US3196001A

US3196001A - Recovery of metal values from complex ores

Info

Publication number: US3196001A
Application number: US222761A
Authority: US
Inventors: Orrin F Marvin
Original assignee: Individual
Current assignee: Individual
Priority date: 1962-09-05
Filing date: 1962-09-05
Publication date: 1965-07-20
Anticipated expiration: 1982-07-20

Description

United States Patent O 3,196,00l RECVERY GE METAL VALUES BRGM CMPLEX @RES @rrin F. Marvin, Cottonwood, Ariz. (1549 VJ. Madison, Phoenix, Ariz.) Filed Sept. 5, 1962, Ser. No. 222,761 7 Claims. (Cl. 75-21) This application is a continuation-impart of my prior application Serial No. 10,210 iiled February 23, 1960, now abandoned.

This invention relates to the recovery of metal values from complex ores. It relates particularly to an improvement in the process for the recovery of copper, zinc, lead, iron, silver and gold metal values from complex ores disclosed in my issued Patent No. 2,927,017, dated March 1, 1960.

The complex ores, here under discussion, are those that contain two or more metal values in chemical union or in such physical union as to prevent normal mechanical separation of the values. The complex ores include metal values such as copper-iron sulfide, zinc-iron polysuliide, zinc-copper polysullide, zinc-copper-iron trisulfide, lead vanadate, lead chromate, lead molybdate, lead tungstate, lead silicate, and lead compounds combined with other acidic radicals, and also possibly other complex suliides. The ores under discussion also include simple compounds such as zinc sulfide, copper sulfide, iron suliide, and lead suiiide uncombined chemically with other suliides, and include also gangue, dirt and other insolubles.

In the past, attempts have been made to extract copper, zinc and lead values from such complex ores by a selective concentration of one of the ores and a subsequent smelting to recover the particular value desired. Such a procedure has several major disadvantages. First, the ability to concentrate a complex ore by standard gravity or oil flotation methods does not meet with much success inasmuch as such processes are advantageous only in separating the simple suliides from each other (such as copper suliide from zinc suliide), but cannot separate each of the metal values in the double suliides, the triple suldes or other complex compounds above mentioned. The concentrate, when shipped to the smelter, thus contains, in addition to the particular metal sought (for example, zinc), substantial amounts of other values such as copper and lead, and since the smelter is designed to process a particular metal, the shipper does not get paid for these other values (with the exception of gold and silver). In fact, if the lead or copper content in a zinc concentrate is above a certain point, penalties for their presence would be exacted by the smelter.

The average assay of so-called zinc and lead selective concentrates for the Western States in recent years, as produced by standard oil otation methods, are given in Table l.

Table 1.-A verage assay of selective concentrates It Will be seen by studying the above table that substantial amounts of copper and lead values are present in the zinc concentrates, while substantial amounts of copper and zinc values are present in the lead concentrates. In

igi Faented July 20, 1965 "Ice addition, generally both zinc and lead concentrates have approximately 10% of iron therein. It can thus be seen that because of the Vother metal values invariably present, a process which involves the steps of rst selectively concentrating an ore for a particular value, and then smelting to obtain that particular values, is both inefiicient and uneconomical.

The process of selectively concentrating such complex ores is in itself a relatively ineiiicient mode of recovery in comparison with that of making a collective concentrate of the values. Thus, it is found that a 15-20% increase n the recovery of lead, zinc, copper, gold and silver values is possible by collectively concentrating the ore for the various values, instead of attempting to selectively concentrate the ore for a particular value.

There are still other disadvantages to the smelting operation itself, chiefiy those of the high cost of processing and the high investment per pound of metal recovered. For all of the above reasons, the winning of copper, zinc, lead, gold, silver, and iron values from such complex ores as above described, has, in general,4 been commercially unsuccessful. Complex ores treated by the orthodox method outlined above, that is, by smelting a selective concentrate of each of the metal values, results, on the average, in recoveries of approximately 40-50% copper and approximately 60-70% zinc.

Attempts at achieving some measure of separation by other means, such as by grinding, have also been unsuccessful since the metal values are usually crystallized in such minute particles that grinding to the degree of lineness required to actually achieve separation, is very costly and economically prohibitive.

Wet processes, that is, those involving leaching operations after a roasting operation, have also not been successful in recovering zinc, copper, lead values from such complex ores regardless of whether the ore is first selectively concentrated or collectively concentrated. By orthodox Wet processes, approximately 60-80% of the total amount of zinc and approximately Litl-70% of the total amount of copper present in such complex ores can be recovered.

It appears that the loss in recovery of zinc and copper in both smelting and Wet processing is due to the formation of certain highly insoluble zinc and copper compounds, principally the ferrites of these metal values, during the treatment of the ore. In my prior patent, above identified, I disclosed a process whereby substantially complete recovery of copper, lead, zinc, silver, gold and iron values can be obtained from complex ores in a manner that is simple in operation and inexpensive in comparison with both Wet and smelting processes of the prior art. The process described in the above-identified prior patent comprises the steps of lirst roasting a complex ore at a temperature suiciently low to prevent the formation of any substantial amounts of acid-insoluble compounds of the metal values capable of forming said acid-insoluble compounds to acid-soluble compounds. Following the irst low-temperature roasting step, the acid-soluble compounds are separated from the remaining part of the metal values left in the ore residue. The ore residue is then roasted at a temperature suiciently high to convert the remaining part of metal values such as copper and zinc to acid-soluble compounds.

The process described in my prior patent avoids the formation of highly insoluble compounds such as the ferrites of zinc, copper or lead. The combination of metal values such as zinc, copper or lead in ferritic compounds substantially reduces the recovery of these metal values from complex ores. The primary cause of formation of the ferritic compounds is due to the conversion in an oxidizing roast of ferrous compounds such as copper by a second leaching step.

encerrar pyrite (CuFeS2) and iron pyrite (FeSZ) into ferrites having a formula MFeZO.,z where M represents a particular metal such as zinc, copper or lead. In the process of my said prior patent, the temperature of the first roast is maintained at a relatively low value so that decomposition and oxidation of the ferrous compounds to ferrie compounds is attained Without decomposition and oxidation of the other sulfides, either complex or simple, of zinc, lead silver `and gold.V While copper pyrite is decomposed and converted primarily to cupric oxide and some cupric sulfate, no copper ferrites are formed under the conditions of the low-temperature roast. It can also .be seen that there is no possibility of zinc and lead ferrite formation because the compounds in which these values rare found are not decomposed and are not, therefore susceptible to ferrite formation.

While the first roasting step of my prior patented process so converted the several compounds in the ore so that the compounds causing formation of ferritic compounds were removed, the first leaching step of such process resulted 'in removal of only about half of the total iron in the ore, and the resulting leachate also con'- -tained copper sulfate which had to be recovered therefrom. I have now discovered that if the first low-temperature oxidizing roast is followed by a reducing roast at a somewhat higher temperature, substantially all of the iron may be leached as ferrous sulfate by means of a vdilute water solution of sulfuric acid, and that the resulting solution contains substantially no copper, zinc or lead. It may, therefore, be employed to recover iron values, as for example for production of agricultural grade ferrous sulfate, without preliminary treatment for separation of copper.

The first low-temperature oxidizing roast of my prior patented process results in the formation of iron oxide, at best not containing more than about fifty percent soluble oxides under operating conditions, arid since only the ferrous iron is freely soluble in dilute sulfuric acid, the ferrie iron remains behind after the first leaching step. By the use of 4the reducing step in accordance with the present invention, ferric iron is reduced to ferrous iron, and the already present ferrous iron is substantially unaffected. The reducing step also has the effect of partially reducing copper sulfate formed in the first low temperature oxidizing roast back to metallic copper and the zinc sulfate to zinc sulfide, so that they are not removed by the dilute sulfuric acid solution, While the reducing step may be continued until some elemental iron is found in the roasted ore to assure that reduction has gone to the point where all of the iron will be removed, in general it is preferred that no elemental iron be formed during the reducing step because its presence is not conducive to best practice of the invention from a metal recovery standpoint.

The ore residue remaining after the leaching step contains substantially all of the identified metal values of interest except y.the iron removed in the leaching step. This residue is then, subjected to a high-temperature roast in which the copper and zinc are converted to acid-soluble compounds without danger of ferrite formation.V The acid-soluble compounds of zinc and copper are separated The residue following the second-leaching step Vis then treated by any of the standard methods to recover the lead, silver, gold or` other acidinsoluble rnetal values contained therein.

The advantave of the reducing roast in the process of my present invention is that the necessity for treatment of the iirstleachate to remove copper, in particular, is eliminated and the process is simplified by reason of the fact that only .the leachate from the second leaching step need be treated to recover copper and zinc.

The process of my present invention will become clearly understood by referring to the following description made in conjunction with the accompanying drawing which is a diagrammatic flow sheet of the process according to my present invention.

ln general, .after the collective or bull: concentration of the metal values, the process comprises a first lowtemperature roasting step whereby some or all of the copper values in the bulk concentrate are rendered acidsoluble and undesirable acid-insoluble compounds of the metal values are prevented from forming. The roasted ore is then subjected to a reducing roast, and the ore from the reducing roast is leached to remove the soluble iron compounds. However, apart from iron, the other metal values of the complex ore remain in the leached residue.

The leached residue is then roasted at a substantially higher temperature than the first roast, inasmuch as the undesirable substances which would formV acid-insoluble compounds such as ferrites have been removed by the rst leaching step. The zinc and copper are converted into the acid-soluble form by the second roasting step and leached out by an acid solution.

The lead is then recovered from the remaining residue by suitable conversions and leaching operations, as will be described. After the removal of the lead, copper and zinc values, the remaining residue is treated to recover the gold and silver by any standar-d method, such as by cyaniding to dissolve the silver and gold therefrom. Any remaining iron values are then removed from the remaining residue by standard reduction of ferric oxide contained therein.

It will be understood that the process of my present invention can also be used for the recovery of cadmium, antimony, bismuth, arsenic, cobalt and nickel from complex ores in addition to the metal values set forth above. The steps for recovery of Cd, Sb, Bi, As, Co and Ni values, while not herein described, will be apparent to thoses skilled in the art from the description of the process of my invention which follows.

Referring now to the flow sheet, complex ore as mined is iirst ground and sent to a bull: concentrating zone lil along the line l2 where the ore is concentrated by standard oil flotation methods. That is to say, la concentrate of the complex zinc, copper, lead iron, and precious metal suliides, as well as all the simple sul-des of the metal, is separated from the gangue, rock and dirt, designated generally by the tailings 14. It is found that the bulk concentrate produced contains approximately 15-20% more of the metal values than are present in the conibined totals of selective concentrates for each of the values. The reason for the increased recovery of the zinc, lead, or copper-zinc values is that standard oil flotation methods for selectively concentrating each of these values separate only the simple suliides of each of the elements, and, of course, cannot separate or decompose the complex disuliides and trisulfides which are'lpresent'in substantial amounts. It is also possible to employ my process Without any ore concentration, i.e., processing the ore directly as mined, but the collective concentration is found much more feasible.

The bulk concentrates, upon analysis, have been found to usually contain the following compounds: Copper-zinc polysuliide Copper-zinc-iron polysuliide Copper-iron sulfide Zinc-iron polysuliide Copper sulfide Y Silver sulfide Zinc sulfide Lead sulfide Lead vanadate Lead chromate VLead molybdate Lead tungstate Lead silicate Lead plus other `acidic radicals l Ferrous suliides, iron disuliides or sulfides or iro Cadmium sulfide Antimony sulfide t Arsenic sulfide i:

eyery .small amounts por traces,

Bismuth sulfide Gold (elemental)* *Very small amounts o1' traces. The proportions of the metal values in the bulk concentrate, of course, vary a great deal. By way of illustration, however, an average assay of a bulk lead concentrate of an Arizona ore is given below.

Metal Percent Zinc (Zn) 9.95 Copper (Cu) 1.16 Lead (Pb) 28.8 Iron (Fe) 21.0 Silver (Ag) (OL/ton) 38.3 Gold (Au) (oz./ton) 1.0 Cadmium (Cd) 0.01 Antimony (Sb) 0.1 Silicon (Si) 2.0

The bulk concentrate is sent along a line 16 to a first low-temperature roasting zone 18, the temperature of which is maintained between certain specified, relatively low limits. The primary purpose of this low-temperature roast is to prevent formation of ferrites which would otherwise combine with the zinc, copper and lead values in the ore concentrate to form acid-insoluble ferritic compounds. The term roast, roasting and the like used herein and in the claims, are used to denote a heating in an oxidizing atmosphere.

Preferably, the temperature in roasting zone 18 is maintained above 340 C., but below 460 C. While the formation of ferrites is minimized in the temperature range from 340 C. to 400 C., no substantial ferrite formation occurs when the temperature of the iirst roast is in the range from 340 C. to 460 C. The upper portion of this latter range has the added advantage that the optimum temperature of the first roast for maximum solubilization of iron in acid appears to be around 454 C. When the temperature of the first roast is significantly increased above this temperature, the acid-soluble iron in the roasted concentrate decreases. For example, at a temperature of about 650 C. only about 10% acidsoluble iron is produced, whereas, at 454 C. from 50% to 80% of the iron present is soluble in 15% sulfuric acid solution. Similarly, when the temperature of the roast is decreased below about 454 C., the percentage of acid-soluble iron in the concentrate is decreased. For example, at about 370 C. approximately 25% of the iron present in the concentrate is acid-soluble. So long as the formation of ferrites is avoided, it is preferable in the practice of the process to take advantage of a iirst roasting temperature at which maximum solubilization of iron results.

At the low temperature of roasting zone 18, while no copper ferrites are formed, copper pyrite is converted to cupric oxide and, to some extent, cupric sulfate. In addition, the ferrous compounds are readily converted to acid-soluble iron compounds and to oxides at these temperatures.

Zinc sulfide, as well as the sultides of cadmium, bismuth, cobalt, nickel and silver remain substantially unchanged at the low roasting temperature of roasting zone 1S. Lead sulfide is partially converted to the sulfate, but the major portion of this sulfide remains unchanged. The compounds of lead with an acidic radical, such as lead vanadate, lead chromate, lead molybdate, lead tungstate, etc., are not converted to the oxide or sulfate form, to any appreciable extent, during the low-temperature roast and remain substantially unchanged.

The decomposition of the simple and complex sulides of iron and copper (principally copper pyrite and pyrite) during the roasting oeration in zone 1S produces sulfur dioxide, which is used as the basis for the making of sulfuric acid for the acid leaching operation to be described. 'I'he sulfur dioxide gas leaves the low-temperature roasting zone along a line 20.

The roasted concentration is sent to a reducing zone 22 along a line 24. To provide a reducing atmosphere, a reducing medium such as carbon monoxide (CO) is added to the reducing zone through a line 26. The compounds of iron in the roasted concentrate are reduced in accordance with the reactions previously set forth. Ferrie oxide (Fe203) is reduced to ferrosoferric oxide (Fe304) which in turn is reduced to ferrous oxide (FeO). The ferrous oxide is reduced to elemental iron. In general, ferrie oxide and ferrosoferric oxide are reduced to ferrous oxide before any appreciable amount of ferous oxide is reduced to elemental iron.

I have already pointed out that while the reducing roast may be continued until some elemental iron is formed to assure maximum reduction of ferrie iron to ferrous iron, the ideal condition is one in which all of the iron is reduced to ferrous iron and no metallic iron has been formed. Such a condition would, of course, permit removal of substantially one hundred percent of the iron during the first leaching step. The degree of completion of the reducing roast, that is, the extent to which all of the iron is reduced to a condition for leaching by dilute sulfuric acid, is generally controlled by the variables of time, temperature, pressure and concentration of reactants and products. Those skilled in the art are familiar with the manner in which these conditions may be controlled to yield a given desired result. In general, a reducing roast in the range of 450 C. to 650 C. in an atmosphere of hydrogen or -other common reducing material commonly used in metal roasting operations has been found satisfactory. The removal of water and carbon dioxide from the reducing zone through line 28 removes the products of the reduction reaction and promotes the rate thereof.

The concentrate from reducing zone 22 is delivered to a leaching zone 30 along line 32, and leached with sulfuric acid of from 1% to 50% concentration. The sul furic acid leach solution is delivered to zone 30 by line 34, and is preferably maintained at a slightly elevated temperature, for example, at 60 C. This sulfuric acid leach dissolves all of the ferrous iron in the ore afected by the previous low temperature oxidizing and higher temperature reducing roasts. In general, this usually comprises almost all of the iron in the ore. Substantially no copper or zinc are removed from the ore by these steps. The resulting leach solution containing principally ferrous sulfate leaves the leaching zone by line 36. Since the first acid leach solution contains principally iron as ferrous sulfate, it may be discarded without recovery of the iron if the cormnercial situation does not justify treatment to recover the iron. 1f, however, recovery of the iron values is advantageous, the solution is passed to an Iron removal Zone A, indicated by reference character 33, in whichl the ferrous iron is oxidized by passing an oxygen-containing gas, such as air, through the solution. The oxygencontaining gas enters Iron Removal Zone A through aline 40. Approximately half of the iron is precipitated as ferric sulfate [Fe2(SO4)3] which leaves Iron Removal Zone A along a line 42.

The remaining ferrie iron passes to an Iron Recovery Zone B, designated by the reference character 44, along a line 46, Where it is precipitated by the addition of zinc oxide as hydrated ferrie oxide (FegOB-HZO). Also, traces of any other base metals that may be present are thereby precipitated, with the exception of possible traces of magnesium, manganese and sodium. The zinc oxide employed in precipitating the ferric iron is obtained by later steps in the process, as will be described, either during the formation of the zinc metal or directly from the calcines formed in the second roasting step.

The final solution remaining comprises essentially zinc sulfate only, and passes from Iron Recovery Zone B along a line 48 to a Zinc Recovery Zone to be described.

The concentrate not dissolved by the iirst leaching solution is sent to a second high-temperature roasting zone 7 50 along a line 52 and is roasted at a temperature above 540 C., preferably within the range of 600`to 800 C. A minimum temperature of 540 C. is required since, While the minimum decomposition temperature of the lead sulfide is 450 C., the minimum decomposition temperature of the zinc sulfide is 540 C. In the second high-temperature roasting step, the zinc sulfides are converted to the oxides, with some being converted to sulfates; the copper is converted to copper oxide, with some beingy converted to sulfatos; the remainder of the complex sulfides are decomposed and oxidized; likewise, cadmium, antimony, bismuth, arsenic, cobalt, and nickel sulfides and the like are similarly converted. The lead sulfide present is convertedto both the oxide and the sulfate. Silver sulfide (AgS) is readily decomposed and reduced to the elemental form. lt should be noted that since all ferrite forming components have been removed by the first leach, the high-temperature roast can be conducted without fear of Vloss of any zinc, copper, lead, iron or other metal values such as Cd, Sb, Bi, As, Co and Ni as insoluble ferrites.

Sulfur dioxide is also produced in the second roasting zone and is passed from zone Sti through a line 54 to an acid plant for use in makingsulfuric acid. The sulfurie acid is then used in first leaching zone 30 or in a second leaching Zone now to be described.

n The concentrate from the second roasting zone 50 is sent along a line o to a second acid leaching zone 58 where a sulfuric acid solution of fairly high concentration, e.g., between l550%, is introduced through a line 60 to leach substantially all of the zinc and copper values in the concentrate. ATraces of ferriev iron values may go into the leach solution at this point. The leaching operation may be conducted in a single or multiple number of stages, as the particular ore concentrate may require.

inasmuch as the concentrate after the second roasting, depending on the conditions used for the roast, may contain essentially all of the zinc as Zinc oxide, an alternative procedure sometimes followed is to takethe requisite portion of the concentrate directly to Iron Recovery Zone B at di. for the removal of iron and metal impurities by substitution of the iron and base metals for the Zinc in zinc oxide. Whether this mode of purification of the first leach solution and recovery of the metal values therefrom is chosen, or whether zinc oxide produced from the oxidation of the zinc pigs during production of zinc in later stages of my process is used, is an economic question.

in following this alternate process Vand employing the calcines from the second roasting zone, a portion thereof is sent along dotted line 62 to Iron Recovery Zone B at 44. The zinc-iron exchange takes place in Iron Recovery Zone B and the resulting zinc sulfate solution leaves along line 48.

The second acid leach solution is separated from the solid concentrates and leaves zone 5S along a line 64, the second acid leach solution containing all of the zinc and copper values and all the cadmium, bismuth, cobalt and nickel (present in trace amounts). The copper, cadmium, bismuth, cobalt and nickel values are removed from the second leach solution by any appropriate means, such as by oxidation-reduction reaction with zinc dust entering a Copper Recovery Zone 65 along a line 68. The copper recovered is sent, along with the traces of other elements aforementioned, to refining operations through a line 70.

The remaining solution is comprised substantially of zinc sulfate from two sources, zinc sulfate produced by the second leaching step in zone 58 and the Zinc sulfate produced by oxidation of added zinc metal in the Copper Recovery Zone at66. The solution is sent along a line 72 directly to a Zinc Recovery Zone '74 or by-passed to Iron Recovery Zones A and B at 38 and 44, respectively, if the presence of iron and other impurities so requires. That is to say, the Zinc sulfate solution may contain traces of ferrous (and ferrie) iron which can be removed @.9 by oxidation of the ferrous to the ferrie formin Iron Recovery Zone A at 3S, and by neutralizing with zinc oxide in the iron Recovery Zone B at 44, to thereby precipitate all the iron as ferrie sulfate and hydrated ferrie oxide (and to precipitate other impurities, if present), as described previously. Thus, if the concentration of ferrous and -ferric iron or other impurities is sufficiently high to possibly interfere with subsequent zinc recovery (a maximum of about 0.007%), the zinc sulfate solution is sent to the aforementioned Iron Recovery Zones A and B along a dotted line '76.

T he zinc sulfate solution, leaving Iron Recovery Zone B along line d8 and/ or leaving Copper Recovery Zone 66, then enters the main zinc sulfate line "i8 to be sent to Zinc Recovery Zone 74. The zinc is recovered by any of a variety of methods, for example, by electro-deposition. `The spent electrolyte, containing some zinc sulfate and sulfuric acid, is then recycled along aline to second leaching zone 53 for use in the leaching operations therein, as previously described. V

The presence of magnesium, sodium calcium and potassium is usually so small as not to interfere with the electro-deposition. This is especially true if the first leach solution is not processed for the small amount of Zinc it may contain. lf, after some time, however, substantial buildup of these elements doesoccur (they can be removed prior to deposition of the zinc by any of a number of standard methods. The presence of manganese in the electrolyte is preferable inasmuch as it will act as a built-in7 depolarizino agent during zinc electrodeposition.

If one were to follow the prior art practices and conduct a single, high-temperature roast, the average recovery of copper is found to be in the neighborhood of t0-70%, whereas by following the two-stage roasting, reducing and leaching operation described above, recovery of copper is almost always 99+%. The high recoveries are thought to be due entirely to the prevention of the formation of copper ferrites which are acid-insoluble compounds and the chief source of loss in orthodox leaching and smelting operations.

Similarly, with respect to zinc, by following the othodox one-stage roasting operation at high temperatures, the average recovery of zinc, regardless of whether it is smelted from a selected concentrate or Whether it is leached, is in the neighborhood of oil-80%. The loss is due primarily to the formation of ferrites or other undesirable compounds formed during the roasting operation. The staging of the low-temperature roast and the high-temperature roast with the intermediate removal Vof ferrous iron, eliminates the undesirable formation of zinc ferrites and substantially increases the percentage of recovery of zinc so that it is substantially a complete recovery.

The Zinc recovered in Zinc Recovery Zone 74- is sent along a line 82 where it is melted into pigs for shipment. A portion of the zinc is ground into dust to be sent through a line'St, to Copper Recovery Zone 66 for the oxidationreduction reaction therein.

`In melting the zinc to pig, there will be some oxidation of the Zinc to zinc oxide (ZnO). The Zinc oxide thus produced is sent from line S2 to Iron Recovery Zone B at 44 through a line do for the precipitation of impurities and iron as hydrated iron oxide, as previously explained.

The leached concentrates containing all of the gold, silver, lead, and the majority of the iron values insoluble in the first leach leave the second leaching zone along a line Sg and enter a Lead Conversion Zone90. The lead values are substantially entirely present in the form of lead sulfate because of the prior sulfuric acid leaching steps in the first and second leaching zones. In order to leach the lead values from the solids and s-tillv regenerate the leaching agent, the lead is first converted 4to the carbonate form in Lead Conversion Zone and isthereafter leached out as lead acetate with acetic acid. The lead acetate is then converted to sulfate by means of sulfuric acid whereby the lead sulfate is recovered as a precipitate and the acetic acid is regenerated for use in leaching the lead sulfate from the solid concentrates.

To this end, a near-boiling sodium carbonate solution enter-s zone 90 along a line 92 at -a concentration of 10 to 40% and converts the lead sulfate in `the solids to insoluble lead carbonate. A sodium sulfate solution thus leaves zone 90 along a line 94 .according to the equation:

The lead-containing solids are then sent from zone 90 yalong a line 96 to a third leaching zone 9S wherein the solids are leached with an acetic acid solution entering zone 98 along a line 100. The lead carbonate is thus decomposed and soluble lead acetate solution is formed according to the equation:

The lead then leaves the third leaching zone along a line 102 lto a Lead Recovery Zone 104. Snlfuric acid is then added Ithrough a line 106 to the solution of lead acetate in zone 104 whereupon the lead precipitates as lead sulfate, leaving the recovery Zone along a line :108.

The -ace-tic acid solution is regenerated according to Athe equation:

leaving zone 104 along a line 110 to be recirculated to third leaching zone 98. Fresh acetic acid solution is added to the recirculating acetic acid line D-110 along a line 112.

It should be noted that lead may be recovered from the system in other forms lthan the sulfate, inasmuch as the addition of any mineral acid which is more highly ionized than lead acetate will precipitate the lead as the lead combined with acidic radical of the miner-al acid employed. It :should be noted valso that lead cau Ibe recovered as a sponge metal from the Lead Recovery Zone by electrolysis, .if it is more desirable to sell the lead in this form.

The solids, Aafter undergoing the acetic acid leaching treatment in the third leaching zone at 98 are sent along a line 114 to a fourth leaching zone 116 for the purpose of leaching out the silver and gold Values from the remaining insolubles in the solids (such as ferrie oxide, and silica) by-sEudard rneans. For example, sodium cyanide enters leaching zone 116 along aline 118 leaching out the silver and gold values as silver cyanide and gold cyanide, respectively. The soluble cyanide values leave the fourth leaching zone along aline 120 to be sent to a Gold and Silver Recovery Zone `122. It is to be emphasized that the gold and silver entering -the fourth leaching zone are pure gold and silver metal, respectively, which are readily converted to the cyanide complex form. Thus, a short leaching time only is required. The gold and silver values are selectively converted to 1.their reduced elemental form by reduction with zinc rnetal or by other standard methods.

The iron remaining in the concentrate, which is in the form of ferrie oxide, as Well yas other insolubles such as silica are sent along a line 124 to appropriate reduction processes where Ithe iron is recovered in an Iron Recovery Zone C 126. For example, the iron oxide may be reduced in an hydrogen atmosphere or by carbon monoxide gas in a blast furnace. The silica-tes and other remaining insolubles are removed from viron Recovery Zone C along a line 12S.

An example which is typical of the process above-described is set forth below.

10 EXAMPLE A ground sulde ore was colle-ctively concentrated with the raw concentrate having the yfollowing assay:

Metal value: Percent by weight Zinc 39.0 Copper 1.0 Lead 12.0 Iron 12.0 Cadmium 0.1

From the above assay, it will be seen that the concentrate was high in zinc and iron.

The concentrate was roasted in a low-temperature roasting zone in an oxidizing atmosphere containing about 8% sulfur dioxide. The roasting time was about two and one-:half (2l/2) hours and the temperature of the roast was approximately 400 C.

The roasted concentrate was then subjected to a reduc- .ing roast where hydrogen gas provided the reducing rnedium. The residence Atime in the reducing roast was about 0.25 hours and the tempera-ture Aof the reducing roast was about 538 C.

rthe concentra-te from the reducing zone was then leached with 4a 15 sulfuric acid solution at a temperature of abou-t 77 C. for one hour. The leach solution was separated from the ore residue and ,the leach solution was Itreated to recover the metal values contained therein. The recovery of metal values from the rst acid leach solution is summarized in the table below:

Percent by weight extracted Mea mercenaire Zinc 1.0 Copper 0.5 Lead 0.0 Iron 95.04- Cadmiuin 0.0

in second leach of total Metal VallJESI in Ore residue Zinc 990+ Copper 99.0-1- Lead 0.0

Iron 50.0 `Cadmium 99.04-

The ore residue was then 'treated for recovery of lead with a 15% solution of sodium carbonate at 100" C. and a 10% acetic acid solution, as previously described in conjunction with the tlow sheet. The Aresidue from the lead recovery zone was not further treated to recover silver, gold or .other residual metal values.

The tot-al recovery of metal values from the ore concentrate is summarized in the table below:

Percent total recovery based on Weight in raw .f Metal values. ore concentrate Zinc 98.0-1- Copper 95.0%- Lead 98.0 Iron 97.5 Cadmium 98.0-1-

It will be seen from the above example that the recovery of metal values is substantially complete based upon the metal values initially in the raw concentrate. Furtherclaims:

more, it will be seen that only iron values are found in the first acid leach solution and that substantially all recovery of copper and zinc values is from the second acid leach solution. Where, therefore, recovery of iron is not sought, the process may be economically practised by treating only the second acid leach solution for recovery of metal values. Y

As the specification makes clear, sulfide ores of the character identified as a'rule are mined for their zinc, copper and/ or lead values and normally iron, while usually present in substantial proportions, is not considered of prime importance in commonly used recovery procedures. Gold and silver, when present in economically significant proportions, are also usually recovered. Such elements as vanadium, chromium, molybdenum, tungsten, cadmium, arsenic, bismuth and selenium are frequently present in relatively small proportions as impurities, but frequently not in sufficient quantity to warrant recovery. The process of the present invention, however, does permit recovery of some ofrthese materials profitably even when present in i relatively small amounts. y

Arsenic, antimony and selenium, for example, when present in the ore will normally be leached and recovered as sulfates during the first leaching step. Removal of antimony at this stage may be important because among the acid insoluble compounds which may be produced are lead antimoniate, wh-ich could be formed by the final oxidizing roast if antimony were not previously removed.

The present invention -is particularly useful in the treatment of complex sulfide ores of the type identified not only from its ability to produce separate concentrates of such values as copper, zinc, lead and the like, but in the removal of substantially all of the iron and other elements such as antimony capable of forming insoluble compounds with copper, zinc, lead and the like, so that such metal values may later be recovered either 'by the process of the present'invention or optionally by known processes of thev prior art.

While a preferred embodiment of this invent-ion has been described, it will be understood by those skilled in the art that changes and modifications may be made that lie within the scope of the invention as defined by the I cla-im:

l. A process for recovery of metal values from a cornplex sulphide ore of the character identified and including copper, lead, zinc and iron, which comprises:

(A) roasting said complex ore at a temperature sufficiently low to prevent any substantial formation of ferrites but sufiiciently high to convert at least part of i2 the iron values to ferrie and ferrous oxides, and leaving other metal sulphides unoxidized,

(B) heating said complex ore in a reducing atmosphere to convert the said ferrie oxide to ferrous oxide (F60),

(C) leaching said complex ore to separate therefrom the ferrous oxide.

2. The process claimed in claim 1 wherein the leached residue is roasted at a temperature sufficiently high to convert at least a part of the remaining metal values to acid soluble compounds.

3. A process for recovery of metal values from a cornplex sulphide ore of the character identified and including copper, lead, zinc and iron, which comprises:

(A) roasting said complex ore at a temperature below 400 C. to convert at least a part of the iron values to ferrie and ferrous oxides, and leaving other metal sulphides unoxidized,

(B) heating said complex ore in a reducing atmosphere above 550 C. to convert said ferrie oxide to ferrous oxide.

4. The process claimed in claim 3, wherein the said ferrous oxide is acid leached to separate the iron from the other constituents.

5. The process claimed in claim 4, wherein the acid leached residue is roasted at a temperature sufficiently high to convert the remaining metal values -to loxide compounds.

6. The process claimed in claim 5, wherein the oxide compounds are acid leached to recover the copper and zinc and leaving an insoluble residue containing all of the lead.

7. The process claimed in'clairn 6, wherein the concentrate remaining after the sec-ond acid leach is treated with sodium carbonate solution to thereby convert the lead in said concentrate to lead carbonate, the resulting lead carbonate is then treated with acetic Vacid to leach the lead values from the concentrate as lead acetate, and re-precipitated by treatment of said lead acetate with a mineral acid whereby said acetic acid is regenerated for reuse in the leaching of lead carbonates from said concentrate.

References Cited by the Examiner UNITED STATES PATENTS 153,573 7/74 Kidwell 75-21 1,637,838 S/27 Simonds 75-21 1,833,686 ll/31 Meyer 75-1 2,123,240 7/38 Hammarberg 75-21 2,927,017 3/60 Marvin 75-101 3,053,651 9/62 McGauley 75-1 BENIAMIN HENKIN, Primary Examiner.

Claims

1. A PROCESS FOR RECOVERY OF METAL VALUES FROM A COMPLEX SULPHIDE ORE OF THE CHARACTER IDENTIFIED AND INCLUDING COPPER, LEAD, ZINX AND IRON, WHICH COMPRISES: (A) ROASTING SAID COMPLEX ORE AT A TEMPERATURE SUFFICIENTLY LOW TO PREVENT ANY SUBSTANTIAL FORMATION OF FERRITES BUT SUFFICIENTLY HIGH TO CONVERT AT LEAST PART OF THE IRON VALUES TO FERRIC AND FERROUS OXIDE, AND LEAVING OTHER METAL SULPHIDES UNOXIDIZED, (B) HEATING SAID COMPLEX ORE IN A REDUCING ATMOSPHERE TO CONVERT THE SAID FERRIC OXIDE TO FERROUS OXIDE (FEO), (C) LEACHING SAID COMPLEX ORE TO SEPARATE THEREFROM THE FERROUS OXIDE.