WO1998036102A1 - Raffinage de minerais contenant du sulfure de zinc - Google Patents

Raffinage de minerais contenant du sulfure de zinc Download PDF

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Publication number
WO1998036102A1
WO1998036102A1 PCT/AU1998/000094 AU9800094W WO9836102A1 WO 1998036102 A1 WO1998036102 A1 WO 1998036102A1 AU 9800094 W AU9800094 W AU 9800094W WO 9836102 A1 WO9836102 A1 WO 9836102A1
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WO
WIPO (PCT)
Prior art keywords
zinc
fume
solution
sulphate
process according
Prior art date
Application number
PCT/AU1998/000094
Other languages
English (en)
Inventor
Graham Alan Buckett
Christopher Rivers Fountain
Roderick James Sinclair
Original Assignee
Buka Technologies Pty. Ltd.
Priority date (The priority date is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the date listed.)
Filing date
Publication date
Priority claimed from AUPO5148A external-priority patent/AUPO514897A0/en
Priority claimed from AUPO7711A external-priority patent/AUPO771197A0/en
Application filed by Buka Technologies Pty. Ltd. filed Critical Buka Technologies Pty. Ltd.
Priority to AU59743/98A priority Critical patent/AU5974398A/en
Publication of WO1998036102A1 publication Critical patent/WO1998036102A1/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
    • C22B19/20—Obtaining zinc otherwise than by distilling
    • C22B19/24—Obtaining zinc otherwise than by distilling with leaching with alkaline solutions, e.g. ammonia
    • 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
    • C22B19/04—Obtaining zinc by distilling
    • 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
    • C22B19/34—Obtaining 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
    • C22B5/00—General methods of reducing to metals
    • C22B5/02—Dry methods smelting of sulfides or formation of mattes
    • C22B5/16—Dry methods smelting of sulfides or formation of mattes with volatilisation or condensation of the metal being produced
    • 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 relates to a process for the metallurgical processing of suitable zinc sulphide-containing material.
  • the process provides for the production of zinc oxide as an end product.
  • the process includes electrolytic refining for the production of zinc metal in which zinc oxide is produced for use as an intermediate for such production.
  • the invention has application in the processing of feed material comprising concentrates of zinc sulphide.
  • the invention also is applicable to the processing of zinc sulphide ores, subject to their availability at a sufficiently high grade for economic recovery through direct smelting.
  • the invention also enables the separation and recovery of other metal values from the feed material, such as lead, silver, and cadmium usually present in zinc sulphide ores and concentrates.
  • the principal source of suitable feed material is zinc sulphide ores.
  • zinc sulphide ores are concentrated at or near the minesite, with flotation the usual method.
  • the concentration can be to provide a single, mixed metal, bulk concentrate containing most of the zinc mineral content of the ore, along with other mineral species.
  • the concentration can produce at least two relatively clean concentrates, including a zinc rich concentrate and a second concentrate rich in another mineral species, usually lead.
  • the production of a single concentrate enables higher metal recoveries to concentrate, but can realise lower net smelter returns. Production of at least two concentrates realises greater net smelter returns, despite some contamination of each concentrate with unwanted mineral species.
  • Flotation alone or in combination with other pre-concentration techniques, necessitates crushed ore being ground to a particle size sufficient to liberate the zinc mineral species from the gangue minerals and from other mineral species.
  • the size reduction required for this varies from one ore body to another, but it is not uncommon with sulphide ores having fine grained, intimately intermixed mineral species for it to be necessary to achieve 80% of ground particles less than 20 microns or finer. Grinding to such fine sizes is costly in both capital and operating costs, while resultant high slimes levels interfere with flotation and necessitate desliming or selective flocculation.
  • the EZ process has a roasting stage, in which zinc sulphide-bearing mineral concentrate is heated to the point where the sulphur contained in the minerals reacts with atmospheric oxygen to form sulphur dioxide. Oxygen also reacts with the metal ions remaining in the solid phase to form calcine comprising oxides of zinc, iron, lead, and other metals. The zinc contained in the calcine is dissolved in sulphuric acid, treated in several purification steps, and finally extracted from the acid solution by electrowinning.
  • An ISP-based smelter uses a sinter plant to oxidise the zinc sulphide- bearing concentrate and to fuse it to form sinter lumps of porous metal oxides.
  • the sinter is placed in a sealed blast furnace in which the atmosphere contains substantial amounts of carbon monoxide.
  • the carbon monoxide causes the conversion of zinc and lead oxides contained in the sinter to their metals, and the high temperatures causes the zinc to form a vapor.
  • This vapor is carried from the furnace in the off-gas and passes through a chamber wherein it is cooled rapidly by splashing with liquid lead at a considerably lower temperature.
  • the zinc vapor is condensed and cooled by the lead, which passes out of the chamber through a weir.
  • separation of zinc, lead and silver is often difficult by conventional flotation methods and there can be significant loss of these valuable metals in tailings or as non-payable components, such as zinc in lead concentrates or lead and a large part of the silver contained in zinc concentrates.
  • a bulk mixed concentrate suitable for an ISP smelter may achieve higher overall metal recoveries but financial returns to the mine are likely to be lower and market outlets are much more restricted.
  • the present invention provides an improved process for the recovery of zinc from zinc sulphide ore and/or concentrate.
  • the invention is directed to providing a process for production of high grade zinc metal by production of zinc oxide product of a grade suitable for use as feed material for the production of metallic zinc by an electrolytic process; that is, by electrowinning.
  • the process enables the disadvantages of conventional processes utilising preconcentration in the production of a zinc from ore to be obviated, while providing numerous practical benefits over the conventional route of preconcentration.
  • the present invention enables a more flexible method of smelting zinc sulphide-containing feed material and a means of maximising the value of individual metals payable to the mining operation.
  • the process can tolerate greater concentrations of minor elements in the feed than can the EZ process.
  • Zinc oxide fume presently has limited use as raw material feed for zinc production by the electrolytic process due to the presence of impurities which are deleterious to that process, particularly halides. It is used more extensively as a concentrate feed replacement for pyrometallurgical processes such as the Imperial Smelting Process and the Retort Process, but usually such oxides are derived from secondary materials at relatively low value, and not from primary ores.
  • the present invention provides a route for the efficient recovery and production of high quality zinc and by-product metals from complex ores using a unique combination of processing steps which include a fuming procedure and subsequent leaching and recovery steps to provide zinc oxide of a grade suitable for primary zinc recovery.
  • zinc sulphide-containing feed material such as detailed above is subjected to a first stage, pyrometallurgical operation to generate crude zinc oxide fume product.
  • the crude fume product then is subjected to a second stage, hydro-metallurgical operation to produce a high purity zinc oxide product and to separate other valuable metals such as lead and silver.
  • the zinc oxide product then preferably is utilised, in a third stage of the process of the invention, as feed material in the production of metallic zinc by the electrolytic process.
  • This can be a conventional electrolytic process for the electrowinning of metallic zinc, but preferably the zinc oxide product is used as feed for a highly simplified electrolytic process using a standard sulphate electrolyte.
  • the feed material is smelted in a high temperature furnace to oxidize sulphidic sulphur to sulphur dioxide, to volatize zinc into the vapour phase and to produce a molten slag containing the gangue minerals.
  • the vaporised zinc is subsequently oxidized to a zinc oxide rich crude fume which is collected for further processing.
  • the fume may contain lead and silver values but can be substantially free of gangue minerals such as iron, manganese and silica.
  • the furnace may be a flash smelting shaft, a high temperature flame reactor, a shaft furnace, a kiln, or a slag bath reactor with either lance or tuyere injection.
  • the feed material is smelted in a molten slag bath.
  • the feed material is added to and dissolved in the slag and may require the addition of fluxing materials such as silica, lime, or iron compounds.
  • the bath is agitated and heated by injecting air or oxygen-enriched air into the slag, below or onto the top surface of the bath.
  • Zinc sulphide in the feed material is decomposed to zinc oxide in the slag and sulphur dioxide.
  • the zinc oxide so produced is reduced to elemental zinc by carbonaceous reductant charged to the slag whereby metallic zinc is volatilized from the bath.
  • the resultant zinc vapor is oxidised in the gas phase above the top surface of the bath to form zinc oxide fume which is removed from the furnace with sulphur dioxide and other combustion products in smelter gases.
  • the fume is separated from the gases as a crude fume product.
  • the crude fume product is leached in a concentrated aqueous solution of ammonia and carbon dioxide having a stoichoimetric excess of ammonia, that is, an ammonia/ammonium carbonate (AAC) solution, to form a soluble zinc.ammonia. carbonate complex.
  • Resultant leach solution then is treated to precipitate basic zinc carbonate product which is able to be recovered and calcined to provide a purified zinc oxide product.
  • AAC ammonia/ammonium carbonate
  • the purified zinc oxide product produced by the second stage preferably is used in the electrolytic process for the production of metallic zinc.
  • the zinc oxide can be utilised in the circuit of a conventional electrolytic process. However, it preferably is used in a modified electrolytic process as disclosed herein.
  • the present invention facilitates attainment of, or provides, significant practical benefits and addresses significant difficulties associated with the currently employed technologies used in the main stream production of zinc by the electrolytic process. These include the following:
  • the invention enables high metal recoveries for each of zinc, lead and silver. This is achievable at both the mine and refinery and applies particularly where the feed material is ore, rather than a concentrate, such that the loss of metal values in flotation failings is avoided.
  • the invention provides flexibility in relation to suitable feed materials.
  • the feed material differs from the recycled residues, principally EAF dust, with which disclosures of the above-detailed patents of Peters and Keegel are concerned, in that the feed material is either zinc sulphide ore or zinc sulphide concentrate.
  • the invention is flexible in relation to the mineral species in which the metal values are present, the relative amounts and mineral species of the gangue constituents, and the fineness and complexity of dissemination of the mineral species.
  • the process of the invention is able to be operated with a high level of environmental acceptability. Substantially all wastes are able to remain at the mine/smelter site, in a slag suitable for use as land-fill, such as for mine backfill. Also, substantially no waste is produced in the leaching and electrowinning stages, since outputs other than the respective end products of zinc oxide and metallic zinc comprise by-products amendable to conventional processing. This particularly addresses the issue of iron residues in current electrolytic zinc technology.
  • the present invention utilises technology known in other contexts.
  • Wendt, W.J. Ammonia: ammonia carbonate leaching of zinc ores
  • Eng. Min. J. 154(9), September, 1953, 84-90 the process steps of the second stage of the present invention are proposed as a potential new process.
  • US patent 4071357 issued on 31 July 1978 to Peters there is disclosed the use of similar steps in the recovery of zinc.
  • U.S. patent 5538532 to Keegel, Jr. there is proposed a two-stage process for the recovery of zinc and other metals from electric arc furnace (EAF) dust.
  • EAF electric arc furnace
  • the first stage of the Keegel process involves roasting the EAF dust in a reducing furnace to vaporize, and to re-oxidize in the furnace, the cadmium, zinc, and lead, and to leave a residual sinter mass containing iron.
  • the second stage of the Keegel process provides for an AAC leach and treatment for separation of cadmium, zinc and lead values.
  • EAF dust a material listed as being a hazardous waste
  • the treatment of EAF dust is to recover a substantial proportion of the zinc in the EAF solids feed.
  • Such proportion is indicated by Example 1 of Peters to be about 55%, with the substantial balance of the zinc remaining in the form of zinc ferrite in a residue remaining after leaching the EAF dust with a concentrated solution of ammonia and carbon dioxide.
  • residue retains about 95% of the lead and about 57% of the cadmium, the principal elements which resulted in EAF dust being listed as a hazardous waste.
  • Peters indicates that some further zinc can be recovered by leaching the residue with acetic acid.
  • Keegel also is applicable to EAF dusts. As indicated in USP 5538532, this material is composed mainly of oxides of iron, zinc, lead, tin, cadmium, chromium, manganese, nickel, copper and molybdenum; while silica, lime and alumina may be present.
  • the process of Keegel is applicable to other materials comprising dusts generated in steel production, recycling of scrap metal and other metallurgical operations, including neutral leach residues, foundry and blast furnace dust, recovered metal powders and metallic ores.
  • the ores do not include sulphide ores since the Keegel process does not take sulphur into consideration and, as is evident from USP 5538532, the process will not work for the recovery of zinc from a zinc sulphide ore or concentrate. This is the case where the end product is to be zinc oxide, but it is even more clearly the case where the end product is to be metallic zinc produced by electrolysis.
  • the smelting may be conducted in a top-submerged lancing reactor system having a reactor vessel containing the slag bath, and at least one top-submerged lance, by which oxygen-containing gas is injected into the slag.
  • the smelting may be in a similar reactor system but providing for top blowing rather than top-submerged injection.
  • the smelting may be conducted in a reactor containing the slag bath, with the reactor having at least one side or bottom blowing tuyere by which the oxygen-containing gas is injected into the slag.
  • smelting may be conducted in a kiln such as a Waelz Kiln, in a shaft furnace or within a high temperature flame such as in a flash smelting furnace or flame cyclone reactor.
  • a top-submerged lancing reactor system is preferred for a variety of reasons, including increased overall control of smelting conditions, avoidance of problems of blockage encountered with use of tuyeres and avoidance of the need for extensive feed preparation or presentation of a molten feed required in some of the other options.
  • the zinc sulphide-containing feed material can comprise zinc concentrate.
  • a zinc ore of a sufficient grade can be treated directly, rather than after pre-concentration, maximising the recovery of valuable metals.
  • the feed material does not require fine grinding and it may be charged to the slag bath as relatively coarse crushed material.
  • the particle size can be set by the need to achieve dissolution of the feed material in the slag in a practical time since, as will be appreciated, finer particles will be dissolved at a faster rate than coarse lumps.
  • fines smaller than two to three millimetres may cause excessive dust carryover to fume and may therefore need to be limited, through the use of a pelletizer or some other agglomeration technique.
  • the zinc sulphide-containing feed material can be charged to the slag bath in a wide variety of feed sizes, it is preferred that fines be charged separately from coarser and/or lump feed material, or agglomerated such as by wetting and balling on a disc and fed with coarser material. This separation may not be necessary if the ratio of fines to coarser particles is sufficiently small.
  • Agglomerated fines and coarser material can be charged through a suitable charging port of the reactor, such as via feed chute.
  • unagglomerated fines may be injected into the slag bath, below the surface thereof, by a lance or tuyere whether this be the same lance or tuyere used for oxygen injection or another injector.
  • the reductant preferably is coal in lump form. Where lump coal is used, it preferably is added to the bath via a suitable charging port of the reactor which substantially precludes the egress of gases from the reactor. However, lump coal can be added via a lance and particulate coal can be used, if required. Also, other carbonaceous reductants can be used, although they generally are less efficient and/or are more costly than coal.
  • the other reductants include, but are not limited to, coke, coke breeze, natural gas, LPG and fuel oil. Such other reductants, or particulate coal reductant, can be supplied to the reactor by blowing, or by submerged injection into the slag via a lance or via a tuyere.
  • the lance or tuyere for supplying the reductant may be the same as that providing for oxygen and fuel injection. Also, where a lance is used for oxygen and fuel injection, a tuyere can be used for supplying reductant and, where a tuyere is used for oxygen injection, a lance can be used for supplying reductant. Where a lance or tuyere for oxygen injection also is used to inject reductant, the reductant may be entrained in a suitable carrier gas, such as nitrogen or air.
  • the lance or tuyere most preferably has respective passages, such as concentric passages, for separate flow therethrough of the oxygen and reductant which isolates them from each other at least until they are close to the discharge end of the lance or tuyere.
  • reductant may be burned to provide part of the energy requirement of the furnace. In this case, sufficient oxygen must be provided to allow for this combustion and the oxidation of sulphur, iron and other compounds in the furnace.
  • the following reactions occur: 2ZnS+30 2 ⁇ 2ZnO + 2S0 2 (1) or
  • the free-oxygen containing gas can comprise air, oxygen or oxygen- enriched air.
  • the feed material need not only comprise sulphide. Thus, it may include sulphate and/or oxide but preferably with sulphide predominating.
  • the oxygen requirement for. reactions (1) and (2) and similar reactions for other metal sulphides will vary with the total sulphide content, and for the requirement, if any, for combustion of part of the reductant as fuel.
  • other volatile metallic base metals or volatile base metal compounds in particular lead and lead compounds as well as silver and silver compounds, generated by smelting are evolved from the slag.
  • the process most preferably is operated to optimise this evolution, while S0 2 , CO, C0 2 , H 2 0 and H 2 generated during smelting also are evolved.
  • Free-oxygen containing gas such as oxygen, air or oxygen-enriched air
  • Free-oxygen containing gas such as oxygen, air or oxygen-enriched air
  • Free-oxygen containing gas is introduced into the reactor, above the bath, to oxidise the vapours of any metal elements and any sulphur vapours to produce oxide fume, as required by the invention, and to prevent reversion of the elemental zinc in the vapour to sulphide.
  • some sulphatizing of the zinc oxide can occur due to the reactions: 2S0 2 + O 2 ⁇ 2S0 3 (8),
  • the smelting process may be divided into two stages.
  • the first stage comprises a sulphur elimination stage under more oxidizing conditions and the second stage comprises a reduction stage wherein the zinc content of slag is reduced to a required final level by fuming.
  • the reducing conditions need not be as strong as with lead-containing feed material.
  • the strength of the reducing conditions, controlled by the level of reductant addition is one factor. Another factor is the temperature prevailing in the bath.
  • the temperature required usually is in excess of about 1200°C up to about 1350°C.
  • the first stage comprises a sulphur elimination stage under more oxidizing conditions, the second stage being a reduction stage under such conditions as to produce a lead silver bullion and a zinc rich slag, and the third stage comprising a reduction stage wherein the zinc content of slag is reduced to the required final level by fuming.
  • lead sulphide in the feed material, there is a tendency for it to evolve as lead sulphide.
  • some of the lead sulphide is reduced to the metal, evolved and oxidised above the bath, principally to lead sulphate but with a proportion of lead oxide formed.
  • the presence of lead sulphate in the crude fume is a problem, and, as detailed later herein, it therefore needs to be eliminated in a pre-treatment of the hydro-metallurgical stage.
  • lead can report in a lead-silver bullion, in which case a possibly minor proportion of the lead is evolved and oxidised.
  • the oxide fume and other evolved gases are extracted from the reactor of each stage via the off-take flue, and processed to recover a fume product.
  • the fume and gases are cooled and passed to a filtration zone, such as a baghouse, in which solid fume particles are retained and filtered from the gases.
  • the extracted gases after removal of the fume particles, can be further treated for recovery of S0 2 for use in acid production.
  • the recovered fume product comprising crude zinc oxide, or zinc oxide with lead sulphate and other base metal oxides such as lead oxide, also can contain silver oxide. Depending on fume collection temperatures some sulphation can occur as well as the collection of halides and other minor elements such as selenium, arsenic, and germanium.
  • the crude fume may also contain slag and feed carried over from the furnaces and such minor elements as cadmium and bismuth. This crude fume product then is subjected to hydrometallurgical refinement according to the invention, for recovery of refined zinc oxide.
  • the extracted gases contain at least one of selenium, arsenic and germanium
  • the gases With cooling of the gases to baghouse temperatures in excess of about 200°C, the elements substantially will report in the fume.
  • the fume product can be substantially reduced in the content of selenium, arsenic and germanium, thereby substantially reducing the requirements for purification with respect to these elements at a later stage. This is particularly beneficial where high purity zinc oxide is to be produced for use in electrowinning metallic zinc by a simplified electrolytic process according to the invention.
  • the smelting can proceed with a single lance or tuyere injecting oxygen-containing gas into the slag to give rise to closely adjacent regions in one of which zinc oxide is taken into the slag and the other in which the oxide is reduced to the metal and evolved. This is facilitated by use of lump coal which floats on the slag.
  • the reactor may utilize two lances or tuyeres operating in respective regions of the bath, one providing for less strongly reducing conditions than the other. The less strongly reducing conditions favouring the smelting to zinc oxide with its take-up in the slag, while the more strongly reducing conditions result in reduction of the zinc oxide to metal vapour.
  • these respective reducing conditions can be generated in successive periods of smelting, with some fume generated in the first period but most generated in the second period.
  • use of two reactors can be beneficial to allow continuous operation. Less strongly reducing conditions can be generated in one reactor, with the zinc oxide-containing slag passing, such as via a launder, to the second reactor in which more strongly reducing conditions are generated.
  • the recovered fume product is leached in a concentrated aqueous solution of ammonia and carbon dioxide, having a stoichiometric excess of ammonia.
  • the leach solution effectively of ammonia/ammonium carbonate (AAC), takes substantially all of the zinc oxide into solution, but only minor amounts of lead and silver which predominantly remain in the leach residue.
  • Oxides of some other metals present in the fume product also are taken into solution, while sulphate and halide ions also can be taken into solution.
  • Other metals can be present in the fume, depending on the volatility of these elements or their compounds present in the feed material subjected to smelting, such as cadmium, arsenic, germanium and selenium.
  • a degree of feed material and carbon dust or slag can carryover into the fume, and thus introduce iron, manganese, copper, nickel, and cobalt, as well as other gangue components such as silica, alumina, lime and magnesia into the leach solution.
  • iron, manganese, copper, nickel, and cobalt, as well as other gangue components such as silica, alumina, lime and magnesia.
  • copper, nickel, cobalt and cadmium are soluble in the leach solution together with traces of iron.
  • the fume product has a very fine particle size, in contrast to the major part of the ground feed proposed in the paper by Wendt, where about 80% of the feed is greater than 325 mesh.
  • the residence time for leaching is substantially reduced in the process of this invention.
  • a leaching time of 30 minutes or less is sufficient, compared with the several hours required by Peters for similarly fine EAF dust, and up to 48 hours indicated by Wendt for substantially full dissolution of zinc oxide from oxidised ores.
  • the solution preferably has about 10Og/l of carbon dioxide or higher, and may be substantially saturated with respect to carbon dioxide.
  • the stoichiometric excess of ammonia may be to provide a concentration of about 120g/l or higher of ammonia for a solution having 10Og/l or higher of carbon dioxide.
  • the excess of ammonia preferably is such that the ratio of moles of ammonia to moles of carbon dioxide is in excess of 3:1.
  • the rate of leaching, and the load of zinc oxide (and other oxides) able to be taken into solution by the leach solution increases with the concentration of ammonia and carbon dioxide.
  • a substantially saturated leach solution is able to take into solution in excess of 150g/l of zinc as the oxide, although this loading will vary with the extent of other oxides taken into solution as well as the extent of sulphate and halides taken into solution.
  • the pregnant solution resulting from leaching has a substantially depressed boiling point, necessitating leaching being conducted below about 60°C unless conducted under pressure.
  • heating is preferred, excessive heating results in an undesirable level of evolution of ammonia and carbon dioxide, limiting the capacity of the solution to dissolve zinc and risking premature precipitation of zinc from solution. Heating most preferably is not at a temperature greater than is necessary for energy minimization and for maximizing leaching rates.
  • the pregnant solution is filtered to separate the solution from the residue. Filtering also separates with the residue any carbon dust particles from reductant and/or fuel, as well as particles of slag and sulphide feed material, which are entrained in the fume and evolved gases extracted from the reactor in the first stage smelting operation.
  • the process of Peters results in a substantial zinc-rich residue remaining after the ammonia/carbon dioxide leach, due in part to zinc and possibly other base metals being present as ferrite, there is substantially less potential for ferrites to be a problem in the present invention.
  • the residue obtained with the process of the invention is low in zinc, and can substantially comprise a lead residue containing most of the silver content of the fume product, and insolubles such as slag and ore particles carried over with the fume.
  • lead contained in the leach residue essentially has been converted to lead carbonate and that the sulphate ions are transferred to the pregnant leach solution.
  • the lead carbonate contained in the leach residues can be readily reduced to metal by known methods, such as the use of a short rotary furnace, without any sulphur emissions. Any residual zinc values contained in the leach residues will report to the lead reduction furnace slag and may be recycled to the primary smelting stage for recovery into crude fume from the smelting operation.
  • the pregnant leach solution after filtering to remove solids, then is treated for recovery of zinc values.
  • the pregnant solution preferably is stripped of ammonia and some carbon dioxide to give rise to precipitation of zinc as its basic carbonate.
  • the precipitate is recovered, and then calcined to produce refined zinc oxide product.
  • the stripping of ammonia from the pregnant solution is effected by heating the solution to drive off the ammonia and thereby remove its stoichiometric excess.
  • the heating may be by application of external heat energy to a vessel containing the solution, to raise the solution to its boiling point.
  • heating is by direct steam injection in which ammonia is removed into the vapour stream.
  • the steam stripping preferably is by a multistage, counter-current operation in which the pregnant solution passes through a series of vessels, with fresh steam injected into the last of the vessels and passing in turn to the others.
  • the vessels may comprise plate columns, such as of the bubble cap form, or packed columns.
  • the vessels preferably are agitated tanks fitted with steam spargers.
  • the ammonia and carbon dioxide stripped from the solution during precipitation of the basic zinc carbonate, together with the carbon dioxide in gases from calcining the basic zinc carbonate, can be recovered in a condenser- absorber to regenerate leachate.
  • leaching reagents can be almost fully recycled and the major net process input is steam. Consequently integration with the smelting/fuming process is a major advantage of this invention since that part of the process will generate more than sufficient steam for the hydrometallurgical process.
  • Stripping of the pregnant leach solution enables substantially complete precipitation of zinc, substantially as its basic carbonate. This precipitation can be relatively selective, such that co-precipitation of some impurity elements such as halogens can be low.
  • the precipitate is recovered by filtering, washed on the filter and then is able to be calcined to provide high purity zinc oxide product.
  • the residual solution may be recycled, after cooling, to the condenser/absorber to reconstitute leachate solution or may be totally or partially discarded from the circuit to provide an outlet for halides and any other soluble elements such as sodium and potassium.
  • the calcining step may be undertaken in any suitable calcining vessel, while the calcining temperature can be at any economically suitable temperature for decomposition of the basic carbonate, such as from 250°C to 600°C.
  • the filtered pregnant solution may be treated with zinc dust to cause those metals lower than zinc in the electrochemical series such as lead, cadmium, copper, nickel and cobalt, to be displaced from solution.
  • a stoichiometric excess of zinc dust, relative to the metals to be displaced is preferred for this operation.
  • the efficiency of this cementation separation can be higher in the filtered pregnant solution than in acid solutions.
  • the smelting of the zinc sulphide-containing feed material to generate the crude fume product also generates sulphur dioxide which results in the formation of sulphates in the fume.
  • the sulphate formation is predominantly associated with lead present in the fume.
  • sulphur in that product usually is substantially present as sulphate.
  • Keegel disclosed in USP 5538532
  • a crude fume product is subjected to a second stage hydrometallurgical leach refinement, based on use of an AAC leach solution.
  • Lead in the crude fume product of the present invention is precipitated as carbonate during the AAC leach, whether the lead is present as oxide or sulphate.
  • lead sulphate or other sulphate it is found that sulphate ions will report in the pregnant AAC leach solution and to give rise to two undesirable results.
  • the sulphate accummulates in the AAC solution circuit and reduces the capacity of the AAC solution to dissolve zinc, by locking-up ammonia as ammonium sulphate. It also means that ammonia cannot be fully recovered by the steam stripping operation and further treatment of any discard solution as an outlet for impurities will be necessary to recover ammonia.
  • basic zinc sulphate is co-precipitated with the basic zinc carbonate.
  • basic zinc sulphate will not be decomposed and will be retained within and degrade the calcined zinc oxide product.
  • the zinc content of the zinc oxide product can, for example, be reduced from about 80% to about 72% as a result of the presence of sulphate.
  • the calcining temperature would need to be increased to about 750°C.
  • the product gases from such a calcining operation would contain sulphur dioxide which must be separated if carbon dioxide is to be recovered and recycled. The higher temperature calcination and presence of sulphur dioxide are complications which are best avoided by removing sulphate from the raw fume prior to AAC leaching.
  • the zinc content of the product zinc oxide still can be reduced from about 80 wt% to about 72 wt% due to co-precipitated zinc sulphate.
  • the problems presented by sulphate can be overcome by a suitable pre-treatment of the crude fume product.
  • the crude fume product is subjected to a suitable aqueous wash, prior to the AAC solution leach, by use of a carbonate wash solution.
  • the preferred wash solution is based on an alkali metal carbonate, preferably sodium carbonate.
  • an ammonium carbonate solution is possible.
  • a soda ash pre-wash treatment the following reactions occur: PbS0 4 +Na 2 C0 3 ⁇ Na 2 S0 4 + PbC0 3 (10),
  • the carbonate pre-wash is found to result in negligible amounts of zinc and lead being taken into the wash solution, such as less than 0.01wt% based on solution analysis. Importantly, the carbonate pre-wash is found also to serve to separate halides in a similar manner. Fume and washed fume analysis have shown 75 to 80% chlorine removal and 85 to 90% fluorine removal. The lower chlorine removal may be attributable to part of the chlorine being bound as highly insoluble silver chloride and will depend on relative levels of silver and chlorine present in the fume. Preliminary testwork has also indicated that the pre-wash can remove at least 30 to 50% of thallium present in the crude fume product. Also, some removal of arsenic, such as at least up to 10% can be achieved.
  • pre-washing As a result of pre-washing, higher zinc concentrations in the AAC leach solution are possible, a high grade zinc oxide at close to 80% zinc content can be produced, and recovery and recycle of ammonia is simplified. As indicated, pre-washing of the crude fume product with ammonium carbonate is possible. This is more compatible with the AAC solution leach, as the ammonia can be substantially fully recovered using the same condensation and absorption equipment, whereas some loss of alkali metal from its pre-wash will occur. However, there is a greater risk of zinc extraction during pre-wash with the use of an ammonium carbonate solution, while the efficiency of sulphate removal can be less than with an alkali metal carbonate such as soda ash.
  • an alkali metal carbonate such as soda ash
  • a pre-wash solution preferably has a minimum excess of carbonate consistent with minimising bleed losses.
  • soda ash as is preferred, an excess of less than 15% above the stoichiometric requirement, such as an excess of about 10%, can be suitable.
  • the resulting sodium sulphate solution may be treated with lime to precipitate gypsum and generate a caustic soda solution, while the latter solution readily can absorb carbon dioxide to reform the pre-wash soda ash solution.
  • Some bleed of solution will be necessary to provide an outlet for halides, although this will result in a loss of sodium requiring a compensating input of fresh soda ash.
  • the high purity zinc oxide product which can be produced by this invention presents a feed to the electrolytic zinc process which essentially is able to generate no residues. Consequently electrolytic refineries will be able to avoid major environmental difficulties associated with the disposal of iron residues such as jarosite or goethite.
  • Zinc rich oxide fumes have not been favoured as a feed to electrolytic zinc refineries largely because of the presence of halides and other minor impurities which are particularly deleterious to the electrolytic process.
  • the latter include volatile elements such as arsenic, antimony, germanium and selenium.
  • high levels of lead are not favoured as the resulting lead sulphate leach residues are not readily marketable as a raw material feed to lead smelters and consequently have relatively low value or the lead contaminates iron residues for acceptable disposal, if the fume is used for iron precipitation.
  • Advantages for the present invention are the separation of zinc and lead, the separation of halides, and the opportunity to separate the full range of deleterious minor elements in or before the second stage leaching operation to present a high purity zinc oxide for electrowinning of metallic zinc.
  • a substantial proportion of halides present in the initial feed material to the first stage smelting operation can be decomposed and eliminated in smelter gases.
  • Halides which may be captured in the crude fume product may be minimised by operating at elevated fume collection temperatures but, in any event, are able to be finally removed in both the pre-washing and the AAC leaching operation.
  • the process of this invention also recovers substantially all lead and silver values in a high grade carbonate residue from the fume leaching stage which may be readily reduced to a lead-silver bullion by known pyrometallurgical methods such as the use of a short rotary furnace.
  • This may be integrated into the total process whereby slag from lead reduction can be returned to the smelting stage to improve overall metal recoveries and provide for only one waste slag product.
  • This avoids or substantially reduces by-product lead-silver residue production at the zinc refinery.
  • These are usually low grade sulphate residues of marginal value, and represent a significant loss of value in the overall metal production process.
  • the high grade zinc oxide product may be agglomerated, such as by granulation, into a dust free form, for example, by using various binding agents such as zinc sulphate or sulphuric acid for transport to an electrolytic zinc refinery.
  • This invention will significantly reduce the quantity of feed materials to be transported between mine and electrolytic refinery which is normally located close to sources of low-cost electric power.
  • a preferred embodiment of this invention would be to incorporate a dedicated zinc electrowinning stage in the overall process which may be located adjacent to, or separate from, the oxide production facilities.
  • a dedicated electrowinning operation may be greatly simplified in comparison with conventional electrolytic zinc refineries. Firstly roasting and acid production facilities are not required. Complex leaching and iron separation processes are unnecessary and substantially all waste residues are eliminated. The processing steps may be confined to the use of facilities for a simple oxide leach in spent electrolyte, the electrolytic tankhouse, and cathode melting and ingot casting facilities.
  • the highly simplified leaching facilities with no or minimal purification requirements, will largely remove constraints on circulating solution volume and will permit operation of the electrolysis stage at higher zinc and lower acid concentrations than is normal practice. This will enable a significant improvement in deposition current efficiency with a reduction in electrical energy consumption as the single largest cost item for this process. Operation at higher temperature is also possible with the lower acid concentrations, and this will reduce cell voltages for zinc electrodeposition and simplify solution cooling requirements.
  • FIG. 3 A possible flow sheet is shown in Figure 3 which essentially consists of zinc oxide leaching, electrowinning, zinc cathode melting and ingot casting facilities.
  • the flowsheet also includes an optional zinc dust purification stage and provision for bleeding a small portion of circuit solution through a basic zinc sulphate precipitation step to serve as an outlet for soluble impurities such as halides.
  • An alternative approach would be to feed the bleed solution to an evaporative crystallizer to produce a zinc sulphate by-product for sale.
  • Dross from cathode melting may be returned to the primary leach after washing with sodium carbonate solution to separate chlorides, or it may be returned untreated to the smelting stage or to the subsequent crude fume pre- wash and leach.
  • the option selected will depend on requirements for halide balance control.
  • the overall impact of a feed of high purity zinc oxide to a dedicated electrolytic zinc refinery is a major reduction in capital and operating costs compared with conventional plants, improved metal quality and the elimination of environmental issues associated with the residues produced from conventional plants.
  • the process of the invention enables operation with substantial environmental benefits.
  • the products of the first stage smelting in addition to generating fume and smelting and combustion gases, can be operated to provide a slag that will not leach heavy metals in quantities that may be damaging to the environment.
  • the slag is low in environmentally hazardous base metals. If the operation is conducted at a minesite the slag may be produced in a form suitable for mine fill.
  • the slag is also free of sulphur, avoiding common problems with acid generation in mine tailings.
  • the overall operation of the process of the invention can limit the production of solid wastes essentially to that slag, substantially eliminating the production of iron and other leach residues from the electrolytic zinc process.
  • Other deleterious elements in minor quantities such as arsenic and thallium, can be produced in such a form that they can economically be rendered harmless, such as mixing with concrete that can be used to add strength to fill in underground mining processes.
  • the overall process of this invention particularly if applied to the direct processing of ore, can greatly enhance the overall recovery of valuable metals. Performance is independent of mineral grain size, mineral type and complexity which often severely affects the effectiveness of metal recovery in physical separation processes such as flotation.
  • the Keegel process of USP 5538532 proposes an integration of the production of zinc oxide product with electrolytic winning of zinc.
  • this is limited to the use of neutral leach residue by-products of zinc electrowinning by zinc smelting plants as feed material, in place of or additional to EAF dust. That is, the proposal is limited to the context of a conventional electrolytic process which utilises conventional calcine from a zinc smelting plant.
  • the zinc oxide product of the Keegel process is suitable for use as feed material, at least in part in place of conventional calcine in a conventional electrolytic process, let alone that it is possible to use the zinc oxide product as the sole source of zinc feed material for a dedicated electrolytic process for electrowinning of zinc.
  • the form of the present invention providing a three-stage operation producing metallic zinc by a dedicated electrolytic process has:
  • Figure 1 is an overall flow-chart for the process of the present invention
  • Figure 2 is a flow-chart of the second stage of the process of Figure 1 ;
  • FIG 3 is a flow-chart of the third stage of Figure 1.
  • the process depicted in Figure 1 is suited for the treatment of zinc sulphide ore of a grade suitable for smelting. However, the flow-chart readily is able to be used with other zinc sulphide-containing feed materials detailed herein.
  • the zinc sulphide ore After primary crushing, the zinc sulphide ore passes to a secondary crusher 10, if necessary via an ore bin 12.
  • the ore from crusher 12 passes to feeder bin 14, with the stockpile at 16 providing surge capacity.
  • bin 14 the ore is passed to a mixer 18, such as a pugmill or pelletizer, in which it is mixed with flux, such as silica from bin 20 and/or limestone from bin 22, and with coal from bin 24.
  • the ore mixture passes to a first top-submerged lancing reactor 26, in which it is melted.
  • Reactor 26 is a refractory-lined steel vessel 28 containing a suitable molten slag bath 30, preferably of an iron oxide containing slag.
  • the reactor 26 has a top-submerged lance 32 having its lower, discharge end submerged in the slag.
  • Oxygen from source 34 is able to be mixed with air and supplied to lance 32 under the action of blower 36.
  • Feed mixture from mixer 18 is supplied to reactor 26, and is charged to the bath 30. Under turbulent smelting conditions in reactor 26, the ore is smelted under mildly reducing conditions causing sulphur elimination as S0 2 , and the taking of zinc and lead values into the slag as the oxide.
  • a significant quantity of lead may be fumed as lead sulphide and a degree of reduction of the lead and zinc oxides can occur, with evolution of the respective metals and formation of lead sulphate.
  • Silver follows the lead values in the absence of formation of a matte phase during the smelting.
  • Resultant fume is oxidised above the slag, to provide crude oxide fume product.
  • This product may contain significant quantities of lead sulphate.
  • the fume product, S0 2 and other smelting gases are drawn from reactor 26 via its flue 38 and is passed to waste heat boiler 40.
  • Oxide containing slag is tapped from reactor 26, such as at required intervals or continuously, and passed to a second reactor 26a which is similar to reactor 26.
  • the slag is subjected to further, similar top-submerged smelting, but under more strongly reducing conditions.
  • Air or oxygen-enriched or nitrogen-enriched air is injected via lance 32a, into the slag bath 30a, with the supply of further air or oxygen-enriched air supplied to the reactor space above bath 30a.
  • Coal from bin 42 is passed to mill 44 in which it is reduced to a suitable size. From mill 44, the coal is entrained in nitrogen or air in injector 45 from which it passes to lance 32a for injection into bath 30a.
  • the zinc and lead oxides are reduced to the respective metal which are evolved from bath 30a.
  • the metal vapours are oxidised above bath 30a to provide further crude zinc oxide fume product containing lead sulphate which product, with residual S0 2 and smelting gases, is extracted via flue 38a and passed to waste heat boiler 40.
  • Slag with low levels of zinc and lead is tapped from reactor 26a, and passed to granulator 46 and then either passed to stockpile 48 or to ball mill 50 in which it is ground, for use such as mine fill.
  • the broken outline from slag granulation system 46, and reactors 26 and 26a indicates the hygiene system ducting designed to carry fugitive metal fumes from these points to the hygiene system baghouse 51.
  • a turbine/condenser system 53 preferably converts waste heat energy to electricity to power the oxygen plant, the air blowers and other items of electrical equipment used in the plant. Any excess electricity generated may be sold onto an electricity grid. However, it is possible to drive blowers directly from the steam produced in the smelters. In addition, some pass-out steam can be taken from the turbine for use in the ammonia/ammonium carbonate leach process described below.
  • Slag generated in furnace 62 can be recycled, such as to the feed for furnace 26.
  • the zinc oxide fume product from bin 54 may pass to a pre-treatment unit 55 before passing to leaching plant 58.
  • the zinc oxide product is washed with an aqueous carbonate solution, such as of sodium carbonate.
  • the washing causes metal values present in the oxide powder as sulphate to precipitate as the carbonate, while sulphate ions are taken into solution and discarded with the wash solution or are discarded as gypsum by precipitation with lime.
  • unit 55 has a vessel 55a in which the fume product is mixed with recycled and/or regenerated soda ash solution.
  • vessel 55a metal values present as sulphate are converted to carbonate which precipitates, while sulphate ions are taken into and remain in solution.
  • the mixture formed in vessel 55a is passed to thickener 55c from which overflow sulphate containing solution passes to vessel 55d and solids pass to filter 55e.
  • the solids, comprising the oxide component and other insolubles of the fume product, with the carbonate precipitate, are washed on filter 55e and then pass to leach tank 70.
  • the sulphate bearing solution is contacted with lime, to precipitate gypsum which is separated as solids passing thickener 55f and retained and washed on filter 55g.
  • the washed gypsum passes to a tailings dump or the like.
  • Wash liquid from filter 55g and overflow from thickener 55f are passed to a carbonation tower 55h, in which the soda ash solution is regenerated.
  • carbon dioxide is passed to tower 55h from kiln 55i in which limestone is heated to generate lime for use in vessel 55d.
  • the crude fume product received into bin 54 is passed, directly or via unit 55, to a leach tank 70 in which it is treated with a concentrated aqueous leach solution 72 of ammonium carbonate having a stoichiometric excess of ammonia.
  • the leach tank 70 may be a single tank or several tanks in series.
  • the fume/solution mixture is agitated by stirrer 74.
  • the zinc oxide is taken into solution as zinc ammonium carbonate complex, along with some impurity metals, such as cadmium, copper and nickel carried over with the fume in the smelting stage. After substantially all of the zinc has been complexed, the solution is passed to a thickener 76.
  • Solids withdrawn from the thickener pass to a filter 78.
  • Pregnant solution removed from thickener 76 and filtrate from filter 78 are recombined for further processing, while solids retained by filter 78 are washed by water added at 78a before being passed to the furnace 62 referred to in relation to Figure 1 , for recovery of lead-silver bullion.
  • an optional purification step can be provided between tank 70 (or the last tank 70 of the series) and thickener 76.
  • the solution from tank 70 is passed to a tank 75 in which it is treated with iron sulphate solution from line 75a and aeration by bottom injected air from line 75b.
  • This iron treatment results in precipitation of ferric hydroxide, with co- precipitation of metals such as arsenic and germanium.
  • the precipitate is removed by filter 78, and then passes with other filtered solids to furnace 62.
  • the recombined solution from thickener 76 and filter 78 preferably are purified in tank 80 by addition of zinc dust. Resultant cementation results in precipitation of metals lower than zinc in the electrochemical series, including cadmium, lead, copper, and nickel.
  • the solution then is filtered by filter 82, to remove the precipitate, following which the solution is treated in a series of preferably four vessels (shown as 84a and 84b) of a multistage steam stripping system.
  • the solution passes first to vessel 84a and then to vessel 84b.
  • Steam via line 85 is injected into the base of vessel 84b and then is passed from vessel 84b and similarly through vessel 84a.
  • the steam strips the excess ammonia from the solution, along with some carbon dioxide, with resultant precipitation of high purity basic zinc carbonate.
  • the mixed gases issuing from vessel 84a are passed to condenser/absorber vessel 86 for use in regeneration of the initial leach solution for tank 70.
  • the stripped solution from vessel 84b is passed to the thickener 88, and the settled solids to filter 90 for recovery of the basic zinc carbonate.
  • the carbonate is washed on filter 90 by water from line 90a and then is passed to an indirectly fuel fired calcining kiln 92 to generate zinc oxide, with recovery of carbon dioxide.
  • the recovered carbon dioxide is recycled for leach solution regeneration in condenser/absorber vessel 86, with make-up C0 2 and water also being provided to vessel 86.
  • Leach solution recovered from thickener 88 and filter 90 also are recycled, via cooling tower 94, to condenser/absorber vessel 86.
  • filtrate from filter 90 and/or overflow from thickener 88 may be discarded to provide a bleed for halides as sodium or potassium salts.
  • the zinc oxide calcine product from kiln 92 passes to installation 60, as shown in Figure 1 and in detail in Figure 3. However, if required, some or all of the high-purity zinc oxide can be recovered for direct sale.
  • the zinc oxide from calcining kiln 92 passes via a fume hopper 102 to a leach tank 100, which may be one of a series of such tanks.
  • tank 100 the zinc oxide is digested in part of spent electrolyte recycled from electrolytic cells 104 used to produce cathode zinc.
  • the latter is suitable for melting in furnace 106 for casting at 108 of high purity zinc ingots.
  • the cathode zinc can be used to provide zinc dust, at least in a quantity useable in the electrolytic refining circuit.
  • Part of the recycled spent electrolyte passes to cooling system 110 and is returned to cells 104.
  • the bulk of the leach solution from tank 100 comprising a zinc sulphate solution on digestion of the zinc oxide and any recycled zinc dust, is passed to a clarifier 112.
  • the overflow solution from clarifier 112 passes to cells 104, if required, via an optional purification step.
  • the latter includes passing the zinc sulphate solution to agitated tank 114 in which it is treated with zinc dust to precipitate minor impurity residuals such as copper and cadmium.
  • the solution then passes to filter 116, and then to cells 104 via cooling system 110.
  • the precipitate from filter 116 can comprise a copper-cadmium cement which can be recovered for sale.
  • the small quantity of residue from clarifier 112 is subjected to a strong acid leach in tank 115, using part of the recycled electrolyte and fresh acid if required for sulphate make-up.
  • the acid solution and resultant residue are passed to filter 120, to separate the residue from the solution which is passed back to tank 100.
  • the residue which will be a relatively minor quantity, can be returned to furnace 26a, but preferably to furnace 26, both shown in Figure 1.
  • a minor part of the solution of tank 100, resulting from digestion of zinc oxide, is passed to a two stage impurity level control system which, while requiring only a single vessel 124, has its stages represented as stage 1 and stage 2.
  • stage 1 the minor part of the solution is treated with zinc oxide from hopper 102, to precipitate basic zinc sulphate.
  • lime is added via line 124a to precipitate residual amounts of zinc from solution as additional basic zinc sulphate.
  • the solution then is passed to filter 126 which separates the precipitate.
  • the liquid from filter 126 containing soluble impurities such as halides is discarded via line 126a, while the basic zinc sulphate precipitate is recycled to vessel 100.
  • the pilot runs consisted of 11 oxidation runs using low-grade zinc sulphide ore obtained from the 4-Level of the Lady Loretta mine in North West Queensland, three oxidation runs using high-grade ore from the mine's 6-Level, two oxidation runs using a bulk concentrate, and four reduction runs. Three of the reduction runs were undertaken using slag produced from the oxidation of the 4- Level ore. One of these was a batch reduction, in which ore was added to a slag under oxidising conditions to raise the zinc level to 5 to 6%, and then reducing conditions were imposed. This oxidation and reduction cycle was repeated during the same test. The remaining reduction run used all the slag produced from the oxidation of the bulk concentrate and 6-Level ore. Table 1 gives average analyses for the oxidation runs; while Table 2 presents the average composition of the slag used in the reduction runs.
  • Table 3 The average compositions of the slag and fume generated from the oxidation runs using 4-Level ore are presented in Table 3 which also gives the average composition of the ore.
  • Table 3 includes the average composition of the slag and fume produced during the reduction runs using slag generated by oxidation smelting of 4-Level ore.
  • the reduction runs were undertaken using only part of the slag generated during the oxidation runs. Consequently, the feed to the reduction process was not identical in composition to the average slag produced during the oxidation tests.
  • Table 3 - Average composition of 4-Level ore feed, slag and fume
  • the average operating conditions for a test run during smelting of the slag were as set out in Table 12.
  • the composition of the coal was the same as detailed in Tables 9 to 11 , while other analyses of materials were as set out in Tables 13 to 15. Difficulties may be encountered when attempting a mass balance on the data of Table 15 through analytical error, sampling error, and hold-up of same elements in the fume collection system. This hold-up can result in some material from previous smelting trials reporting to the fume of the current trial, and some of the material from the current trial reporting to the fume of future trials.
  • Zinc Dust Cementation to separate elements more electropositive than zinc, such as lead, copper, cadmium, nickel and cobalt.
  • the prewash also served to separate halides.
  • the indications from fume and washed fume analyses indicated 75 to 80% chlorine removal and 85 to 90% fluorine removal.
  • Tests using an ammonium carbonate prewash showed a significantly lower sulphate removal efficiency by about 30%, and up to 3.5% of the zinc was extracted.
  • the first series of tests also showed similar zinc and lead extractions at 99.2% and 0.4% respectively, giving a high degree of separation of the two major components.
  • Copper extraction was found to be low at around 27.5%. This was believed to be due the fact that the copper was present in carryover slag or feed material from fuming operations and may be as a sulphide in a matte phase of the slag rather than as a copper oxide.
  • Halides were efficiently extracted during the leaching operation which may imply that improved filter cake washing of the pretreated fume could improve the separation of halides at that point.
  • the standard lead refining operations of copper drossing, softening and desilvering could be used to produce high grade lead.
  • Other elements are likely to have been predominantly contained in reduction furnace slag (eg. iron, nickel, cobalt, germanium) or in fume (eg. cadmium).
  • Table 27 gives the results of the first and second series ferrous sulphate addition tests. Table 27 - Results of ferrous sulphate addition.
  • Zinc dust cementation had minimal impact on the level of arsenic in solution, but had a variable effect on the removal of germanium, reducing concentration down to 1.5 mg/l in one case. Thus, cementation may be a useful additional step to ensure low levels of germanium in the final zinc oxide product.
  • Zinc was precipitated from solution as a basic zinc carbonate by removal of ammonia and carbon dioxide from solution by steam stripping. In practice this would be conducted in a continuous multistage system (4 or 5 stages). However in testwork, this step was conducted using a batch approach, in which steam was continuously passed through a fixed volume of pregnant solution until all the zinc has been precipitated. Two steam rates were used at 10% and 20% of the solution volume in litres as kg/h of injected steam. Solution samples taken at various times were analysed and provide useful information on equilibrium solution compositions.
  • Tests showed minimal difference in zinc oxide purity at calcination temperature between 300, 450 and 600 °C.
  • the zinc oxide analyses for the same material were 71.4%, 73.5% and 73.8% respectively. Consequently it was considered that calcination at 300 °C was adequate.
  • the residual sulphate was high at 2.7% (8.1 % SO4) due to the absence of pretreatment.
  • Final zinc oxide product can be broadly grouped into material produced with and without purification by zinc cementation.
  • Table 31 gives average analyses for these two groups produced from the later calcination tests, including pretreatment for sulphate removal:
  • the zinc oxide product was clearly high grade with adequate control of most elements of concern to subsequent metallurgical operations, with the possible exception of arsenic and germanium. If the zinc oxide was required as feed to a dedicated and simplified electrowinning zinc plant without solution purification, then additional measures to control these elements would be necessary. Higher temperature collection of smelter fume will enhance the separation of arsenic from fume. Iron precipitation is also likely to be effective for control of these elements as well as silica, but will require optimization of iron addition and precipitation procedures. Alternatively neutral leaching of the zinc oxide with some iron addition in the electrolytic zinc circuit, should also be effective in control of these impurities and may be preferable since it is a well established procedure.
  • any electrowinning plant should at least include a zinc purification stage to act as a safeguard against accidental impurity input and to enable poor solution quality to be recovered. If this is the case, then it would seem preferable to carry out all such cementation purification at the electrolytic plant where the zinc dust is readily available and the bulk of the contained zinc is automatically recycled.
  • Halides are the primary elements of concern in the use of fume product in a conventional electrolytic zinc plant, and these are adequately controlled by the fume separation process, meeting target levels of 50 ppm for chlorine and 5 ppm for fluorine. Soluble elements such as magnesium, potassium and manganese are also well controlled by this process.

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Abstract

On récupère des valeurs de zinc depuis une charge de matière première contenant du sulfure de zinc au moyen d'un procédé consistant à fondre la matière première en présence d'un agent de réduction carboné afin de réduire pratiquement le sulfure de zinc en zinc métallique et de vaporiser le zinc, ce qui produit des gaz de combustion contenant de la vapeur de zinc et du dioxyde de soufre; à mettre en contact les gaz de combustion avec un gaz contenant de l'oxygène libre afin d'oxyder le zinc métallique et de produire une fumée brute riche en oxyde de zinc dans lesdits gaz de combustion. On sépare la fumée brute des gaz de combustion, on la recueille et on la mélange avec une solution concentrée contenant de l'ammoniac aqueux et du dioxyde de carbone et présentant un excédent stoechiométrique d'ammoniac, de manière à produire une liqueur d'attaque contenant du zinc et à laisser un résidu de solides. On sépare la liqueur d'attaque du résidu et on supprime l'ammoniac de la liqueur d'attaque afin de précipiter le carbonate de zinc basique et de laisser une liqueur d'attaque exempte d'ammoniac. On sépare cette dernière du carbonate de zinc basique et on calcine celui-ci afin d'obtenir un produit purifié d'oxyde de zinc.
PCT/AU1998/000094 1997-02-17 1998-02-17 Raffinage de minerais contenant du sulfure de zinc WO1998036102A1 (fr)

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AU59743/98A AU5974398A (en) 1997-02-17 1998-02-17 Refining zinc sulphide ores

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AUPO5148A AUPO514897A0 (en) 1997-02-17 1997-02-17 Processing of zinc-containing feed material
AUPO5148 1997-02-17
AUPO7711A AUPO771197A0 (en) 1997-07-07 1997-07-07 Improved processing of zinc-containing feed material
AUPO7711 1997-07-07

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US7815708B2 (en) 2003-09-29 2010-10-19 Umicore Process and apparatus for recovery of non-ferrous metals from zinc residues
WO2012040829A1 (fr) * 2010-09-30 2012-04-05 Yava Technologies Inc. Récupération sélective du zinc dans le lixiviat à partir d'un dépôt de minerai sulfuré composite, de résidus, de minerai concassé ou de boue de mine
CN103146927A (zh) * 2013-04-02 2013-06-12 长沙有色冶金设计研究院有限公司 一种用氧浸渣混合处理锌浸出渣的冶炼方法
CN103182344A (zh) * 2011-12-30 2013-07-03 北京有色金属研究总院 一种处理高泥铅锌氧硫混合矿的选矿组合工艺
US8641921B2 (en) 2007-03-19 2014-02-04 Umicore Room temperature single phase Li insertion/extraction material for use in Li-based battery
US20140205519A1 (en) * 2010-11-23 2014-07-24 Metallic Waste Solutions Pty Ltd Process for recovering zinc and/or zinc oxide ii
CN104419826A (zh) * 2013-08-26 2015-03-18 四川宏达股份有限公司 氨浸氧化锌制电积锌的方法
CN105296769A (zh) * 2015-11-27 2016-02-03 中南大学 一种湿法炼锌的方法
CN105316493A (zh) * 2015-11-27 2016-02-10 中南大学 一种湿法炼锌工艺
CN105483394A (zh) * 2015-11-27 2016-04-13 中南大学 一种促进锌精矿氧化的方法
US10400306B2 (en) 2014-05-12 2019-09-03 Summit Mining International Inc. Brine leaching process for recovering valuable metals from oxide materials
CN112176199A (zh) * 2020-09-27 2021-01-05 衡阳市大宇锌业有限公司 一种从炼锌废渣中高效提纯氧化锌的工艺
CN112708777A (zh) * 2020-12-23 2021-04-27 江西睿锋环保有限公司 一种由含锌废料回收硫酸锌的方法
CN113145615A (zh) * 2021-04-23 2021-07-23 中国恩菲工程技术有限公司 含氟磷废渣的处理方法和处理装置
EP4012058A1 (fr) * 2020-12-09 2022-06-15 S.A. Lhoist Recherche Et Developpement Procédé pyro-métallurgique dans un four rotatif
CN115896466A (zh) * 2023-01-05 2023-04-04 昆明理工大学 一种处理复杂有色金属二次物料的方法
CN116173703A (zh) * 2022-12-07 2023-05-30 西南科技大学 电解锰渣矿化co2协同固化金属离子的方法和矿化产物
CN117663743A (zh) * 2023-12-14 2024-03-08 高邑县永昌锌业有限责任公司 一种氧化锌冶炼炉及冶炼方法
CN114182097B (zh) * 2021-12-08 2024-03-12 西安建筑科技大学 一种含铜锌氧化物与硫化锌协同资源化的方法

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CA1200701A (fr) * 1982-10-27 1986-02-18 Kornel R.V. Szarmes Dehalogenation des matieres oxydes a teneur de zinc
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US7597740B2 (en) 2003-07-04 2009-10-06 Umicore Recovery of non-ferrous metals from zinc residues
WO2005005674A1 (fr) * 2003-07-04 2005-01-20 Umicore Recuperation de metaux non ferreux a partir de residus de zinc
US8557174B2 (en) 2003-09-29 2013-10-15 Umicore Process and apparatus for recovery of non-ferrous metals from zinc residues
US7815708B2 (en) 2003-09-29 2010-10-19 Umicore Process and apparatus for recovery of non-ferrous metals from zinc residues
US8641921B2 (en) 2007-03-19 2014-02-04 Umicore Room temperature single phase Li insertion/extraction material for use in Li-based battery
WO2008135574A1 (fr) * 2007-05-07 2008-11-13 Zincox Resources Plc Procede de production de poudre d'oxyde de zinc et poudre ainsi obtenue
US7939037B2 (en) 2007-05-07 2011-05-10 Zincox Resources Plc Method of producing zinc oxide powder
AU2011308038B2 (en) * 2010-09-30 2015-10-01 Yava Technologies Inc. Selective leach recovery of zinc from a composite sulphide ore deposit, tailings, crushed ore or mine sludge
US8961911B2 (en) 2010-09-30 2015-02-24 Yava Technologies Inc. Selective leach recovery of zinc from a composite sulphide ore deposit, tailings, crushed ore or mine sludge
CN103237909A (zh) * 2010-09-30 2013-08-07 亚瓦技术有限公司 从复合硫化矿沉积物、尾矿、碎矿石或者矿泥中选择性沥滤回收锌
WO2012040829A1 (fr) * 2010-09-30 2012-04-05 Yava Technologies Inc. Récupération sélective du zinc dans le lixiviat à partir d'un dépôt de minerai sulfuré composite, de résidus, de minerai concassé ou de boue de mine
CN103237909B (zh) * 2010-09-30 2015-09-30 亚瓦技术有限公司 从复合硫化矿沉积物、尾矿、碎矿石或者矿泥中选择性沥滤回收锌
USRE47673E1 (en) 2010-11-23 2019-10-29 Steel Dynamics Investments Llc Process for recovering zinc and/or zinc oxide II
US20140205519A1 (en) * 2010-11-23 2014-07-24 Metallic Waste Solutions Pty Ltd Process for recovering zinc and/or zinc oxide ii
US9284624B2 (en) * 2010-11-23 2016-03-15 Metallic Waste Solutions Pty Ltd. Process for recovering zinc and/or zinc oxide II
CN103182344A (zh) * 2011-12-30 2013-07-03 北京有色金属研究总院 一种处理高泥铅锌氧硫混合矿的选矿组合工艺
CN103146927B (zh) * 2013-04-02 2014-11-05 长沙有色冶金设计研究院有限公司 一种用氧浸渣混合处理锌浸出渣的冶炼方法
CN103146927A (zh) * 2013-04-02 2013-06-12 长沙有色冶金设计研究院有限公司 一种用氧浸渣混合处理锌浸出渣的冶炼方法
CN104419826A (zh) * 2013-08-26 2015-03-18 四川宏达股份有限公司 氨浸氧化锌制电积锌的方法
CN104419826B (zh) * 2013-08-26 2017-03-29 四川宏达股份有限公司 氨浸氧化锌制电积锌的方法
US10400306B2 (en) 2014-05-12 2019-09-03 Summit Mining International Inc. Brine leaching process for recovering valuable metals from oxide materials
CN105483394A (zh) * 2015-11-27 2016-04-13 中南大学 一种促进锌精矿氧化的方法
CN105316493A (zh) * 2015-11-27 2016-02-10 中南大学 一种湿法炼锌工艺
CN105296769A (zh) * 2015-11-27 2016-02-03 中南大学 一种湿法炼锌的方法
CN112176199A (zh) * 2020-09-27 2021-01-05 衡阳市大宇锌业有限公司 一种从炼锌废渣中高效提纯氧化锌的工艺
EP4012058A1 (fr) * 2020-12-09 2022-06-15 S.A. Lhoist Recherche Et Developpement Procédé pyro-métallurgique dans un four rotatif
CN112708777A (zh) * 2020-12-23 2021-04-27 江西睿锋环保有限公司 一种由含锌废料回收硫酸锌的方法
CN113145615A (zh) * 2021-04-23 2021-07-23 中国恩菲工程技术有限公司 含氟磷废渣的处理方法和处理装置
CN114182097B (zh) * 2021-12-08 2024-03-12 西安建筑科技大学 一种含铜锌氧化物与硫化锌协同资源化的方法
CN116173703A (zh) * 2022-12-07 2023-05-30 西南科技大学 电解锰渣矿化co2协同固化金属离子的方法和矿化产物
CN115896466A (zh) * 2023-01-05 2023-04-04 昆明理工大学 一种处理复杂有色金属二次物料的方法
CN117663743A (zh) * 2023-12-14 2024-03-08 高邑县永昌锌业有限责任公司 一种氧化锌冶炼炉及冶炼方法

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