Patent Application: US-91431301-A

Abstract:
a process for treating a metal sulphide concentrate which includes the steps of : a ) roasting the concentrate to reduce the sulphide content of the concentrate , to a negligible value and b ) melting the concentrate , under reducing conditions , in an electrically stabilized open - arc furnace , in particular a dc arc furnace .

Description:
zinc concentrates of the kind encountered in the gamsberg deposit have a manganese level which is up to 10 times higher than normal . this high manganese level causes problems and additional costs when recovering the zinc , after leaching , in a conventional electrowinning plant for the electrowinning route much research has been carried out on means of removing the manganese from the electrolyte , or on electrolytic processes which enhance the production of mno 2 at the anode in a zinc cell . the former technique is expensive , and the latter approach , which is directed to the production of high quality electrolytic manganese oxide , appears to be problematical . it is advantageous to remove sulphur from metal sulphide concentrates before smelting . for example existing pgm and base metal pyrometallurgical processes have a number of limitations , particularly in the converting stage . it is difficult to achieve environmentally acceptable levels of sulphur capture , especially in view of the problem of fugitive emissions from feirce - smith converters . converting is a batch process which has inherent scheduling problems , losses and spillages from the crane transport of ladles , and high labour costs . another consideration with conventional furnaces is that there is a limit to the amount of pgm - containing ug2 concentrate that can be treated . a problem with chromite in ug2 concentrate is that it cannot easily be solubilised in slag during normal smelting . a spinel forms , builds up in the furnace , and needs to be dug out frequently . the invention is described hereinafter firstly with reference to a generalized treatment process , as exemplified in fig1 and thereafter with reference to three particular forms of the process shown respectively in fig2 and 5 . fig1 of the accompanying drawings illustrates a generalized process for the treatment of metal sulphide wherein a concentrate 10 of the metal sulphide or metals sulphide is fed to a fluidized bed roaster 12 which , preferably , produces a steady stream of high strength so 2 bearing gas 14 which can be used as feedstock for example in a sulphuric acid plant . this is not essential to the process though , for gas from the reactor could alternatively be subjected to gas scrubbing and neutralization . the calcined product from the roaster is fed to a dc arc furnace 16 together with a reductant 18 which , for example , is in a form of coke . in the fluidized bed reactor 12 the sulphur content of the concentrate is reduced substantially , to approximately 10 % by mass in the case of high roasting , or to approximately less than 1 % of the initial value , or even lower , in the case of dead - roasting . the elimination of sulphur from the concentrate results in the valuable metals being collected in an alloy 20 which is produced by the furnace 16 , rather than as a matte . alloys have a much greater pgm collection efficiency than matte . the furnace also produces a slag 22 which is depleted in metal values . the nature of the process thereafter depends on the nature of the concentrate which is being treated . if the concentrate is a zinc - bearing concentrate such as zinc sulphide then the zinc is reduced to metal in the dc arc furnace and fumed off in a gas stream or vapour 24 which is mainly zinc and carbon monoxide . the gases are led directly to a lead - splash condenser 26 for absorption , or condensing , and subsequent recovery as a product 28 . for the treatment of copper , nickel or cobalt sulphide , or pgm sulphide , the alloy produced by the dc arc furnace may be atomized ( step 30 ) and then subjected to a hydrometallurgical recovery process 32 to produce a product 34 . alternatively the alloy is first fed to a converter 36 such as a peirce - smith converter and then to the atomizer 30 and through to the hydrometallurgical process 32 . the converter slag 37 may be returned to the dc arc furnace 16 . the process 32 may be of any appropriate type and a particularly suitable process 32 , which is intended to fall within the scope of the invention , is described hereinafter with reference to fig5 . fig2 is a particular example of the process of fig1 for the recovery of zinc from a high manganese ore such as for example the gamsberg lead - zinc deposit . it is assumed that the mining of ore from the gamsberg deposit , followed by grinding and flotation , yields a concentrate 10 which contains about 48 % zinc , 29 % sulphur , from 4 % to 5 % manganese as oxide , and 5 % moisture . the concentrate 10 is fed from a suitable store to fluidized bed roasters 12 where the sulphur content of the concentrate is reduced to approximately 0 , 75 %. the gases from the roasters are cooled in a waste heat boiler , cleaned in cyclones , subjected to electrostatic precipitation , and are then passed to a sulphuric acid plant 14 . it is assumed that the final exhaust gases can be discharged to atmosphere without the need for scrubbing out the last traces of sulphur dioxide . the calcine which contains about 58 % zinc as oxide is fed with dry coke 18 and a small amount of lime to the dc arc furnace 16 , with a sealed freeboard . the calcine may optionally be agglomerated in a step 38 before being fed to the dc arc furnace . this step does however involve additional capital and operating costs . in the dc arc furnace the zinc oxide is reduced to metal and fumed in a gas stream 40 which principally contains zinc and carbon monoxide . these gases are led directly to the lead splash condenser 26 where the zinc and any lead are removed from the gas stream by absorbing or condensing these metals in a curtain of lead droplets . the gases 42 exiting the condenser are burnt in a combustion chamber , cooled in a waste heat boiler and are cleaned in a bag filter 44 before being exhausted to atmosphere . the maximum concentration of sulphur dioxide in the exhaust gases is estimated to be less than 100 parts per million which does not pose an environmental problem . the dust collected in the bag filter , which consists mainly of zinc oxide , is washed with water to remove any halides before being returned to the roasters . slag 46 which is produced by the furnace is granulated before being removed to a waste storage dump . a small amount of metal 48 produced in the furnace is periodically tapped from the furnace and is run into rough moulds . dross 50 from the lead splash condenser 26 is collected and batch treated in a small furnace to separate out lead . the dry dross is then recycled to the dc arc furnace . impure zinc 52 from the lead splash condenser is transferred by ladle to a zinc distillation plant holding furnace . a plant of this type requires a constant feed and consequently provision is made for casting ingots or reheating ingots to balance the flow from the condenser so that the requirements of a zinc distillation plant 54 are met . the distillation plant produces shg ( special high grade ) zinc 56 , an impure lead and cadmium - zinc alloy 58 which is treated according to requirement , and a small amount of hard zinc 60 , in the form of a zn — pb — fe — cu alloy , which is recycled to the dc arc furnace . fig3 illustrates the steps of a particular form of the process shown in fig1 used for the treatment of pgm and base metal concentrates 10 . the feed material 10 is dead - roasted in a fluidized bed roaster 12 which effectively reduces the sulphur content of the concentrate to zero . this limits later sulphur emissions from the concentrate . the reactor 12 produces a steady stream of so 2 - bearing gas which is used as feedstock for a sulphuric acid plant 14 . it is to be noted though that the so 2 - bearing gas may alternatively be subjected to gas scrubbing and neutralization instead of being used as feedstock . the concentrate , after being roasted , is fed to a dc arc furnace 16 to produce an alloy 20 , and a slag 22 which is depleted in metal values and which is discardable . it may possibly be beneficial to add a base metal collector 62 , as is indicated in fig1 and 3 in dotted outline , to the fluidized bed reactor 12 . for example by adding nickel ( e . g . in the form of nickel sulphate ) or copper ( e . g . in the form of copper sulphate ) to the fluidized - bed roaster , along with the concentrate , a greater quantity of nickel or copper ( as oxide ) is established in the furnace feed , and this decreases the requirement to reduce a large quantity of iron which would otherwise be required to produce sufficient alloy for effective collection of the valuable metals . any base metal oxide , sulphate or sulphide , which is compatible with the process and which is in a fine form which can react with the feed , could be used as a collector . the alloy 20 may directly be passed to an atomizer 64 which makes the alloy suitable for subsequent leaching in a hydrometallurgical step 32 , for the recovery of metal values . alternatively the alloy is fed to a converter 36 which removes most of the iron from the alloy in oxide form . the resulting slag 37 may be returned to the dc arc furnace 16 . thereafter the alloy is atomized and subjected to the aforementioned leaching step to enable the metal values to be recovered . disadvantages or problems which are overcome or reduced are the difficulty of hot ladle transportation from the furnaces to the converters which create scheduling problems in the converter aisle , losses due to spillages , skull formation in the ladles , high labour and maintenance costs as well as pollution problems . advantages include the elimination or reduction of the cr - spinel problem in the furnace , the tolerance for higher cr - levels in the feed , with resulting higher pgm recoveries , the avoidance of the matte breakout problem , a lowering of energy consumption and , in one variation of the process , the elimination of the converter . the process is also able to accommodate a wide range of feed materials , up to 100 % ug2 ( in the case of pgm concentrates ), with a higher chromite content . this provides significant advantages with respect to pgm recovery in the mining and concentrating operations . fig4 illustrates a general flowsheet of high roast - reduction smelting of base - metal sulphide concentrate which incorporates a dc arc furnace . it is understood that there are a number of variations of this flowsheet . concentrate slurry 66 is continuously fed to fluid bed roasters 68 . the degree of concentrate sulphur elimination ( degree of roast ) may vary from 70 % all the way to 100 % ( i . e . dead - roast ). roaster off - gas 70 is cooled , cleaned , and directed to an acid plant for so 2 fixation . calcine 72 premixed with flux and coal 74 is smelted in a dc arc furnace 76 . the smelting furnace would produce discard quality slag 78 . the grade , iron and sulphur content of the alloy or matte 80 produced will depend on the degree of roast and the ratio of reductant to calcine . the furnace off - gas 82 would be of a low volume and high co concentration . a high grade , low sulphur , highly metallized matte from the furnace would require minimal treatment in a pyrometallurgical converting process 84 . the following is a list of variations of the general flowsheet in fig4 : the dc arc furnace can easily accommodate the high degree of oxide reduction required . the dc arc furnace also allows for the production of more alloy as more reductant is added to the furnace . fig5 illustrates a particular form of the technique of fig1 wherein , in a continuous process , sulphur is removed by the dead - roasting of pgm - containing matte followed by a smelting stage under reducing conditions . a six - in - line or a three - electrode furnace is used for the production of pgm - containing matte from concentrate . the green furnace matte is then granulated and milled , or water atomized , to produce a feed 10 which is fed to the fluidized bed roaster 12 . a steady stream of gas containing so 2 is fed to a sulphuric acid plant 14 . the roasted material is then subjected to a two - stage reduction smelting process which makes use of a first furnace 86 and a second furnace 16 , which is a dc arc furnace . the first furnace may be of any appropriate type and may for example be a dc arc furnace . the first furnace allows for the settling , under slightly reducing conditions , of some of the copper and nickel as an alloy 88 . a large fraction of the pgms partitions to this alloy which is then treated in an atomizer 90 before being directed to a hydrometallurgical process 32 . slag 92 from the furnace 86 , and a reductant 18 are fed to the second furnace 16 which operates under highly reducing conditions in order to remove virtually all of the nickel and pgms contained in the slag , along with most of the cobalt , to produce an iron - based alloy 94 which may also contain some copper . this alloy passes to a converter 95 and is then atomized in a step 96 in preparation for leaching in a hydrometallurgical process 32 . slag from the converter may be returned to the furnace 16 . slag 98 produced by the dc arc furnace is discardable and is sufficiently devoid of valuable metals that it can either be discarded or used in applications such as road construction or shot blasting . again it should be noted that a base metal collector 62 could be used , as has been described hereinbefore , to decrease the capability which would otherwise be called for , of the furnace to reduce a large quantity of iron . a simpler and potentially more cost effective process to the aforementioned two - furnace process involves the single - stage smelting of the roasted furnace matte in a dc arc furnace . this is essentially the technique which is shown in fig3 where the feed material 10 is a green furnace matte . the simpler process requires fewer process units but it has the disadvantage of capturing all base metals and some iron together with the pgms . a number of examples of the invention have been described hereinbefore . in each case the concentrate , which optionally is in green furnace matte form , is dead - roasted and thereafter is smelted under reducing conditions . the essentially complete removal of sulphur means that later sulphur emissions are limited . the spine problem ( in the smelting of high chromite containing pgm concentrates ) is reduced and discardable slags are produced . the feed to the dc arc furnace is pre - heated in the fluidized bed reactor . it is to be noted that a dc arc furnace is well suited to handling fines . the alloy which is produced by the dc arc furnace can be water atomized ( step 64 ). although it is possible to make use of a converter it is also possible to eliminate the need for a converter and in this way the likelihood of spillages is reduced and scheduling problems are also reduced . the nature of the hydrometallurgical process 32 depends on the major elements of interest in the alloys which are produced by means of any of the aforegoing pyrometallurgical techniques . typically these elements are iron , nickel , copper , cobalt and pgms . the hydrometallurgical processing of these alloys depends on case - specific factors . the unit operations that could be applied in the treatment of the alloy include ambient - pressure leaching , pressure leaching , precipitation , solvent extraction , electrowinning and crystallisation . the principles of the individual unit operations are generally known in the industry . they may be used in a wide variety of combinations , and a person skilled in the practise of hydrometallurgy will be able to devise an appropriate circuit for any specific case . each of the examples shown in fig6 to 10 embodies a general approach , and is not meant to limit the applicable hydrometallurgical option for processing the alloy in question . process steps for the removal of impurity elements such as selenium are omitted for the sake of brevity , but it should be understood that they would be incorporated as necessary , as known to those skilled in hydrometallurgy . where acid addition is shown , it may be either fresh acid or acid recycled from a metal recovery stage such as electrowinning . the copper solvent extraction and electrowinning stages are as conventionally practised in the industry . the iron precipitation can be done at elevated pressure and temperature , such that hematite is precipitated and acid is regenerated for recycle . it could also be done by means of neutralisation with an appropriate alkali ( an example is limestone , but a number of others exist ) such that goethite , jarosite , basic ferric sulphate or other similar compound is precipitated . the solutions containing cobalt and / or nickel ( shown as proceeding to ni / co separation and recovery ) would be treated in the same way as is done in conventional base metal refining , for the recovery of the cobalt and / or nickel . this could entail the precipitation of cobalt ( iii ) hydroxide or the solvent extraction of cobalt , and the electrowinning of nickel and / or cobalt . alternatively , it could entail the crystallisation of mixed or separate cobalt and / or nickel salts , or the precipitation of hydroxides , sulphide or carbonates . ion exchange could also be used in some cases . the examples in fig1 to 5 have been described with reference to the use of a dc arc furnace . this is non - limiting for , as has been indicated hereinbefore , a dc arc furnace is a particular form of a stabilised open arc furnace . although use of a dc arc furnace is preferred and the operation of a furnace of this type is well established it is possible to make use of an ac open arc furnace which has been stabilised , using suitable control techniques , to confine the arc in the furnace so that it extends vertically from an overhead electrode and does not diverge to side walls of the furnace . this example applies to those situations in which the alloy contains pgms and valuable base metals . in the first step , the iron and base metals are dissolved , leaving a residue that comprises a pgm concentrate that can proceed to a pgm refinery . oxidative leaching would normally be used , but non - oxidative leaching may also be used ( in which case the air / oxygen supply to the leach would be omitted ). elevated temperature and pressure may be used , either alone or in combination with ambient - pressure leaching . in some cases , elevated pressure may not be necessary . the resulting solution could be passed directly to a copper solvent extraction and electrowinning sequence for copper recovery , or it could be passed to an iron - precipitation stage and then to the copper solvent extraction and electrowinning stage . the raffinate from copper solvent extraction would be neutralised and any remaining iron precipitated , to produce a solution containing mainly nickel and / or cobalt , from which these metals can be recovered . this example applies when pgms are not present in the alloy , for example when the alloy comes from the reduction of converter slag , for the recovery of base metals . often , this will entail cobalt as the major metal value . in this case , the oxidative leach would be operated so as to solubilise the copper , nickel and cobalt while rejecting all or most of the iron as a hematite or goethite residue . after copper solvent extraction and electrowinning , the cobaltinickel solution would proceed to conventional treatment for the recovery of nickel and / or cobalt . in this example an atomised alloy from a smelting plant is fed to an atmospheric leach , where the bulk of the iron and nickel is leached in the presence of oxygen and sulphuric acid , at a temperature between 30 ° c . and 95 ° c . the copper from the electrowinning spent recycle is cemented in the atmospheric leach and assists in the leaching of the iron and nickel . conditions for the atmospheric leach were optimised during a laboratory scale test programme . a pilot - scale ( 100 l ) batch atmospheric leach , based on the optimised conditions , was performed on 5 . 5 kg of atomised alloy . the performance of the pilot - scale batch leach is summarised below . the leach residence time is set according to the material to limit the leaching of copper while still maintaining high iron and nickel recoveries . leach residence times of between 5 and 10 hours are required . the optimum leach residence time was exceeded in the test above , such that some copper leaching was observed . the residue from the atmospheric leach is then subjected to a two - stadium pressure leach to remove all the copper and the residual iron and nickel in the presence of sulphuric acid . the pressure leach was tested in laboratory - scale batch autoclaves . the pressure leach operates at temperatures between 110 ° c . and 170 ° c . with no oxygen in the first stadium and 0 . 1 to 6 bar oxygen in the second stadium . residence times of 60 to 180 minutes are required in the first stadium and 5 to 60 minutes in the second stadium . the pressure leach residue contains high levels of pgms and is suitable for further processing . pgm loss to the leach liquor can be minimised to less than 5 % while achieving a pgm concentrate of greater than 60 % precious metals . the composition ( mass %) of the pgm concentrate produced from pressure leaching of the atmospheric leach residue is shown below . the solution from the atmospheric and pressure leach is treated in pressure vessels to oxidise the iron and precipitate it as hematite . acid is produced during the hematite precipitation , and the bulk of the solution following the hematite precipitation from the atmospheric leach liquor is recycled back to the leach . batch hematite precipitation tests were performed on a laboratory scale to test the removal of iron from the atmospheric leach liquor . the pressure oxidation operates at temperatures between 140 ° c . and 200 ° c ., with oxygen overpressures of 1 to 10 bar . the performance of the laboratory - scale batch pressure oxidation is summarised below . a bleed stream is taken from the solution following hematite precipitation . this bleed is neutralised with lime and any residual iron is precipitated as goethite . the neutral solution is then crystallised to produce nickel sulphate . the gypsum / goethite residue is disposed of . the copper sulphate solution from the pressure leach is also treated in a pressure vessel to remove iron as hematite . selenium is removed in an additional unit operation to produce a purified solution from which copper is electrowon . the spent copper electrolyte is recycled back to the atmosphere and pressure leaches to utilise the acid generated during electrowinning . the copper in the solution is cemented as copper metal and aids in the leaching of the iron and nickel . in this variation the first leach ( at ambient and / or elevated pressure ) is operated so as to dissolve only nickel and cobalt . the iron and copper are dissolved and then re - precipitated as goethite and antlerite , respectively . this requires an alloy that is reactive enough to raise the ph of the leach solution sufficiently for the copper to hydrolyse and precipitate as antlerite . the solution proceeds to nickel / cobalt recovery . the goethite / antlerite is re - leached to selectively dissolve the antlerite without co - dissolving more than a small part of the goethite . the copper - rich solution is passed to copper electrowinning , and the spent electrolyte returned to dissolve more antlerite . the remaining goethite is then redissolved under more aggressive conditions , leaving the pgms as a concentrate that is sent to a pgm refinery . the solution leaving the goethite dissolution stage is passed to a high - temperature autoclave to precipitate the iron as hematite and regenerate acid for recycle to the goethite dissolution stage . this is similar to example 4 , but in this case no pgms are present , therefore the goethite re - dissolution stage is omitted because it is not needed . using a dc arc furnace with an internal diameter of 1 . 0 m , connected to a 5 . 6 mva power supply , approximately 26 tons of calcine (‘ dead - roasted ’ concentrate ) was processed over a period of 9 days , during which time 83 slag taps were carried out . the metallurgical data presented here is a weighted - averaged summary of the operation during 22 taps under the preferred conditions for producing good metallurgical performance , i . e . just over a quarter of the campaign . these taps cover a wide range of operating conditions , but the overall average is considered representative of the steady operation of the furnace during this campaign . the anthracite addition was approximately 12 % based on the mass of calcine fed . ( actual additions were 12 . 7 %, 11 . 6 %, and 12 . 0 % during the three periods summarised here .) metal was produced at a rate of 250 kg per ton of calcine fed . typical operating conditions included feedrates of around 220 kg / h of calcine , power levels around 300 kw ( including losses of about 150 kw ), voltages between 175 and 250 v , and total power fluxes around 400 to 500 kw / m 2 . the energy requirement of the process was 760 kwh / t of calcine , excluding losses from the furnace . the recoveries of the valuable elements were calculated based on the following analyses . the rest of the compositions and flowrates were calculated on the basis of these numbers . the actual recoveries obtained on this campaign were calculated using both the typical and the best analyses obtained . approximately 30 tons of pgm - bearing sulphide ore concentrate was treated in a fluidized - bed reactor , then smelted in a pilot - scale dc arc furnace . the resulting alloy was refined using a blowing operation , then treated hydrometallurgically to produce a high - grade pgm concentrate . the fluidized bed was operated at approximately 1000 ° c ., and the concentrate was fed at about 140 kg / h . gas velocities of about 0 . 4 m / s were used . the residence time was rather low , at approximately 20 seconds per pass . most of the material underwent two passes through the reactor , with a small quantity passing through three times . the sulphur level decreased from 4 . 55 % s to 0 . 5 % after the first pass ( 96 % elimination of s ), and to 0 . 24 % after the second pass ( 98 % elimination ), and to 0 . 13 % s after the third pass . during roasting , the impurities were diminished as follows : smelting was carried out in a pilot - scale dc arc furnace . 24 tons of ( mostly double pass ) dead - roasted concentrate ( including 1 ton of triple - roasted material ) was processed in a week - long campaign . the furnace was operated at a power level of 300 to 500 kw , which translates to a power flux of 290 to 480 kw / m 2 . the average operating temperature was 1650 ° c . calcine was fed to the furnace at feedrates of 200 to 300 kg / h , and approximately 5 % coke addition was used . no additional fluxes were added . an energy requirement of 650 kwh / t of calcine was required ( neglecting energy losses from the furnace shell ). ( obviously in a full - scale plant operating with hot feeding of calcine to the furnace , this figure would be less .) the process was operated consistently with less than 1 g / t pgm in slag , and values as low as 0 . 3 g / t in the slag were demonstrated . the average pgm loss to the slag over the entire campaign was 2 . 9 g / t . the analyses of the original concentrate , roasted concentrate , and slag are shown below ( mass %). impurity removal overall ( including roasting and smelting ) is shown below , as a percentage of the amount originally present in the unroasted concentrate . approximately 109 kg of alloy per ton of roasted concentrate was produced in the furnace . over the campaign , about 2 . 6 tons of alloy was produced in total . most of the alloy was tapped in two large batches . ( the first alloy tap was diluted somewhat by the initial metal heel in the furnace .) shown below is the composition of the alloy , together with the composition of the alloy produced in a laboratory - scale preliminary test ( all in mass %). also shown is the composition of the refined alloy produced by blowing the molten alloy with air , as discussed below . the alloys produced during the furnace campaign had the following ranges of composition . the alloy with the worst composition ( i . e . from the 836 kg batch ) was selected to demonstrate the downstream process on the most conservative basis . in order to lower the quantities of carbon and silicon ( and chromium ) prior to leaching , it was necessary to blow air into the molten alloy ( using a top - blown rotary converter , to simulate the operation of the proposed ladle furnace to be used for this operation ). the composition of the resulting refined alloy is shown in the table above . this alloy was water - atomized to a particle size less than 100 μm . the atomized alloy was then used for the leaching tests . after hydrometallurgical processing , a final pgm concentrate was produced with the composition below ( mass %). small - scale laboratory tests were carried out on pgm - containing furnace matte . the matte was either milled as a solid , or water - atomized from the liquid state , then dead - roasted in either a fluidized bed or a rotary kiln . ( no difference was found between the roasting behaviour of the milled and the atomized matte .) it was shown that matte can be roasted to extremely low levels of sulphur . the dead - roasted matte was then smelted in a two - stage process . the first - stage of smelting produces a small quantity of copper - nickel alloy that contains almost no iron or sulphur . the pgms essentially all report to the copper - nickel alloy . the first - stage alloy has a pgm content around 2 %. ( this can be upgraded by leaching the cu and ni to produce a pgm concentrate .) the slag from the first stage is then smelted , using a carbonaceous reductant , to produce a second alloy containing most of the remaining base metals , as well as the residual precious metals . small - scale fluidized - bed roasting tests were carried out on 20 g samples in a 25 mm silica tube fluidized bed . successful roasting was achieved using a particle size range of 250 - 300 μm , and a temperature of 800 - 850 ° c . the sulphur content of the matte decreased from 28 . 7 % to 0 . 03 % in 3 hours . good results were also obtained after 1 . 5 hours at 950 ° c . temperatures above 900 ° c . are recommended for complete desulphurization . in order to provide larger samples for smelting tests further roasting was carried out in a laboratory - scale rotary kiln . a 20 kg sample of furnace matte was crushed to a 100 - 600 μm particle size range . the roasting was accomplished in 4 days of operation , with 9 passes of 12 hours each , with a stepped increase in temperature from 675 ° c . to 1000 ° c . over this period . the sulphur content of this material decreased from 26 . 7 % to 0 . 04 % by mass . a mass balance shows that 1 . kg of dead - roasted matte is produced from 1 . 05 kg of furnace matte as originally supplied . during roasting , the impurities were diminished as follows : s from 27 . 4 % to 0 . 035 %; se from 244 ppm to 14 ppm ; te from 96 ppm to 32 ppm ; and as from 54 ppm to 46 ppm . there is no significant loss of pgms , except for some os . a crucible test in a laboratory - scale furnace was performed using a feed comprising 1050 g of dead - roasted furnace matte ( derived from 1098 g of unroasted furnace matte ), 450 g of silica , and 31 . 5 g of carbon . this produced 1517 g of slag , and 38 g of a copper - nickel alloy containing the vast majority of the precious metals . the metal button that was produced was equivalent in mass to 12 % of the cu — ni content of the original furnace matte . this alloy quantity is comparable to the amount of pgm - containing alloy produced in the traditional slow - cooling process . the recovery of the precious metals was 99 . 0 %, expressed as ( pgm + au in alloy )/( pgm + au in alloy and slag ). it is clearly quite possible to treat the slag from the first smelting stage according to standard slag - cleaning practice in a dc arc furnace . very high recoveries of the base metals and the residual precious metals would be expected in the second - stage collection . calcined zinc concentrate was fed , together with coke as a reductant , to a pilot - scale dc arc furnace , fuming off zinc vapour . ( other works , has demonstrated the production of prime westem grade zinc by further treatment of the zinc vapour in a lead splash condenser . it is also possible to use distillation to refine this zinc even further .) a total of 56 tons of calcine was processed during the test work , with coke and lime additions averaging approximately 13 % and 3 %, respectively . approximately 16 tons of discard slag and 38 tons of zinc oxide - rich bag - plant dust ( fume ) was produced by operating the dc arc furnace at a power level between 500 and 700 kw . in this series of tests , the zinc vapour leaving the furnace was combusted with air and collected in a bag plant . the feed materials included unagglomerated calcine , pellets dried to 150 ° c ., pellets dried to 350 ° c ., and pellets indurated at 1300 ° c . the sulphur content of the feed materials varied between 1 and 2 . 4 %. the addition of between 12 and 13 % coke resulted in an overall zinc extraction efficiency of 95 . 4 %. fuming rates of up to 170 kg zn / h per m 2 of bath area were obtained . ( in the case of feeding unagglomerated calcine , a zinc extraction efficiency of 98 . 7 % was obtained , and the average zinc fuming rate was 164 kg / h · m 2 .) the specific energy requirement was found to be approximately 1 . 17 mwh / ton feed at an average operating temperature of 1490 ° c . iron production varied between 2 and 21 kg per ton of calcine . the fume produced during the test work was of an even better quality than that for previous test work during which the condenser was successfully coupled to the furnace . the ratio of cao , mgo , sio 2 , and feo to zno was found to be approximately 0 . 04 in this test work , compared to a value of 0 . 14 found previously . therefore it is reasonable to expect good condenser performance . 1 . h fabian , copper , ullmann &# 39 ; s encyclopedia of industrial chemistry , fifth edition , vol . a7 , vch vertagsgesellschaft mbh , new york , 1986 , pp . 471 - 523 . 2 . r w bartlett , r j mcclincy , & amp ; r j wesely , smelting copper without converters , jom , vol . 37 , no . 5 , may 1985 , pp . 17 - 19 . 3 . r r odle , a e morris , & amp ; 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