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Publication numberUS3655538 A
Publication typeGrant
Publication dateApr 11, 1972
Filing dateMay 19, 1969
Priority dateMay 19, 1969
Publication numberUS 3655538 A, US 3655538A, US-A-3655538, US3655538 A, US3655538A
InventorsRenken Howard C, Zegers Theodoor W
Original AssigneeTexas Gulf Sulphur Co
Export CitationBiBTeX, EndNote, RefMan
External Links: USPTO, USPTO Assignment, Espacenet
Process for electrowinning zinc from sulfide concentrates
US 3655538 A
Abstract
This disclosure is directed to leaching zinc sulfide concentrates to facilitate the production of zinc and sulphur. solution. To provide for increased recovery the copper sulfate leach The zinc sulfide is leached with a copper sulfate solution, preferably under oxidizing conditions, to produce a copper sulfide and a zinc sulfate solution. The elemental sulphur can be obtained from the copper sulfide by a subsequent oxidizing leach and the zinc metal can be obtained by electrolyzing the zinc sulfate leach may be accomplished in a staging operation to establish an excess of zinc sulfide concentrate in the initial leach step.
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United States Patent Renken et a1.

[54] PROCESS FOR ELECTROWINNING ZINC FROM SULFIDE CONCENTRATES [72] Inventors: Howard C. Renken; Theodoor W. Zegers, both of Columbus, Ohio [73] Assignee: Texas Gulf Sulphur Company, New York,

[22] Filed: May 19, 1969 [21] AppLNo; 825,578

[52] U.S.Cl ..204/114,23/114 [51] C22d 1/22, COlb 17/06 [58] Field ofSearch ..204/119, 114, 108;75/117, 75/120; 23/125, 135, 114

[5 6] References Cited UNITED STATES PATENTS 3,095,363 6/1963 Ruckwardt et a1 ..204/119 1,937,634 12/1933 Christensen ....204/119 1,890,934 12/1932 Carson ..204/114 15] 3,655,538 [451 Apr. 11, 1972 1,553,414 9/1925 Van Arsdale ..204/108 FOREIGN PATENTS OR APPLICATIONS 107,982 7/1943 Sweden ..204/1 14 Primary Examiner-John I-I. Mack Assistant Examiner-R. L. Andrews An0rneyKenyon & Kenyon Reilly Carr & Chapin [5 7] ABSTRACT This disclosure is directed to leaching zinc sulfide concentrates to facilitate the production of zinc and sulphur. solution. To provide for increased recovery the copper sulfate leach 8 Claims, 5 Drawing Figures PATENTEDAFR 'l 1 I972 Beam/auras &

1 N VEN 'TORS BY m Zc-aaes 1 'PROCESSFORELECTROWINNINGZINCFROM SULFIDE CONCENTRATES BACKGROUND acid.

3. Dissolution of the zinc oxide in dilute sulphuric acid to form -zinc sulfate.

l. Purification ofthe zinc sulfate solution. 5. Electrolysisof the zinc sulfate solution to produce zinc metal withth'e regeneration of sulphuric acid.

Modificationsof this process have been sought toprovide elemental sulphur while maintaining a high yield of zinc metal.

Manytheoretical chemical equations can be drawn to illustrate methods by which zinc metal and elemental. sulphur couldbe extracted from zinc sulfide. Foremost among these are the usesof sulphuric acid, under oxidizing and non-oxidizing conditions, to leach zincsulfide'concentrates. However these proposals-have not proven to be entirely satisfactory,

-due'either to low leaching rates, the requirements of complex process equipment or unattractive cost factors.

It is therefore an object of this invention to provide an improved method of leaching zinc sulfide ores to recover values therefrom. p

lt is another object of the present invention to provide an improved method for the recovery of elemental sulphur and -'zinc metal-from zinc sulfide concentrates,

modification of the electrolyticzincprocess.

Qther objects andadvantages of the invention are set forth inconju'nction with the following detailed description.

THE INVENTION This invention is directed to the use of copper sulfate as a leaching agent for zinc sulfide concentrates, preferably under oxidizing conditions, to produce elemental sulphur and zinc metal.

The method of this invention includes first leaching the zinc concentrate under pressure by an exchange reaction with copper-sulfate solution to give a zinc sulfate solution for electrolysis in accordance with conventional methods and a copper-sulfide precipitate. Recovery of elemental sulphur and rege'ne'ra'tionof the copper sulfate solution is accomplished by a second pressure leach reaction of the copper sulfide precipitate with spent electrolyte-under oxidizing conditions. In this process, the copper is precipitated as cuproussulfide with the concurrent production of sulfuric acid. To alleviate the-difficulties in purification procedures caused by the sulfuricacid, the copper sulfate leach may altematively'be carried out-under oxidizing conditions. Furthermore, by staging the-operation to establish an excess of zinc concentrate in the initial copper sulfate leach, the amount-of acid production and soluble sulfate formation can'be minimized, with consequent higher recovery. of elemental sulphur at'high zinc extraction.

The invention is described in greater detail in conjunction with thefollowing figures and working examples in which:

FIG. -1 -is a process flow chart of the non-oxidizing copper sulfate-leach of zinc sulfide to produce-elemental sulphur and zincmetal.

FlG52 isa process flowchart of the oxidizing copper sulfate leach of zinc sulfide to produce elemental sulphur and zinc metal.

FIG. 3 is a graph of the production of acid sulphur (based on total sulphur) as a function of time in the process of FIG. 2.

F IG. 4 is a graph of the acidsulphur to zinc ratio as a functionoftime in the process of FIG. 2.

FIG. '5 is a graph of the calculated zinc extractions as a function of time in the process of FIG. 2.

particularly by a In the process of FIG. 1 a zinc concentrate and copper sulfate solution are fed into a pressure leaching vessel 1. The zinc concentrate comprises predominantly zinc sulfide with minor amounts of other naturally occurring constituents, such as iron sulfides. The copper sulfate solution is largely regenerated in thisproces's as referred to in greater detail below.

The mole ratio of copper to zinc feed to vessel 1 is from about 1:1 to 2:1, greater'amounts of copper may also be used. The preferred ratio of copper to zinc is about 1.5: l.

The temperature in the vessel is from about to 300 to 550F., preferably 400 to 450F. The vessel itself is a corrosion resistant vessel, having external temperature control means and an internal stirrer (neither shown). The reaction is carried out continuously with a normal residence time of from about 1 to 7 hours.

The reaction product, which includes a copper sulfide and zinc sulfate solution is withdrawn from the vessel 1 and passed through filtration means 2. The filtration means is a conventional unit for the continuous separation of a liquid-solid slurry. The impure zinc sulfate solution is sent to a conventional electrolytic zinc process plant starting with the purification procedure. The electrolytic zinc process is well known and described in numerous texts, for example, Zinc The Science and Technology of the Metal, Its Alloys and Compounds, C.

H. Mathewson, ed., pp. 174-225 (1959) and by Allen C.

Jephson et al. in Journal of Metals (18, 947-956 (Aug, 1966)). In unit 3 the solution is purified, clarified, and electrolyzed to produce cathode zinc. The cathode zinc is smelted and cast into slabs which are stored as slab zinc in ware house 4 for shipment. Precipitates are kept separated in storage means 5. The spent electrolyte or acid from the electrolysis is sent to the pressure oxidation unit 6. Air, and the precipitate from filtration means 2, which is largely a copper sulfide, are

concurrently fed to the pressure oxidation unit. This unit may be similar in structure to vessel 1. The reaction products from the means 6 are sulphur and a mixture of sulfates, predominantly copper sulfate with some zinc sulfate formed from zinc sulfide. The sulfates are recycled to the pressure leaching vessel l. The sulphur is purified and recovered in a conventional flotation means 7. 7

In this process copper sulfate solution wasfound to be an effective agent for leaching zinc sulfide concentrate. Experiments set forth below established that a temperature of 400F gave satisfactory zinc leaching rates. At this temperature, zinc extractions of about percent after 4 hours were obtained when using 1.5 moles of copper in the leach liquor for every mole of zinc in the concentrate. At 450F, an extraction of percent was obtained under the same conditions. The presence of zinc sulfate in the leach solution did not affect the reaction rate.

It was found that copper precipitates principally as cuprous sulfide (Cu S), and not as cupric sulfide (CuS). Thermodynamic calculations also predict that cuprous sulfide is the stable form. The precipitation of Cu S is accompanied by the formation of appreciable'amounts of sulfuric acid, which, however, is detrimental to maximum sulphur recovery. The following process minimizes this problem of acid formation.

FIG. 2 illustrates a preferred embodiment in which air or oxygen is fed to the pressure leach vessel 8 in addition to the zinc concentrate and copper sulfate. The oxygen partial pressure may be from about 5 to psig, preferably from 10 to 50 psig. The reaction mixture is filtered in means 9 and the solution therefrom is purified in means 10 with lime and zinc dust in accordance with conventional procedures. The purification precipitates are stored in 11. The purified solution is electrolyzed in means 12, from which zinc metal is sent to storage 13. The spent electrolyte, air, or oxygen, and the solids from the filtration means 9, are reacted in vessel 14. The pressure oxidation in 14 yields the regenerated sulfate solutions which are recycled into the pressure leach vessel 8. The residue from the pressure oxidation includes sulphur which is purified in flotation unit 15 and zinc sulfide which is separated in flotation unit 16. The zinc sulfide is returned to the pressure leach vessel 8.

The above processes are illustrated further by the following examples.

EXPERIMENTAL PROCEDURE NON-OXIDIZING CONDITIONS The conditions under which copper sulfate solutions can be used effectively for the leaching of zinc sulfide concentrate are shown in the following experiments.

These experiments were conducted in a Pyrex vessel contained within a S-gallon autoclave and equipped with electrical resistance heating and a cooling coil for automatic temperature control. A glass or tantalum stirrer was used, the latter being more suitable for higher pulp densities.

The zinc concentrate used in the experiments was analyzed as follows:

Cu 0.73% Fe 8.2% S 32.1%

The copper in the concentrate was present as chalcopyrite (CuFeS The major part of the iron was found in a solid solution with zinc sulfide.

The main variables in these experiments were temperature, copper content of leach solutions, and the presence of zinc in the leach liquor.

EXAMPLES 1-6 The effect of temperature on the reaction of zinc sulfide with copper sulfate is shown in Examples l-6. In all runs, 50 g of zinc concentrate was leached with 500 ml aqueous copper sulfate solution of indicated strength at various temperatures. The results of these experiments are listed in Table I.

n Zine contents of solutions were not determined because almost no copper had precipitated W The results indicate that a minimum temperature of 300F is required and that 400F is preferable to achieve acceptable zinc extractions in a reasonable time. The upper temperature limits are in part dictated by the equipment and economies. A suitable temperature range is from about 300F to about 550F.

The results indicate that more copper is consumed than necessary for the stoichiometric formation of CuS. For the formation of Cu S, the molar ratio of Cu precipitated to Zn dissolved is 1.6. The iron incorporated in the ZnS lattice will react similar to zinc. Some of the iron in the concentrate is combined with copper in chalcopyrite, CuFeS This mineral cent of the concentrate, is then incorporated in the sphalerite lattice. For every mole of zinc we therefore have 7.56/54.3 65.4/55.85 0.163 moles of iron, reacting similarly. Every mole of zinc requires then 1.163 X l.6=l.86l moles of 5 copper. Thecopper precipitate of the leach conducted at 400F (Table I, No. 6) should therefore not contain CuS. The same conclusion was arrived at from thermodynamic considerations. Examination of polished sections of residues from other leaches carried out at 400F confirmed the absence of CuS.

EXAMPLES 7-15 In a series of leaches at 400F, the effects of time and initial copper to zinc ratio are shown. Again 50 g of zinc concentrate was leached with 500 ml copper sulfate solution in most experiments; exceptions are clearly identified in Table II which indicates the results.

TABLE II M01 ratio Mol ratio of of Cu to Zn dis- Cu preeip- Cu precipi- Time, Zn in solved, itated, tated to Zn In charge percent percent dissolved l 1. 63. 2 05. 5 1. 84 2 1. 51 73. 1 88. 8 I. 85 3 1. 51 77. 5 06.1 1. 86 4 1. 51 70. 3 90. 2 1. 87 4 1. 25 77. 8 100. 0 1. 4 1.38 75.6 100.0 1.83 4 2.0 81.7 76.0 1.87 4 1.25 71.5 100.0 1.74 4 2.0 82.3 76.0 1. S5

' 5.3 grams H2504 added. 155 grams Zn-eoneentrate used.

Generally copper to zinc ratios of about 1.85 were obtained, indicating precipitation of Cu S. Exceptions are No. 11 and No. 14. In these examples the solution assays were confirmed by zinc assays of the residues. In both cases copper was completely precipitated. It is possible that under these conditions small quantities of CuS have been formed.

Comparison of Nos. 13 and 15 shows no effect of increased pulp density. This was to be expected because the concentration of copper ions per unit surface area was not changed in these experiments by increasing the pulp density.

EXAMPLES 16-19 Actual process liquors contain zinc sulfate carried over from spent electrolyte through the pressure oxidation (see FIG. 1). Some experiments were conducted which illustrated that this initial zinc content of the leach solution does not have any effect on the extraction. In addition, a pulp density was chosen to produce pregnant solutions suitable for normal zinc electrolysis practice. Previous examples had illustrated that a change in pulp density does not affect the rate of extraction.

In the earlier experiments at 400F, a molar ratio of copper in solution to zinc in concentrate of 1.51 was used most frequently. For reason of comparison, the same ratio was used in these experiments, while zinc sulfate was added to the leach solutions.

The data pertaining to these experiments are listed in Table 60 III.

These experiments show that the presence of zinc in the TABLE III Run N0 1G 17 1g 19 Solid e11arge 10 5;. Zn cone 100 5;. Zn eonc 100 g. Zn cone Residue from No. 18. Leaeli solution 4Synthetic 4Synthetic Pregnant liquor b from No 17 Synthetic. Time, hr 1 Temperature, 400... 450.

Zn extraction, per 79.4 85.

Fe extraction, percentot analyzed Cu in pregnant liquor, g.p.l 14.8 7.87 Nil. 6.68. Molar ratio Cu precipitated to Zn dissolved 1.75 1.73 1.26 1.82.

a 500 m1. liquor containing 163 g.p.l. Cu and 57 g.p.l. Zn. b 407 ml. liquor containing 7.87 g.p.l. Cu and 149.0 g.p.l. Zn.

does not react with copper sulfate, and polished sections ofresidue do not show a sign of attack. Since the concentrate leach solution has no noticeable effect on the extraction. An extraction of 79.4 percent was obtained in experiment No. 16

contains 0.73 percent Cu, the amount of iron combined with it which compares with 79.3 percent for run No. 10 in Table 11,

in CuFeS, is 0.64 percent. The balance of the iron, 7.56 perwhere no zinc was present in the initial leach solution.

.cess of copper in theleach solution is desirable. In experiments No. 17, No. 18 andNo. 19, it was shown that excess 1copper could be recovered in aleach in two stages. The excess copper in leach solution No. 17 was fully utilized in a l-hour leach withfresh zinc concentrate. As pointed out below, CuS

'is likely to be formed under these conditions. This is conifirmed by the copper to zinc ratio obtained in experiment No.

18. If then all the copper from the pregnant solution of run No. 17 is precipitated as CuS, an equal amount of copper will be required in run No. 19 to convert CuS, without dissolving any .zinc. This accounts for the higher copper to zinc ratio obtained inrun No. 19.

In the use of copper sulfate solutions for the leaching of zinc sulfide concentrate, the followingreaction may occur to produce cupric sulfide.

Alternatively the precipitation of cuprous sulfide can take place according to:

Measurements of the pH of leach liquors showed values generally below 0.7 and pH values of 0.2 were common. This indicated that the reaction (2) takes place to a considerable extent. Under these conditions the following reaction becomes important.

80,, H H50 3. Calculations show that at pH l, HSO, dominates strongly over SO.

' The second reaction then should actually be written:

8Cu SZnS 4H O 3S0, 4Cu S 5Zn 4HS0 4H". 2'. The relative importance of reactions (1) and (2) is shown by the following reaction:

8CuS 3Zn 4H O 350, 4Cu S 4HSO 4H 3ZnS 4. AF F -1 3,090 cal/mole formula, and

(HSO4-) (H+) 5 ZnH), one 1o These values show that at pH below about 0.5, CuS is stable relative to Cu S provided ZnS is present. During the copper sulfate leach, Cu S will first be precipitated on the ZnS particles. By the time the acidity of the solution is such that CuS could be formed, all ZnS particles will be covered with Cu S. This. inhibits the formation of CuS according to equation 1 Therefore, when the leaching operation is carried out in two stages, where in the second stage fresh zinc concentrate is contacted with pregnant liquor containing excess copper sulfate, the CuS would be formed.

EXAMPLE 20 The cuprous sulfide formed in thepressure leach can be leached'l with. acid under oxygen pressure to yield elemental sulphur. In: Canadian Pat. No. 712,989 granted to Sherritt GordonaMines Limited,andincorporated herein by reference, a'processis described based onthe reaction:

Although. this equation adequately described the overall reaction, calculations of thethermodynamics of the system indicateithatCus has to be formed as an intermediate product.

Inu-Example20 a 100 g sample of copper sulfide precipitate (residue: run. No. 16, Table III) was leached with simulated spentaelectrolyteunder 250 psig.air'at 225F for 2 hours. A residue'of. 74.1 g was obtained containing a traceof elemental sulphur. Av polished section'of the residue revealed that the ori'ginalxcuprous sulfide was almost completely converted to cupric sulfide. lnthis stage of the oxidation process no elemental sulphur can .befonned.

In a continuation of this experiment, 42.8 g of the firsttreated residue was mixed with 52.2 g fresh chalcocite (Cu S) and leached under the same conditions. A residue of 62.0 g was obtained with 4.83 percent elemental sulphur.

This experiment shows that the oxidation of cuprous sulfide takes place in two steps. In the first step, cuprous sulfide decomposes into cupric sulfide-and copper ions. The formation of elemental sulphur from cupric sulfide takes place only after virtually complete conversion of Cu S.

As noted in Canadian Pat. No. 712,989 about 1 mole of sulphuric acid may be used for each mole of copper. The oxygen pressure is from about 5 to about l00'psig, and the temperature from about 175F to 225F. The Cu S is selectively oxidized at a rapid rate until about 95 percent of the copper is extracted and the sulphur bound to it is converted to elemental sulphur. Present results indicate that about percent of the sulphur from decomposed cupric sulfide reports as elemental sulphur.

The copper sulfate formed in this leaching step is recycled to react with further zinc sulfide. The zinc sulfate which is formed is similarly recycled and as noted above has a negligible effect in the first leach step in 1, from which it passes to the electrolysis unit.

The results obtained in Examples l-l9 by leaching zinc concentrate with copper sulfate solutions were generally much better than the results obtained by leaching with sulfuric acid. Extractions of about 80 percent were obtained after 4 hours at 400F with a copper sulfate deficiency. A slight excess of copper resulted in about 82 percent extraction. An extraction of 85 percent was achieved in a 4-hour leach at 450F in spite of the copper deficiency. Thus extractions of about 80 percent are readily obtained under proper conditions. Above this point, the rate of extraction decreases because of the smaller surface area available for reaction.

Several alternatives are available to enhance the dissolution of zinc. One is to physically separate and regrind the coarse fraction prior to leaching. Alternatively, the unreacted zinc sulfide particles may be left behind in the residue and recovered during subsequent treatment of the residue. In the copper sulfate regenerating step (pressure oxidation, FIG. 1), the copper sulfide layer will be removed from the zinc sulfide particles. By selectively removing the unreacted zinc sulfide from the final residue by flotation, the zinc sulfide can be recycled, to result in high zinc recoveries. An additional advantage of this method is a beneficial effect on the sulphur recovery. The presence of zinc sulfide in the leach residue would result in milder oxidizing conditions towards the end of the copper sulfate regenerating step. This inhibits the formation of sulfuric acid and this improves the yield of elemental sulphur. Thus the presence of unleached zinc sulfide in the leach residue does not constitute a drawback for the overall process. This alternative is depicted in FIG. 2, but of course could also be employed in the method of FIG. 1.

A disadvantageous factor in the process of FIG. 1 is inherent formation of sulfuric acid which was noted to occur during the precipitation of cuprous sulfide. Because conventional procedures for purification of the pregnant solution prior to electrolysis require essentially neutral conditions, the use of a neutralizing agent might be required. A practical solution is a change in the leaching procedure to ensure production of CuS rather than Cu S, as illustrated in FIG. 2. The parameters for this process are illustrated in the following sequence of examples.

Zinc concentrate was leached with copper sulfate solutions at temperatures ranging from 300 to 450F. under oxygen pressures from 20 to psig. Copper precipitated initially as cuprous sulfide and leach solutions contained considerable quantities of acid. Immediately after precipitation of cuprous sulfide, an oxidation reaction converted cuprous sulfide to cupric sulfide under consumption of acid. This second reaction was strongly promoted by high oxygen pressures while the first reaction was promoted by high temperatures. Leach solutions obtained at 400F. and 20 psig oxygen contained about 20 percent of the sulphur in solution as acid after 4 hours. Solutions from experiments at 300 F. and 100 psig oxygen were essentially free of acid at any time.

With time cupric sulfide dissolved, especially under high oxygen pressures. Because of this reaction, from 33 to 65 percent of the sulphur in the zinc concentrate was solubilized. The results indicate the advantages of a staging operation to combine high zinc extraction with low sulphur dissolution and an essentially acid free solution. Thus with .a low zinc extraction, obtained by using an excess of zinc concentrate, the resulting leach solution is well suited for solution purification prior to zinc electrolysis. Elemental sulphur is produced from cupric sulfide in the residue, iron is rejected and excess zinc concentrate is recycled. This process provides for higher zinc recoveries than possible in the conventional roasting-electrolysis process and is suitable for treating middlings and other copper-zinc sulfide products.

EXPERIMENTAL PROCEDURE OXlDlZlNG PROCEDURES The zinc concentrate used in the following examples was analyzed with the following results: Zn 51.2 percent; Cu 0.96 percent; Fe 8.3 percent; and S 32.0 percent.

The experiments were performed in an Autoclave Engineers S-gallon autoclave that was modified by placing a tantalum reaction vessel within a steel sleeve holder. The space between the steel sleeve holder and the autoclave wall was filled with water. A cooling coil was immersed in the water. This arrangement made it possible to cool the reaction mixture rapidly after termination of an experiment without exposing to coil to the corrosive leach liquors. The autoclave was heated to the desired reaction temperature by electric'resistance heating; constant temperature was maintained by automatic control. During the experiments the slurry was vigorously agitated by a tantalum stirrer. Pressurized air was introduced into the solution through stainless steel tubing with its opening approximately 2 inches above the stirrer blades. Pressure was maintained at the desired level by a pressure regulator on the air cylinder. To avoid oxygen depletion during the reaction, gas was constantly removed from the autoclave at a rate of approximately 1.25 liters per minute (atmospheric pressure and room temperature). In each experiment the solid charge consisted of 200 grams of zinc concentrate which would require about liters of oxygen to react completely according to Reaction (1). Consequently sufficient oxygen for complete reaction was provided every hour.

Slurry samples were taken at regular intervals. For this purpose a valve in the air supply line was closed momentarily and a second valve between the first one and the autoclave was opened to the atmosphere. The pressure in the autoclave forced a slurry sample up through the air inlet tube into a graduated glass cylinder. in most experiments two samples of 100 ml volume each were removed during a run.

EXAMPLES 21-26 A series of six leaching experiments was performed in which the temperature was either 300F, 350F, or 400F, and the partial pressure of oxygen above the solution was 50 psig or 100 psig as indicated below. In all experiments, 200 grams of zinc concentrate was leached with one liter of solution containing 95.5 grams copper as cupric sulfate. The amount of copper present in the leach solutions was 105 percent of the amount required to dissolve all of the zinc and the iron associated with it in the zinc lattice as calculated from the reactron:

5ZnS 4Cu 5Zn 4CuS S0,," and the analogous reaction for dissolution of iron.

The first experiment was terminated after 3 hours. The five other experiments each lasted 6 hours and slurries were sampled after 1 hour and after 3 hours.

Solution samples were analyzed for Zn, Cu, Fe, and S.

After the experiments were terminated, the slurries were filtered, the residues washed and dried and the filtrates combined with washings made up to 2 liters. Solutions and residues were both analyzed for Zn, Cu, Fe, and S. The results of these experiments are presented in Table IV.

TABLE IV 200 grams of concentrate, 1 liter solution, (1 hours except N0. 1

Experiment No 21 22 23 2t 25 21;

Temperature, F 300 300 350 400 300 350 Oxygen partial pressure,

p.s. .g 100 100 100 100 50 50 Solution after 1 inn, g.p.i

Zn 18. 3 27. 1 39. 5 16. 2 27. (i 63. 7 41.7 14.0 57.5 35.7 2. 07 1. 74 3. 26 2. 24 2. 38 42. 8 40. 7 40. 3 39. 8 39. 7

40.7 35.0 41.4 39. 7 33. 8 29. 6 20. 1 27. 7 17. 5 17. 8 2. 96 2. 24 1. 58 3. 15 3. 88 S 38. 4 32. 2 40. 9 32. 2 30. 8 Final residue weight, g 156.0 98. 7 132. 7 109. 3 137. 8 150.0 Residue analysis, percent:

Zn 33. 8 17. 2 19.1 18. 8 15. 3 20.3 21.0 29. 0 36.0 36. 8 39. 3 37. 7 7. 69 11.0 9.29 12.6 7.52 7. 09 S 30.1 25.9 27.1 24.1 29.0 28.4 Zinc extraction, percent. 49. 5 83.8 75. 1 76. 5 80. 3 71. 1

QLBased n B liters solution. N.S.=no sample.

Observation of solution samples and residue weights indicated that the best elimination of copper had been achieved within the first 3 hours of Run No. 24, at 400F, and that some benefit could be expected from lower oxygen pressures. Accordingly further experiments were conducted at relatively high temperatures and low partial pressures of oxygen. The leaching time was reduced to 4 hours and the solution was sampled after 1 and after 2 hours. The results are discussed in further detail below.

EXAMPLES 27-30 Three experiments were performed at 400F and 20 psig oxygen, and one at 450F, and 20 psig oxygen. In all experiments 200 grams of zinc concentrate was used. initial solution volume was one liter in three experiments, and 400 ml in the other run.

Results of these experiments are shown in Table V. Zinc extractions for all experiments were based on the distribution of zinc between solutions and final residue.

TABLE V 200 grams concentrate, 20 p.s.i.g. oxygen, 4 hours, 1 liter solution except as noted Experiment N o 27 28 29 30 Temperature, F 400 450 400 400 First sample b solution, g.p.1.:

Zn 25.2 40.7 41.4 N.S.

" 400 ml. solution. After 1 hour for Runs 7 and alter hour for Run No. .1. e After 2 hours for Run No. 7; after 1% hours for Run No. 9. Based on distribution of zinc between solids and solutions. N.S.=no sample.

Most leach solutions contained acid, evidence of formation of cuprotm xulfide. This was also confirmed by sulphur deficicncicn in the residues if all the copper was assumed to be present as cupric sulfide. The amount of sulphur as acid inbeen formed by Reactions (2) and (7). Hence, the increase of sulphur in leach solutions was initially determined by the stoichiometry of Reaction 2). Because iron reacted proportionally to zinc, the increase in sulphur under these conditions creased wiiii increasing p l r and ranged from about 3 was directly proportional to the amount of zinc dissolved. In percent at 300 F to about 20 percent of the total amount of ll experiments h amount f copper in solution had sulphur m so utlon at 4 F- decreased sharply by the time the first sample was taken. Thus most Predominant i'eaciioiis f i Piace dumig Reaction (9) did not occur to any measurable extent in the ii s i gcg gga tigg sgi g under Conditions p y in first hour. On this basis minimum values for the solution i volumes at the time of first sampling were calculated. SZnS 18 $3236 2 2:; f g 'zg 'g Similarly maximum values were calculated assuming Reaction 5ZnS z+ ICu 20 q SZX 1' 46:8 S Z- (9) did proceed at a constant rate throughout the entire run. 2 4 The zinc extraction curves are shown in FIG. 5. It appeared I m F3 a g i A S t R that Reaction (9) could indeed be neglected for the first hour a l s i e e c e en y a Ogou o of each experiment. Approximate zinc extractions at subt sampling times, were calculated by assuming Reac- SFeS 8cu++ 4H,o 5Fe+ 401 s 1+ are 50,-. f

tion (9) to proceed at a constant rate. 10. Part of the dissolved iron subsequently precipitated according It has Shown that by choice of leaching condl to one or more of the followin reactions tlons, cupric sulfide can be obtained as the condensed reac- 4Fe++ O 10" 4Fe(O'H) 1+ 8H+ H tion product without fon'nation of appreciable quantities of 4Fe+Aq +022+ 45042: 2H2O 4"Fe(OI nso4 l acid. To achieve adequate zinc extractions however it was the reaction time which caused redis- 4Fe A? 411+ 0 660 2Fe (SO 2n,o. ,3. necesary 9 In View of the Small quantities involved any acid formation solution of cupric sulfide. ThlS is equivalent to loss of sulphur due to precipitation of iron would be negligible. The presence from the a i it a gt iz resissoiigion of acid in the leach solutions is thus attributed to a cupric i e on Y Occur? a er a Coiisi era 6 P predominance of the acid forming Reaction (2) over the acid Zinc had b'eeii dissoivedy Providing sufficient Zinc consuming Reaction 7 fide area, 1.e., by using sufficlent excess ZlllC concentrate, the

The amount of sulphur found in final leach solutions in- 3O i q i y 0f redissoiuiioii of cupric Sulfide would be dicated that Reaction (9), dissolution of cupric sulfide, took mlmmlled- The molar w o n to Cu Should be greater than place. If Reaction (2), and (7), only take place, only 16 to 20 iii and p y from 1251 to i a P to percent of the sulphur in zinc concentrate should have been Alternatively when lower Yams of Zn 9 Cl! are Used, as dissolved during leaching. However, 33 to 65 percent was acfrom 1:1 to 1:2, the reaction should be carried to about to tually dissolved. 35 80 percent conversion, preferably from to percent con- Table VI shows the dissolution of sulphur from zinc concenversion, based upon the extraction of Zn from the ZnS. The trate for the various experiments. Also shown are the percentpregnant solution of such a leaching operation is essentially ages of sulphur converted to acid. N A i free of copper and acid, and suitable for solution purification TABLE Iv Actual S Zinc Theoretical B dissolution, S conversion Oxygen cxtrac- S dissolution, percent of S to acid, percent pressure, tion, percent of S in in conof total S in p.s.i. percent concentrate centrate solution 100 83. a 19. a 60.9 2. 6 100 75. 1 17. a 43.8 9. 2 100 76. 6 17.6 62. 6 1a. 1 50 80.3 18.5 40.4 3.6 60 71. 1 16. 4 35. a s. s 20 72. 4 16. 6 33. 0 20. 2 20 86. 6 19. 9 64. 6 16. 7 20 71.6 16. 5 as. 1 19. u 20 77. 9 17. 9 47. 5 6. 5

' Assuming no redissolution of 0118.

in onclitersolutiou.

It can be seen that up to 400F, dissolution of sulphur was largely controlled by oxygen pressure. Best sulphur retentions were obtained at 400F and 20 psig oxygen pressure. The experiments at higher oxygen pressures indicate a minimum sulphur extraction at 350F. Thus sulphur extraction would have been lower at 350F. and 20 psig than at 400F and 20 psig oxygen. Zinc extraction, however, was relatively low in both experiments at 350F, and quite large quantities of acid were formed. In FIG. 3, acid formation is shown as a function of time for all experiments where samples were taken. In all cases the amount of acid formed was at a maximum at a time prior to termination of the experiment. This indicates that Cu S was the primary copper precipitate and that CuS was formed by subsequent oxidation. FIG. 4 shows the ratio of acid sulphur versus zinc as a function of time. The general similarity of the corresponding curves in FIG. 3 and 4, indicates that the decline in acid content of the solutions was caused by the acid consuming Reaction (7) and not by an increase in total sulphur content according to Reaction (9). The data presented in Table VI show, however, that Reaction (9) did occur. Reaction (9) can only take place after CuS has prior to zinc electrolysis. The leach residue, containing cupric sulfide, oxidized iron and excess concentrate can be leached with return acid under conditions conducive to the formation of elemental sulphur. The resulting solution is suitable for leaching additional zinc concentrate. The solids can be separated by physical methods e.g., flotation, into three fractions; elemental sulphur, zinc concentrate, and iron reject. Zinc concentrate is of course returned to the copper sulfate leach. This shows the importance of excess of zinc concentrate in the initial stage to avoid dissolution of cupric sulfide. The use of excess concentrate in the first stage of the process is therefore considered a preferred method of operation.

This process, outlined in FIG. 2, provides higher zinc extractions than possible with the conventional roasting-electrolysis process, because the roasting step with its inherent formation of insoluble zinc ferrites is eliminated. The process can also be integrated with a conventional installation for treatment of middling fractions from the flotation circuit. Another application is the treatment of bulk flotation concentrates or treatment of concentrates from ores that cannot be effectively separated by selective flotation.

This invention has been described in terms of specific embodiments set forth in detail. Alternative embodiments will be apparent to those skilled in the art in view of this disclosure, and accordingly such modifications are to be contemplated within the spirit of the invention as disclosed and claimed herein.

We claim:

1. The process which comprises mixing zinc sulfide ore and a copper sulfate solution at an elevated temperature and pressure to form a copper sulfide and a zinc sulfate solution,

separating said copper sulfide from said zinc sulfate solution,

purifying said zinc sulfate solution, electrolyzing said zinc sulfate solution to form metallic zinc and sulfuric acid, mixing said copper sulfide with oxygen and said sulfuric acid at elevated temperatures and pressures to form copper sulfate solution and sulphur, separating said copper sulfate solution therefrom and mixing therewith fresh zinc sulfide ore to repeat said process.

2. The process of claim 1 wherein the mole ratio of zinc to copper is from 2:1 to 1:2.

3. The process of claim 1 wherein the mole ratio of zinc to copper is 1:1 to 1:2 and the conversion of zinc sulfide to zinc sulfate is from 40 to 80 percent.

4. The process which comprises mixing zinc sulfide ore and copper sulfate solution and an oxygen containing gas in a first pressure leaching vessel, discharging the reaction mixture to a filtering means, withdrawing the pregnant solution from the filtering means and treating said pregnant solution in a purification and electrolysis means to produce zinc metal and a spent sulfuric acid electrolyte, charging said sulfuric acid, air and the solids from said filtering means to a second pressure leaching vessel, regenerating copper sulfate solution in said second leaching vessel, recycling said copper sulfate solution to said first pressure leaching vessel, discharging the residue from said second leaching vessel to a sulphur flotation means, withdrawing sulphur from said flotation means, withdrawing unreacted zinc sulfide from said flotation means and recycling said zinc sulfide to said first pressure leaching vessel.

5. The process of claim 1 in which the amount of oxygen in said first leaching vessel is about 20 to 100 psig.

6. The process of claim 1 wherein the ratio of zinc sulfide to copper sulfate in said first pressure leaching vessel is in excess of the molar quantities needed for reaction.

7. The process of claim 1 wherein the ratio of Zn to Cu in said first pressure leaching vessel is 1:1 to 1:2 and the conversion of zinc sulfide to zinc sulfate therein is from 40 to percent.

8. The process of claim 7 wherein the charge to said second leaching vessel is unreacted zinc sulfide, a copper sulfide, and about 1 mole of sulfuric acid per mole of copper.

2 g UNITED STATES PAT ENT omen CERTIFIC ATE 0F CORRECTION Patent No. v Dated April #1972 I Inv n C's) Howard C. Renken and Theodoor 'W. Ze'gers- It is certified that error appears in the above-identifies patent and that said Letters Patent are hereby corrected as shown below:

' In the Abstract, line 2, delete "solution."

In the Abstract, line 3, delete the entire line 3 V In the Abstrzatct, line 9, after "sulfate" insert solution. To providefor increased recovery the copper suifate-e- Col. 6, lihe.50,' change "this" ,to thus Col. 7, line 33, after "ing" change "to" to the Col. 9",""iW8fi Tnsert 2+7j Col. 9, line 22, change "new" to we Col. I 9, line 23; change "l +Fe*TA?" to 4Fe Signed and sealed this 25th day of July 19-72.

(SEAL) Attest:

EDWARD M.FLETCHER,JR. ROBERT GOI'TSCHALK Attesting Officer 1 Commissioner of Patents

Patent Citations
Cited PatentFiling datePublication dateApplicantTitle
US1553414 *May 18, 1923Sep 15, 1925Inspiration Cons Copper CompanMethod of leaching sulphide and mixed ores
US1890934 *Apr 19, 1930Dec 13, 1932Campbell Carson GeorgeMethod of and means for producing sulphates from mixed sulphide materials and the recovery of values therefrom
US1937634 *Sep 23, 1931Dec 5, 1933Christensen Niels CProcess of treating zinc ores
US3095363 *Feb 17, 1960Jun 25, 1963Anaconda CoCalcination of zinc sulfide concentrates
SE107982A * Title not available
Referenced by
Citing PatentFiling datePublication dateApplicantTitle
US4049770 *Mar 2, 1976Sep 20, 1977Sherritt Gordon Mines LimitedRecovery of copper and zinc as sulfides from copper-iron sulfides
US4260588 *Sep 10, 1979Apr 7, 1981Duisburger DupferhutteProduction of sulphidic copper concentrates
US4439288 *Jul 11, 1983Mar 27, 1984Exxon Research & Engineering CompanyProcess for reducing Zn consumption in zinc electrolyte purification
US5651947 *Nov 7, 1995Jul 29, 1997698638 Alberta Ltd.Recovery of zinc from sulphidic concentrates
US5711922 *Mar 26, 1996Jan 27, 1998R & O Mining Processing LtdPreferential hydrometallurgical conversion of zinc sulfide to sulfate from zinc sulfide containing ores and concentrates
EP0214324A1 *Sep 5, 1985Mar 18, 1987Cheminvest A/SMethod for separation and leaching of the valuable metals in metal sulphide materials
WO1985003952A1 *Mar 8, 1985Sep 12, 1985Cheminor AsMethod for separation and leaching of the valuable metals in metal sulphide materials
WO1996005329A1 *Aug 9, 1995Feb 22, 1996R & O Mining Processing LtdHydrometallurgical conversion of zinc sulfide to sulfate from zinc sulfide containing ores and concentrates
Classifications
U.S. Classification205/607, 423/37, 423/109
International ClassificationC25C1/00, C25C1/16
Cooperative ClassificationC25C1/16
European ClassificationC25C1/16
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