CA1076367A - Process for the treatment of complex lead-zinc concentrates - Google Patents

Abstract

PROCESS FOR THE TREATMENT OF COMPLEX LEAD-ZINC CONCENTRATES

ABSTRACT

Sulfide concentrates containing metals such as lead, zinc, copper and silver are treated to selectively recover these metals in high yields in a process wherein the concentrate is sequentially leached or extracted with sulfuric acid in the presence of oxygen; lime in the presence of sulfide ion; and ferric chloride. The leach solutions resulting from each extraction may then be treated to recover metals in their elemental state. The process does not produce sulfur oxides.

Description

J~

This invention relates to n~etallur~y, and ~
more particularly, to the selective recovery of _ metals such as lead, copper, zinc and silver by :~t hydrometallurgical techniq~es. :~
Pyrometallurgical processes for recovering metals such as lead from ore deposits are well-known in the art and used extensively throughout the industry. However, since the lead in such ores ;
exists primarily in combination with sulfur or sulfur-bearing materials~ pyrometallurgy leads to the oxidation of these sulfur materials and the formation of sulfur oxides.
Air pollution regulations and standards 7 ~
promulgated by both federal and state agencies _ in recent years, have placed severe restrictions on permissible sulfur oxide emission levels from such processes. While the industry has developed numerous methods for recovering sulfur oxides emitted from pyrometallurgical processes, such techniques are ~
20 highly expensive. Indeed, the anticipated advent of ~ , yet more stringent pollution standards may impose 9~"
technologically impossible, as ~lell as economically unfeasible, demands on the industry.
Hydrolnetallurgical techniques for re-25 covering metals, however, offer an attractive A',' alternative e~nission control approach since they do not result in the forrnation of sulfur oxides.
Rather, hydrometallurgy permits the recover of sulfur in its elefllental form. Despite this seem-in~ly simplc solution to thc emission problem, ~., .

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however, hydrometallurgy is economically competitive with pyrometallurgical processing only if near complete recovery of the desired metal from the ore is e~fected, a condition difficult to achieve on a commercial scale utilizing current hydrometallurgi-cal techniques. To offset these costs it is further desired to effect complete recovery of all other metals present within the ore.
The foregoing problems are particularly pronounced where the desired metal exists in a low grade deposit since the costs per unit of metal produced are extremely high. In particular9 low-grade lead concentrates containing copper, zinc and silver cannot be economically processed unless a near complete recovery of all these metals is obtained. Heretofore, hydrometallurgy has not been successful in selectively recovering the variety of metals contained in such concentratesO
It is accordingly an object of this invention to selectively recover, in high yields, a number of valuable metals from ores containing such metals.
A further object of this invention is to accomplish the foregoing recovery utilizing hydrometallurgical techinques.
A more specific object of this invention is to selectively recover lead, copper, zinc and silver from low grade concentrates containing such metals.
These and other objects will become more r~

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apparent upon reading the more detailed description which follows.
In accordance with this invention, lead concentrates are treated to selectively recover high yields of lead, copper, zinc and silver therefrom by a process comprising leaching the concentrate Witil sulfuric acid in the presence of oxygen, leaching the ..
residue therefrom with lime, and leaching the remain~ :~
lng residue with ferric chloride. .~
The present invention describes a process _ for treating sulfide concentrates containing lead, copper, zinc and silver to selectively recover said .
metals therefrom,comprising the steps of: a) contacting said concentrate with sulfuric acid in the presence of oxygen at elevated temperature and pressure to extract copper and zinc materials from said concentrate; b) thereafter contacting said concentrate with lime to remove elemental sulfur therefroln; and c) thereafter by contacting said concentrate with a mixture of calcium chloride and ferric chloride to extract lead and silver materials ~ :~
therefrom;
In particular, the present invention ls directed to a process for treating sulfide concen-trates containing lead, copper, zinc and silver to selectively recover said metals theref`rom, comprising .
the steps of: a) contacting said concentrate with sulfuric acid in the presence of oxygen at a pressure ~

of about 60 to 80 psig and a temperature of at least _ about 100C; b) dividing the concentrate treated in

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~:37636~ It.. -a) into a first residue fraction and a first super- `
natant fraction; c) treating said first supernatant fraction to recover metallic copper and zinc therefrom;
d) contacting said first residue fraction with lime at a temperature of about 95C; e) dividing the material resulting from d) into a second residue fraction and a second supernatant fraction; f) treating said second supernatant fraction to recover , elemental sulfur therefrom; g) contacting said second residue fraction with a mixture of ferric chloride and calcium chloride and removing a third supernatant fraction from the material resulting from said contacting; and h) treating said third supernatant to .
recover metallic lead and silver therefrom. ~$~
In the initial leaching step, hereinafter refer-red to as "oxygen leaching," the lead concentrate is contacted with sulfuric acid in the presence of oxygen. In this manner, copper and zinc are extrac-ted from the concentrate as sulfates, lead sulfide is converted to lead sulfate, and a portion of the sulfur converted to the elemental state. In a preferred ~ s embodiment of this invention, chloride ion is also present to aid copper dissolution. ~$
The oxygen leach is preferably conducted under elevated temperature and pressure. Preferred condi-tions in this respect are temperatures above about 100C and pressures in the range of about 60-80 p~ig.
The temperature is preferably maintained below the melting point of sulfur to prevent molten sulfur from coating the mineral particles. The tirne or extent of _5_ ~`

~L~76367 I
i.`,,~,`-~, the oxy~en leach will necessarily vary ~epending upon the pressure, concentration of reactants and metals 3 and other lil<e factors. The requisite time period should be sufficient to effect complete dissolution i

5 of zinc and copper, yet not so long as to result in :
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overleaching of iron.
As a result of the oxygen leach, there ~ ,~
remains a residue of the so~treated concentrate and a solution containing the extracted zinc and copper compounds~ This solution is then treated, as described h~reinafter, to isolate and remove metallic copper and zinc.
The residue from the oxygen leach is then contacted with lime. This lime leach results in the removal of elemental sulfur and is attended by the reconversion of lead sulfate to lead sulfides. This leach is preferably conducted in the presence of added sulfide materials. As will be described in more cletail hereinafter, critical conditions are believed 23 to exist f`or both the quantity of lin)e employed in the lime leach, and the temperature at which the ~?,',`~
leach is conducted. The primary purpose of this le~ch is to ~nhance the subsequent extraction and recovery of silv~r.

Reconversion of lead sulfate to lead sulfide in the lime leach is very important to the process reagent balance.
Reconversion allows chlorine produced during lead chloxide electrolysis to be recyc].ed to the leach as ferric chloride produced in the oxidation stage.
The residue remaining from the lime leach is then further treated to remove silver and lead while the ~. , ~t ~
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solution resulting from the lime leach is treated to remove sulfur. The former recovery involves contacting the residue with ferric chloride and calcium chloride at a re-duced pH. By virtue of this leach, lead and silver are dis-solved and further processed to selectively recover the metallic forms thereof.
As may be seen frorn the diagrammatic flow-sheet, a sulfide concentrate containing, inter alia, ~.

lead, eopper, zine and silver is admitted to an appropriate autoclave apparatus designed to withstand ~`
pressures of at least 100 psig. The concentrate is eontaeted with sulfurie aeid and ~he pressure within the autoelave brought to between about 60-80 psig with oxygen. The temperature in the autoclave is main- L
tained at greater than 100C and the leaching is eondueted for two hours.
It is preferred to include chloride ion in the autoclave from an appropriate source to aid the dissolution of copper and to prevent silver solubilizing ~ as a sulfate. Generally, a coneentration of from about 100-200 parts per-million chloride is utilized based upon the solids weight in the concentrate.
In this oxygen leach, the following princi-pal reaetions are believed to occur:

2 2 S4 + 2 ~CuSo4 ~ FeS04 -~ 2S -~ 2H 0 ZnS + H S0 ~ 1/20 -~ZnS04 + S + H20 PbS + H2S04 + 1/202--~ 4 2 2FeS04 + H2S01~ + 1/202 ~ Fe2(S04) ~ H20 Thus, zinc and copper are extracted from the concen- ~
trate as sulfates while a portion of the sulfur ~_ present is converted to elemental form.

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The leach solution removed from the auto- I
clave is then processed to remove copper and zinc, as well as other metals such as iron or cadmium. Iron may be removed as either jarosite, gocthite or helnatite.
Copper is cernented with finely divided zinc metal and zinc recovered by electrolysis.
The sulfuric acid utllized in the oxygen leach may be provided, in whole or part from spent zinc electrolyte solution resulting from the electro- ~
lytic recovery of zinc metals from zinc sulfate, `
sulfuric acid being regenerated by the anode reaction.
Such zinc spen~ electroly~e typically has a con- ~.
c~ntration of f.rom 100-200 grams/liter of sulfuric ac id .
lS The oxygen leach residue, containing lead sulfate and elemental sulfur, is next leached with lime and material capable of providing sulfide ion such as polysulfides. While it is possible to forego this treatment and pass the residue from the oxy~en leach directly to the chloride leach, described in !`
more detail hereinafter, to remove silver and lead, it has been determined that silver recoveries may be '~
enhanced by first extracting the oxygen leach residue ~
~itll lime to remove elemental sulfur thereform. This ~ -2~ lime leach adclitionally results in the reconversion of lead sulfate to lead sulficie. It has been deter-mined that at least a ten percent excess of lime over thc stoictliometrically requir-~ed amount is needed to L
effect the conversion of lead sulfate and the leaching of sulfur. Sulfide ion is pref`erably provided as r~-~
leach solutior. recycled before acidification to remove sulfur, as described hereinafter.

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~lternatively~ hydrogen sulfide may be added to the lime leach solution. Since the conversion of lead sulfate to lead sulfide consumes polysulfides, suffici~nt polysulfide to convert about ten percent of lead sulfate to lead sulficle is generally required.
The temperature at which the lime leach is conducted must be at least 95C to insure a proper reaction. The principal reactions believed to occur are as follows: ~ r~

3Ca(OH2) ~ 12S ~ 2CaS5 ~ CaS203 + 3H20 ~
Pb~0l~ + CaS5 + 2~120 --~PbS + CaS04.2H20 + 4S _ The purification of the leach solution is effected by addition of sulfuric acid to a pH of about 2.0 to precipitate gypsum sulfur according to the following reactions:

CaS5 ~ H2S04 + 2H20--~CaS04.2H20 +H2S ~4S
CaS23 -~ H2S04 ~ H20 -~ CaS04 2H2o 2 Should thiosulfate levels within the leach liquor become unacceptably high, the thiosulfate can be converted to gypsum and sulfur by using pressure oxidation with oxygen.
The residue remaining from the oxygen and lime leaches is then treated with ferric chloride and calcium chloride to dissolve lead and silver contained 25 therein according to the following reactions: `;~
PbS + 2FeC13 ->PbC12 + 2FeC12 +S i PbS0l~ + CaC12 -~ 2H20 -/PbC12 + CaS0~.2H20 Ag2S ~ 2FeC13 `~2AgC1 -~ 2FeC12 -~ S
In this leach it has been determined that it is necessary to maintain the system at a pH below _9_ ~C~7 Ei3~7 a~out 1.5. Suitable quantities of any acidic material such as hydrochloric acid may be added to attain this conaition.
Insufficient acidity results in the hydrolysis o~ ferric ion which, in turn, reduced the filterability of the residue and decreases silver recovery. The temperature at which this leach is performed is preferably in the range of about 70 to 90C.
Additionally, separation of the leach solution from the resulting residue should be conducted at temperatures above that at which lead chloride will crystallize.
Silver is preferably recovered from the resulting leach liquor by cementation on lead according to the reaction:

2AgCl ~ Pb ___~ PbC12 + 2Ag Prior to cementation any ferric iron is reduced using stoichiometric quantities of lead powder. Additionally, such a process will recover any copper not recovered during the oxygen leach.
The low silver loadings in solution generally neces-sitates the use of an extremely fine lead powder to effect cementation. It is preferred to employ an excess of lead in the cementation process. A lead silver weight ratio of about 5:1 generally provides a slight excess over stoichiometric requirements for both copper and silver. Temperature is held at about 8C which is sufficient to maintain lead solubility during cementation.
Alternatively, silver may be recovered by chemical precipitation with hydrogen sulfide produced during the earlier mentioned acidification of lime lead liquors.

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1, Lead is recovered from the leach liquor by crystallization of lead chloride and subsequent electrolysis. i~
Crystallization of lead chloride occurs at solution temperatures ~s of about 30C. It is preferred to perform the crystallization af~er s,ilver recovery to avoid losing silver during the crystallization.
The lead chloride is then fed to an electrolysis cell along with a suitable electrolyte wherein it is reduced '.
to elemental lead.
Chlorine produced in the electrolysis of lead chloride is used to regenerate ferric chloride for recycle to the leach. ,~
2 FeC12 ~ C12 2 FeC13 ;
Iron and zinc loadings in the leach solution are controlled by a small bleed from the circuit. Hydrogen sulfide from the acidification step is employed with lime addition ~q~
to remove a lead precipitate for recycle to the brine leach and a zinc-iron precipitate for recycle to the autoclave leach.
~i.
The stripped solution is returned to the circuit as residue b wash solution.
The use of the hydrogen sulfide generated in the acidi~
~i..'`'' fication step and the chlorine generated in the fused salt electrolysis provides a self-sustaining regeneration system, protects the environment adjacent the plant, and reduced opera-ting costs.
The following examples are presented to further ¦, describe and illustrate the process of this invention.
EX~MPLE I
A series of tests were performed wherein the pressure and time of the oxygen leach were varied in order to - 11 - ~, , ~ `

3~7636~

determine optimum conditions for the extraction of copper and ;~,.,"~
zinc from a low grade lead concentrate. Each test utilized 250 grams o~ concentrate and 500 milliliters~of a leach solution containing 200 grams/liter sulfuric acid and 148 grams/liter of ~zinc sulfate. The leach was conducted at 100C in a stirred Parr autoclave. The results o these tests are summarized in Table I.
TABLE I ;~

Test Conditions Autoclave Leach ~esults Pb No~ Time Oxygen % Extraction Conversion (psig) Ag Cu Zn Fe to PbSO
~ 0 0~02 ~10 43.3 7~. lO.~ i 153 2 40 0.02 0.11 85.6 91.316.7 72.3 1S5 l~ 40 0.02 0.16 85.9 94.119.7 77~6 166 7 40 0.3 0.10 89.2 97.3 - 88.7 160 l 80 0.02 0~ lO 87.8 87.118.2 70.7 159 2 80 0.02 0.10 90.9 92.221.3 75.2 ,.162 4 80 0.02 0.10 95.7 97.724.3 80.0 163 7 80 0.02 - 96.2 98.0 - 81.8 125 2 40 - - 86.1 91.3 - - ~ ~ .~,~127 2 50 0.02 0.12 91.9 92.6 - 57.2 ~ ~;
128 4 50 0.02 0.11 97.2 98.6 - 65.6 ~ l F
137 2 40 0.03 0.15 93.2 95.7 - 74.8 A further series of tests were performed using a lead concentrate containing by weight 12.33% lead, 118 ounces per ton o~ silver, 4.3~% copper and 14.04% zinc.
The leach solution contained 200 grams/liter of sulfuric acid and 148 grams/liter zinc sulfate. Pressure ~!
,~ .,, during the extraction was maintained at 40 psig : ~
(oxygen) and the temperature was 100C. The results of these tests are summarized in Table II.

TABLE II

Test Weight Amount Leaching Free Acid %Extraction %Pb ~io. Conc. Leach Time Equiv. in Conver- -S~lution (hrs.) Pregnant Pb Ag Cu Zn sion _ _ ~ (g/l) to PbS0 ~ A ~ g ~OO~)n~ -T ~ -~ 0 110A 500 g 1000 ml. 4 60 0.02 0.01 96.5 98.2 85.8 111A 500 g 1000 ml. 2 60 0.02 0.01 93.3 95.1 70.2 112A 500 g 1000 ml. 1 60 0.04 1.33 93.6 93.5 62.0 113A 750 g 1000 ml. 3 12 0.02 1.21 53.9 85.1 55.5 - l2 - L;
d~ ~ ' ~. j, ~76367 EXAMPLE II
.
Low grade lead concentrate containing by weight approximately 8% copper, 8% zinc, 12~ lead, 6% water and 130 ounces of silver per ton, is admixed with spent æinc electrolyte (210 grams/liter of sulfuric acid; 48 grams/liter zinc~ and 150 parts per million chloride ion in an autoclave. The con-centrate is leached for 2 hours at an oxygen pressure of 60 psig and a temperature of lOO~C.
As a result of this leach, the feed concen-trate is extracted of 98~ of its original copper and zinc concentrations. Ninety-two percent t92%) of the lead sulfide contained in the concentrate is converted to lead sulfate.
To the concentrate residue resulting from the above oxygen leach is added lime in an amount 10%
greater than required according to the following stoichiometric equation :
3Ca(OH)2 + 12S~ 2CaS5 ~ CaS203 + 3H20 Additionally, there is added sulfide ions (CaS5) by recycling the product stream produced in the above reaction. The concentrate is leached for 90 minutes at 95C resulting in the removal of 98% of the original elemental sulfur content and conversion of 82% of the lead sulfate to lead sulfide.
The resulting concentrate residue is then leached with ferric chloride and calcium chloride at a pH of 1.0 and a temperature of 800C for 90 minutes.
The foregoing resulted in the extraction of 99% of the lead content of the residue, 20% of residual copper, 40% of residual zinc, and 98% of the original silver content.

~76367 The leach liquor resulting from the above-described lime leach is acidified with sulfuric acid to a pH of 2.0 and reacted at 9GC for 1 hour to precipitate gypsum and sulfur Silver is recovered from the final leach liquor by cementation on very fine lead powder at 80~C for 30`minutes, the following principal reactions taking place: !

2FeCl3 + Pb ~ 2FeCl~ + PbCl2 2AgCl ~ Pb ,PbCl2 ~ 2Ag CuC12 + Pb~--~7PbC12 ~ Cu The substantially silver-free liquor is then treated to recover lead therefrom by cooling to below about 30C to crystallize lead chloride. An electrolyte solution is then prepared containing 50%
lead chloride, 25.5% lithium chloride, 21.7% potassium chloride and 2.8% calcium chloride. Electrolysis is conducted at 425C, a current density of 700 amperes/square foot, and an electrode spacing of 3 centimeters (resistivity 0.5 ohm-centimeters).
As will be apparent to those skilled in this art, the process of this invention may be applied to the recovery of metals from a variety of metallurgical products such as smelter dusts, metal drosses, middling concentrates from floatation processing, and other like sources of lead, zinc, copper-and zinc metals. Further, in the treatment of mineral con-centrates, the valueless and sulfur-rich mineral pyrite (FeS2) remains substantially unattacked in all states of leaching.

~76~7 SUPPLEMENTARY DISCLOSURE

A combination of laboratory and pilot plant tes~s were conducted as indicated below to explore and define the process parameters of the present invention.
1. A~TOCLAVE LEACH
a) LA~ORATORY
Effect of Time Head: 8.48% copper, 7.12% zinc, 26.85% iron (by weight) Charge: 200 grams of lead concentrate 0.75 liters of zinc plant spent electrolyte:
1~ 48 grams/liter of zinc 158 grams/liter of sulfuric acid approximately 250 parts per million chloride Conditions: 60 psig oxygen pressure 100 C temperature Results:

Test Time Percent Extraction Percent Lead Potential End Hrs. Copper Zinc Iron Sulfation Millivolts Acid Grams/Liter Al 0* 16.2 25.3 6.1 46.2 615 132 A2 0.25 47.1 59.416.4 57.4 655 94 A3 0.50 67.7 81.922.6 65.3 667 66 ~O A4 1.0 94.2 95.130.6 65.9 710 40 A5 2.0 99.3 98.835.4 89.5 716 30 A6 4.0 98.8 99.339.2 91.0 724 14 A7 8.0 99.2 99.451.3 93.1 722 33 A8 16.0 99.4 99.476.4 98.6 742 38 * 0 time denotes that the autoclave had reached operating temperature.
b) PILOT PLANT

Pilot Plant tests were performed continuously in a three compartment horizontal autoclave of 680 liter capaclty. Operation was carried out three shifts per day, five days per week. Samples were cut routinely at one hour intervals and composited daily for assay.

1C~7~367 Leaches w~ere performed using synthetic spent electrolyte prepared by diluting Zinc Plant neutral solution with water, Hydrochloric acid and 93% sulfuric acid were added to achieve the required concentra-tions of chloride and acid.
Effect of Pressure Head:

(% by weight) copper 10.07 7.64 zinc8.31 8.82 lead13.85 15~48 iron27.68 24.75 Conditions: 4 hours nominal retention time 101 C temperature 147 grams/liter of sulfuric acid Results:
PRESSURE PERCENT EXTRACTION PERCENT LEAD END PARTS PER MIL-psig COPPER ZINC IRONSULFATIONACID LION CHLORIDE
GR~MS/LITER START END

Series 1 95.5 96.0 37~2 85.1 34.0 325133 96 98 36.1 86.0 40.0 293103 96.7 98.4 38.7 78.8 28.5 25234 96.5 98.2 38.1 85.1 42.0 333117 Series 2 94.8 96.9 33.3 87.7 34.8 341119 96.9 98.3 34.1 88.8 40.0 369133 96.4 97.4 27.0 88.1 42.0 342105 96.2 97.1 28.0 84.1 44.0 351113 ~6367 Effect of Retention Time Head:

(% by weight) copper 9.71 8.57 7.94 zinc7.93 9.19 7.72 lead13.52 13.18 13.23 iron25.91 26.23 25.7 Conditions: 60 psig oxygen pressure 100 C temperature 147 grams/liter of sulfuric acid Results:
RETENTION PERCENT EXTRACTIONS PERCENT LEAD END PARTS PER MIL-TIME (HOURS) COPPER ZINC IRON SULFATION ACID LION CHLORIDE
GRAMS/LITER START END

Series 1 4 96.9 98.2 41.6 90.2 50.6 321 37 4.5 96.9 98.7 39.1 91.2 50.8 299 103 5.0 97.0 97.5 45.2 93.7 44.0 310 115 5.5 97.4 98.6 44.0 85.4 32.0 312 131

6.0 97.4 98.8 40.7 91.0 36.0 318 122 Series 2 4 96.2 96.7 35.3 78.0 34.0 511 444 96.7 98.0 40.7 86.4 40.0 334 133 6 97.1 97.8 39.6 87.6 46.8 344 190 Series 3 4 96.3 97.4 38.6 89.4 54.0 342 110 97.1 98.3 43.9 84.7 40.0 291 131 6 96.7 97.8 37.1 87.9 43.2 257 132 Effect of Temperature Head: 8.18% copper, 9.44% zinc, 13.83% lead, 23.53% iron (by weight) Conditions: 4 hours nominal retention time 60 psig oxygen pressure 147 grams/liter of sulfuric acid Results:
TEMPERATUREPERCENT EXTRACTIONPERCENT LEAD END PART PER MIL-C COPPER ZINC IRON SULFATION ACID LION CHLORIDE
GRAMS/LITER START END
, 94.6 96.5 38.8 79.7 46.0 312 137 100 95.2 95.4 38.8 87.3 37.6 371 142 105 96.6 96.9 44.3 87.7 39.0 367 151 110 97.2 97.1 41.0 80.4 38.8 312 119 Effect of Chloride Concentration Head: 5.79% copper, 7.39% zinc 14.88% lead, 26.8% iron (by weight) Conditions: 5 hours nominal retention time 70 psig oxygen pressure 110 C temperature 194 grams/liter of sulfuric acid Results:
PARTS PER MIL-PERCENT EXTRACTION PERCENT LEAD END
LION CHLORIDECOPPER ZINC IRON SULFATION ACID

5* 3 79.0 96.7 29.2 77.7 64.0 112 40 97.0 98.0 23.3 89.5 54~0 309 71 97.9 98.6 26.6 95.4 40.0 1000 421 . 95.6 97.5 31.9 84.4 50.0 * water and acid only 2. LIME LEACH
Effect of Recycled Solution Head: 40.95% sulfur, 17.36% lead, 15.68% lead as sulfate (by weight) Charge: 125 grams of autoclave residue 0.6 liters of water 15 grams of lime ~;Y6~6'7 Results:
No Recycled Solution TEMPERATURE C 95 _ BOILING
TIME (HOURS) 2 HOURS 4 HOURS 6 HOURS 2 HOURS 4 HOURS 6 HOURS
Sulfur Extraction2.3 3.2 4.5 18.1 3.1 15.3 Percent Lead Sulfate Reconversion .7 4.3 15.1 11 12.2 4.2 Percent With Recycled Solution TEMPERATURE RECYCLE VOLUME (Milliliters) Sulfur Extraction Percent 3.6 4.5 .8 21 22.9 Lead Sulfate Reconversion Percent 38 40 49 76 76 Sulfur Extraction Percent 19 19.6 18.9 23 24.1 Lead Sulfate Reconversion Percent 78 79 77 77 78 0 * NOTE: Sulfur extractions only were monitored during the leach. Twenty-four percent sulfur extraction from this sample represents an elemental sulfur extraction of approximately 98 percent.
Effect of Retention Ti Head: 39.02% sulfur, 11.8% elemental sulfur, 15.8%
lead, 14.8% lead as sulfate (by weight) Charge: 1000 grams of autoclave leach residue lime 0.4 liters of recycle (22.0 grams/liter sulfur) Condition: 95 C temperature ~7S31~

Results:
A. 144 Grams lime (120 percent stoichiometric) PERCENT
TIME RESIDUE ASSAY PERCENT PERCENT SULFUR EXTRACTION LEAD SULFATE
HOURS TOTAL SULFUR ELEMENTAL SULFUR TOTAL ELEMENTAL RECONVERSIONS

0 39.02 11.8 - - -.25 33.12 3.6 15.1 70.7 61.7 .50 33.06 .38 15.3 96.9 72.6 .75 32.12 .45 17.7 96.3 72.7 1.0 32.43 .10 16.9 99.2 74.0 2.0 32.67 .25 16.3 98.0 71.9 3.0 33.06 .14 15.3 '98.9 72.2 4.0 32.37 .14 17.0 98.9 72.2 5.0 31.80 .10 18.5 99.2 70.9 _ B. 160 Grams lime (133 percent stoichiometric) PERCENT
T~IE _ RESIDUE ASSAY PERCENT PERCENT SULFUR EXTRACTION LEAD SULFATE
llOURS TOTAL SULFUR ELEMENTAL SULFUR TOTAL ELEMENTAL RECONVERSIONS
0 39.02 11.8 - - -34.70 .11 11.1 - 71.6 33.0 .35 15.4 97.2 76.0 .75 32.65 .16 16.3 98.7 74.8 1.0 32.51 .28 16.7 97.2 73.0 2.0 32.36 .31 17.1 97.5 73.4 3.0 32.18 .28 17.5 97.7 73.2 4.0 32.18 .1 17.5 99.2 70.8 5.0 31.63 .1 18.9 99.2 64.7 3. LI~IE LEACH PURIFICATION
A. Acidification Head Solution: 11.96 Grams/Liter of calcium, 27.55 Grams/Liter of sulfur Charge: 0.5 Liters of solution 10.0 milliliters of 96 percent sulfuric acid Results:
PERCENT UNEXTRACTED MATERIAL
TE~lPERATURE TIME NON-SULFATE
TEST C HOURS CALCIUM TOTAL SULFUR SULFUR
Al 23 1 11.2 49.7 35.4 A2 23 4 9.2 56.9 40.2 Bl 70 1 7.2 28.0 18.4 - B2 70 4 6.9 27.8 19.8 ~L~7t~i3~;7 Cl 90 1 8.1 28.2 17.5 C2 9O 4 6.9 26.7 16.3 B. Pressure O_idation Head Solution: 12.9 Grams/Liter oE calcium, 26.75 Grams/Liter of sulfur Charge: 0.4 Liters of solution Conditions: 90 psig oxygen pressure 1 hour retention time Results:

T~IPERATURE FINAL PERCENT UNEXTRACTED MATERIAL
TEST C pH CALCIUM SULFUR

110 3.6 96.1 65.0 B 115 3.5 72.1 59.9 C 120 3.4 47-7 37.8 D 125 3.0 17.4 16.2 E 130 2.4 8.9 9.2 Head Solution: 11.6 Grams/Liter of calcium, 23.4 Grams/Liter of sulfur Charge: 0.4 Liters of solution Conditions: 135 C temperature 90 psig oxygen pressure Results:
TIME FINAL PERCENT UNEXTRACTED MATERIAL
TEST HOURS pH CALCIUM SULFUR
0.25 3.6 75.6 44.6 B 0.50 2.9 21.1 18.0 C 1.0 2.4 10.8 8.4 D 2.0 1.7 6.7 6.7 4. BRINE LEACH
Metal Extractions With Time and Temperature -Head: 0.21% copper, 0.17% zinc, 16.2% lead, (by weight), 66.77 ounces per ton of silver

7~367 Charge: 95 grams of lime leach residue 0.65 liters of leach solution (35 Grams/Liter of ferric chloride, 3 Grams/Liter of ferrous chloride, 42 Grams/Liter of zinc chloride, pH 1.5) Results:

PERCENT EYTRACTION
TESTTEMPERATURETIME FINAL
C HOURS pHCOPPER LEAD ZINC SILVER
Al 70 0.5 1.516.7 98.9 31.2 96.2 A2 70 1.0 1.517.6 98.9 31.2 95.9 A3 70 1.5 1.517.6 99.1 37.5 96.1 Bl 80 0.5 1.325.0 99.3 37-5 96.0 B2 80 1.0 1.322.2 99.1 35.0 95.2 ~3 80 1.5 1.422.2 99.1 29.3 96.0 Cl 90 0.5 1.233.3 99.3 37.5 96.1 C2 90 1.0 1.327.7 99.3 39.8 96.0 C3 90 1.5 1.235.3 99.3 41.8 96.2 Dl 70 1.5 1.021.1 99.3 31.2 95.8 D2 80 1.5 1.412.5 99.2 43.7 96.8 D3 90 1.5 - 47.4 99.3 43.7 97.0 * Head for B2, B3 and C2, C3 - 0.20% copper, 0.19% zinc, 16.6% lead (by weight), 61 14 ounces per ton silv~r Ferric Chloride Levels -Head: 0.46% copper, 0.25% zinc, 17.96% lead~0 (by weight), 84.09 ounces per ton of silver Charge: 95 grams of lime leach residue 0.65 liters of leach solution Conditions: 80 C temperature l hour retention time Results:
FERRIC LEACH ENDPERCENT EXTRACTION
TESTCHLORIDEPOTENTIAL
CONC. MILLIVOLTS COPPER LEAD ZINC SILVER
GRAM/LITER
A 20 203 2.1 93.9 19.8 7.6 B 30 252 2.1 98.1 23.1 26.0 C 40 470 13.6 98.9 30.5 96.2 D 50 652 19.6 99.3 39.0 96.8 ~76311~7 5. FERROUS CHLORIDE OXIDATION
Head: 40 Grams/Liter of ferrous chloride9 400 Grams/Liter of calcium chloride, pH 1.5 Apparatus: 50 millimeter ID column, packed with 8 millimeter O.D. Raschig rings Conditions: 60 C temperature, 0.18 liters per minute liquid flow Results:
COLUMN LIQUID CHLORINE FLOW SOLUTION PERCENT
HEIGHT RETENTION MILLILITER/ PERCENT POTENTI~L CHLORINE
10METERS TIME sec. MIN.CHLORINEMILLIVOLTS UTILIZATION
0.61 17 550* 100 726 74.4 0.61 17 550 94 726 74.4 0.61 17 550 68 721 69.2 0.61 17 550 42 714 61.3 1.22 46 550 100 743 91.8 1.22` 46 550 94 742 90.8 1.22 46 550 68 738 87.4 1.22 46 550 42 734 83.3 1.52 60.5 550 100 758 1.52 60.5 550 94 751 98.7 1.52 60.5 550 68 743 91.8 1.52 60.5 550 42 741 89.8 1.95 82.5 550 100 753 100 1.95 82.5 550 94 754 1.95 82.5 550 68 756 1.95 82.5 550 42 745 93.5 1.83 75 600** 100 764 98.3 1.83 75 600 70 757 94.2 201.83 75 600 44 7~i2 83.2 * 80~ oxidation at 100 percent chlorine utilization ** 87~ oxidation at 100 percent chlorine utilization 6. TREATMENT OF BRINE LEACH BLEED STREAM
Head Solution: 48.0 grams/liter of lead, 21.0 grams/liter of zinc, 16.1 grams/liter of iron Charge: 1.5 liters of solution, reagent grade hydrogen sulEide (160 milliliters per minute), lime as required to maintain pH
Conditions: 70 C temperature, pH 3~0 ~3'76367 Results:

PRODUCTS ADDITION GRA~IS IEAD ZINC IRON
Lead Precipitate 25.8 93.1 1.0 1.7 Zinc Precipitate 33.2 6.5 86.9 8.9 Recycle Solution 0.4 12.1 89.4 7. FUSED SALT ELECTROLYSIS
FIVE TESTS (AVERAGED) Electrolyte: 25.5% (by weight) lithium chloride 2.8% calcium chloride 21.7% potassium chloride 50.0% lead chloride Conditions: 425 C temperature 3 centimeter electrode spacing Results: electrolyte melting poin~: approximately 340C
electrolyte resistivity: 0.5 ohm-centimeters at 425C
decomposition voltage for lead chloride: 1.35 volts Current Efficiency: 99.1% @ 0.80 amps per square centimeter current .. :
density 99.2% @ 0.43 amps per square centimeter current density 99.5% @ Q.43 amps per square centimeter current denslty Electrolysis can, thus, be conducted at 425C, a current density of 700-1000 amperes/square foot, and an electrode spacing oE 3 centimeters (resis~ivity 0.5 ohm-centimeters).

~7636'7

8. EFFECT OF LIME LEACH ON SILVER EXTRACTIONS
Head Sample: Al - 6.21% copper (by weight), 9.2% lead, 6.55% zinc, 126.6 ounces per ton silver Head Sample: A2 - 6.75% copper (by weight) 9 9 . 96% lead, 7.10% zinc, 127.6 ounces per ton silver Charge: autoclave leach: 188 grams of concentrate 0.75 liters of spent electrolyte ~153 grams/liter oE sulfuric acid, 40 grams/liter of zinc) lime leach: 124.4 grams of concentrate 18 grams lime 60 milliliters of recycle solution 150 milliliters of water brine leach: 90 grams solids (Al), 100 grams (A2) 0.65 liters of leach solution (42 grams/liter of ferric chloride, 400 grams/liter of calcium chloride) Conditions: autoclave leach: 100 C, 60 psig, 2 hours lime leach: 95 C, 1.5 hours brine leach: 80 C, 1 hour Results:
PERCENT UNIT EXTRACTION
ELEMENTAL
TEST COPPER ZINC SULFUR LEADSILVER
Al - Lime Leach Autoclave Leach 97.1 96.7 Lime Leach 97.9 Brine Leach 37.5 69.0 99.298.1 A-2 - No Lime Leach Autoclave Leach 95.4 95.3 Brine Leach 16.1 50.6 97.789.1 ~LCI 7~i3~;~

As have well been seen from the foregoing description, the present invention also provides a process Eor treating sulfide concentrate containing lead, copper, zinc and silver to selectively recover said metals therefrom, comprising the steps of: a) contacting said concentrate with sulfuric acid in the presence of oxygen at elevated temperature and pressure to extract copper and zinc materials from said concentrate;
b) thereafter contacting said concentrate with a reagent selected from the group consisting of lime, sulfide containing solution and mixtures thereo~
to remove elemental sulfur therefrom; and c) thereafter by contacting said concentrate with a mixture of calcium chloride and ferric chloride to extract lead and silver materials therefrom. In one aspect of this lnvention, there is provided such a process as described above in which gangue mineral pyrite remains substantially inert at all stages of the process and in step b) the reagent is lime. In another aspect of this invention, there is provided such a process as described above in which gangue mineral pyrite remains substantially inert at all stages of the process and in the step b) refractory silver-containing minerals in the concentrate are activated.

Claims (15)

WHAT IS CLAIMED IS:

1. A process for treating sulfide concen-trates containing lead, copper, zinc and silver to selectively recover said metals therefrom, comprising the steps of:

a) contacting said concentrate with sulfuric acid in the presence of oxygen at elevated temperature and pressure to extract copper and zinc materials from said concentrate;

b) thereafter contacting said concen-trate with lime to remove elemental sulfur therefrom; and c) thereafter by contacting said con-centrate with a mixture of calcium chloride and ferric chloride to extract lead and silver materials therefrom.

2. The process of claim 1 wherein said extraction of copper and zinc is conducted at an oxygen pressure in the range of about 60 to about 80 psig a temperature of at least about 100°C.

3. The process of claim 1 wherein said contacting with lime is conducted at a temperature of about 95°C.

4. The process of claim 1 wherein said extraction of lead and silver is conducted at a pH of less than about 1.5.

5. The process of claim 3 wherein said contacting with lime is conducted in the presence of added sulfide ions.

6. The process of claim 1 further com-prising treating the extracted copper and zinc materials to recover metallic copper and zinc therefrom.

7. The process of claim 6 wherein the treatment of said extracted copper material to recover metallic copper comprises contacting said copper materials with finely-divided zinc metal to cement copper metal on said zinc.

8. The process of claim 6 wherein the treatment of said zinc material to recover metallic zinc therefrom comprises electrolytically reducing zinc contained in said materials to its elemental state.

9. The process of claim 1 further compris-ing treating the extracted lead and silver materials to recover metallic lead and silver.

10. The process of claim 9 wherein said treatment of extracted silver materials to recover metallic silver therefrom comprises contacting said silver materials with finely-divided metallic lead to cement silver metal on said lead.

11. The process of claim 9 wherein said treatment of extracted lead material to recover metallic lead therefrom comprises recovering lead chloride and thereafter electrolytically reducing lead chloride in a fused salt electrolyte to produce lead metal and chlorine gas for the regeneration of ferric chloride.

12. A process for treating sulfide con-centrates containing lead, copper, zinc and silver to selectively recover said metals therefrom, comprising the steps of:

a) contacting said concentrate with sulfuric acid in the presence of oxygen at a pressure of about 60 to 80 psi and a temperature of at least about 100°C;

b) dividing the concentrate treated in a) into a first residue fraction and a first supernatant fraction;

c) treating said first supernatant fraction to recover metallic copper and zinc therefrom;

d) contacting said first residue fraction with lime at a temperature of at least about 95°C;

e) dividing the material resulting from d) into a second residue fraction and a second supernatant fraction;

f) treating said second supernatant fraction to remove elemental sulfur therefrom;

g) contacting said second residue fraction with a mixture of ferric chloride and calcium chloride and removing a third supernatant fraction from the material result-ing from said contacting; and h) treating said third supernatant to recover metallic lead and silver therefrom.

"CLAIMS SUPPORTED BY THE SUPPLEMENTARY DISCLOSURE"

13. A process for treating sulfide concentrate containing lead, copper, zinc and silver to selectively recover said metals therefrom, comprising the steps of:
a) contacting said concentrate with sulfuric acid in the presence of oxygen at elevated temperature and pressure to extract copper and zinc materials from said concentrate;
b) thereafter contacting said concentrate with a reagent selected from the group consisting of lime, sulfide containing solution and mixtures thereof to remove elemental sulfur therefrom; and c) thereafter by contacting said concent-rate with a mixture of calcium chloride and ferric chloride to extract lead and silver materials therefrom.

14. The process of claim 1, wherein gangue mineral pyrite remains substantially inert at all stages of the process.

15. The process of claim 13, wherein gangue mineral pyrite remains substantially inert at all stages of the process and in the step b) refractory silver containing minerals in the concentrate are activated.