WO1990013679A1

WO1990013679A1 - A novel process for the treatment of zinc sulphide containing ores and/or concentrates

Info

Publication number: WO1990013679A1
Application number: PCT/CA1990/000130
Authority: WO
Inventors: Murry C. Robinson; Donald R. Spink; Kim D. Nguyen
Original assignee: Materials Concepts Research Ltd; University of Waterloo
Current assignee: Materials Concepts Research Ltd; University of Waterloo
Priority date: 1989-05-03
Filing date: 1990-04-25
Publication date: 1990-11-15
Anticipated expiration: 1991-11-03
Also published as: CA2055427A1; AU5546690A; GB8910083D0

Abstract

A process is described for the partial desulphurization roasting of a variety of zinc bearing sulphide ores or concentrates by adjusting the roaster temperature and residence time in the roaster thus providing an oxygen deficient atmosphere so that the required amount of sulphide retention is maintained. The obtained calcine or calcines are subsequently subject to various chemical and/or physical separation process steps to separate the unreacted sulphides which are, according to certain embodiments of the process, then dead roasted and treated for zinc recovery. In another embodiment, the separated unreacted sulphides are again partially desulphurized in an oxygen deficient atmosphere for subsequent treatment to recover zinc, lead and precious metals. The chemical separation process steps include aqueous sulphur dioxide treatment, dilute sulphuric acid solution treatment and treatment with dilute sulphuric acid containing sulphur dioxide in solution, all conducted in the temperature range of 50 to 75°C. Physical separation process steps include one or more of the flotation and magnetic separation techniques although other physical separation steps may also apply in some instances.

Description

Title: A Novel Process For The Treatment Of Zinc Sulphide Containing Ores And/Or Concentrates

FIELD OF THE INVENTION This invention deals with recovering zinc from zinc and iron-bearing sulphides which are either in the form of conventional zinc sulphide concentrates, low-grade zinc sulphide concentrates, or in the form of bulk zinc sulphide concentrates, the latter which consist of zinc, iron, and non-ferrous metal sulphide minerals including lead sulphides and precious metals in complex form. Additionally, the accompanying valuable non-ferrous metals are recovered, which include lead, copper, cadmium and silver. Throughout the world, conventional processing of ores containing zinc sulphide as the most valuable constituent but including substantial amounts of iron as well as lesser amounts of other base metal sulphides involves the concentration of the zinc sulphide component by flotation techniques thus resulting in the formation of zinc concentrates containing some 45% to 65% zinc as zinc sulphide with the chief impurity being iron in the form of iron sulphides. Such conventional zinc concentrates are generally treated in a dead roasting step wherein the resulting calcine contains zinc oxide as a major constituent; however, a substantial portion of the zinc will generally be in the form of zinc ferrites.

This calcine is then leached in what can be referred to as a neutral sulphuric acid leach wherein the zinc oxide content is readily dissolved but without dissolving the zinc present as zinc ferrite. The separated zinc solution is then purified and electrolyzed to produce zinc metal with the major portion of the spent electrolyte recycled to the neutral leaching steps. The undissolved leach residue containing substantial amounts of zinc as zinc ferrite was in earlier days disposed of, usually to a stockpile .

In more recent years, the leach residues have been directed to a hot concentrated sulphuric acid solution to dissolve both the zinc and iron present in the zinc ferrite. Processes have been developed and are now used in practice to treat the co-dissolved zinc and iron sulphates in order to precipitate the iron rather than the zinc. Two such processes are the jarosite and goethite processes, both of which provide means to recover the zinc as a zinc sulphate solution which then can be treated to produce metallic zinc. Unfortunately the iron-containing residues are quite voluminous and can contain toxic substances which result in environmental as well as economical problems. In certain circumstances, zinc sulphide concentrates may be too poor in grade to warrant conventional roasting followed by conventional zinc refining steps. In these instances, the zinc concentration might be too low, say in the range of 30%-45% zinc or the iron concentration might be too high, say in the range of 10%-20% iron and in many instances both of these factors may apply. Some examples of the possibility of producing such types of zinc concentrates follows. In one case, an orebody may have been mined and milled to the extent that only a lower grade uneconomical zinc concentrate could be produced. In another instance, an orebody might be such that only a lower grade zinc concentrate could be produced in substantial quantities. In a third situation, the flotation processing might be such that a middlings product could be produced that would be of an unsatisfactory grade so that according to present techniques, it would have to be discarded, thus resulting in zinc losses.

On the other hand, our novel technology can reduce the formation of zinc ferrite in the produced calcine to such a small fraction that the need for such expensive, troublesome and environmentally undesirable treatment steps as in conventional processing can be eliminated; other beneficial results are also obtained in the overall zinc refining processing.

The novel technology also has a great deal of flexibility in that it can be used beneficially for treating a wide variety of zinc concentrates.

For example, optionally, our novel technology can be applied to upgrade low-grade zinc concentrates so that a zinc concentrate could be produced of equal or better grade to that now being economically processed by conventional means. Alternatively, a lower grade zinc sulphide flotation concentrate might be produced at a much higher yield; pretreatment of such a concentrate using our novel technology would readily upgrade the concentrate to conventional levels with a consequently greater production of zinc.

In still other circumstances, zinc sulphide containing ore bodies can be so complex in physical nature that these are either incapable of being treated by conventional means to produce a saleable product or are of such a nature that a zinc sulphide containing orebody is treated in a manner that results in the production of a bulk concentrate that is low in zinc, say in the 30%-40% range, high in iron, say in the 10%-20% range, high in lead, say in the 10%-20% range and sulphur in the range of 30%-36%. Many of such bulk concentrates are shipped to Imperial Smelting Furnaces resulting in low revenues for recovery of the contained zinc, lead, other base metals and precious metal values contained therein. Depending upon the economical climate at any given time, such bulk concentrates might be rejected as being impractical to treat.

Our novel process technology has the potential to circumvent the difficulties described in the foregoing paragraphs and will be described later in this disclosure.

Our process involves a two stage roast wherein the first calcine can contain part or nearly all of the iron in an easily acid soluble iron oxide form leaving most of the zinc and other base metals present as sulphides. Such a first stage roast also produces a higher purity sulphur dioxide-containing gas than that conventionally produced which may be useable as such or more easily treated to produce sulphuric acid. The iron oxide component in the first calcine can readily be dissolved along with any base metals co-oxidized in a warm aqueous sulphur dioxide solution or a warm dilute sulphuric acid solution or a combination of a warm dilute sulphuric acid solution and aqueous sulphur dioxide, thus leaving a leach residue that is higher in zinc content and much lower in iron content. The leaching temperature would normally be in the range of 50°C and 75°C and preferentially between 60°C and 70°C. The aqueous sulphur dioxide solution would be preferentially at or close to saturation and, if sulphuric acid solution were used it would be in the range of 2-5 wt% H₂S0₄ and preferably 3-4 wt% H₂S0₄. Where a sulphuric acid and aqueous S0₂ solution are combined, the sulphuric acid solution would still be between 2-5 wt% H₂S0₄ and the aqueous S0₂ dissolved therein would range from a minor addition to close to saturation. Spent electrolyte may serve as a substitute for sulphuric acid in whole or in part in some instances. In some instances the leaching pulp density would be betwee_.n 60 and 120 gpl and the leaching period would be three hours or less. However, staged leaching could result in broader pulp densities and longer leaching times.

In some instances, particularly when the zinc sulphide is predominantly sphaleritic rather than marmatitic in nature, optionally physical separation techniques such as flotation and/or magnetic methods might be employed either alone or in combination with the chemical dissolution methods already described above to separate the oxides from the unreacted sulphides and thus provide a useful separation technique. These physical separation techniques might be applied to the partially desulphurized calcines and/or to upgrade the leach residue produced.

In any event, the leach residue or physical residue remaining after the iron oxide fraction has been separated, provides the feed to the second stage roasting step.

In one embodiment using conventional zinc sulphide concentrates or low grade zinc sulphide concentrates, the leach residue or physical residue containing the unreacted sulphides present in the first calcine would be subjected to a second stage conventional dead roast to produce a calcine which would contain less zinc ferrite than that which would be conventionally produced. The first stage roast would be conducted in the presence of an oxygen bearing gas wherein a degree of sulphide sulphur retention in the calcine is maintained by controlling the oxygen flow rate and/or the residence time of the feed material in the roaster thus resulting in an oxygen deficient atmosphere. The percentage of sulphur removed will be a function of the iron content in the concentrate to be treated and the degree of iron removal desired using the partial desulphurization roast and leaching and/or physical separation steps. The partially desulphurized roast will normally contain between 15% and 27% sulphur and preferentially between 20% and 25% sulphur by controlling the retention time of the ore or concentrate in the oxygen deficient atmosphere. The first stage roast would be conducted in the temperature range of 650°C to below the sintering temperature and preferably between 700°C and 1050°C and more preferably between 850°C and 1000°C. Under these specific conditions, zinc ferrite formation is reduced in the dead roasting step to the degree desirable for any given application. In this context, the words "partially desulphurized" refers to partial oxidation of the contained metal sulphides.

As an alternative embodiment, in the case of bulk concentrate treatment, two oxygen deficient roasts might be conducted prior to the final sulphide conversion operation. In this instance, the first stage roast would be conducted under the conditions described in the previous paragraphs to produce a calcine wherein most of the iron sulphides are converted to an easily soluble iron oxide form such that all the soluble oxides can be leached using one of the leaching and/or physical separation techniques hereinbefore described leaving the bulk of the zinc sulphides, lead sulphides, other base metals sulphides and contained precious metals in the separated sulphide containing residue.

This separated sulphide-containing residue is then conceived to be roasted in a second stage roaster also using an oxygen deficient atmosphere to the extent wherein more than 80% and preferably more than 90% of the contained zinc sulphide is oxidized to form zinc oxide leaving the lead sulphides essentially unreacted. This second calcine would then be leached using a neutral leach or one of the leaching techniques previously described which after liquid-solid separation would leave a leach residue that is primarily lead sulphide but rich in precious metals which could be fed directly to a conventional lead smelter. Flotation techniques might alternatively be used to separate a relatively high grade zinc oxide from the lead sulphide fraction. In this instance the preferred roasting temperature range would be 650°C to 850°C but preferably between 675°C to 750°C for each of the two stage roasting operations. Prior to feeding the lead sulphide concentrate to the lead smelter, chemical or physical separation techniques might be employed to separate the precious metals from the lead sulphide. These techniques will be described later in this disclosure.

By the use of the techniques described above, it should be possible to use mining and milling techniques on complex metal sulphide deposits to recover a much higher percentage of metal sulphides such as zinc, copper, lead. cadmium, silver and gold from the mineral deposits than is presently practiced and could, in addition, make the treatment of some of the now dormant orebodies of such a nature into profitable operations. This is because lower grade bulk concentrates could be upgraded.

Some of the advantages discussed above can be achieved by using the process for the recovery of zinc bearing ores or concentrates comprising the steps of:

(a) roasting the ore or concentrate in the presence of an oxygen-containing gas wherein sulphur retention is maintained at a level of 15% or higher by controlling the residence time of the ore or concentrate in an oxygen deficient atmosphere in a roaster using a temperature in the range of 650^CC to 1050°C in order to preferentially oxidize the iron rather than the zinc and other base metal constituents and thus reduce or eliminate zinc ferrite formation

(b) leaching the partially desulphurized roasted ore or concentrate, using a lixiviant selected from the group consisting of aqueous sulphur dioxide solution, a 2-5 wt% sulphuric acid

.solution, and a 2-5 wt% sulphuric acid solution containing solubilized sulphur dioxide, at a temperature between 50°C and 75°C in a pH range of 1.0 to 2.5 for one to three hours in order to preferentially dissolve the iron oxide thus formed as well as base metals that have co- oxidized to obtain a leachate and a leach residue

(c) subjecting the leach slurry to liquid-solid separation

(d) treating the leachate to separate the dissolved iron and other dissolved oxides of base metals from the remaining sulphides, thus leaving a leach residue which is richer in zinc sulphide content and depleted in iron content which

(i) is dead roasted as in conventional practice with the resultant reduction or effective elimination of zinc ferrite formation or

(ii) if the original concentrate is rich in lead, such as in certain bulk concentrates, subjecting said leach residue to a second stage partial desulphurization roast using a controlled retention time in the temperature range of 650°C to 850°C such that the retained sulphur content is 8% or higher depending on the percent of lead present, to produce a second stage calcine which has a very high zinc oxide content and substantially all of the lead in the sulphide form, subjecting said second calcine to physical separation techniques or to a neutral leach or to an aqueous sulphur dioxide leach under similar conditions as earlier described, in order to separate the zinc oxide fraction from the lead sulphide fraction, the sulphide-containing residue is then fed directly into a lead smelter or is further treated to separate the lead sulphide from the precious metals. According to the above description, the process may be applied to low grade zinc sulphide concentrates or bulk zinc sulphide concentrates or to conventional zinc sulphide concentrates. In the case of conventional zinc concentrates these would contain in the range of 45 to 65% zinc, 3 to 15% iron and lesser amounts of copper, cadmium and lead, all predominantly in their sulphide form with a variety of other minor impurities present. In the case of low grade zinc concentrates, these would contain in the range of 30 to 45% zinc, 10 to 20% iron and smaller amounts of copper, cadmium and lead all predominantly in their sulphide form with a variety of other minor impurities present. In the case of bulk zinc concentrates, these would contain in the range of 25 to 40% zinc, 10 to 25% iron and 10 to 25% lead and smaller amounts of copper and cadmium all predominantly in their sulphide form with a variety of other minor impurities present.

The following paragraphs describe detailed techniques for treating the leachate, referred to in item (d) above, obtained between the first and second stage of roasting. Also to be described are leaching steps that could be employed in the case where leaching is desirable after a second stage oxygen deficient roast, such as in the case of one option for the treatment of zinc sulphide bulk concentrates. Also a method of leaching a lead sulphide concentrate product for selectively separating precious metals from the predominant lead sulphide species will be described.

In the first instance, we will describe one option for treating the calcine or, if you will, a partially desulphurized concentrate, after a conventional zinc concentrate has been subjected to the first stage of a two stage roast. In this instance, we will concern ourselves with an ore or concentrate wherein the zinc sulphide is primarily present in the marmatitic form. In instances where the zinc sulphide is present primarily in the sphaleritic form, it is to be expected that the conversion of iron sulphides to readily soluble iron oxides will be accomplished with less conversion of zinc sulphide to zinc oxide. In order to get the iron level of a zinc sulphide concentrate down from 8% to 11% iron to 0% to 3% iron prior to dead roasting, a quantity of zinc oxide will also be formed during the partial desulphurization roast step as previously described, resulting in zinc dissolution into the leachate of about 5% to 25% of the total zinc dependent upon the type of zinc concentrate treated.

An embodiment of the invention is shown in flowsheet form by combining Figure I with Figure IVA for treating a conventional zinc concentrate. In one instance a conventional zinc concentrate containing 49.0 wt% Zn, 9.10 wt% Fe, 0.70 wt% Cu, 0.24 wt% Cd and 32.4 wt% s was given a partial desulphurization roast to the extent that a partially desulphurized concentrate was produced analyzing 53.4 wt% Zn, 9.87 wt% Fe, 1.06 wt% Cu, 0.27 wt% Cd, and 25.70 wt% S. (See Example No. 1) According to this particular example an aqueous

S0₂ leach was employed at a temperature of 65 ± 5°C and a pH of 1.8 to 2.1 for two hours at an initial pulp density of about 80 gl^"1.

After liquid-solid separation, the resultant leach residue was reported to analyze 59.3% zinc, 1.16% iron, 1.30% copper, 0.31% cadmium and 30.8% sulphur and would provide an excellent feed material to a dead roast step due to its low iron content.

The leachate analysis showed that 90.8% of the iron, 14.2% of the zinc, 2.52% of the copper and 5.47% of the cadmium had been dissolved from the partially desulphurized roasted product.

In this example, the treatment of the leachate involved thermal decomposition to drive off S0₂-containing gas for recycle thus precipitating a solid consisting chiefly of iron sulphite and zinc sulphite.

The solid mixture was treated by an ammonia leach whereby most of the zinc dissolved and all of the iron was left as a residue. The iron residue was separated by liquid-solid separation. In one instance wherein the starting material was a much lower grade of concentrate analyzing 34.5% Zn, 15.7% Fe, 1.40% Cu, 0.23% Cd, 32.6% S, - li ¬ the iron residue produced was reported to analyze 61.1% Fe, 4.31% Zn, 0.03% Cu, 0.07% Cd and 1.30% S.

The leachate was steam stripped to remove ammonia for recycle thus precipitating basic zinc sulphite which after liquid-solid separation would be treated with spent electrolyte to produce zinc sulphate solution for feeding to a zinc refinery and S0₂ gas for recycle to the aqueous S0₂ leaching step. In the experiment conducted, starting with a much lower grade of zinc concentrate, the recovery of zinc from the S0₂ leachate was reported as 89.4%. After calcination of the basic zinc sulphite in this experiment, a zinc oxide product was produced which was reported to contain 77.6% Zn, 0.005% Fe, 0.03% Cu, 0.34% Cd and 0.61% S. The impurity level of the produced zinc oxide was remarkable considering that no purification steps such as zinc dust cementation were carried out.

In this embodiment the basic zinc sulphite would be dehydrated if necessary, and then treated with spent electrolyte liquor to dissolve it and feed the resultant zinc sulphate product directly to a conventional electrolytic zinc refinery. The reaction between zinc sulphite and H₂S0₄ will give off pure S0₂ for recycle when and as needed.

The spent liquor from the thermal decomposition step where basic zinc sulphite is produced would be treated with lime or calcium hydroxide in order to free and recycle its ammonia content in liquid form and to recover its contained zinc in the precipitate thus formed. After liquid-solid separation, a sulphuric acid treatment step on the precipitate would be required to dissolve the base metal compounds for recovery from the insoluble calcium compounds. However, the circuitry would be quite small and therefore quite inexpensive.

Other embodiments exist for treating the leachate from the aqueous sulphur dioxide solution leaching step.

These include solvent extraction of the leachate to separate the zinc fraction from the iron fraction by methods similar to those disclosed by Clitheroe. (U.S. Patent 4,053,552) with modifications (See Figure IV-B) Another method is to use hydrogen sulphide at a pH range of 3 to 6 to convert the zinc present as bisulphite in the leach liquor to insoluble zinc sulphide which is then fed to either the first stage roaster or the second stage roaster dependent upon its iron content. With proper control of pH during the zinc sulphide precipitation step, virtually complete recovery of the zinc constituents should be achieved with little or no iron co-precipitated. (See Figure VII) .

Another embodiment for treating the partially desulphurized zinc concentrate or, if you will, calcine produced after the first stage partial desulphurization roast of a conventional zinc concentrate is to leach this calcine in warm dilute sulphuric acid solution containing aqueous sulphur dioxide followed by liquid-solid separation techniques as previously described, whereby the leach residue is fed to a conventional dead roast and the leachate is subjected to a solvent extraction technique to selectively separate the dissolved zinc and iron and thus produce a zinc sulphate solution. The zinc sulphate solution is then fed to an appropriate place in a conventional zinc refining circuit while the raffinate is treated by one of the three options shown in Figure III.

Another embodiment for treating the leachate produced after the first stage desulphurization roast is described herewith. The leachate is treated with oxygen (air) to oxidize the dissolved ferrous iron to ferric and by hydrolytic action to precipitate the ferric iron as goethite. To enable iron precipitation to go to completion, lime is required to neutralize the acid released by the hydrolytic reaction while holding the pH in the range of 3-6. After liquid-solid separation, the resulting solution is fed into the zinc refining circuit evolving S0₂ for recycle, while the iron-containing solid is sent to disposal.

Other methods of treating the partially desulphurized concentrate or calcine in a two stage roasting process when a conventional zinc concentrate is to be treated in order to retard zinc ferrite formation may be found in the public domain or be obvious to one normally skilled in the art.

Referring back to the leach residue produced after the conventional neutral leach of the dead roasted product, the leach residue would be of very much smaller volume than that conventionally produced because of its very low zinc ferrite content and thus would be enriched in lead and precious metals content. This might be treated by flotation techniques to separate the lead as well as the precious metals from other gangue material and also to separate the silver from the lead component. An alternative method would be to use sodium cyanide or thiourea to leach and then separate the silver sulphide from the other leach residue materials. (See Figure VI in this later instance)

Another embodiment is in the case of the treatment of a low grade zinc sulphide concentrate not suitable for conventional roasting, where the low grade zinc concentrate would be given a first stage partial desulphurization roast to the extent that it would bring it up to a grade equivalent or better than a normal concentrate suitable for dead roasting after the intermediate leaching step. In this case, the preferred leachant might be a mixture of dilute sulphuric acid solution containing sulphur dioxide as previously described. (See Figure III which shows an aqueous S0₂ leachant variant, also Figure V) .

In one instance, a low grade zinc concentrate reported to contain 34.52% Zn, 15.95% Fe, 1.15% Cu, 0.23% Cd and 32.68% S was given a partial desulphurization roast to the extent that the calcine analysis was reported as 43.86% Zn, 16.23% Fe, 1.57% Cu, 0.28% Cd, 0.082% Pb and 24 . 05% S .

This partially roasted concentrate was then leached at an initial pulp density of 80 gl^"1 in a 3% H₂S0₄ solution containing dissolved S0₂ at a temperature of 65°C to 69°C for approximately three hours. After leaching and liquid-solid separation, the leach residue was reported to contain 50.8% Zn, 7.11% Fe, 1.86% Cu, 0.29% Cd, 0.062% Pb and 30.12% S, with only 3.72% of the zinc extracted into the leachate. As can be seen a non-useable zinc concentrate had thus been converted to an equivalent or superior grade of conventional zinc concentrate with only a slight loss of zinc, which would otherwise be lost in any event.

It is expected that the leachate being high in iron content and very low in other dissolved base metals could be treated with lime and disposed to a tailing pond or be oxidized to precipitate goethite and then limed for disposal.

Other methods of disposal or recovery of the iron might also be devised depending on circumstances and the surrounding environment.

The above described examples illustrate the flexibility of the process to upgrade low-grade zinc concentrates or conventional zinc concentrates to concentrates having any desirable level of iron content.

In the embodiment for the treatment of zinc sulphide bulk concentrates produced from complex massive sulphide mineral deposits, several options exist. For example, a conventionally produced zinc sulphide bulk concentrate could be upgraded to produce a superior grade of bulk concentrate having the attractive features such as a higher level of zinc and lead and a lower level of sulphur and a much lower level of iron thus reducing both transportation and treatment charges, the latter substantially. One example of this type of processing is shown in Figure II.

With conventional methods of mineral processing of such massive deposits, zinc sulphide bulk concentrates of relatively low value are produced. Example No. 3 shows a method of increasing the value of a zinc sulphide bulk concentrate. Also with our novel process it is conceivable that higher recoveries of all valuable metals could result. This is because lower grades of bulk concentrates could be upgraded to acceptable levels by using more of the orebody and discarding less mineral processing tailings. Also our process is adaptable to orebodies that contain less lead but are not suitable for conventional processing (See Example 4).

A flowsheet is provided for one method of treatment of zinc sulphide bulk concentrates produced or to be produced from complex massive base metal sulphide concentrates containing substantial levels of zinc, lead and iron sulphides. This flowsheet is presented in Figure II.

Example No. 3 provides laboratory results for one method of treating a zinc sulphide bulk concentrate. (See also Figure II combined with Figure IVA) .

The following description gives one embodiment of a method of using partial desulphurization roasting techniques for treating bulk concentrates to extract high values of zinc, lead and precious metals and to provide an environmentally satisfactory iron fraction.

A zinc sulphide concentrate consisting principally of zinc sulphides, iron sulphides and lead sulphides with lesser amounts of other metallic sulphides, usually in the form of complex sulphide compounds or solid solutions thereof is treated in a roaster using an oxygen containing feed gas, presumably but not necessarily ordinary air, in a manner that results in an oxygen deficient atmosphere at all times by controlling the retention time of the solid feed material at temperatures between 650°C and 1050°C in order to selectively convert its iron-containing constituents into readily soluble iron oxide, leaving unreacted the major portion of all the remaining sulphides resulting in a calcine or, if you will, a partially desulphurized concentrate.

This calcine is then treated with a medium temperature (50°C-75°C) dilute sulphuric acid solution containing dissolved sulphur dioxide or a medium temperature (50°C-75°C) aqueous sulphur dioxide solution as previously described, to leach any soluble oxides, which includes the major portion of the total iron and a minor portion of the converted base metal oxides. The resulting slurry is subjected to liquid-solid separation to separate the soluble oxides portion from the insoluble remaining sulphides.

The iron containing solution is then treated in one or more of the methods previously described in order to dispose of the iron fraction in an environmentally satisfactory manner.

The remaining sulphides which contains chiefly zinc sulphide and lead sulphide but also other metallic sulphides is subjected to a second stage partial desulphurization roasting operation using an oxygen containing gas for the conversion of the bulk of the contained zinc sulphides to zinc oxide under oxygen deficient roasting conditions as previously described and preferably in the temperature range of 650°C to 850°C.

The resulting second stage partial desulphurization roasting operation would be designed to produce a calcine which is chiefly composed of zinc oxide and unreacted sulphides, chiefly lead sulphide plus a concentrated amount of precious metals. This second stage calcine is either leached and subjected to liquid-solid separation techniques or treated by flotation or other physical separation techniques in order to selectively separate the zinc oxide and other base metal oxides fraction, containing the bulk of the zinc, from the remaining sulphides which would be chiefly composed of lead sulphide but would also contain almost all of the precious metals.

The separated oxide fraction which contains the bulk of the zinc as zinc oxide would then be sent to a zinc refinery operation for producing zinc metal or might be sold as a zinc oxide product. If the aqueous S0₂ leaching roast were used the resulting leachate containing chiefly zinc in the bisulphite form would be treated with spent electrolyte in order to convert the zinc to its soluble sulphate form for feed to a zinc refinery, thus regenerating sulphur dioxide for recycle.

The remaining sulphide fraction which would contain chiefly lead sulphide but would also contain a concentrated amount of precious metals and perhaps some gangue material could be sent directly to a lead smelter. Alternatively, the precious metals fraction would be separated from the remaining sulphide fraction by flotation techniques or by leaching with a cyanide or thiourea solution before the lead sulphide fraction is sent to a lead smelter for the production of lead. Some bulk sulphide concentrates might contain significant portions of arsenic and perhaps other elements not specifically mentioned in the description thus far. These types of bulk concentrates may be adapted to the techniques hereinbefore described but may complicate the roaster exit gas treatment system which would normally contain a high purity S0₂-laden gas stream which might be used as a source of sulphur dioxide gas for make-up purposes where aqueous sulphur dioxide leaching is employed but would mainly be used for conversion to sulphuric acid.

Other modifications to the circuitry may also be necessary to take into account the presence of amounts of other extraneous impurities.

In fluid bed dead roasting of conventional zinc sulphide concentrates, from about 50% to as much as 90% of the product may be collected from the gas stream leaving the top of the roaster. Under any such conditions, it may be difficult to control the residence time in the roaster and thus of the sulphur content of the product. Because this is our goal, i.e. to control the level of sulphur of the partial desulphurization roast product through control of the residence time or oxygen availability in the roaster, it may be necessary to agglomerate the concentrate such that a maximum of <5% of the product is elutriated from the roaster bed. Preliminary studies showed that this goal was achieved when the concentrate was agglomerated to -40 to +60 mesh size range. This also led to very smooth operation of the fluid bed roaster.

If the concentrate to be treated is ground to the -40 to +60 mesh size prior to flotation, this could affect the yield of zinc (and other base and precious metals). In practice, concentrate analyses as a function of grind (mesh size) would be determined to see whether this approach was desirable. If the yield would be greater but the tenor of the concentrate lower, a partial desulphurization roast would be conducted without the need for an agglomeration step while the leach residue after the partial desulphurization roast could be equal to or superior to most zinc sulphide concentrates available today and could conceivably be less than 1% iron.

If a fluid bed roaster were used for the partial desulphurization roasting, control of the particle size distribution to the fluid bed roaster may be desirable to achieve a desired residence time in the roaster and thus a controlled degree of sulphide sulphur removal, in order to provide an improved operation. Figures I through IX by means of flowsheets, show a variety of embodiments for the treatment of zinc sulphide concentrates including conventional zinc concentrates, low grade zinc concentrate and bulk zinc concentrates using partial desulphurization roasting techniques. Some of the embodiments that have been described have not been shown in flowsheet form.

Figures I through VII provide a variety of treatments of zinc sulphide concentrates including conventional concentrates, low grade zinc concentrates and bulk concentrates. Additional methods of treatment are included in the text. These all depend on two stage roasting wherein at least the first stage is conducted in an oxygen deficient atmosphere. Other methods of treatment using such two stage roasting may become apparent to those normally skilled in the art because of the wide range of flexibility. EXAMPLE NO. 1

A conventional zinc sulphide concentrate of the following analysis was partially desulphurized in air at a temperature of 850°C in a fluidized bed roaster to produce a partially desulphurized calcine of the analysis given below:

Concentrate Analysis Calcine Analysis Zn 49.0 Wt% 53.40 Wt%

Fe 9.10 Wt% 9.87 Wt%

Cu 0.70 Wt% 1.06 Wt% Cd 0.24 Wt% 0.27 Wt%

S 32.40 Wt% 25.70 Wt%

The S0₂ laden off-gas was reported to contain 19 vol% S0₂ and less than 0.1 vol% oxygen.

Forty grams (40.0 g) of the resulting partially desulphurized calcine was leached in 503 ml of an aqueous solution of sulphur dioxide. The leach was conducted at a temperature of 65 ± 5°C and a pH of 1.8 to 2.1 for two (2) hours. The leach pulp was filtered from the leachate. Analysis of the leachate showed a 14.20% zinc extraction, 90.77% removal of iron, 2.52% removal of copper, and 5.47% removal of cadmium. The resulting leach residue analyzed 59.30% zinc, 1.16% iron, 1.30% copper, 0.31% cadmium, and 30.80% sulphur.

Further treatment of the leachate to recover its 14.20% zinc content can be accomplished by various routes. EXAMPLE NO. 2

A low-grade zinc sulphide concentrate was partially desulphurized at a temperature of 850°C in a fluidized bed roaster to produce a partially desulphurized calcine. The concentrate and calcine analysis were reported to be as follows: Concentrate Analysis Calcine Analysis

Zn 34.5 Wt% 43.86 Wt%

Fe 16.5 Wt% 16.23 Wt%

Cu 1.15 Wt% 1.57 Wt%

Cd 0.23 Wt% 0.23 Wt% S 32.7 Wt% 24.05 Wt%

The off-gas was reported to contain 19 vol% S0₂ and less than 0.1 vol% oxygen.

Forty grams (40.0 g) of the resulting partially desulphurized calcine was leached in 500 ml of aqueous 3% H₂S0₄ plus S0₂ in solution. The leach was conducted at a temperature of 65°C ± 5°C and a pH of 1.1 for three (3) hours. The leachate was then filtered from the leach pulp. Analysis of the leachate showed an extraction of 3.72% Zn, 63.54% Fe, 0.35% Cu, and 5.32% Cd. The resulting leach residue analyzed 50.80% Zn, 7.11% Fe, 1.85% Cu, 0.28% Cd, and 30.12% S. EXAMPLE NO. 3

A complex New Brunswick zinc sulphide bulk concentrate of the analysis shown below was given a partial desulphurization roast in air at 750°C in a fluidized bed roaster to produce a partially desulphurized concentrate. Figure X provides a temperature profile during the continuous partial desulphurization roast along with S0₂ and 0₂ off-gas concentrations. Products removed during constant operating conditions are shown as PI, P2 and P3. A lower temperature was employed on the roast because of the high lead content of the complex concentrate. The partially desulphurized concentrate

(sample P3) was given a warm S0₂ leach as described in Example 1. After filtration and washing, the leach residue had the analysis shown below:

Concentrate Analysis Leach Residue Zn 34.6 Wt% 40.8 Wt%

Fe 12.6 Wt% 1.87 Wt%

Cu 0.9 Wt% 1.23 Wt%

Cd 0.1 Wt% 0.1 Wt% Pb 16.4 Wt% 28.9 Wt%

S 33 Wt% 24.2 Wt%

Two additional partially desulphurized concentrates were produced from the same New Brunswick bulk zinc concentrate. In the first (sample PI), the % S was reduced from 33 wt% to 24.9 wt% and in the second (sample P2), the % S was reduced to 18.8 wt%. The partially desulphurized concentrates from each roast were given identical aqueous S0₂ leaches. The extraction of zinc and iron for both leaches was determined as a function of leaching time and the results plotted in Figure XIII. For the roast that had less sulphur removed (PI)_/ less than 2% of the zinc dissolved along with about 60% of the iron; for the roast that had a much greater amount of sulphur removed (P2), some 20% of the zinc was extracted along with up to 90% of the iron.

This provides two distinct possibilities: an upgrading roast to remove part of the iron along with a very small amount of zinc or a roast carried out to permit dissolution of nearly all of the iron and a minor portion of the zinc while leaving the more stable sulphides untouched. EXAMPLE NO. 4

A second New Brunswick zinc sulphide concentrate of the analysis shown below was given a partial desulphurization roast as profiled in Figure XII. The temperature at the end of the roast was raised to above 900°C to determine whether the 1.2% lead content would result in a defluidized bed. No operational problems were noted. Concentrate Analysis

Zn 51.1 Wt% Fe 10.2 Wt% Cu 0.24 Wt%

Cd 0.10 Wt%

Pb 1.20 Wt%

S 34.4 Wt% Four samples of the partially roasted concentrate were removed during the roast (i.e. PI, P2, P3 and P4) . Similar S0₂ leaches to that in Example No. 1 were conducted on the first three samples which were taken when the roaster was operating at about 750°C. Slight changes in the feed rate to the roaster and thus the residence time of roasting resulted in slightly different sulphur contents in each partially desulphurized roasted product.

The leaching profiles for each sample are presented in

Figure XIII. These results illustrate the ability to control the roasting operation by residence time alone to achieve just about any level of iron removal desired. The amount of zinc co-leached with the iron is also shown.

There are a number of approaches that can be used to recover the zinc, and recycle the S0₂ and dispose of the iron.

Claims

THE EMBODIMENTS OF THE INVENTION IN WHICH AN EXCLUSIVE PROPERTY OR PRIVILEGE IS CLAIMED ARE DEFINED AS FOLLOWS:

1. A process for the recovery of zinc from iron- containing zinc sulphide ores or concentrates comprising the steps of:

(a) roasting the ore or concentrate in the presence of an oxygen-containing gas wherein sulphur retention is maintained at a level of 15% or higher by controlling the residence time of the ore or concentrate and/or the flow of oxygen in the roaster to provide an oxygen deficient atmosphere in the roaster using a temperature in the range of 700°C to 1050°C in order to preferentially oxidize the iron rather than the zinc and other base metal constituents and thus reduce or eliminate zinc ferrite formation,

(b) leaching the partially desulphurized roasted ore or concentrate using a lixiviant selected from the group consisting of aqueous sulphur dioxide, a 2-5 wt% sulphuric acid solution, and a 2-5 wt% sulphuric acid solution containing solubilized sulphur dioxide, at a temperature between 50°C and 75°C in a pH range of 1.0 to 2.5 for one to three hours in order to preferentially dissolve the iron oxide thus formed as well as base metals that have co- oxidized thus obtaining a leach slurry

(c) subjecting the leach slurry to liquid-solid separation techniques

(d) treating the leachate to separate the dissolved iron from the dissolved base metals, thus leaving a leach residue which is richer in zinc sulphide content and depleted in iron content which

(ii) if the original concentrate is rich in lead, such as in a zinc sulphide bulk concentrate, subjecting said leach residue to a second stage partial desulphurizationroast using a controlled retention time in the temperature range of 650°C to 850°C such that the retained sulphur content is 8% or higher which relates to the amount of lead present to produce a second stage calcine which has a very high zinc oxide content with substantially all of the lead remaining in the sulphide form, subjecting said second calcine to physical separation techniques or to an aqueous sulphur dioxide leach under similar conditions as earlier described in order to separate the zinc oxide fraction from the lead sulphide fraction, the sulphide- containing residue is then fed directly into a lead smelter or is further treated to separate the lead sulphide from the precious metals.

2. A process according to Claim 1 wherein the sulphur retention is between 15 and 27%.

3. A process according to Claim 1 wherein the sulphur retention is between 20 and 25%.

4. A process according to Claims 1, 2 or 3 wherein the first stage roasting temperature is between 850°C and

1000°C.

5. A process as described in Claims 1, 2, 3 or 4 wherein a 3-4 wt% sulphuric acid solution is used as the leachant.

6. A process as described in Claims 1, 2, 3, 4 or 5 wherein the leaching temperature range is between 60°C and 70°C and the pulp density is between 60 and 120 gpl and the leaching period is two hours or less.

7. A process as described in Claims 1, 2, 3, 4 or 5 wherein the feed material is a conventional zinc concentrate and there is a first stage oxygen deficient roast followed by a leach step and a dead roast of the leach residue.

8. A process as described in Claim 7 wherein the feed material is a conventional zinc concentrate or a low- grade zinc concentrate containing less than 2% by weight of lead sulphide.

9. A process as described in Claims 1, 2, 3, 4 or 5 wherein the leachate produced as a result of the leach step is thermally decomposed to drive off sulphur dioxide for recycle and to produce a precipitate consisting principally of a mixture of zinc sulphite and iron sulphite and a depleted liquor solution. The solid sulphite mixture is then treated by an aqueous ammonia leach to selectively dissolve the base metal components and leave effectively all of the iron content in the residue. The iron-containing residue after liquid-solid separation is then subjected to thermal and physical treatment steps to produce a high purity iron oxide product. The solution portion is then thermally decomposed to drive off ammonia for recycle producing a basic zinc sulphite precipitate and a partially depleted ammonia liquor. The basic zinc sulphite precipitate, after liquid- solid separation, is dehydrated to produce a zinc sulphite product suitable for treatment with a portion of the spent electrolyte stream being recycled to a conventional zinc refinery neutral leach circuit thus to convert the zinc sulphite to soluble zinc sulphate and thus drive off sulphur dioxide for recycle and the partially depleted ammonia liquor is then treated with calcium oxide or calcium hydroxide to precipitate its sulphite content as

" calcium sulphite and after liquid-solid separation, the dissolved ammonium hydroxide is recycled to the ammonia leaching circuitry. The calcium sulphite precipitate is treated with a sulphuric acid solution to dissolve any base metal components of the precipitate and fed to a liquid-solid separation step to separate the insoluble calcium sulphate thus formed and permit the base metal constituents to be recovered as well as the generated sulphur dioxide.

10. A process as described in Claims 1, 2, 3, 4 or 5 wherein the aqueous sulphur dioxide leachate is subjected to solvent extraction by a method similar to Clitheroe's U.S. Patent No. 4,053,552 to recover the contained zinc as zinc sulphate whereas the iron bisulphite raffinate is further treated before disposal.

11. A process as described in Claims 1, 2, 3, 4 or 5 wherein the aqueous sulphur dioxide leachate is treated with hydrogen sulphide at a pH range of 3 to 6 to precipitate the zinc as zinc sulphide so that it can be separated from the iron component remaining in its soluble form.

12. A process according to Claims 1, 2, 3, 4 or 5 wherein the feed material is a low-zinc high-iron content zinc concentrate wherein the first stage partial desulphurization roast leads to the formation of a conventional or superior zinc concentrate after leaching and a solution containing a relatively high iron level with only a relatively small amount of dissolved zinc, wherein the solution is treated with calcium hydroxide or calcium oxide and the resultant iron-containing residue is fed to a tailings pond.

13. A process according to Claim 1 wherein a zinc sulphide bulk concentrate containing substantial quantities of zinc, lead and iron sulphides as well as some precious metals is treated so that the product of the second stage partial desulphurization roast contains little iron but large quantities of zinc oxide, lead sulphide and perhaps some precious metals wherein the zinc oxide is separated from the lead sulphide and precious metals by described leaching techniques followed by liquid-solid separation techniques leaving an insoluble residue containing essentially lead sulphides, enriched precious metals and perhaps some gangue material which is then treated with a thiourea or cyanide solution by a leaching process to selectively separate the precious metals from the lead sulphides. The lead sulphide concentrate is then fed to a lead smelter.

14. A process wherein the bulk concentrate is treated by a partial desulphurization roasting technique by controlling the residence time in the roaster to the extent that the sulphide sulphur retention is between 20 and 25% sulphur and the roasting temperature is in the range of 650°C to 850°C in order to convert the contained iron into its oxide form in preference to the contained base metals and precious metals and after separating the iron oxides from the remaining sulphides by leaching and liquid-solid separation as described in Claim 1 by the use of leachants consisting of aqueous S0₂ solution, sulphuric acid or a mixture thereof, the remaining sulphide fraction is treated by a second partial desulphurization roast in the temperature range of 650°C to 850°C to reduce the sulphur content in the second stage roast to a level where 80% of the contained zinc or more is converted to its oxide form, and preferably 90% of zinc or more is thus converted, and the resulting calcine is then treated by selective flotation techniques to separate the contained zinc oxide and gangue material from the remaining sulphides which will consist primarily of lead sulphides which are enriched with precious metals, so that the lead sulphide fraction can be fed to a lead smelter either before or after separating the precious metals.

15. A process wherein the bulk concentrate is treated by a partial desulphurization roasting technique by controlling the residence time in the roaster to the extent that the sulphide sulphur retention is above 8% and the roasting temperature is in the range of 650°C and the 850°C in order to convert the contained iron and contained zinc in preference to the contained lead and precious metals into their oxide form but short of converting lead sulphides or precious metals sulphide to their oxide form. The oxidized components are selectively dissolved with aqueous sulphur dioxide solution under the conditions described in Claim 1 and after liquid-solid separation, the leachate is treated by solvent extraction to separate the zinc from the iron components as in Claim 9. The leach residue is treated by thiourea or cyanide leaching to separate the precious metals from the lead sulphides, after which the silver is recovered and the lead concentrate is fed to a lead smelter.

16. A process as described in Claim 15 wherein a dilute sulphuric acid solution is used to dissolve the oxidized components and after liquid-solid separation the leachate is treated with air and lime to a pH of 5 to 6 to separate the zinc and iron components.

17. A process as described in Claim 16 wherein the original feed material is either a conventional zinc concentrate or a low grade zinc concentrate rather than a bulk concentrate.

18. A process according to Claim 1 wherein the off- gas from the partial desulphurization roasting step is used as sulphur dioxide make-up in the aqueous sulphur dioxide leaching process and the balance of high quality sulphur dioxide-containing gas is converted into sulphuric acid.

19. A process according to Claim 1 wherein the concentrate feed material contains substantial amounts of arsenic wherein the arsenic is oxidized and is emitted with the off-gas and is then separated from the off-gas stream prior to feeding the S0₂-laden gas to the sulphuric acid plant .

20. A process according to each claim where one or more of the roasters are fluid bed roasters.

21. A process wherein the size distribution of concentrate particles is maintained at >100 and preferably

<20 mesh (>200 Aim; <900 Aim) where it is desired to control the residence time of the solids fed to a fluid bed roaster, and thus to reduce the elutriation of small particles from the bed.

22. A process as in Claim 21 wherein the preferred size distribution is approximately <60 mesh; >40 mesh

23. A process for recovering zinc from zinc bearing ores or concentrate substantially as disclosed herein.