h of leach).

TABLE 3 Metal recoveries vs additions (g), temperature and time

NaC'IFec'3

T°C

th

"'%

Pb

Iz°

12.88

61.23

90

1.0

67.5

100

6.0

83.8

13.2

12.88

61.23

90

1.0

70.0

98.6

6.0

81.3

11.5

12.88

61.23

90

1.0

57.5

94.3

5.0

77.5

10.6

0

81.07

90

0.5

30.0

60.0

5.0

35.6

4.8

0

81.07

80

1.0

44.7

n.d.

n.d.

n.d.

n.d.

12.88

61.23

100

0.5

33.8

65.0

3.0

18.8

6.2

39.70

19.84

100

1.0

50.3

75.3

5.0

55.3

7.0

39.70

19.84

90

1.5

42.5

65.0

4.0

32.5

5.7

12.88

61.23

80

1.5

65.0

85.7

6.0

73.8

8.8

i

n.d. = not determined In order to quantify the need of oxidant, 3 tests were carried out at 80°C with 3 M [CI-] solutions with different additions of FeC13.6H20. Figure 6 shows the lower efficiency of the most oxidant solution without NaC1.

100

-

-

% 80 A g r e (.

o

60

40

V

e r e d

20

0

-

0

-

I

I

0.5

1

19.84 g

~

. - L _ _

1.5 Time, h 61.23 g

~

2

2.5

3

81.07 g F e C I 3 a d d e d

Fig.6 Silver recovery as a function of FeC13.6H20 addition in 3 M [C1-] solutions

Leaching of a silver beating sulphide by-product

267

Figure 7, where experiments were conducted at 80°C with addition of 19.84 g of FeCI3.6H20, shows the influence of CI- concentration, highlighting the need for using solutions with a minimum CI- concentration of 3 M. 100 % 80

A g r e c o v e

60

40

r e

20

d 0 0.5

0

•

1M

1

~

2 M

1.5 Time, h --~-3

M

,D

2

4 M

2.5

×

5 M

3

¢'

6 M

Fig.7 Influence of chloride concentration and time on Ag recovery for 61.23 g FeCI3.6H20/300 m~ solutions The influence of temperature is reported in Figure 8 where the similarity of the tests conducted at 80 and 90°C is evident, both producing the best results.

100 % 80

A g r

60

e c

o

40

V e r

e d

2O

0

0

0.5

60

1

C

~

1.5 Time, h 70

C

[]

2

80

C

2.5

-~-90

3

C

Fig.8 Influence of temperature and time on Ag recovery for 61.23 g ferric chloride/300 mQ, 3 M [C1-] solutions

268

M . F . ALMEIDA and M. A. AMARANTE

Figures 6, 7 and 8 all indicate that after 3 hours silver is still being dissolved, meaning that this period of time is not sufficient to complete silver recovery. Figure 9 shows lead and bismuth dissolution behaviour. Bismuth shows better solubility than lead, this meaning an easier liberation of the silver from mathildite (AgBiS2).

100

~100

80

18o

% P b r

B i 60

60

e

•

c

o

.~. . . . . . . . . . . . . . . . ..p . . . . . . . . . . . . . .

.t.- . . . . . . . . . . . .

-F

40

e

r

2O

d

..............

Oil'"

0

Jl- . . . . . . . . . . . . . . .

- 20

Pb

....

• . . . . . . . . . . . . . .

I

. . . . . . . . . .

- ....

L

I

I

I

I

0.5

1

1.5

2

2.5

r e C

40

v

e

%

0

v

e r e d

-0 3

Time, h • "IM

i 2 M

~

3M

+ 4 M

Fig.9 Lead and bismuth recovery as a function of time and chloride concentration for 61.23 g FeC13.6H20/300 m~ solutions Several tests were performed with 3--6 M C1- solutions with no addition of ferric chloride. Figure 10 refers to 4 of those tests that confirmed the high recoveries obtained for silver. In such tests bismuth and lead had the highest recoveries of 72 to 100%, as opposed to the low values for copper and zinc, in the range of 5-10%. In general, towards the end of the leaching tests, all the dissolved values decreased. The influence of particle size on the recovery was verified with the Type A sulphides milled at minus 150 prn. Experiments carded out with addition of ferric chloride have the general tendency shown in Figure 11, the increase in Zn and Cu dissolved being remarkable. As depicted in Table 4, no addition of ferric chloride lowers the level of both Zn and Cu, especially the latter, and losses of silver already dissolved are evident. It can be concluded that for leaching the milled Type A sulphides it is necessary to add Fe 3÷ in order to keep silver in solution, this also increasing zinc and copper dissolution. The behaviour of the unmilled Type B sulphides was also tested but using only solutions with no addition of ferric chloride. It was verified that recoveries of silver, bismuth and lead are 80-100% of those obtained with Type A sulphides under the same leaching conditions. Copper and zinc are dissolved in much less amounts, i.e., in the range 2--4%.

Leaching of a silver bearing sulphide by-product

269

100 % 80

A g r

60

e c o

40

v

e r e d

20

0

-

0

I

_I

0.25

r

0.5

"3M

1

J

I

1.5 Time, h

J4M

~

5M

___L

2

[]

2.5

3

6M

Fig, 10 Silver recovery as a function of time and chloride concentration in solutions with no addition of ferric chloride

120 %

100

r e c o v e r e d

80

Test 1 2 3 NaCI, g 12.88 39.70 39.70 FeCI3, g 61.23 19.84 19.84 T, C 100 100 90 t,h 0.5 1 1.5

60

40 20 0 Ag

Bi

Cu

Pb

Metal 1 Fig.ll

HE 8 : 3 - C

~

2

~

3 test

Metal recoveries in Type A sulphides milled at <150 ~ n

Zn

M. F. ALMEIDA and M. A. AMARANTE

270

T A B L E 4 M e t a l r e c o v e r i e s w i t h n o ferric a d d i t i o n on T y p e A s u l p h i d e s m i l l e d at < 1 5 0 p m

Leaching conditions

Ag

Bi

Cu

Pb

Zn

t, h

60.0

90.0

5.0

89.4

17.2

0.5

45.0

90.0

2.0

83.1

21.0

1

30.0

90.0

<1.0

79.4

19.8

1.5

52.65 g NaCI,

25.0

94.3

< 1.0

73.1

20.3

2

90 °C

25.0

94.3

< 1.0

70.6

22.1

2.5

20.0

98.6

<1.0

70.0

22.5

3

18.8

100

<1.7

75.0

22.1

F3

31.3

38.6

4.1

37.5

8.4

0.25

47.5

55.7

3.9

55.6

12.8

0.5

65.0

85.7

3.9

92.5

18.1

1

65.0

94.3

3.0

96.3

19.4

2

60.0

85.7

2.5

90.0

18.1

3

40.5

61.7

1.6

61.9

12.7

F3

105.3 g NaCI, 80 °C

CONCLUSIONS For hydrometallurgical recovery of the silver contained in two types of sulphides from the mine of Vale das Gatas, several screening leaching tests were carried out, further followed by the development of the most promising leaching routes: cyanidation, thiourea leaching and chloride leaching. Type A sulphides being older and therefore more weathered than Type B sulphides, show higher recoveries at similar leaching conditions. Their alteration is confirmed by microscopic observation, this facilitating the contact of the solution with mathildite and galena, the silver bearing minerals. In spite of the above, only after milling is it possible to dissolve a significant amount of silver by cyanidation. About 20% of the silver is dissolved in the early part of the reaction, which possibly corresponds to the A g in the altered minerals. The reactions taking place in sulphide oxidation make the solution acidic, and then the cyanide reacts with sulphates with formation of thiocyanate, which is favoured by the presence of Fe 3+. Therefore, the high consumptions of cyanide and lime, together with the precipitation of silver as AgSCN, would make cyanidation of the milled product uneconomic. Although recommended for the treatment of refractory ores, thiourea has been shown to be even less viable than cyanide. Thiourea is required in higher concentrations than cyanide and even without addition of Fe 3+ decomposes very rapidly. No advantages were shown in the use of thiourea. Chloride leaching seems to be the most promising process for these sulphide residues, because it dissolves more silver than any of the other methods and is more efficient in recovering by-product metals such as bismuth and lead. The feed does not require milling, and even without Fe 3+ addition it is possible to achieve very reasonable recoveries of the most valuable metals, i.e., silver and bismuth.

Leachingof a silver bearingsulphideby-product

271

Under these conditions small quantities of zinc and copper also dissolve, but this does not significantly affect silver recovery. A flowsheet including chloride leaching followed by cementation of dissolved metals (except zinc), in one or more steps, would therefore be appropriate to the conditions of this mine. The low zinc dissolution can be quite interesting, making it possible to reuse the solution a large number of times before being removed for zinc precipitation. However, further work is still necessary to optimize the operating conditions in this respect.

ACKNOWLEDGEMENTS

The authors thank Ms. Elisa Fernandes for the analysis and to the Ges~o de Minas for providing the samples from Vale das Gatas.

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