WO1991008317A1 - Production de zinc par fusion - Google Patents

Production de zinc par fusion Download PDF

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Publication number
WO1991008317A1
WO1991008317A1 PCT/AU1990/000578 AU9000578W WO9108317A1 WO 1991008317 A1 WO1991008317 A1 WO 1991008317A1 AU 9000578 W AU9000578 W AU 9000578W WO 9108317 A1 WO9108317 A1 WO 9108317A1
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WO
WIPO (PCT)
Prior art keywords
zinc
bath
vapour
gas stream
ferric
Prior art date
Application number
PCT/AU1990/000578
Other languages
English (en)
Inventor
Roger Leo Player
Steven Paul Matthew
Original Assignee
Mount Isa Mines Limited
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
Application filed by Mount Isa Mines Limited filed Critical Mount Isa Mines Limited
Publication of WO1991008317A1 publication Critical patent/WO1991008317A1/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
    • C22B19/04—Obtaining zinc by distilling
    • 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

  • the present invention relates to a method of producing zinc from a composition containing zinc sulphide, for example from a zinc concentrate or a bulk concentrate containing both zinc sulphide and lead sulphide.
  • the electrolytic zinc process consumes a large quantity of electrical energy and produces a toxic heavy metal waste (jarosite sludge) as a by-product.
  • the by-product presents a problem in waste disposal and it would be advantageous to produce zinc from zinc concentrate without the toxic by-product.
  • a flash smelting process has been proposed by Davey in which a bone dry finely divided zinc sulphide concentrate and a fuel are mixed with, and subjected to combustion in, gaseous oxygen (reaction I).
  • reaction II The chemical equilibrium for reaction I requires that very high temperatures are employed in order to transform a high percentage of the zinc from zinc sulphide to zinc vapour.
  • Davey proposed to combine oxygen with a diluent in a selected ratio as a means for obtaining an acceptable yield at temperatures as low as 1500°K.
  • Shock chilling of the reaction gas stream was proposed as a means of suppressing reversion (reaction II).
  • the zinc oxide bearing slag is transferred to a second furnace where the zinc oxide in the bath is subjected to strongly reducing conditions (reaction IV) to produce a fume of metallic zinc which is collected in a splash condenser.
  • Zinc sulphide is injected with an oxidizing gas by lance into the first of two zones in a furnace to produce zinc oxide in a slag and remove sulphur dioxide from a first flue.
  • the slag layer containing the dissolved zinc oxide is then transferred to a second zone wherein the zinc oxide is reduced to zinc metal which is removed as a vapour from a second flue.
  • the zones are separated by a fluid cooled wall which extends into the slag layer and divides the gas space above the bath.
  • the two zone smelting apparatus and method creates, by its own design, problems associated with circulation and mixing of the slag between the two zones. It is essential to the efficiency of the two zone smelting apparatus and process that proper circulation and mixing of the slag occurs between the two zones as proper circulation and mixing ensures transfer of zinc oxide from the first zone to the second zone. It is not clear how the control of this circulation would be achieved.
  • Zinc is usually found in nature as a sulphide. The ore is commonly concentrated prior to further treatment. Zinc is often found in combination with lead and an increasing amount of concentrate is becoming available which contains both zinc and lead in a form which renders difficult the separation of zinc and lead from the ore.
  • An object of the present invention is to provide a process for the recovery of zinc from a composition containing zinc sulphide which avoids or at least ameliorates disadvantages of the prior art.
  • An object of a preferred embodiment is to provide an improved process whereby zinc and lead can be recovered from a bulk concentrate as elemental metals.
  • the present invention consists in a method for producing zinc from a composition containing zinc sulphide comprising the steps of:
  • the composition may be a concentrate containing both zinc sulphide with lead sulphide. In that case both sulphides dissolve in the bath and are oxidized by the ferric oxide. Both zinc metal vapour and lead metal vapour leave the bath with the sulphur dioxide gas. Some lead may remain as a molten lead layer below the slag depending upon the initial lead 'content of the composition.
  • the zinc and, if present lead, metal vapours may be collected directly as metals by rapid quenching or by suitable means.
  • the metal vapours may be oxidized in the gas stream for example by addition of an oxygen containing gas such as air to precipitate a fine metal oxide powder which may be separated from the gas stream by conventional methods from which the valuable metals may be recovered by conventional electrochemical means.
  • the ferric to ferrous iron ratio in the bath is maintained within the range of 0.01 to 0.10 at 1200°C or 0.10 to 0.25 at 1300°C and by linear interpolation at intermediate temperatures by adjusting the rate of injection of oxygen and or of addition of carbon.
  • the bath is vigorously mixed (by the injection of the oxygen bearing gases at below the surface of the slag bath) to ensure a substantially uniform distribution of ferric iron in the slag.
  • Preferred embodiments of the invention permit zinc containing concentrate (or a combined zinc and lead bulk concentrate) to be fed into a molten slag bath continuously and enable a high yield of zinc, or lead and zinc, to be obtained at a commercially viable temperature.
  • the method eliminates the sintering step of previous chemical reducing agent methods and, advantageously, coke is not required.
  • the process can be conducted efficiently in a single furnace of relatively simple construction. In preferred embodiments of the invention the process is conducted at from 1150°C to 1250°C to reduce energy costs and minimize refractory wear.
  • Figure 1 is a schematic diagram showing an apparatus for conducting the method of the invention.
  • Figure 2 is a schematic diagram showing another apparatus for conducting the method of the invention.
  • An important aspect of the present invention is the control of the oxygen activity of a one step oxidation of ZnS to give metallic zinc vapour and sulphur dioxide as follows:
  • the oxygen activity for this reaction can be calculated from chemical thermodynamics and is less than 10 ⁇ atmospheres.
  • reaction (V) can be readily controlled in a molten ferrous silicate slag bath, the slag bath having three functions:
  • the slag acts as a buffer to control the oxygen activity, via the redox reaction:
  • the slag also acts as an oxygen transfer medium from the point of gaseous oxygen injection into the melt to the point of addition of zinc sulphide containing concentrate to the melt, so isolating the sulphides from direct contact with oxygen gas.
  • the slag fluxes any high melting point phases that may be formed e.g. ZnO or.
  • the ferric to ferrous iron ratio of the slag is determined by, and indicative of, the balance between reactions VIII to XI. That is to say the ratio is determined by the rate of supply of oxygen to the slag forming ferric oxide, and the rate of reaction of the ferric oxide with the concentrate plus the rate of transfer of oxygen from the slag (via the ferric iron) to the reductant.
  • the ratio is determined by the rate of supply of oxygen to the slag forming ferric oxide, and the rate of reaction of the ferric oxide with the concentrate plus the rate of transfer of oxygen from the slag (via the ferric iron) to the reductant.
  • the reaction is controlled to produce the metal vapour in high yield direct from the bath.
  • lump coal floating on the surface of the slag is used as a reductant to control the ferric to ferrous iron ratio, it has been found that, for coal with 65% fixed carbon, it is desirable to use a minimum specific coal addition rate
  • FIG. 1 there is shown schematically by way of example a furnace or reactor 1 containing a molten slag bath 2 containing ferric and ferrous iron.
  • the slag in the present examples is a ferrous silicate slag but can be a calcium ferrite or other' ferrite slag.
  • Zinc concentrate is added continuously at feed port 3 together with lump coal. The concentrate dissolves in bath 2 while the lumps of coal tend to float on the surface of the bath in a layer 4.
  • Air or oxygen enriched air is injected to below the surface of the slag bath 2 via a submerged smelting lance 5.
  • Introduction of the oxygen below the surface of the slag minimizes direct reaction of oxygen with the carbonaceous material floating in layer 4 on top of slag 2 and vigorously agitates the bath.
  • Fuel is added to the reactor by either injecting a suitable fuel (gas, oil or pulverised coal) together with the injected oxygen via lance 5 or by adding lump coal via feed port 3 to the top surface of the slag bath as described above.
  • a suitable fuel gas, oil or pulverised coal
  • the process is not sensitive to the fuel rate and excessive amounts of fuel do not result in a loss of yield.
  • the process can be operated with acceptable yield at bath temperatures in the preferred range of 1150°C to 1300°C, more preferably 1200°C to 1250°C. This temperature range is just above the melting point of the ferrous silicate slag but below the temperature at which the usual magnesite chrome refractories lose strength thus minimising furnace refractory wear.
  • the gases are shock cooled to avoid or minimize reversion.
  • Shock cooling is achieved for example by bringing the hot gases into direct contact and heat exchange relationship with a fluidized bed 8 of cooled particles for example silica or zinc particles.
  • Bed 8 is selected to be of a ' large thermal mass in comparison with that of the gas stream in heat exchange relationship with the bed.
  • Bed 8 may be cooled by water injection or by other cooling means not illustrated in Figure 1 to below the condensation temperature and preferably to below the melting point of zinc in a rapid quenching action.
  • the metal vapour is thus rapidly condensed and captured in the fluidized bed as zinc metal and the sulphur dioxide is separated from the solids, leaving via offtake 15. A portion of the fluidized bed containing captured zinc is drawn off from seal 13.
  • FIG. 2 With reference to Figure 2 there is shown schematically another system for shock cooling the zinc vapour. Parts of Figure 2 corresponding in function to parts illustrated in Figure 1 are identified by corresponding numerals.
  • Figure 2 differs from that of Figure 1 in that the fluidized bed 8 is recirculated between a first chamber 9 and a second chamber 10.
  • Particles 8 enter the furnace offgas stream at port 11 of chamber 9 and are carried by the gas stream into chamber 10 where the gas is cyclonically separated from the solids and removed at 12.
  • the bed 8 particles are recovered, a proportion removed via seal 13 and the balance recirculated to chamber 9 via seal 14.
  • the particles are cooled prior to entry into the gas stream at opening 11 by direct addition of coolant and/or by heat transfer devices in chamber 9 or. in chamber 10 or both.
  • the metal vapour of the gas stream is rapidly cooled, from at or above the bath temperature (above 1,140°C) to below the metal condensation temperature and preferably to below the metal melting point (419°C in the case of zinc, 327°C in the case of lead). Moreover, the cooling takes place rapidly, typically in less than 100 milliseconds and preferably in 10 milliseconds or less.
  • the fluidized bed of figures 1 and 2 has a particle loading in excess of 200 Kg/m 3, preferably greater than 400 Kg/m 3 and as much as 1600 Kg/m3 or more.
  • the thermal mass (mass multiplied by specific heat) of particles in the fluidized bed is such that if the bed were uncooled the rate of heat temperature rise of the bed in heat exchange relationship with the furnace gas would be less than 100°C/sec. and desirably less than 20°C/sec.
  • the particles of the bed have an average grain size of less than 2 mm in diameter and preferably 90% of the particles of the fluidized bed have a grain size of less than 0.5 mm in diameter.
  • Oxygen concentration in the vapour stream at the entrance to the fluidized bed can be controlled by introducing inert gases to the furnace, or by maintaining a positive pressure to prevent air ingress, or by adjusting feeds to consume a predetermined ratio of oxygen, or by use of after burners located between the bath and the fluidized bed.
  • the temperature of the vapour entering the fluidized bed may be controlled at an elevated level by use of after burners or other heaters and/or by furnace design to minimize heat loss prior to quenching.
  • the metal may be recovered by oxidation.
  • air is admitted to the space above the bath or to the furnace offgas stream so as to oxidize the zinc and lead metal vapours, whereby to precipitate the zinc as zinc oxide and lead as lead oxide.
  • the metal oxides may then be separated from the gas stream by means of cyclones, bag house filters or the like.
  • the preferred range for the ratio is a function of temperature. At 1200°C the ratio should be between 0.01 and 0.10. At 1300°C the ratio should be within 0.10 and 0.25. Intermediate ratios apply at intermediate temperatures by linear interpolation. Provided the ratio is within a predetermined range having regard to the bath temperature the process is within control. If necessary, the rate of injection of oxygen via the lance is increased or decreased to control the ferric to ferrous iron ratio in the slag bath or the rate of carbon addition is adjusted. Alternatively, the "magnetite" content or "magnetic coefficient" of the slag can be measured and used as an approximate control.
  • ferric to ferrous ratio in the slag falls below the preferred range, the sulphur is inadequately oxidized and thus inadequately removed from the bath.
  • the ferric to ferrous ratio exceeds the preferred range, the zinc oxide content of the slag becomes too high to maintain the fluidity of the slag at the temperatures indicated.
  • the carbonaceous material added to the surface of the bath is low volatile coal having a high percentage of fixed carbon. It is believed that char resulting from fixed carbon in the coal floats on top of the surface of the slag and reacts with the slag while the volatiles leave the system substantially unreacted.
  • Zinc containing concentrate is added to the surface of the bath, together with sufficient additional oxygen in stoichiometric quantities to oxidise all the sulphur in the concentrate to sulphur dioxide and the iron to iron oxides while maintaining the ferric to ferrous oxide ratio referred to earlier.
  • the metallic zinc vapour is rapidly cooled so that zinc metal may be recovered in the presence of sulphur dioxide or is oxidized to facilitate separation.
  • the furnace is therefore operated in a cyclic manner with a 2 to 3 hour total cycle time, involving a continuous smelting stage followed by batch stripping and tapping.
  • concentrate, fluxes and coal are fed continuously to the furnace.
  • Smelting air is also added to contr.ol the ferric to ferrous ratio in the slag, as well as fuel oil and combustion air as required to maintain furnace temperatures of 1200 to 1250°C.
  • zinc in slag concentrations of 10 to 20% are maintained and approximately 85% to 95% of the zinc in the concentrate is fumed off.
  • the furnace After smelting for from 1 1/2 hours to 2 1/2 hours the furnace is filled with slag and at this point the concentrate, fluxes and smelting air additions are stopped. Coal addition continues, together with fuel oil and combustion air injection to maintain 1200 to 1250°C furnace temperatures, and the slag is batch reduced to fume the residual zinc from the slag to a level suitable for disposal. Finally the furnace is tapped leaving sufficient slag in which to resume smelting and smelting is then resumed.
  • a zinc concentrate feed mixture containing 66% coal and 6% limestone flux by weight of dry concentrate was prepared by mixing 200 Kg of zinc concentrate containing 6.8% water, 123 Kg of coal containing 0.2% free water and 11.2 Kg of limestone containing 0.2% water with 8 Kg of additional water and pelletising.
  • the furnace Prior to the run, the furnace was heated to 1200°C and 200 Kg of ferrous silicate was added and melted over 4 hours. During this time, the furnace temperature was raised to 1300°C and thermally equilibrated. Air and oil were injected into the bath through a 40 mm diameter submerged lance. The ferrous silicate slag was then partially tapped out, leaving a 150 mm deep layer of slag at the bottom of the furnace as a starting bath. The above pelletised feed mix was then smelted at a rate of 165 Kg/hr, this being equivalent to a feed rate of 90 Kg/hr dry concentrate.
  • a zinc concentrate feed mixture containing 50% coal and 6.5% limestone flux by weight of dry concentrate (assays given in Table 1) was prepared by mixing 200 Kg of zinc concentrate containing 7.3% water, 93 Kg of coal containing 0.2% free water and 12 Kg of limestone containing 0.2% water with 8 Kg of additional water and pelletising. This run followed on immediately from another run which was only partially tapped to leave a 150 mm deep slag layer as the starting bath.
  • the above pelletised feed mix was smelted at a rate of 100 Kg/hr, this being equivalent to a feed rate of 60 Kg/hr concentrate.
  • the operating temperature was 1200°C.
  • Smelting air was injected into the bath through a 40 mm diameter lance at a rate of 145 Nm /hr, this being equivalent to 3 mole oxygen per mole of zinc in the feed.
  • An additional 92 NmVhr of combustion air and 8.7 Kg/hr of fuel oil were also injected down the lance and burnt to maintain the operating temperature at 1200°C.
  • the reaction chemistry was monitored by periodic slag samples taken from the furnace, giving the results shown in Table 3.
  • a zinc concentrate feed mixture containing 32% coal and 5.4% limestone flux by weight of dry concentrate was prepared by mixing 200 Kg of zinc concentrate containing 7.3% water, 60 Kg of coal containing 0.2% free water and 10 Kg of limestone containing 0.2% water with 8 Kg of additional water and pelletising.
  • the furnace Prior to the run, the furnace was thermally equilibrated and a 150 mm deep ferrous silicate slag starting bath was established in the same manner as in Example 1 above except that the operating temperature was 1200°C.
  • the above pelletised feed mix was then smelted at a rate of 90 Kg/hr, this being equivalent to a feedrate of 60 Kg/hr dry concentrate. Smelting air was injected into the bath through a 40 mm diameter lance at a rate of 169 NmVhr, this being equivalent to 3.5 mole oxygen per mole of zinc in the feed.
  • the smelting air rate was 4900 Nm /hr, which was equivalent to 3.0 mole oxygen per mole of zinc in the feed.
  • coal rate was kept constant at 700 Kg/hr throughout the trials to maintain a specific coal rate of 250 Kg per hour per square meter of bath surface at both 2 and 3 tph zinc concentrate.
  • the silica and limestone flux rates were controlled to give a SiO_/Fe ratio in the slag of 0.6 and Si0 2 /Ca0 ratio of 3.

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  • Engineering & Computer Science (AREA)
  • Chemical & Material Sciences (AREA)
  • Manufacturing & Machinery (AREA)
  • Materials Engineering (AREA)
  • Mechanical Engineering (AREA)
  • Metallurgy (AREA)
  • Organic Chemistry (AREA)
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Abstract

On peut récupérer le zinc contenu dans des concentrés de sulphure de zinc en établissant un bain de laitier en fusion à une température comprise entre 1150 °C et 1300 °C contenant des oxydes ferreux et ferriques, en dissolvant le sulphure de zinc dans le bain, en oxydant le sulphure de zinc par réaction avec les oxydes ferriques, afin de produire de la vapeur de zinc et du dioxyde de soufre, et en régulant le rapport entre les oxydes ferriques et les oxydes ferreux dans le bain. Le zinc est séparé du dioxyde de soufre par une trempe rapide destinée à empêcher tout phénomène de réversion ou par oxydation de la vapeur de zinc. Le rapport Fe(II)/Fe(III) peut être régulé dans des limites prédéterminées par injection d'oxygène et/ou par addition de charbon.
PCT/AU1990/000578 1989-12-05 1990-12-04 Production de zinc par fusion WO1991008317A1 (fr)

Applications Claiming Priority (2)

Application Number Priority Date Filing Date Title
AUPJ772589 1989-12-05
AUPJ7725 1989-12-05

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WO1991008317A1 true WO1991008317A1 (fr) 1991-06-13

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AU (1) AU632650B2 (fr)
IE (1) IE904367A1 (fr)
WO (1) WO1991008317A1 (fr)

Cited By (4)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US5443614A (en) * 1994-07-28 1995-08-22 Noranda, Inc. Direct smelting or zinc concentrates and residues
WO1998036102A1 (fr) * 1997-02-17 1998-08-20 Buka Technologies Pty. Ltd. Raffinage de minerais contenant du sulfure de zinc
WO2014046593A1 (fr) * 2012-09-21 2014-03-27 Valeas Recycling Ab Évaporation induite par plasma
WO2015024073A1 (fr) * 2013-08-19 2015-02-26 Glencore Technology Pty Limited Traitement de matières solides à haute teneur en soufre

Families Citing this family (1)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
EP0557312B1 (fr) * 1990-11-14 1997-03-05 Minproc Technology Pty. Ltd. Distillation par sulfidisation directe de zinc

Citations (6)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
GB893875A (en) * 1958-04-24 1962-04-18 Mansfield Huetten Kom Wilhelm Process for the volatilization of metals or metallic compounds
AU5605873A (en) * 1972-06-26 1974-11-28 Borax Consolidated Limited Improvements in or relating to zinc and lead smelting
AU5950580A (en) * 1979-06-26 1981-01-08 Commonwealth Scientific And Industrial Research Organisation Producing zinc from sulphide concentrate
AU5551386A (en) * 1985-04-03 1986-10-09 Cra Services Limited Smelting process
AU6731187A (en) * 1985-11-19 1987-06-02 Ausmelt Pty Ltd Oxidation-reduction smelting of zn ores
AU7918687A (en) * 1986-08-27 1988-03-24 Commonwealth Scientific And Industrial Research Organisation Recovery of lead, zinc and other metals from ores concentrates or residues

Family Cites Families (1)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
AU634309B2 (en) * 1989-11-08 1993-02-18 Commonwealth Scientific And Industrial Research Organisation Condensation of metal vapours in a fluidized bed

Patent Citations (6)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
GB893875A (en) * 1958-04-24 1962-04-18 Mansfield Huetten Kom Wilhelm Process for the volatilization of metals or metallic compounds
AU5605873A (en) * 1972-06-26 1974-11-28 Borax Consolidated Limited Improvements in or relating to zinc and lead smelting
AU5950580A (en) * 1979-06-26 1981-01-08 Commonwealth Scientific And Industrial Research Organisation Producing zinc from sulphide concentrate
AU5551386A (en) * 1985-04-03 1986-10-09 Cra Services Limited Smelting process
AU6731187A (en) * 1985-11-19 1987-06-02 Ausmelt Pty Ltd Oxidation-reduction smelting of zn ores
AU7918687A (en) * 1986-08-27 1988-03-24 Commonwealth Scientific And Industrial Research Organisation Recovery of lead, zinc and other metals from ores concentrates or residues

Cited By (9)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US5443614A (en) * 1994-07-28 1995-08-22 Noranda, Inc. Direct smelting or zinc concentrates and residues
WO1998036102A1 (fr) * 1997-02-17 1998-08-20 Buka Technologies Pty. Ltd. Raffinage de minerais contenant du sulfure de zinc
WO2014046593A1 (fr) * 2012-09-21 2014-03-27 Valeas Recycling Ab Évaporation induite par plasma
KR20150076168A (ko) * 2012-09-21 2015-07-06 발＇이아스 리싸이클링 솔루션즈 에이비 플라스마 유도 퓨밍
JP2015535883A (ja) * 2012-09-21 2015-12-17 ヴァル’イース・リサイクリング・ソリューションズ・アクチボラグ プラズマ誘起蒸散法
US10006100B2 (en) 2012-09-21 2018-06-26 Val'eas Recycling Solutions Ab Plasma induced fuming
KR102176989B1 (ko) * 2012-09-21 2020-11-10 발＇이아스 리싸이클링 솔루션즈 에이비 플라스마 유도 퓨밍
WO2015024073A1 (fr) * 2013-08-19 2015-02-26 Glencore Technology Pty Limited Traitement de matières solides à haute teneur en soufre
US9650694B2 (en) 2013-08-19 2017-05-16 Glencore Technology Pty Limited Treatment of high sulphur solids

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IE904367A1 (en) 1991-06-05
AU6893491A (en) 1991-06-26
AU632650B2 (en) 1993-01-07

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