WO1990013679A1 - Nouveau procede de traitement de minerais et/ou de concentres contenant du sulfure de zinc - Google Patents

Nouveau procede de traitement de minerais et/ou de concentres contenant du sulfure de zinc Download PDF

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Publication number
WO1990013679A1
WO1990013679A1 PCT/CA1990/000130 CA9000130W WO9013679A1 WO 1990013679 A1 WO1990013679 A1 WO 1990013679A1 CA 9000130 W CA9000130 W CA 9000130W WO 9013679 A1 WO9013679 A1 WO 9013679A1
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Prior art keywords
zinc
iron
sulphide
concentrate
lead
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PCT/CA1990/000130
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English (en)
Inventor
Murry C. Robinson
Donald R. Spink
Kim D. Nguyen
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University Of Waterloo
Materials-Concepts-Research Ltd.
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Application filed by University Of Waterloo, Materials-Concepts-Research Ltd. filed Critical University Of Waterloo
Publication of WO1990013679A1 publication Critical patent/WO1990013679A1/fr

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Classifications

    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B19/00—Obtaining zinc or zinc oxide
    • C22B19/02—Preliminary treatment of ores; Preliminary refining of zinc oxide
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • C22B1/10—Roasting processes in fluidised form
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

Definitions

  • 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.
  • 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 .
  • 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.
  • 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.
  • 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.
  • 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.
  • 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 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% H 2 S0 4 and preferably 3-4 wt% H 2 S0 4 . Where a sulphuric acid and aqueous S0 2 solution are combined, the sulphuric acid solution would still be between 2-5 wt% H 2 S0 4 and the aqueous S0 2 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.
  • 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.
  • physical separation techniques might be applied to the partially desulphurized calcines and/or to upgrade the leach residue produced.
  • the leach residue or physical residue remaining after the iron oxide fraction has been separated provides the feed to the second stage roasting step.
  • 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.
  • the words "partially desulphurized” refers to partial oxidation of the contained metal sulphides.
  • 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.
  • 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.
  • chemical or physical separation techniques might be employed prior to feeding the lead sulphide concentrate to the lead smelter. These techniques will be described later in this disclosure.
  • the process may be applied to low grade zinc sulphide concentrates or bulk zinc sulphide concentrates or to conventional zinc sulphide concentrates.
  • 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.
  • 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.
  • 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.
  • FIG. 1 An embodiment of the invention is shown in flowsheet form by combining Figure I with Figure IVA for treating a conventional zinc concentrate.
  • 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.
  • an aqueous 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%
  • S0 2 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 .
  • 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.
  • the treatment of the leachate involved thermal decomposition to drive off S0 2 -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.
  • 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 S0 2 gas for recycle to the aqueous S0 2 leaching step.
  • the recovery of zinc from the S0 2 leachate was reported as 89.4%.
  • 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.
  • 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 H 2 S0 4 will give off pure S0 2 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.
  • a sulphuric acid treatment step on the precipitate would be required to dissolve the base metal compounds for recovery from the insoluble calcium compounds.
  • the circuitry would be quite small and therefore quite inexpensive.
  • 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.
  • 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.
  • oxygen air
  • lime is required to neutralize the acid released by the hydrolytic reaction while holding the pH in the range of 3-6.
  • the resulting solution is fed into the zinc refining circuit evolving S0 2 for recycle, while the iron-containing solid is sent to disposal.
  • 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.
  • 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 S0 2 leachant variant, also Figure V) .
  • 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% H 2 S0 4 solution containing dissolved S0 2 at a temperature of 65°C to 69°C for approximately three hours.
  • 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.
  • 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.
  • 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.
  • 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) .
  • 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 S0 2 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.
  • 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.
  • circuitry may also be necessary to take into account the presence of amounts of other extraneous impurities.
  • concentrate analyses as a function of grind 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.
  • FIGS I through IX 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:
  • the S0 2 laden off-gas was reported to contain 19 vol% S0 2 and less than 0.1 vol% oxygen.
  • 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
  • the off-gas was reported to contain 19 vol% S0 2 and less than 0.1 vol% oxygen.
  • FIG. 1 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 S0 2 and 0 2 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 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 S0 2 and 0 2 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 was given a partial
  • Example P3 was given a warm S0 2 leach as described in Example 1. After filtration and washing, the leach residue had the analysis shown below:

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Abstract

Le procédé décrit permet le grillage de désulfuration partielle d'une variété de minerais ou de concentrés de sulfure contenant du zinc, en réglant la température de l'appareil de grillage et le temps de séjour dans l'appareil de grillage formant ainsi une atmosphère pauvre en oxygène, de sorte que la quantité requise de rétention de sulfure est maintenue. La matière calcinée ainsi obtenue est ensuite soumise à diverses étapes de séparation chimique et/ou physique, pour séparer les sulfures qui n'ont pas réagi, lesquels, selon certains modes de réalisation dudit procédé sont ensuite grillés à fond et traités en vue de l'extraction du zinc. Dans un autre mode de réalisation, les sulfures ainsi séparés qui n'ont pas réagi sont à nouveau partiellement désulfurés dans une atmosphère pauvre en oxygène, en vue d'un traitement ultérieur destiné à l'extraction du zinc, du plomb et des métaux précieux. Les étapes de séparation chimique se composent d'un traitement au dioxyde de soufre aqueux, d'un traitement en solution d'acide sulfurique dilué et d'un traitement au moyen d'acide sulfurique dilué contenant du dioxyde de soufre en solution, tous ces traitements étant effectués à une température comprise entre 50 et 75°C. Les étapes des séparation physique se composent d'une ou de plusieurs techniques de flottation et de séparation magnétique, bien que d'autres étapes de séparation physique puissent également s'appliquer dans certains cas.
PCT/CA1990/000130 1989-05-03 1990-04-25 Nouveau procede de traitement de minerais et/ou de concentres contenant du sulfure de zinc WO1990013679A1 (fr)

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WO1997033007A1 (fr) * 1996-03-07 1997-09-12 N.V. Union Miniere S.A. Procede de recuperation de zinc a partir de minerais ou de concentres contenant de la sphalerite
CN102912147A (zh) * 2012-11-15 2013-02-06 昆明冶金研究院 锌氧压浸出渣浮选硫磺后尾渣中回收铅锌、银、铁的工艺
US20130291684A1 (en) * 2010-12-14 2013-11-07 Outotec Oyj Process and plant for treating ore concentrate particles containing valuable metal
CN104014420A (zh) * 2014-06-10 2014-09-03 李锦源 一种低品位氧硫混合铅锌矿多金属回收的方法
CN104258981A (zh) * 2014-09-15 2015-01-07 中冶北方（大连）工程技术有限公司 一种锌铁矿选别工艺
CN109467119A (zh) * 2018-12-18 2019-03-15 兴化市万润锌业有限公司 一种可降低污染物的高纯度氧化锌制备工艺及其制备方法
CN114657372A (zh) * 2022-03-01 2022-06-24 中国恩菲工程技术有限公司 从低品位硫化铜钴精矿中提取铜元素和钴元素的方法

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WO1997033007A1 (fr) * 1996-03-07 1997-09-12 N.V. Union Miniere S.A. Procede de recuperation de zinc a partir de minerais ou de concentres contenant de la sphalerite
US6162346A (en) * 1996-03-07 2000-12-19 N.V. Union Miniere S.A. Process for recovery of zinc from sphalerite containing ores or concentrates
CN1066203C (zh) * 1996-03-07 2001-05-23 联合矿业有限公司 从含闪锌矿的矿石或精砂中回收锌的方法
KR100496320B1 (ko) * 1996-03-07 2005-09-15 우미코르 광석또는정광을함유하는섬아연광으로부터아연을회수하는방법
US20130291684A1 (en) * 2010-12-14 2013-11-07 Outotec Oyj Process and plant for treating ore concentrate particles containing valuable metal
US9200345B2 (en) * 2010-12-14 2015-12-01 Outotec Oyj Process and plant for treating ore concentrate particles containing valuable metal
CN102912147A (zh) * 2012-11-15 2013-02-06 昆明冶金研究院 锌氧压浸出渣浮选硫磺后尾渣中回收铅锌、银、铁的工艺
CN104014420A (zh) * 2014-06-10 2014-09-03 李锦源 一种低品位氧硫混合铅锌矿多金属回收的方法
CN104014420B (zh) * 2014-06-10 2016-03-02 李锦源 一种低品位氧硫混合铅锌矿多金属回收的方法
CN104258981A (zh) * 2014-09-15 2015-01-07 中冶北方（大连）工程技术有限公司 一种锌铁矿选别工艺
CN109467119A (zh) * 2018-12-18 2019-03-15 兴化市万润锌业有限公司 一种可降低污染物的高纯度氧化锌制备工艺及其制备方法
CN114657372A (zh) * 2022-03-01 2022-06-24 中国恩菲工程技术有限公司 从低品位硫化铜钴精矿中提取铜元素和钴元素的方法

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