Internatmnal Journal of Mineral Processing, 8 (1981) 177~193 177 Elsevier Scmntffm Pubhshmg Company, Amsterdam -- Printed in The Netherlands

THE AQUEOUS OXIDATION OF COMPLEX SULFIDE CONCENTRATES IN HYDROCHLORIC ACID

TADAAKI MIZOGUCHI* and FATHI HABASHI

Department of Mining and Metallurgy, Laval Unwers~ty, Quebec City, Que GIK 7P4 (Canada)

(Received September 22, 1980, rewsed and accepted December 10, 1980)

ABSTRACT

Mlzoguchl, T and Habashl, F, 1981 The aqueous omdatlon of complex sulfide concen- trates m hydrochlorm acid Int J Miner Process, 8 177--193

Complex sulfides containing sphalemte, galena, chalcopyrlte, and small amounts of silver in a matrix of pyrite can be decomposed at 120°C and oxygen pressure of 1000 kPa in 1--2 N HCI for 90 min to yield > 97% of the zinc and > 95% of the copper in solution whlle about 83% of the lead remains m the residue as PbCl 2 and PbSO 4 and 85% of the silver, together with most of the pyrite The recovery of elemental sulfur m nearly 100% wlth respect to ZnS, PbS, and CuFeS2 Leaching m HCI m faster than in H2SO 4 at the same acid normality, and the process IS diffusion-controlled (strongly dependant on agltatmn speed and having an achvatmn energy of 3 6 kcal/mole) Lead jaroslte, Pb 0 sFes (SO4) 2- (OH)~, ]s mainly formed when H2SO 4 is used as a leaching agent

INTRODUCTION

There are large deposits of massive pyrite containing sphalerlte, galena, and chalcopynte m d~ssemmated form. In many cases ~t ~s not possible to prepare f lo tahon concentrates of the individual mmerals because to hberate these minerals, extremely free grinding is necessary, and under such conditions of very free partmle size, f lo tahon becomes meffmmnt. Treatment by pyrometal- lurgmal methods is usually not attractive because large amounts of SO2 will be produced and an outlet for H2SO4 m the Immediate vmmlty should be avail- able Also, a large amount of iron omde must be slagged, or used for making iron which is usually not econommally possible

Hydrometallurgmal treatment of such ores seems to be a convement chome ff it would be possible to selectively solubfllze sphalente, galena, and chalco- pymte leaving pyrite unattacked Further, condltmns could be so chosen that during the treatment, the sulfide sulfur assocmted with these minerals could be transformed into the elemental form thus minimizing waste disposal prob- lems m the aqueous effluents due to sulfate mn

*Present address Chemical Research Institute of Non-aqueous Soluhons, Tohoku Umver- slty, Sendal, Japan

0301-7516/81/0000--0000/$02 50 © 1981 Elsevier Scientific Pubhshlng Company

178

Recently there has been a great interest m the treatment of these complex sulfide ores as can be judged from the two conferences devoted to this prob- lem (Complex Metallurgy, 1978, Complex Sulfide Ores, 1980) as well as numerous other pubhcatlons (Ono et a l , 1962. Tarabaev, 1963, Yazawa et a l , 1964, Mackm and Veltraan, 1967, Ermllov et a l , 1968, Dutrlzac and Mac Donald, 1977, Georgeaux et a l , 1977, Demarthe and Georgeaux, 1978. Bolton et a l , 1979, Dutnzac, 1979) A varmty of leaching agents has been proposed, e g , H:SOa + 02, FeC13 solution, Fe:(SO4)3 solution, and CuC1. solution Of these, the process developed by Sherntt-Gordon in Canada (Macklw and Veltman, 1967. Bolton et a l , 1979) and that developed by Mmemet Recherche m France (Georgeaux et a l , 1977, Demarthe and George- aux, 1978) have recewed the greatest at tention In the Shernt-Gordon Process (Macklw and Veltman, 1967, Bolton et a l , 1979), the concentrate is leached by H:SO4 and air under pressure at 150°C and 350 kPa oxygen partial pressure for one hour Zinc and copper go into solution while lead is transformed into PbSO4 and lead ]arostte, eb0 sFe3(SO4):(OH)6, and is retained m the residue together with the pyrite, and any iron minerals that are attacked and repreclp- ltated The reactions taking place may be represented as follows

ZnS + H:SO4 + %0: ~ ZnSO4 + S + H:O

PbS + H2SO4 + 1AO2 ~ PbSO4 + S + H:O

CuFeS: + H:SO4 + 5/4 O: ~ CuSO4 + FeOOH + 2S + 1/~H~O

and to a minor extent

2FeS~ + 70: + 2H:O -~ 2FeSO4 + 2H2SO4

4FeSO4 + 2H:SO4 + O: -~ 2Fe:(SO4)3 + 21-120

3Fe:(804)3 + PbSO4 + 12H:O -~ 2Pb0 ~ [Fe3(SO4):(OH)6] + 6H2SO4

Fe2(SO4)3 + 2H20 -* 2Fe(SO4)(OH) + H:SO4

Fe:(SO4)3 + 6H:O -~ 2Fe(OH)3 + 3H:SO4

After flltratmn, the solutmn is purffmd, and then electrolyzed to recover the zinc, and the acid generated during electrolysis xs recycled to the leaching step Pyrite and elemental sulfur can be floated from the residue, but the recovery of lead from the jaroslte poses a problem since lead m thxs form is insoluble m the commonly used reagents, e g , NaC1 or ammonmm acetate solutmn

In the Mmemet Process (Georgeaux et a l , 1977, Demarthe and Georgeaux, 1978) the concentrate xs leached near the boiling point and at atmospheric pressure by a solutmn of CuC12 (40 g/1 Cu :÷) The nonferrous metal sulfides react according to

ZnS + 2CuClz -* ZnC12 + 2CuC1 + S

PbS + 2CuCI: -, PbCI: + 2CuC1 + S

179

CuFeS2 + 3CuCl2 -* 4CuC1 + FeC12 + 2S

Ag2S + 2CUC12 -* 2AgC1 + 2CuC1 + S

To keep CuC1, PbC12, and AgC1 m solution, NaC1 m a concentratmn of 250 g/1 ,s added to the leaching agent Under these condltmns, pyrite is unattacked The solutmn ,s punfmd by precipitating FeOOH at pH 2.6 with mr mjectmn according to

2FeC12 + 4CuC1 + 3/202 + H20 -* 2FeOOH + 4CuC12

Lead and silver may be recovered by cementation with Zn and Cu respectively Thus, the final solution will be composed of ZnC12, CuC1, CuC12 and NaC1 Zmc is then extracted by dmthyl hexyl phosphoric acid (RH) accordmg to

ZnC12 + 2RH -~ R2 Zn + 2HC1

Zmc is stripped from the orgamc phase by H2SO4 (spent electrolyte) and the strip solutmn electrolyzed m the conventmnal way The leaching agent is re- generated from the acldffmd CuC1 by mr oxidation according to

2CuC1 + 2HC1 + IAO 2 -* 2CuC12 + H20

In this process no lead jaroslte is formed, but there are some other draw- backs, namely" (1} the necessity of oxidizing CuC1 to CuC12 for recycle, (2) the necessity of addmg a large amount of NaC1 to keep CuCI m solutmn, (3) the necessity of separatmg the complexed CuC1 from ZnC12, and (4) the preclpltatmn of Lron is done in an extra step, and due to the presence of a large amount of copper runs m solutmn, the goethlte precipitated contains ap- precmble amounts of copper (1 7 to 12 6%)

In the present work it was at tempted to leach the complex sulfide m acid under oxygen pressure to extract the nonferrous metals and precipitate the iron m a smgle step as m the Sherntt-Gordon Process, the only difference be- mg the use of HC1 instead of H2SO4 to mlmmlze the presence of SO~- runs in solutmn. The high reaction rates observed when HC1 is used m leaching chal- copynte as reported earlier (Habashl, 1978, Habashl and Toor, 1979) also suggested using this method

EXPERIMENTAL

Two complex sulfide concentrates were studmd, one was obtained from New Brunswmk m Canada and the other from the other coast of the Atlantm m Breton, France. Their chemmal analyses are glven m Table I. Both con- centrates are mamly of a pyrite matrLx contammg sphalente, galena, and chal- copynte . Since pyrite and sphalente have the same crystalhne structure it was not possible to distinguish between them by X-ray analysis (both have the same pattern). The samples were 97% minus 325 mesh. Leachmg tests were conducted m a 1 hter t i tanmm autoclave supphed by Parr Instrument Com- pany, Mohne, Ilhnols. A few tests were also conducted m a 2 hter t l tanmm

180

TABLE I

Analysis of complex sulfide concentrates

Element New Brunswick (Canada) Breton (France) %

Fe 22 08 13 65 Zn 26 53 22 0 Pb 8 54 2O 5 Cu 0 84 8 36 Cd 0 O4 S 37 15 27 79 Ag 0 023 0 072

autoclave (supplied by Autoclave Engineers, Erie, Pennsylvania) using large amoun t s of the concen t r a t e When the 1 li ter autoclave was used, each test was c o n d u c t e d with 60 g o f the concen t r a t e unless o therwise mdmated To this amoun t , 400 ml o f leaching agent was added The mix tu re was hea ted to the desired t e m pe ra tu r e and then pressurized by o x y g en to the desired pres- sure F r o m this m o m e n t on the t ime of reac t ion was c o u n t e d The oxygen pressure r epo r t ed here is the d i f fe rence be tween the to ta l pressure exer ted af ter oxygen in t roduc t ion and tha t before . At the end o f each test , hea t ing o f the autoclave was s topped , the oxygen supply was shut off , and the reac t ion mix tu re was quenched to r o o m t empe ra tu r e by admi t t ing cold water in the cool ing coils

The autoclave was empt ied and its con ten t s f i l tered The so lu t ion and wash- lngs were adjusted to one h te r t hen analyzed for lead, copper , Iron and silver by a tomic absorp t ion spec t rome t ry . Zinc was de t e rmined by c h e l a t o m e t r y wath EDTA, using xy leno l orange as ind ica tor Sulfate Ion was d e t e rm in ed gravimetr lcal ly as BaSO4 and the pH of the so lu t ion was measured with a Fisher pH me te r

The leach residues were dried for 2 hrs at 80°C and ana lyzed for e lementa l sulfur by the Soxhle t m e t h o d using ca rbon dmulfide as a solvent Acetate- soluble lead in the leach residue (non-sulf ide lead) was de t e rmined by heat ing 3 g of residue with 50 ml of 20% a m m o n i u m aceta te solut ion fo l lowed by c h e l a t o m e t r y with EDTA, using xy leno l orange as ind ica tor

In some cases, usually at high acidi ty or in the early stages o f leaching, crystals o f PbC12 fo rmed in the leach solut ion on standing. Th e de termina- t ion of lead in so lu t ion was t he re fo re c o n d u c t e d in the fol lowing way the leach solut ion was coo led to abou t 10 ° C, and the crystals f o rm ed were fil- tered off , dissolved in 20 ml 20% a m m o n i u m ace ta te and the lead was deter- mined in the f i l t ra te and in the a m m o n i u m ace ta te so lu t ion , the sum rep- resents lead in solut ion.

Residues were examined by the opt ical mic roscope and by X-ray diffrac- t ion In some cases the residue was ex t r ac t ed by CS2 to r emove e lementa l sulfur, t he n by a m m o n i u m ace ta te solut ion to r emove PbSO4 and PbC12, t hen

181

subjected to decantation and gravity separation m water to separate the Iron oxide fraction from the unreacted sulfide; such techmque was found to be simple and effective in isolating the products of the reaction

RESULTS AND DISCUSSION

Preliminary tests on the New Brunswmk concentrate had shown that PbS is more easily decomposed by either HC1 or H2SO4 than ZnS. But the fraction d~ssolved tended to form crystalhne precipitates wl~en the leach solution was allowed to stand leawng behind a small fraction m solutmn, i.e., In a single leaching test, lead was distributed m three fractmns In the residue, in the crystalhne precipitate, and m solution. It was, therefore, considered that from a practmal point of view, it would be preferable If the lead IS collected In a single fractmn to facilitate its recovery at a later stage, e g , by leaching with a hot saturated solutmn of NaC1. In the present work non-sulfide lead PbSO4 and PbC12 was completely solublhzed by an ammonmm acetate solution X- ray powder diffraction showed that no diffraction lines of PbC12 and PbSO4 remained when the sample was leached with ammonium acetate thus lndtcat- mg that this reagent is very effective for extracting these salts However, it was ineffective in extracting silver from the residue.

Another objective of the work was to avoid the dissolution of the pyrite matrix. If this were not possible then conditions should be found such that the mlmmum amount of iron be present in solution to facthtate the recovery of zinc Further, the formatmn of elemental sulfur should be maximum while SO~- should be mlmmum.

In some leaching tests, especially those at high acidity, the concentrate ag- glomerated In hard lumps and as a result the recoveries were low (see for ex- ample Fig 3} This phenomenon was observed dunng the leaching of other sulfides and will be reported m detml in a later publication

Effect of HCl concentratmn

Fig la, b shows the effect of acid concentration on leaching the New Brunswick concentrate at 120°C. It can be seen that at 1 N HC1 the recovery of zinc and elemental sulfur are high, lead and sliver m solution are low but the non-sulfide lead In the residue is high, iron and SO~- ions in solution are also low. The residue was red m color but under the microscope black par- tmles were also visible thus Indicating that some pymte underwent oxidation and precipitation to ferric hydroxide while the bulk remained unattacked. The final pH of the solution was about 2.

Above 1 N HC1, the recovery of zinc and elemental sulfur did not improve, but Iron, lead, and silver m solution mcreased greatly. The residue was black indicating that no precipitation of ferric hydroxide took place. The final pH of the solution was less than 1

The results show also that most of the sulfide sulfur associated with zinc,lead,

182

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and copper is transformed into the elemental form while most of that asso- ciated with the pyrite that was attacked was transformed mto sulfate as can be seen from Table II 9 11 g sulfur are associated with ZnS, PbS, and CuFeS~ m 60 g concentrate while the total sulfur m 22.29 g. The theoretmal recovery of elemental sulfur from these three minerals is (9 .11122.29) X 100 = 40 9% which m very near to the value'actually found experimentally as for example m Fig. 1. A mmor amount of elemental sulfur m also formed from pyrite whmh accounts for the shght increase m the elemental sulfur recovery as com- pared to the theoretmal amount

It was therefore considered that leaching thin concentrate with 1 N HC1 gave favorable results and most of the tests conducted later were under these condit ions

183

TABLE II

The a m o u n t o f e lemental sulfur fo rmed during the aqueous ox ida t ion of ZnS, PbS, and CuFeS 2 m the New Brunswick complex sulfide

E lement A m o u n t m 60 g feed

% g mole

Theoret ical e lemental sulfur

Zn 26 53 15 92 0 244 7 81 Pb 8 54 5 12 0 025 0 79 Cu 0 94 0 504 0 008 0 51

S 37 15 22 29 Total 9 11

Theore tmal S recovery -- (9 11/22 29) × 100 -- 40 9% Exper imenta l S recovery ~ 40% (see Fig 1)

The fact that zinc (and elemental sulfur) recovery becomes constant after a certain acid concentratmn is in agreement with previous work on the leaching of pure ZnS (Habashh 1966}, Le., at high acidity the rate will be independent of the acid concentration

Fig. l a also shows that m absence of HC1, about 25% of the zinc goes into solutmn This is apparently due to the presence of pyrite because work m progress showed that pure zmc sulfide concentrate is practmally unaffected by water under similar conditions. It seems that a small amount of pyrite is first decomposed and the iron that goes into solutmn ~s hydrolyzed generat- ing acid whmh initiates the attack of ZnS This is supported by the fact that the residue from such treatment was reddish m color and contmned some Fe203

Comparison between HCI and H2S04

Leaching New Brunswick concentrate with 1 N HC1 was faster than with 1 N H2SO4 as shown m Fig. 2 for the recovery of Zn, Pb (after extraction with ammonmm acetate from the residue), Cu, Ag, as well as elemental sulfur Fig. l b shows further the behavior of sliver at different acid concentrations from whmh it is also evident that leaching with HC1 is faster than with H2SO4

The decrease m lead recovery with mcreased leachmg time (Fig. 2) was found by X-ray diffraction analysis to be due to the formation of lead jaroslte (Pb05Fe3(SO4)2(OH)6, ASTM card No 18-698) whmh is yellow in color and is insoluble m ammonmm acetate. It can also be seen from Fig. 2 that more Jaroslte formed m H2SO4 medmm than m HC1. Jaroslte was completely ex- tracted from the sample by boiling with 2 N HC1 for about 2 hours, this was confirmed by X-ray diffraction analysis. Fig. l a also mdmates that no lead jaroslte is formed at high acid concentration

184

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Fig 2 Comparmon between HC1 and H2SO 4 as leaching agents, 75 g New Brunswmk con- centrate, 400 ml of 1 N acid, 120° C, 1000 rpm, Po2 1000 kPa

Figs. 3 and 4 s h o w the leaching o f the French c o n c e n t r a t e m HC1 and m H2SO4, r e s p e c t w e l y It can be seen that its behavior is similar to the N e w B r u n s w m k c o n c e n t r a t e wi th the e x c e p t i o n that a large a m o u n t o f lead jaroslte is f o r m e d m H2SO4 m e d m m as indicated by the decreased lead recovery This is due to the presence o f a large a m o u n t o f lead m this c o n c e n t r a t e The de- creased recovery above 2 N HC1 is due to the agg lomera t ion p h e n o m e n o n m e n t i o n e d earher The increased rate o f dmso lu t lon m HC1 as c o m p a r e d to H2SO4 m a y be at tr ibuted t o the f o r m a t i o n o f ch lor ine according to (Habashl , 1 9 7 8 )

2 HCI + 1A 02 -* C12 + H 2 0

Even if f o r m e d in trace a m o u n t s , ch lor ine is a very act ive reagent towards sul- h d e s , it reacts i m m e d i a t e l y as s o o n as it is f o r m e d thus the e q u l h b r m m is c o n t i n u o u s l y shi f ted to the right

185

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Fig 3 Effect of hydrochloric acid concentrahon, 75 g French concentrate, 400 ml HC1, 1 hr, 120°C, Po2 1000 kPa, 1000 rpm

Fig 4 Effect of sulfurm acid concentrahon, 75 g French concentrate, 400 ml H:SO4, 1 hr, 120°C, Po2 1000 kPa, 1000 rpm

Effect of temperature

Fig. 5 shows the effect of temperature on leaching It can be seen that the recovery of Zn, elemental sulfur, and sulfate ions, gradually Increases until 120°C is reached, then from 120--140°C it remmns practmally constant, silver behaves similarly except that the constant recovery is achmved at 90 ° C On the other hand, lead recovery in solution decreases with increasing temper.

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Fig 5 Effect of temperature, 60 g New Brunswmk concentrate, 1 N HC], 400 m[, 1 hr, Po 1000 kPa, agltatmn 600 rpm 2

186

ature because it is t rans formed from the so luble ch lor ide to the inso luble sul- fate as m d m a t e d by X-ray d l f f rac tmn analysis o f the res idue Iron m so lu t ion is a b o u t 12% but there is a sl ight m i n i m u m at 1 0 0 ° C

Further tests s h o w e d that at 2 N HC1 the recovery o f sdver increases with increasing t emperature m the range 8 0 - - 1 4 0 ° C

The apparent actzvatmn energy for z inc sulf ide d l s s o l u t m n as ca lculated f rom the Arrhemus p lot in the t empera ture range 90 to 1 2 0 ° C was 3 6 kcal / m o l e which indicates a d i f fus ion-contro l l ed process through a b o u n d a r y layer

Effect of oxygen

Leaching in absence o f o x y g e n is s l ower than w h e n o x y g e n is present (Fig 6) Also, m absence o f o x y g e n lead recovery decreases wi th increasing t emperature (Fig 7) X-ray d i f fract ion analysis o f the res idue s h o w e d that at 90°C pract ical ly all the PbS dissolved whi l e ZnS did no t , but at 1 5 0 ° C the reverse was observed Thls can be expla ined o n the basis o f an Interact ion at high t emperature and l o w acidi ty b e t w e e n Pb 2+ ion and ZnS as f o l l o w s

Pb=* + ZnS -~ PbS + Zn 2*

This IS reasonable to as sume s ince the so lubi l i ty produc t o f PbS is m u c h less than that o f ZnS, it is 1 0 × 10 -29 w h d e that for ZnS is 4 5 × 10 -24 On the o ther hand, w h e n leaching is c o n d u c t e d in the presence o f o x y g e n the recov- ery o f z inc, silver, and e l ementa l sulfur increases rapidly wi th increased o x y g e n pressure up to a b o u t 7 5 0 kPa then s l o w l y f r o m 750 to 2 0 0 0 kPa as s h o w n in Fig 8 The recovery o f lead m s o l u t i o n decreases wi th increased o x y g e n pres-

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Fig 7 Leaching m the absence of O~, 60 g New Brunswmk concentrate, 2 N HCI, 400 ml, 1 hr, agitation 600 rpm

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sure, at moderate oxygen pressure (500- -1000 kPa), PbSO4 mmnly precipitates, but at high oxygen pressure mmnly lead jaroslte precipitates as indicated by X-ray diffraction

Effect of speed of agitation

The recovery of zinc and elemental sulfur increases with increased speed of agitation (Fig 9) which suggests that the process is dlffuslon-controUed. Cop- per recovery showed anomalous behavior it increases suddenly when agita- tion speed exceeds 600 rpm Further, at low agitation speed there was a ten- dency of copper to repreclpltate as shown in Fig 10. This anomalous behav- ior could not be explmned at present.

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Fig 9 Effect of speed of agitation, 60 g New Brunswmk concentrate, I N HCI, 400 ml, 120°C, hme 1 hr, PO= 1000 kPa

Fig 10 Anomalous behavior of copper (New B r u ~ w m k concentrate)

18b

Effect of tzme

Fig 2 showed that zinc recovery increased gradually with increased time, after about 90 mm at 120°C and in 1 N HC1 more than 97% of the zinc was solublhzed At low acid concentration (1 N HC1), some pyrite goes Into solu- tion at the beginning of leaching, but due to the acid consumption, and the subsequent increase in pH, iron precipitates as can be seen from Fig 11 Also, it was found that lead m solution and the non-sulfide lead in the residue de-

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Fig 11 Behavior of Fe and pH change during leaching 75 g New Brunswick concentrate, 1 N HC1, 400 ml, Po2 1000 kPa, 1000 rpm

Fig 12 Change m lead recovery as a functton of leaching hme 60 g New Brunswmk con- centrate, 1 N HC1 400 ml, 600 rpm, 120°C Total Pb recovery = Pb m leach solution + Pb extracted from residue with a m m o m u m acetate

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189

crease with mcreasmg t ime (Fig. 12). X-ray d f f f rac tmn analysis showed tha t the longer the reac t ion h m e the more lead jarosl te would be formed.

Figure 13 shows the effect o f t ime on leaching the French concen t r a t e m HCI f rom which it can be seen tha t high recoveries for Zn, Pb, Cu, and ele- menta l sulfur can be achmved, while the a m o u n t o f SO~- and Fe in so lu tmn are very small Also, mos t of the silver remmns m the residue

Effect of the sohd-hqu~d ratio

Tests were c o n d u c t e d using cons tan t weight o f New Brunswmk concen t r a t e and variable acid concen t ra t ion , also cons tan t acid concen t r a t i on and variable weight of concen t r a t e as shown in Table I II , t empera ture , oxygen pressure, speed of agitat ion, react ion h m e , and vo lume o f acid were kept cons tan t . In this way it was possible to s tudy the variat ion o f the rat io HC1/concentra te (expressed m mill1 mole /g) at d i f ferent acid concen t ra t ions and dif ferent pulp densities The results o f these tests are p lo t ted m Fig. 14. I t can be seen tha t when the rat io HC1/concentra te is 5 mllh mole /g , the recovery of zinc and elemental sulfur m 1 hour is m a x i m u m , iron m s o l u h o n is m l m m u m , and the

TABLE III

Effect of sohd-hquld ratio, Po~ 1000 kPa, 120 ° C, 1 hr, 600 rpm, HCI solutlon 400 ml

Concn Welght of HCl/sohd Recovery (%) Final pH of HCI concentrate ratlo Zn Pb Fe Cu S O

(M) (g) (mmol/g) (m soluhon)

0 60 0 238 01 17 62 79 217 05 60 333 477 04 07 22 195 259 075 60 500 823 08 05 413 316 209

10 30 133 810 50 435 595 397 099 10 60 667 828 08 83 180 382 160 1 0 75 5 3 3 829 0 7 3 3 3 2 2 3 4 8 1 71 10 90 444 710 03 02 91 274 -- 1 0 120 333 507 02 09 4 6 194 104 10 160 250 429 0 06 04 17 1 301

125 60 833 844 18 155 317 -- 142 15 60 10 O0 818 19 359 434 412 129

2 0 30 267 775 766 362 1000 400 - - 2 0 60 133 805 37 5 340 891 390 -- 20 90 889 791 164 248 282 378 1 23 2 0 120 6 6 7 813 8 7 7 5 2 2 6 3 5 6 1 59 20 160 500 855 0 8 O1 792 320 280

30 60 200 793 545 356 1002 394 020

1.90

I00

+° t 6O

40

~o~

i i [ . . . . ~ - - .

' Zn

[]

o

5 IO 15 20 I 0

25

a~

HCI/SOLID RATIO, mmol/g

Fig 14 Effect of sohd-hquld ratio, 120°C, Po: 1000 kPa, 1 hr, 600 rpm, leaching agent 400 ml, New Brunswmk concentrate

COMPLEX SULFIDE CONCENTRATE

O2

~EACH~NO ~ Z-- I /~LTRAT,ON ~

PRECIPITATION OF COPPER

FILTRATION

PRECIPITATION OF CADMIUM

FILTRATION ~.

L CADMIUM J RECOVERY

EXTRACTION ,~

FRESH ORGANIC LOADED ORGANIC

H?S04 ELECTROLYSIS (SPENT 1 ELECTROLYTE)

Zn

SOLIDS

ILEACH,NG OP LE~

+ ~ ILEACHING ~ COPPER OPS'LVER I RECOVERY LEAD

RECOVERY RESIDUE

~ ---~ FeS 2 ,

FeO OH,S ° GANGUE

SILVER RECOVERY

Fig 15 Flowsheet of proposed process

191

fmal pH = 2 whmh mdmates complete acid consumption. This ratm represents the stomhmmetrm amount of acid needed to selectively extract zinc

Proposed process

The results of the present work suggest a process similar to the Shemtt- Gordon Process m which H2SO4 is replaced by HC1. The advantages gamed would be increased reactmn rate and ~mproved lead recovery since the forma- tmn of lead jaroslte is minimized. However, smce the direct recovery of zinc from chloride medmm is not technically advanced yet, it is suggested to transfer the chloride to the sulfate system from whmh zinc is recovered by electrolysis usmg conventmnal methods. The transfer from chloride to sulfate can be done using an organic solvent, e .g , dmthylhexyl phosphorm acid as adopted commercially m Spam (Noguelra et al., 1979, 1980} and tested by Mmemet (Georgeaux et a l , 1977, Demarthe and Georgeaux, 1978) With such a system {Fig. 15), the acid generated during electrolysis will be used for stripping, and that generated during extractmn will be recycled for leachmg. The solvent extraction step will be also a punficat lon step smce the sulfate electrolysis system necessitates a tedmus punflcatmn of the solutmn before electrolysis

CONCLUSIONS

The conclusions regarding the t reatment of a complex sulfide concentrate composed mainly of pyn te contmnmg sphalente, galena, chalcopynte, and a small amount of silver can be summarized as follows

(1) Two distract leaching conditions can be identified (a) At low acidity (<~ 1 N HC1 for the New Brunswmk concentrate and

~< 1 75 N for the French concentrate) ZnS is readily solubillzed, PbS is mainly transformed into PbSO4, a small amount of pyrite is transformed m- to ferrm hydroxide while the rest remains unattacked. Most of the silver remains m the residue.

(b) At high acidity (/> 1 N HC1 for the New Brunswmk concentrate and /> 1 75 N for the French concentrate), the solubillzatlon of ZnS is high, PbS is completely transformed into PbC12 and PbSO4, and an appreciable amount of pyrite is solubillzed with no precipitation of insoluble iron compounds. Most of the silver remmns m solutmn

(2) Leaching under low acidity is more advantageous than at high acidity because ZnS dissolves selectively

(3) Sulfide sulfur associated with Zn, Pb, and Cu is completely transformed to the elemental form. The small amount of sulfate ion formed is mmnly a result of the attack of pyrite.

(4) Some lead precipitates as lead jaroslte, Pb0 sFe3(SO4):(OH)6, whmh is insoluble m ammonmm acetate solutmn Precipitation Is favored by

(a) The presence of excess SO~- runs, thus more jaroslte is precipitated from H2SO4 than from HC1 solution

192

(b) Long reaction time (c) High oxygen pressure (> 1000 kPa}

(5) Copper shows anomalous behavlour during leaching, at low speed of ag;tatlon there is indication that reprec~p~tatmn takes place leading to low recovery High copper recovery ~s only obtained at h~gh speed of agitation

(6) Leaching with HC1 m the presence of oxygen is faster than when oxygen is absent In the first case elemental sulfur is formed while in the second case hydrogen sulhde

(7) Leachmg with HC1 is dltfumon-controlled through a boundary layer, the process is strongly dependent on the speed of agltatmn and also the activation energy is 3 6 kcal/mole -- a value that is typical of a dlffusmn-hmlted process

(8) Leaching with HC1 is faster than with H2SO4 under the same conditions of acid normality

(9) When leaching of New Brunswmk concentrate is conducted at 1 N HC1, 120°C, PO 1000 kPa, 1000 rpm, 90 mm reactmn time, and sohd-hquld ratm of 75 g m 400 ml leach solutmn, the following data was obtained

Zinc recovery in solution Copper recovery m solution Lead recovery in solution Lead m residue, soluble m ammonium acetate Lead in remdue, insoluble m ammonium acetate Silver in solution Silver in residue Iron In solution Sulfur as sulfate in solution Sulfur m elemental form Sulfur as PbSO4 Sulfur as basic sulfate (jaroslte) m the residue Sulfur as sulhde m the residue

%

97 4 9 5 2

0 3 82 9 1 6 8 1 5 0 85 0

4 0 183 3 8 5

3.0 2 4

3 7 8 Similar results were obtained for the French concentrate when conducted with 2 N HC1

leaching was

ACKNOWLEDGEMENT

The authors acknowledge with thanks the help given by the following orgamzatlons. Tohoku Umverslty, Sendal, Japan, Canada Center for Energy, Mineral Resources (CANMET), Ottawa, National Research Council for Scmnces & Engmeermg, Otta~va, and Bureau des Recherches Gdologlques et Mml~res, Orl~ans-La-Source, France

REFERENCES

Bolton, G L, Zubryckl, N and Veltman, H, 1979 Pressure leaching process for complex zlnc-lead concentrates In J Laskowskl (Editor), International Mineral Processing Congress, Warsaw Preprmts, pp 581--607

193

Complex Metallurgy '78, 1978. Instltute of Mining and Metallurgy, London Complex Sulflde Ores, 1980 Joint conference orgamzed by Instltute of Mlnmg and Metal-

lurgy (London) and Conslgho Nazlonale delle Rlcerche (Rome) Demarthe, J M and Georgeaux, A, 1978 Hydrometallurgmal treatment of complex sul-

fldes In Complex Metallurgy '78 Inst Mln. Metall, London, pp 113--120 Dutnzac, J, 1979 Ferrm sulfate percolation leaching of a pyntm Zn-Pb-Cu ore CIM Bull,

72 109--118 Dutrizac, J and Mac Donald, R J C , 1977 Percolation leaching of pyntm Zn-Pb-Cu ores

with ferrm chloride solution Can Metall Q , 16 186 -194 Ermllov, V V et a l , 1968 Hydrochloric acid method for processing a copper-lead-zinc

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m~talhques avec ou sans concentrat ion globale pr~alable Ind Mm~rale, Mm~ralurgm, May 86--93

Habashl, F , 1966. The mechanism of oxidat ion of sulfide ores m nature Econ Geo l , 61 587--591

Habashl, F , 1978 Chalcopyr l te - - Its Chemistry and Metallurgy McGraw-Hill, New York, London

Habashl, F and Toor, T , 1979 The aqueous oxldatmn of chalcopynte m hydrochlorm amd Metall Trans , 10(B) 49--56

Macklw, V N and Veltman, H , 1967 Recovery of zinc and lead from complex low-grade sulfide concentrate by acid pressure leaching Can Mm Metall Bull , January 80--85

Noguelra, E D , Reglfe, J.M and Arcocha, A M , 1979 Winning of zinc through solvent extractmn Eng Mm J , 180(10) 92--94

Noguelra, E D , Reglfe, J.M and Blythe, P M , 1980 Zmcex -- The development of a secondary zinc process Chem I n d , January 19, p 63

Ono, K et a l , 1962 Effect of acid concentratmn and temperature on the selective sul- fatlzatmn of complex sulfide ores Tohoku Dalgaku Senko Selren Kenkyusho Iho, 18(2) 147--157 CA 62, 8729C

Tarabaev, S I , 1963 Recovery of zinc from bulk concentrates by the hydrochloric acid method Vestn Akad Nauk Kaz SSR, 19(3) 41--45 CA 59, 4824C

Yazawa, A , Kolke, K and Eguchl, M , 1964 Selective sulfatlzation of complex sulfide ores by dilute sulfuric acid under normal atmospheric pressure Tohuko Dalgaku Senko Selren Kenkyusho lho, 20(1) 51--58 CA 63, 5283e