CA1218238A - Method of processing sulphide copper and/or sulphide copper-zinc concentrates - Google Patents

Abstract

METHOD OF PROCESSING SULPHIDE COPPER- AND/OR
SULPHIDE COPPER-ZINC CONCENTRATES
ABSTRACT
A method of processing sulphide copper- and/or sulphide copper-zinc concentrates comprises flash smelting of said concentrates and basic silicate fluxes in the presence of oxygen with the resultant formation of a dis-persed mixture of highly basic molten slag containing cop-per and zinc oxides, metallic copper and white matte. The initial concentrates and silicate fluxes are taken in ratios permitting production of highly basic slag, where-upon copper and zinc oxides contained in the highly basic molten slag are reduced by means of a solid carbonaceous material with the resultant formation of a vapour-gas mixture containing vapours of zinc, metallic copper and depleted highly basic molten slag containing not more than 18 percent by weight of silicon dioxide. The resultant vapour-gas mixture is then subjected to oxidation with a view to producing and collecting zinc oxide.

Description

1823!3 ~ he present invention relates to non~errous metal-lurgy, and more particul~rly, to the production of non-ferrous metals by a pyrometallur~ical process. ~ore speci-~ically, it is concerned with processing 3~ sulphide cop-per- and/or sulphide copper-zinc concentrates.
Field of Application ~ he i~vention is applicable to the processing o~ a low-grade polymetallic material used for producing metal lic copper~ oxidized zinc sublimates and highly basic slags suitable ~or use as the starting material in other branches of industryo Background Art In modern pyrometallurgical practice sulp~ide con-centrates are treated by using the difference i~ chemical a~finity of metals to oxyge~ and sulphur contained in the raw material. To increase this difference, the smelting process i~ carried out in the presencs of silico~ dio-xide. The silicate melt is then reduced. It should con-tain basic oxide~ required to ensure high rate o~ reduc-tion. ~hus, in this case use is made of the-dif~erence between element~ in their affinity to oxyge~.
Further progress in pyrometallur~ical production of heavy no~ferrou~ metals depends on the development of e~-~ecti~e techniques ~cr their recovery on the b~sis o~

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autogenou~ ~melting of sulphide ma-terial. Advantage~ o~
autogenous proce~e~ are known to permit high production rate (~hort time of holding the material in the oxida-tion zone); a sharp decrease in the amount of process gases; utilizing heating capacity of the concentrate~
and thu~ ~ubstantially diminishing the use o~ external heat sources; the possibility of effective trea-~ment of the raw material which is r21atively poor in non~errous metals. In general, there is known a wide variety of autogenou~ processesO ~owever, a feature most common to most of them is the u~e of highly developed sur~ace of the sulphide material with a ~iew to en~uri~g autogenou~
nature of the roastin~ and smelting proce~3es.
~ or example, the Ontokumpu company, ~inland, has developed a method o~ processing sulphide copper con-centrates. The method in que~tion i~ carried out by u~ing a flow o~ po~sibly preheated air ~that can be o~ygen-enriched) ~or flash ~melting of finely divided sulphide copper concentrate~. The smel-ting process i~ effected in the presence of subYtantially sillcate flUx0~ with the re~ultant dispersed molten mi~ture of silicate ~lag and matte containing not more than 65 percent by weight o~
cop~er~
~ hi~ dispersed mi~ture of ~lag and matte i~ ~eparat-ed in accordance with ~pecific weight~. The matte i~
further treated until met~llic copper is recovered there-~rom, while ~ilicate slag is depleted in an electric ~LZ~8Z3~il furnace by ~ettling or flotation after being crushed ~nd divided~
The gase~ re~ultant from fla~h ~melting of the ini~
tial material contain from 2 to 20 percent of ~ulphurous anhydride~ The lo~ of partially oxidized powder concen-trate during ~melting i~ in the range of 8 to 10 percent by ma~æ of the material ~ed for ~mel-ting (see, for e~ample, 5hein Y.P. I'Non~errou~ Metal~", No 8, 1980, ppo25~29 ~Sympo~ium on Non~errou~ ~etallurgy in ~inlandJ; Engineer-ing and Mining Journal, 1973, 1741~o 11, p. 103-108;
M~J. Ethem "Eræmetall't, ~o 4, 2~ 76, pp. 182-186).
~ here exi~t other modificationæ of this method~ which are directed at enhancing its technical-and-economic cha-racteri~tic~ For instance, the content o~ copper in the matte i~ brought up by increasing the concentration of o~ygen in ths flow o~ air supplied for ~melting, the tem-perature of air or that of o~ygen enriched air being rais-ed. It has been attempted to increase the rate of recovery o~ copper ~rom raw material by arranging electrode~ in the zone where ~lag~ qettled down to become free ~rom matte.
Ho~ever, the above-mentioned improvement~ failed to obvia~e seriou~ di~advantage~ inherent in the prior-art method, namely:
- the impo~ibility of proces~ing raw material with higher-than-average content of zinc, or combin;ng the pro-ce~ing of copper-$inc concentrate~ with the ~taga of pro-ducing blister copper;

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- a low rate of recovery of copper and zinc from ~ilicate slag~.
To o~ercome difficulties that might possibly occur in the proce~s of producing rich matte (with over 65~ by weight of copper) 9 blister copper, or during treatment of a low-grade material containing difficulty oxidizable zinc ~ulphide, and in order to bring down the amount of magnetite (and the conte~t of copper in slag), it has been attempted to di~ribute the supply of oxygen required for oxidation of sulphide sulphur, and to perform smelting along with preroasting and subsequent utilisation of fuel.
According to Pat. ~o~2,515,464 ~Federal Republic of ~ermany, cl. C 22 B 15/04) the initial material is first partially roa~ted and then smelted to produce matte, A~
this happen3, a flow of hot o~ygen is introduced into the roa~t ga~e~ which are then used a~ auxiliary air for com-bustion of fuel in the cour~e of ~melting However9 the above method is rather di~ficult to perform apart from being ine~fec-tive to prevent the ~ormation o~ magnetite.
In addition, it fail~ to ~olYe the problems a~sociated with the proce~sing o~ copper-zinc or copper zinc-bearing concentrate~ and with the depletion of ~lags to permit good recovery of nonferrou~ metals therefrom.
There i9 known a KIVCET proceqs ~hich comprises roasting and smelting oP flotation concentrates, contain-ing nonferrous metals, in the atmo~phere of commercial o~ygen mixed with reusable du~t re~ultant from cleaning 12 3L8Z3~

of flue gaqes, and ~ubstantially by with 3ilicate ~luxes thereby obtaining matte and reducing zinc from the ~ilicate melt in an electric furnace (see, -Eor e2ample, Japane~e Pat. No.16362/76, cl. C 22 B 15/00; V.V.Vylegzha-nin et al~ ~Nonferrou~ Metals", No.l, 1976, pp. 26-28;
I~M. Cherednik et al. "Nonferrou~ Metal~'t, No.7, 1974, pp~ 24-27; Melcher G. et al. Er~metall, 1975, 2~, ~o.7/8, p. 313-322, I, II 9 III).
~he XIVCET proce~s is a~ follow~.
A finely divided copper- or copper-zinc concentrate i~ submitted to flash ~melting in the presence of sub~tan-tially silicate fluxe~ in the atmo~phere of commercial oxygen, follo~ed by the formation of a disper~ed mixture of ~ilicate ~lag and copper matte. ~he di~per~ed melt i~
further ~eparated in accordance with ~pecific weight~ into ~ilicate ~lag and matte. ~e~t, the ~lag and matte are admitted into an electothermic furnace, wherein a ~olid carbonaceou~ material (coke breeze) is charged onto the ~urface of the molten ~ilicate ~lag. Under the in~luence of high temperature the zinc oxide contained in the ~lag i~ reduced to metal which, when evaporized~ pa~qe~ into vapour-gas pha~e while copper oxide~ al~o pre~ent in the ~lag, ars reduced to metal and metallic copper i~ ~ettled down to form matte. The vapour-gas mixture i~ removed from the electric furnace, the vapour of metallic zinc i9 oxidiz-ed therein to form ~inc o~ide by me~ns of air ~upply. The resultant zinc o~ide i~ collected to be fed for further -- 6 ~

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treatment. The matte i~ tapped from the elec-tric furnace for further treatment to be blown through and o~idi~ed ~ith oxygen. As a re~ult o-E ~his treatment, metallic copper is produced to undergo subqequent refining.
In the KIVCET proces~, due to the use of commercial o~ygen for oxidizing finely divided sulphide material with a high ~pecific surface, and owing to smelting carried out to produce matte (up to 55~o by weight of copper and about

2~to by weight of sulphur), it take3 only about 0.1 g for the oxygen fed ~or smelting to be almost completely assi-milated by the sulphide material in su~pension. A~ a result, a high temperature i~ developed during 3melting to ensure sufficiently high rate of oxidation of zinc sulphide to it~ oxide with ~inc oxide passing into silicate 31ag.
Howe~er, the above method of proce~sing sulphide copper- or ~ulphide copper-zinc concentrates ~ith the production of silicate slags also suffers from a number of ~erious disadvantages, namely:
1. The specific production rate at the ~tage of slag depletion i3 low by reason of ~he nature of ~lag used tfrom 4 to 12 kg of zinc per day per sq. meter o~ area in the electric furnace); hence are appreciably high losses of heat (power input) for 31ag depletion, as well as labour input per unit of commercial productO
2~ The degree of recovery of zinc from raw materi~l to produce zinc o~ide is not higher than 80 percent.

3. The impossibility to treat sulphide concentrate~

with a high content of zi~c (about 20~o by weight o~ zinc) because o:E ~ 10W rate and degree of o~idation o.~ ~inc sulphide and its conversion into ~lag ~o a~ to enable further reduotion of zinc oxide ~rom qlag with a ~ati~-factory rate of zinc recovery from the initial material (more than 60-70~)~

4. ~he proce~qsing of the re~ultant copper matte to produce metallic copper requires qubstantial capital in-~e~tment and labour input apart from the neces~ity to develop a complicated emi~qsion gas control _ystem.

5. The impossibility to prevent the formation of con-siderable amount~ of magnetite during *la~h ~melting o~
concentrate~; therefore, complete o~idation of the afore--merltioned sulphide material leeds to the fo~nation of high-viscosity di~persed mi~ture o~ qilicate slag, metal-lic copper or white matte, which is rather difficult to di~integrate. In addition, the recovery of copper and zinc from high-viqcosity silicate slag iq fairly intricate procedure hindering comprenensive treatment of low grade and, in particular, copper-zinc concentrates.
From the above it follows that new method~ ~rereq.uiIed to qub~tantially increa~e the rate of production at the ~tage of slag depletion along with corre~ponding decrea~e in the labour and power input~; to bring up the rate o~
zinc recovery from the initial material; to rai~e the rate of o~idation o~ ~ulphide concentrate~q during flaqh smelting, thereby permitting the production of metallic copper to be effected directly at this stage, as well as to bring down capital investments and labour force in-volved in the production of metallic copper' to prevent the formation of considerable amounts of magnetite during flash smelting of sulphide concentrates, which will make it possible to introduce a wider variety of initial mate-rials into the copper and æinc industxy by making use of low-grade copper and copper-zinc sulphide concentrates along with direct production of metallic copper.
It is an object of the present invention to provide such a method of processing sulphide copper- or sulphide copper-zinc concentrates or mixtures thereof that will make it possible to enhance the production capacity and the rate of smelting of the initial material.
Another object of the invention is to provide such a method of processing sulphide copper- or sulphide copper-zinc concentrates or mixtures thereof that will make it possible to increase the production capacity at the stage of slag depletion along with a higher rate of zinc reco-very from the initial material, while permitting produc-tion of metallic copper or white matte.
The foregoing objects are attained in a method of processing sulphide copper- or sulphide copper-zinc iron-bearing concentrates, which comprises flash smelting of said concentrates in the presence of fluxes selected from the group consisting of basic fluxes and combinations 9 _ :''"

~L2~8Z~3 thereo~ with silicate fluxe~, and oxygen with the re~ult-ant formation of a disper~ed mixture of ~lag, metallic copper or white matte, subsequent reduction of copper and zinc oxide~, contained in the molten ~lag with a solid carbonaceou~ material, followed by the formation of a vapour-ga~ mixture containing zinc vapour, and of metallic copper and slag poor in nonferrous metals, subjecting to oxidation the vapour-ga~ mixture contain-ing zinc vapour, and collecting the resultant zinc o~ide, wherein, according to the invention7 the initial concent-rates and fluxe~ are fed for smelting in amounts ~uffi-cient to en~ure the production of a highly ba~ic molten slag which~ on being reduced to give o~f copper and.zinc oxide~, co~tain~ not more than 1~% by weight of ~ilicon dioxide Owing to ~uch conditions, calcium oxide form~
low-melting o~ysulphide eutectic~ with zinc and iron sulphides at the intermediate stage~ of oxidationl With high-melting and difficulty oxidizable zinc ~ulphide pa~sing into a low-melting liquid phase (aust as the pro-duct of it~ oxidation - zinc o~ide), the rate of the charge de~ulphurization (the xate of oxidation) is markedly increaqed. Further, in the proce~s of smelting carried out to produce highly ba~ic slag with the afore~
-mentioned ratio o~ the ba~ic ~lag-forming component~, there take no place or hardly take~ place any formation of magnetite which otherwi~e greatly increase~ visco~ity 8~3a of conventional silicate slags ~nd thus hinders the pro~
ce~s of oxidation. The present invention perm-its low--melting calcium ferrite to be formed instead of magnetite so as to dilute the products of oxidation and in this way to ensure the transportation of oxygen in the volume of droplet~ of the dispersed oxysulphide melt. Owing to the~e two main factors, that is - the formation of low-melting oxysulphide pha3es by means of calcium oxide and ~ins and iron sulphides, and the ~ormation of low-melting calcium ferrite instead of magnetite 9 it became possible to sub stantially intensify the process of desulphurization so that the copper sulphides present in copper are passed into metallic copper directly at the stage of oxidizirlg--flash smelting. In other words, copper is now produced directly from sulphide copper- or sulphide copper-zinc concentrates with the effect that the conversion o~ matte is ruled out to permit substantial cut-down of e~penses a~d labour involved. Moreover, a considerable increa~e in the rate o~ oxidation of ~inc sulphide has made it pos-~ible to perform effecti~e treatment o~ zinc-bearing sulphide concentrates. ~hi~, in turn, greatly expand~ the variety of initial materials in the copper and ~inc in-dustries by making use of` a low-grade sulphide material from which valuable component~ (copper, zinc and sulphur) are successfully extracted. ~urthermore, since highly ba~ic slag has a lower viscosity than silicate ~lag, the .

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metallic copper or white matte formed durlng smelting is easily separated from the highly ba~ic molte~ ~lag and pas3e3 into the bottom phase~ ~hi~ brings down the amount of copper lost with ~lag in the form of mechanical ~u~-pension (about 10 rel.%). The highly basic molten slag obtained during fla~h ~melting contains the whole of zinc a substantial amount of lead, with the content of copper in the form of oxide con~tituting about 5% by weight. The reduction of o~ides of nonferrou3 metals (copper, ~inc, lead) from highly basic molten slags proceed~ at much higher rate than in the case of ~ilicate ~lags and appro-aches tne ~peclfic rate of the fuming proce~. In part, thi~ takes place due to a higher intensity o~ the follow-ing reaction.
(n + 2m) MeO ~ ~n ~ m)C = (n + 2m~e + nCO + mC02 (1) with a corre~ponding increa~e in the activity of oxide~
of nonferrous metals contained in highly ba~ic ~lags.
Another reason ~or this are key changes in the coordina-tion ~tructure of metals in the highly ba~ic 31ags a~ com-pared to silicate 31ags In highly basic slag~, the struc ture-determining component is calcium (coordination num-ber - six) instead of æilicon in the silicate slag~ (coor-dination number - four)~ Such structural tran~formation~
lead to a decrea~e in the amount of free energy and eqpecially i~ that of free energy resultant from the activation of the following reaction:
~eO si = Fesl ~ ~e2 3 sl (2) 2,3~3 equivalent to the increaqe of it~ velocity by several orders. A~ a result, zinc, copper and lead oxide~ are reduced at a much higher rate in accordance with the following reactions:
ZnO + ~e 1 - Zn~ + ~eO (3) Cu20 ~ ~e~ 2Cu + ~e~ (4) ~hus, during carbon reduction of nonferrou~ metal~ ~rom highly basic slags it is more difficult, as compared to silicate ~lags9 for metallic iron to be separated into an independent phase or into blister copper. ~herefore, with the content of nonferrous metals in the reduced slag amounting to 1-1.5% by weight (0~5 to 0.7~0 by weight of zinc, 0.3 to 0.5% by ~eight of copper, O.Ol/G~ by weight of lead), metallic copper containing 1 to 3~o by weight of iron is found in equilibrium therewith. In contrast, the content of zinc in silicate slags under similar condi-tionq of the melt reduction by means of solid carbonaceous material can not be lower than 3.5 to 4% by weight of zinc, which i~ due to the fact that substantial amount o~ iron pa~e~ into the bottom phaqe with its melting temperature sharply raising to render impossible deep depletion of silicate slag because of solidification of the bottom phase. ~hus~ for the total content of nonferrou~ metals to be in the range of 1-1.5% by weight, it takes 10 to 15 times more time for the depletion of silicate slags than for the depletion of highly basic molten slags, ~urther-. . :

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more, such deep depletion result~ in that copper i~ con-siderably contaminated with metallic iron (not less than 20~o by weight of iron) with the melting temperature there-of being over 1400C. ~rom the above it follows that in comparison with the Japanese Pat. No.16362/76 clo C 22 B
: 15/00, it become~ po~sible ~ot only to improve the pro~
ces~ characte~istics at the stage of flash smelting of sulphide concentrates to produce highly basic slags in the presence of oxygen, but to improve the process cha-racteristics at the stage of slag depletion by using a solid carbonaceous material. ~o develop the point, the amount of zinc reco~ered from the initial material i9 substantially increased (from 75-80~o to 94-97~o), the rate of sublimation of volatile zinc is enhanced (from 4-12 to 40-60 kg/m2 per day)~ with power and labour inputs per unit of commercial product (zinc oxide) being propor-tionally reduced.
According to the invention~ the conte~t of silicon dioxide in the depleted highly basic molten slag is pre-ferably maintained within the range of 3 to 16% by weight with the US8 of silicate fluxa whereafter the melt i~
cooled do~n to a temperature of it~ complete ~olidifica-tion at a rate of Q.5 to 60 degrees per minute with the resultant production of self-disintegrating material from which nonferrous metals are finally recovered~ On the one hand, this will ensure moderate melting temperature of ~2~Z3~3 the melt, and the ~ormatio~ of ~ufficient amounts oY di calcium silicate during cooling and solidification of the melt, on the other. As the depleted highly basic molten slag is cooled down to a temperature of its complete soli~
dification at a rate of 0~5 to 60 degrees per minute9 this ~ilicate is produced in the form of fairly pure large-size crystals (with average diameter of 40 - 80 mcm). At a temperature of ~bout 675C9 the crystals of dicalcium 8ili-cate undergo polymorphic transformation9 which is followed by a 12 to 15% increase in their volume. This, in turn, brings about interior strain in the slag monolith. Under the influence of this interior strain the monolith is spontaneou~ly disintegrated to form particles of a size fit for flotation (the yield of this fraction is 75-85%).
In addition, the temperature decrease at the above-indi-cated rate makes possible the separation of copper left in the slag in the form of droplets with a di~meter of 80 mcm (about 70~0 of copper is recovered from the slag).
Due to spontaneous disintegration of slag and with separa-tion of copper therefrom in the ~orm o~ droplets, the sur-faces of which become exposed during spontaneous disintegra tion of the slag monolith, final extraction o~ copper from the depleted highly basic slag is simplified. ~hus5 with the content of copper in the depleted slag of 0.3 to 0.5G~o by weight, it is possible to obtain a concentrate by means o~ flotation with the content of copper ranging from 10 ;i 15 12~8~3 !3 to 15% ~ld with the discarded materi~l containing from 0.1 to 0.15% by weight of copper (depending upon the initial content o~ copper)~ ~t the ~ame time, following depletion, comminution and treatment of silicate slags by flota-tion, they still contain, as a rule, 0,5% by weight of copper (at best 0.3% by weight of copper), Therefore~ when ~melt-ing is carried out to produce highly basic slags to undergo ~urther treatment with carbonaceous material in order to obtain ~elf-disintegrating depleted slag, the most arduous opera-tions of the separating stage, such as crushing and dividing the solid monolith, are ruled out; the loss of copper with the discarded material is reduced from 1-1.5 rel.% to 0.~-0.6 rel.% (in contrast to the ~melting fol-lowed by the formation o~ silicate slag).
According to the invention, the resultant metallic copper i9 preferably submitted to refining in the presence o~ silicate ~luxes for the purpose of producing refined copper and silicate slag. It has been found that the metallic copper produced during treatment o~ copper- ox copper-zinc sulphide concentrate~ contains impurities of lead, æinc, iron and other elements to be removed from metallic copper. As molten met~llic copper is blown through with air in the presence of silicate fluxe~, the above-me~tioned impurities are ozidized to pa~s into sili-cate slag where oxides o~ the metallic impurities are com-bined to form strong low-volatile silicate~ and thus are -~2~ 3~3 removed from metallic copper. Simultaneously, copper i~
partially ozidized to form copper oxide which also pa~es into silicate slag and i~ partially dissolved in metallic copper to oxidize metallic impurities. Since during refin-ing of metallic copper the melt is intensively mixed with air, the dissolving of silicate flux proceeds at a rapid rate even when large-size lump~ of silicate flux are used.
Therefore~ with a view to reducing the loss of heat dur-ing transportation of metallic copper for refining~ the latter i~ preferably delivered to this stage in m~lten state.
According to the invention, the metallic copper ob-tained during flash smelting of concentrates in the pre-sence of oxygen in the course of treatment of the dispersed mixture of highly basic molten slag, metallic copper or white matte ~ith a solid carbonaceous material, and the metallic copper obtained during treatment o~ highly basic molten slag with a solid carbonaceous material, are dis-charged separately.
It has been found ~hat during flash smelting of cop-per- or copper-zinc concentrate in the presence of basic fluxes, the resultant metallic copper contains up to 90~
of copper contained in the initial material fed for smelt-ing. The metallic copper produced at the ~-tage of flash ~melting is preferably withdrawn from the process with only highly basic oxide melt being fed for carbon treat-~2~8;23~

ment. This permits the bulk of copper (up to 90 relO%) to be obtained in the form of commerciRl product, the pure~t possible for the given type of initial material, whereas the copper residue in the oxide melt is preferably extract-ed in the form of less pure metal at the stage of carbon treatment of slag. The foreign-metal impurities such as lead, zinc and iron pass into metallic copper mostly in the form of metals which are obtained as intermedi~te products resultant from the reaction between corre~ponding sulphides and oxide~, and according to reactions (2) and (3). Since in the process of flash smel-ting, due to high oxidation potential of the medium, the concentration of such inter-mediate products cannot ~e high (it has ma~i~um value for lead), as also the degree of transition of lead, zinc and iron into metallic copper. At the stage of carbon treat-ment of the high]y basic melt, the oxides of copper~ lead, zinc and, in part, of iron are reduced to metals. It should be observed that with the residual total content of copper, zinc and lead in the 02ide melt being 1 to 1.5/o by weight, metallic copper containing up to 8~ by ~eight o~ lead, up to lo 5h by weight of zinc and up to 2% by weight of iron will be in equilibrium there~ith. However, since the bulk of copper contained in the initial material is recovered prior to the stage of carbon treatment in the purest possible form, the overall contam~a~on of metallic copper with the above-mentioned impurities turns out to be ~.Z~L~3238 2 to 3 times lower during separate tapping in contrast to the practice when all copper is tapped following the stage of slag treatment~ In addition, during ~eparate dis-charging of metallic copper, direct recovery of zinc and lead in the form of sublimates inside the furnace is in-creaqed by 6 to ~ rel~O~ which brings down the losses of these metals with metallic copper.
According to the invention, the silicate slag pro-duced during refining of metallic copper i~q preferably used as the silicate flu2, which may contain up to 50~o by weight of copper, up to lO~o by weight of zinc, up to l~o by weigh-t of lead) since these metals are present in the slag basically in the form o~ oxides, the slag should be subjected to treatmen~ with carbonaceous materi~l. On the other handg the ~ag resultant from refining of metallic copper contain~ silicon dioxide which is preferably in-troduced for processing copper- and copper-zinc sulphide conce~trates so as to obtain a~ self-disintegrating ~lag monolith required for sub~equent final recovery of non~
ferrous metals. ~oth of the~e tasks are successfully fulfilled if the above-mentioned ~lag is used a~ the ~ilicate flux.
According to the invention, the silicate flu2 is pre-ferably i~troduced for u~e in flash sm~ting of the con-centrate Due to the fact that high temperature ~over 1500C) is developed in the course of flash smelting of ~Z~L8;:~8 sulphide concentrate in the presence of oXygen~a refrac-tory material, such as quart~ sand reduced to ~ size of minus 0.5 mm, i9 preferably used as the ~ilicate flux.
The presence of silicon dioxide in the furnace burden brings down the effect of formation of oxysulphide phases on the base of calcium oxide and copper and zinc sulphides.
However, since silicate flug is fed for smelting in the form of a refractory material, the dissolving of silicon dioxide in the resultant highly basic melt takes place mainly when the proces~q of oxidation of metal sulphides in suspension ha~ been completed. Therefore, adver~ary effect of silicon dioxide on the flash smelting of sulphide con-centrate turns out to be insignificant, with refractory silicate material being successfully used in the process.
According to the present invention, the silicate flux is preferably introduced into the molten highly basic slag produced during flash melting o~ the concentrate.
Thi~ will make it possible to reduce the load with regard to the inert material (~ilicon dioxide) at the stage of flash smelting of sulphide concentrate. At the same time, thi~ allows for the heat liberation processes during oxida-tion o~ metal sulphides and the heat absorption processes during dissolution of silicate flux in the highly basic melt to be run at minimum intervals. This technique is ad-vised when the ~ilicate ~lag resultant from refining metal-lic copper is used as the silicate flux, which containts ~2~3Z3~

all component~ predominantly in o~idized state~ Since this type of silicate slag has relatively low temperature (about 1200C), it is readily dis301ved in the dispersed highly basic slag having a temperature of more than 1500C.
However~ -to avoid oversaturation of the resultant melt with silicon dioxide and settling of refractory dicalcium silicate, the silicate slag from re~ining of metallic cop-per should be fed at regular intervals, According to the invention the silicate flux is pre-ferably introduced into the depleted highly basic molten slag, which permits proportional reduction in load with regard to an inert material at the stage of fla~h smelting of sulphide concen-trate and at the stage of treatment of the highly basic molten slag with a solid carbonaceous material. In addition, the presence of ~ilicon dioxide at the stage of treatment of the highly basic melt somewhat brings down the process rste and leads to an increase in the melting temperature of the melt (up to 1330C, with the content of silicon dioæide in the melt being 20~o by weight), ~herefore, when treatment of the highly basic melt is carried out with the content of silicon dio~ide (deter-mined by the composition of sulphide concentrate) being at its lo~e3t 9 the rate of flash smelting for the given type of concentrate is the highest and the rate of recovery of nonferrous metal~ from the highly basic sl~g is ma~imal.
According to the present invention, the depleted - .

~18;~3~3 ~ilicate slag poor in nonferrou~ metal~ i~ pre~erably u~ed a~ the ~ilicate Ylux. Thi~ ie adui~ed when silicate flllx i9 introduced into the highly ba~ic ~lag poor in nonfer-rous metals. Howevers the content of nonferrou~ metals (copper, lead) should be more than 0~5% by weight~ Such a low content o~ nonferrou~ metals in silicate ~lag permits their extrac-tion to produce the concentrate after obtain-ing self-disintegrating ~lag monolith and final reco~ery of nonferrou~ metal~, for example, by flotation method~. ~he current pyrometallurgical techniques used for ~melting cop-per with the re3ultant formation of silicate slag make it pos~ible to bring the content of copper therein up to 0~5%
by weight, thi3 being much higher than with the method of the pre~ent invention (0.1 - 0.15~ by weight o~ copper).
~herefore, such depleted silicate slags may be used in the given method.
~ urther obj2cts and advantages o~ the present in-vention will become more apparent to tho~e skilled in -the art upon a further reading of this disclosure, particularly ~vhen viewed in the light of illustrative example~.
Detailed Description of the Preferred Embodiment Sulphide copper conce~trate normally containing 15 to 30 percent by weight of copper, 20 to 35 percent by weight o~ iron, 27 to 40 percent by weight of ~ulphur, up to 4 percent by weight of silicon dioxide, or sulphide copper-zinc concentrate containing 6 to 20 percent by ~2~8~,38 weight of copper, 4 to 24 percent by weight of zinc, 25 -to 30 percent by weight o-~ iron, up to 3.5 percent by weight of lead, 30 to 35 percent by weight of sulphur, up to 6 percent by weight of silicon dioxide~ as well as aluminium and magnesium oxides7 are produced by means o~ flotation from copper- and copper-zinc ores reduced to a size frac-tion of minus 74 mcm. Where complex ores are subjected to dressing by flotation9 preference may be given to the pro-duction of either monometallic (copper or zinc) concentra-tes or polymetallic (copper-zinc) concentrate3, depending on the morphology o~ ores. Thi~ being done to increase the content of nonferrous metals in the resultant concentrates.
According to the invention, the copper- or copper-zinc con-centrate with a size particle of 74 mcm is mixed with lime flux (calcium carbonate, hydrated lime, calcium oxide) reduced to a size of minus 1 mm. The resultant mixture of concentrates and fluges is dried to a residual humidity of about 1% by weightO ~he dried mixture prepared ~rom f1nely di~ided concentrates and fluxes is fed into a mixer which is continuously supplied with the dust collected in an emigsion gQS control system. ~hen, by means of a charge--oxygen burner, arranged in the melting chamber, the mixture of concentrates, fluxes and reusable dust is con-tinuously injected in a ~low of oxygen into the meltlng chamber wherein the mixture is pulverized to become sus-pended. Under the influence of high temperature, which .

~LZ~L~323~

ha~ ri~en in the shaft in the course of burning of the preceding portions of similar mixture, the sulphide mate-ri~l is ignited in the presence of o~ygen, the rate of it~
oxidation is accelerated andS under the influence of liberating heat, calcium carbonate i9 decompo~qed to pro-duce calcium oxide. ~he resultant calcium oxide and non--oxidized particles of sulphides of nonferrou~ metals and iron, as well as zinc7 copper, iron and lead oxide~ resul-tant from their oxidation~ are dissolved in one another to ~orm low-melting oxysulphide phases which continue to interact with oxygen to produce sulphurous anhydride and metallic copper or white matte. At this stage, silicon dioxide and aluminium oxide present in the mixture, pa~s into the melt of oxides. Thus, a dispersed mixture of metallic copper or white matte and highly basic molten slag with æinc, lead and copper oxides dissolved therei~ , descend to the shaft bottom to settle on the 3urface of the melt dispo~ed therein. Thi~ di~per~ed mi~ture is ~eparated according to ~pecific weight~q in two layers, of which one layer is highly basic molten ~lag and the other is metallic copper or white matte. At this ~t~ge, the rate of recovery of copper to obtain the above-mentioned pro-ducts (copper or white matte) reaches ~0~0 of the amount of copper contained in the initial material ~upplied for smelting, with the rate of recovery o~ such forsign-metal impurities as lead, zinc and iron to obtain metallic cop-~L2~ '3 !3 per or white matte being minimal. Therefore, metallic copper or white matte is withdrawn from the process after the stage o~ flash smelting and ~eparation into layers of the di~persed mixture~ while highly basic slag is fed for treatment to extract therefrom zinc, lead and the residue of copper dissolved in the form of oxide (3 to 6 percent by weight of copper). The ga~eou~ products resultant from smelting of ~ulphide concentrates 9 which contain more than 99~ of sulphide sulphur fed for smelting (to produce metal-lic copper) or about 85% of sulphide sulphur ~ed for smelting (to produce white matte) are directed together with dust to a dust collecting system wherein hard partic~
les are separated from gases and continuously returned for flash smelting of the initial sulphide material. ~ree from hard particle~ and having a high content of sulphurou~ an-hydride (up to 80% by volume), these gases are then used for the recovery of sulphur. To perform depletion of highly basic molten slag, the latter is tapped into an electric furnace while a solid carbonaceous material (coke breeze or coal) is charged onto the ~urface of the melt. ~he car-bonaceous material is introduced in an amount su~ficient to provide for the reduction of copper, lead and zinc o~ides~ Under the influence of high temperature 9 copper oxide is reduced to metallic copper settled down at the ~otto~ of the electric furnace, and zinc oxide passes into metallic zinc vJhich i~ volatilized to be removed together with carbon oxide and carbon dioxide from the electric furnace in the form of a vapour-gas mixture.
Under these conditions, lead oxide is also reduced to metal and metallic lead is partly (by 30 to ~0/O) passed into metallic copper and is partly volatilized to be removed from the electric furnace in the form of the above-mentioned vapour-gas- mixture. This mixture may be either cooled with zinc and copper condensed in the form of metal-lic alloy, or else it can be mixed with air. At the same time, the vapours of metals will pass into corresponding oxides, with CO oxidized to C02. The oxides of ~inc and lead are collected to be fed for further treatment. The metalllc copper obtained during treatment of highly basic molten slag is removed from the furnace to undergo refin-ing. After treatment, the slag is tapped from the furnace either continuously or intermittently. It is suitable for further use in other branches of industry, for example, in the production of cement.
In order to obtain impoverished slag in the form of a self-disintegrading material, it is necessary that the impoverished molten slag with the amount 3 to 16 percent by weight of silicon dioxide contained therein should be produced by adding silicate flux. Thereafter, the highly basic molten slag is cooled to a temperature of its comp-lete solidification (about 1000C) at a rate of 0.5 to 60 degrees per min. Where quartz is used as the silicate flux, ~iL2~3238 it is pre~erably introduced in the form of quartz sand with a particle size of minus 0.5 mm in mixture with lime ~lux at the stage of flash smelting of ~ulphide concentr~-te. Since high temperature (over 1500C) is developed dur-ing flash smelting of sulphide concentrate in co~mercial oxygen, the crushed refractory silicate flux rapidly passes into the resultant highly basic molten slag directly at the stage of flash smelting. The amounts o~ lime and silicate to be added to the given sulphide concentrate are determined by appropriate calculations normally used in metallurgical practice. ~or instance, the composition of ~lag with regard to its main componentæ /CaO, SiO2, MgOg A1203, ~eO, ~e203 / is determined either experimantally or on the strength of chemical analysis. This type of slag could be produced during smelting of the concentrate in the absence of fluxes.It isto be assumed for the sake of simp-licity, that nonferrous metals are removed cbmpletely, and the overall content of the main components i~ calculated in such a slag. With a prescri~ea content o~ calcium o~ide and silicon dioxide in the highly basic slag, thè amounts of components (~iO2 and CaO~ to be introduced into the pro-cess are determined in accordance with constitutive rela-tion. Then, correction~ are introduced with regard to the contents of lime and silicon dioxide in the sele~ted lime and silicate materials. Since the calculation i~ made for the depleted slag, it is of no importance at what stage a ~2~8Z31~

given type of flux is introduced.
The depleted highly basic melt containing from 3 to 16% by weigh-t of silicon dioxide, used to produce a qlag monolith prone to ~pontaneous disintegration with the re-sultant particles having minus 74 mcm in size, ~hould be cooled down to a temperature of its complete solidifica-tion at a r~te of 0.5 to 60 per min. To maintain the rate of cooling of the molten slag within the above range, the depleted highly basic slag is poured into ladles so as to permit the melt to be cooled in the air during its natural convection. ~or e~ample ? with the ladle external ~urface -to the ladle volume ratio being 4 m2/m3, the melt i~ cooled to a temperature of its complete solidification at the rate of 3 deg. per min. The effect of the slae spon-taneous disintegration is observed at a temperature of ~00-500C. In tha proce~s of cooling and ~olidifica-tion of the meltg about 80 rel.~0 of the residual amount of copper is separated in the form of regulus of metallic copper with a particle size being 40 to 50 mcm. After comminution of the obtained powder slag, the regulus--formed me-tallic copper iq treated by flotation methods to yield copper concentrate with the content of copper not lower than ~wt~C~c (the yield of concentrate being le~s than

6% by v~eight of slag), so -that the amount of copper in the tailing~ of flotation, with the initial content of copper in the depleted hi~lly basic slag of 0~5wt,%, remains in ;

- 2~ -3~3 the range of 0.1 to 0.15 percent by weight of copper. ~he controlled comminution does not require much power input, since after cooling the melt at a rate of 0. 5 to 60 deg.
per minute, coarse particles of slag (amounting to 15-2~o by wei~ht of slag) have very low mechanical strength~ The tailing~ of flotation are suitable for other commercial applications wi-th iron and calcium oxides contained therein as basic components (the overall amount being 60 to 70wt.%).
According to the present invention~ the metallic cop-per obtained during proce~ing o* sulphide copper- or sul-phide copper-zinc concentrates i~ subjected to refining preferably in the presence of silicate fluxes with the resultant production of purified copper and silicate slag, Thus, the properties of silicate flux can be used to com-bine 3inc, lead and iron oxides, resultant from refining of metallic copper9 into respective silicates. Thi~ makes it po~sible to bring down the waste o~ metal~ carried away with dust during refining of metallic copper. ~urthermore, in the course of refining of metallic copper in the pre-sence of silicate ~luxes~ the rate of their dissolution i3 enhanced due to intensive stirring of molten copper during its refining. The flows of reusable materials are al~o dimini~hed due to the fact that the slag resultant from refining of metallic copper contains several compo-nents required for the process, namely, silicon dioxide, copper, zinc and lead.

~2~ 3Z3~

According to the invention, the metallic copper obtained during flash ~melting of the concentrate in the presence of oxygen and the metallic copper obtained during treatment of the highly ba~ic molten slag with a solid car-bonaceous material, are preferably discharged separately.
As a result, the bu-lk of copper (up to ~0 rel.~o) is obtain-ed in the form of commercial product, the purest possible for the gi~en type of initial material ~he residual amount of copper i~ recovered from the oxide melt in the form of more contaminated me-tal at the stage of carbon treatment of slag. Such foreign-metal imporities as lead, zinc and iron pass into metallic copper mainly in the form of metals resultant as intermediate products from the interaction of corresponding sulphides and oxides, and in accordance with reactions ~2) and (3). Since oxidation potential of the medium is fairly high during flash smelting in the presence of oxygen, the concentration of such intermediate products cannot be high (it is at it~ maximum for lead). Hence is relatively small degree of transition of lead9 zinc and iron into metallic copperO At the stage of carbon t~eat-ment of the highly basic melt, copper, lead, zinc and partly iron o~ides are reduced to metals With the residual amount of copper, zinc and lead in the melt being 1-1.5% by weight, metallic copper containing up to 8% by weight of lead, up to 1.5~o by weight of zinc and up to 2;~o by weight of iron, will be in equilibrium therewith. Since, however, - 3~ -~aZ~32~

the bulk of copper fed with the initial materi~l i8 e~tracted before the stage of carbon treatment in the purest form, the total contamination of metallic copper with the above-mentioned inclusions during separate di~-charging turns out to be 2-3 times lower than at a time of discharging the whole of copper following the stage of slag treatment. In addition3 during separate discharge of metallic copper, the direct recovery of zinc ~s well as lead in the form of sublimates in the electric furnace is raised by 6 to 8 rel.~O due to a decrea~e in the waste of these metals lost with metallic copper. Thus, separate di~charging and separate refining of metallic copper make it possible to bring do~r~ the`overall duration of the copper re~ining cycle and the amount of resultant slags with zinc and lead oxides concentrated therein and with copper oxide passing thereinto.
The silicate slag obtained during copper refining process is preferably used as the silicate flu~ in the process of treatment of sulphide copper- and sulphide copper-zinc concentrates. Thus, the process of transi-tion of silicate flug into the hi~hly basic molten flux is accelerated, since the slag melting temperature is at least 500C lower than the melting temperature of ~ilicon dio~ide. Therefore, this type of slag is precrushed to a size of minus 10 mm before it is into the charge fed ~or flash smelting of the sulphide concentrate, or into the ~2~ 2~

disperse~ hi~hly basic slag obtained during flash ~melt-ing of the concentrate. The amount of ~ilicate ~lag to be introduced into the charge so as to obtain depleted slag with the content of silicon dioxide therein ranging from 3 to 16wt% is calculated in a manner similar to that des-cribed above~ Silicate flux can be introduced into the highly basic molten slag resultant from flash smelting of sulphide concantrate. The introduction of silicate flux after the melting process is conducive to maintaining high rate of interaction between metal sulphides and the gas-phase 02ygen in the course of flash smelting, insofar as silicon dioxide brings down the degree of mutual solu-bility of sulphides and calcium oxide with the formation of oxysulphide low-meltin~ phases which act to step up interaction of the sulphide concentrates with the gas--phase oxygen. This procedure is advisable when the sili-cate flux used is a low-melting material (with a melting temperature of aboùt 1200C) containing nonferrous metals which could be extracted during subsequent treatment of highly basic molten slag with a solid carbonaceous mate-rial. Such silicate ~lag is first precrushed (with frac-tion siæe of minus 10 mm) and then charged into the melt-in~ chamber. It is fed substantially continuou~ly so as to a~oid local supersaturation and supercooling of the melt and to step up the process of assimilation of the ~ilicate flux by the highly basic molten slag, For examp-le, the slag resultant from refining of metallic copper ~2~323Y~

may be used as the silicate flux in question.
Silicate flux can also be introduced into the charge used for flash smelting of the sulphide copper- or sulphi-de copper-zinc concentrate, with the material used as the silicate flux having a high content of silicon dioxide and, consequently, high melting temperature, as well as containing sulphide sulphur. For example, the tailings resultant from treatment of sulphide ores, which along with silicon dioxide (over 60 wt.%) contain ferric sul-phides, are frequently used as the silicate flux. In this case, owing to the fact that the introduction of ferric sulphides provides for additional liberation of heat in the course of their oxidation in suspension, the oxida-tion of sulphide concentrates is not impaired despite the introduction of silicate flux into the charge. Such silicate flux is preferably ground to a particle size of up to minus 0.5 mm, which permits silicon-dioxide to be dissolved in the dispersed highly basic molten slag ~t the final stage of flash smelting of sulphide concentrate in mixture with fluxes.
In addition, the silicate flux is preferably in-troduced into the highly basic molten slag after the lat-ter is depleted. This permits the highly basic molten slag to be treated with a carbonaceous material while preserving a minimum content of silicon dioxide in the melt for the given type of initial material. This permits .~

oxides of nonferrous metals to be reduced at a high rate by means of carbonaceous material and leads to a higher production rate at the stage of depletion~ While making use of this techni~ue, it is advisable to employ silicate ~lux with a low melting temperature (about 1200C) In addition, it should be precrushed to have a par-ticle ~iæe of minus 10 mm before being fed for use. While the slag treating proces~ i~ carried out intermittently with regard to its tapping from the electric furnace, the silicate slag i9 discharged prior to tapping the depleted slag.
With the depleted slag being discharged continuously, the silicate ~lux is likewise charged continuously into the slag tapping in -the electric furnace.
The silicate ~lag poor in nonferrous metals is pre-ferably used as the silicate flux. Silicate sïags obtain-ed at present during pyrometallurgical smelting of copper contain copper and other metals in an amount of about 0.5% by weight of copper, which significantly surpasses the minimum (0.1 - 0.15% by wei~ht of copper) obtainable with the method o~ the invention. Therefore, by using depleted silicate slags as the silicate flux to be in-troduced into the highly basic molten slag after its depletion, it is possible to recover from 60 to 80 rel.% of copper introduced together wîth the silicate slag.
The method of the invention can be performed by oxidizing the initial material in a solid-phase state 12~823~3 until it~ complete desulphurization (or until copper to sulphur weight ratio is about 4:1, ensuring formation of white matte). Then, necessary fluxes are added to the re-sultant ro~sted product, the charge is melted to produce highly basic molten slag9 metallic copper or white matte, whereupon the highly basic melt is treated with a carbo-naceous reducing agent.
The method o~ the invention can be succes~fully per~
formed by means o~ conventional metallurgic equipment. ~or example, highly efficient fluidized-bed roasting furnaces now in use are well suited for roasting and 02idizing pro-cesses. There~ore, there is no need in constructing special metallurgical unit~ for flash smelting of ~ulphide con-centrates.
In addition, the pre3ent invention pe~mits the use of a method as described in Japane~e Pat. ~o.16362/76, ac-cording to which zinc vapour, obtained during treatment o~
highly basic molten ~lag with a carbonaceous material, are cooled and conden~ed to produce metallic zinc.
~rom the above it follows that the method of proce~-ing copper or copper-zinc sulphide concentrate~ is in many ways ~uperior to the prior-art methods, its advantages being the following:
- the conversion of copper matte is expelled ~rom the operating process;
- the recovery of sulphur contained in ga~es rich in ~ 35 -3~

sulphurous anhydride iq effected in a single ~tage;
- full utilization is made of the heatin~ capacity of the ~ulphide material;
- a complex zinc-bearing material is o~idized at a ~airly high rate;
- the degree of recovery of zinc in oxidized zinc sublimates is ~ubstantially enhanced;
- the rate of zinc evaporation is upped several -times during electrothermic treatment of slagsi - the process of treatment is characterized by a low residue of nonferrous metals in slag ~the total amount of 1~ by wèight);
- the depleted slag may be obtained in the form of a self-disintegrating product, which facilitates final reco-very of nonferrous metals therefrom;
- power inputs are brought down approximately in pro-portion with the increase in the production rate at the ~age of electrothermic treatme~t.
~ he method of the in~ention also allows processing of sulphide copper- Qr copper-zinc concentrates with the cor~
tent of lead substantially exceeding 4 wt.%9 in particular~
copper~lead-zinc concentrates (10-20wt.% copper, 6-20wt.%
lead, 10-15wt.% zinc), as well as lead-zinc and lead sul-phide concentrates tup to 70wto~O lead, about 3wt.~o copper, up to 16wt.% zinc), since the behaviour of lead compounds at the stage of flash smelting in the presence of oxygen and at the stage of reduction of the highly basic melt is ;23~3 fairly similar to the behaviour of analogous copper com-pounds, while the presence of lead hQs almost no e~fec-t on the above-described process of obtaining zinc in the fo~ o~ a commercial product.
The invention will be further described by the fol-lowing illustrative Examples.
Example 1 A sulphide copper concentrate with a particle size .
of minus 74 mcm, produced by flotation from copper ore and ha~ing the following composition, in percent by weight:
copper . . . . . , . . . 23.25 iron . . . . . . ~ . . . 30,5~
sulphur . . . . . . . , 35.49 zinc . . . . . . . . . . 0.15 lead . . . . . . . . . . 0.01 silicon dioxide . . . . . 1.8 calcium ogiae , . . . . . 2.57 aluminium oxide . . ~ . . 1.44 magnesium oxide . . . . . 0.99 was u~ed as ~he initial material. A precrushed lImestone (with a fraction size o~ minu 1 mm) was added to the initial material as calculated in terms of pure calcium o~ide taken in an amount of 20% by ~eight o~ the initial ~ulphide material. The charge wa3 then dried until its humidity was about 1% by weight.
The resultant dried charge was continuously fed ~ 37 -3;23~

into a mixer, with the dust resultant from cleaning of`
gases produced in the course of -flash smelting being con-tinuously introduced into the charge in the mixer~ The charge mi~ed with dust was suspended in a flow of commer-cial oxygen (380 nm3 per -ton of concentrate) by means of an oxygen burner so as to be intermittently introduced into the mel-ting charge, Specific load in terms o~ the charge in the melting zone was 50 tons per 1 m2 per day. Under the influence of high temperature, the charge was ignited in the presence of oxygen, then melted with calcium carbonate passing into oxide; sulphide sulphur wa~ o~idi~ed to sulphurous anhydride. Thus 9 a dispersed mixture of metallic copper and highly basic molten slag descended to the bottom of the shaft onto the surface of the melt. Directly under the flame in the zone o~ fla~h smelting, the dispersed mixture was separated into two layers in accordance with specific weights, of which one wa~ metallic copper a~d the other molten slag. The metal-lic copper obtained during flash smelting was periodically discharged (because o~ insufficient volume of the metal-lurgical unit), while slag was continuously fed into the electric furnace. The gases resultant from flash smelting were directed together with dust for treatment. The col-lected dust was continuously returned into the charge fed for flash smel-ting of sulphide concentrate. The highly basic slag incoming to the electric furnace contained~ in -3~3 percent by weight:
copper . . . . . . . . . . 3.23 (in the form of oxide) zinc . . O . . . . . . . , 0.1 calcium oxide . . . . . ~29,1 iron . . . . . . ~ . O . .39.47 silicon dioxide . . . . . 2,97 The yield of slag was 77.5% by weight o~ the con-centrate, and 25.03 kg of copper in the form of copper oxide was introduced together with the slag into the elec-tric furnace per ton of concentrate. According to stoichio-metric requirements, 4.7 kg of carbon is needed for copper oxide to be reduced to metal. Coke breeze was charged into the electric furnace onto the melt surface in an amount of 9.7 kg per ton of concentrate as calculated in terms of pure carbon. Metal oxide from the molten slag was reduced by means of carbon to metallic copper which was deposited at the bottom of the electric furnace. Specific load of the electric furnace in term~ of ~lag during its treat-ment with co~e breeze was 10 tons per sq.meter per day.
The carbon oxide resultant from the reduction of slag was removed from the electric furnace with a 3mall amount o~
zinc vapour, whereupon it wa~ mixed with air and oxidized to C02 and zinc oxide which wa~ collected in the form of dust. Metalli~ copper and depleted highly ba~ic ~lag were periodically discharged from the electric furnace.
Metallic copper was subjected to refining in a manner ~ 39 -3LZ~ILI!3~38 that qilicon dioxide was introduced onto the ~urface of molten copper in the form of quartz in an amount of 4-5 kg of silicon dioxide per kg of iron, Then a flow of air was blown through the molten copper to produce refined copper (99.2wt,% copper) and silicate slag containing 36.6wt,%
copper, 0,3wt,% zinc, O.lwt.% lead, 9,2wt.% iron, 0,~wt,%
calcium oxide, 1.5~/t.% aluminium oxide,43.5wt.% silicon dioxide, ~he total amount of copper extracted Irom the con-centrate in metallic copper was 99,33%; the amount o~ cop-per recovered during flash smelting of the concentrate ~as ~9,33%, and during depletion of -the highly basic melt in the electric furnace - 10.10% of copper fed with the co~-centrate, The content of copper in the depleted slag was 0~21wt~% and that of zinc 0,09wt,%, The metallic copper produced during flash smelt;ng of the concentrate had the following compo~ition, in per cent by weight:
copper . . . . . . . . . . 98~68 iron , . . . . . , . . . . 0.07 zinc . . . . . . . . . . . 0.001 lead . . . . . . . . ~ . . 0,005 sulphur . . ~ . . . . . . 0.15 ~he metallic copper obtained during treatment of slag in the electric furnace contained, in percent by weight:

23~3 copper . . . . , , . . . . 95.~8 iron , . . . . . . . , . . 1.35 zinc . . . . . . . . . . . 0.001 lead . . . . . . . . . , . 0.039 sulphur . . . . . . ~ . . 0,16 Average quality characteristics of metallic copper obtained at the both stages, in percent by weight:
copper . . . . , . . . . . . . 97.18 iron . . . . . . ~ . . . . . . 0,22 zinc . . . . . . . . . . . . . 0.001 lead . . . . . . . . . . . , . 0,019 sulphur . . . . . . . . . . . 0.16 The refined metallic copper contained, in percent by weight:
copper . . . . . . . . . . . . 99.2 iron . . . . . , . . . . . . . 0.01 zinc less than . . . . . . , . 0.001 lead . . . . . . . . . . . , . 0.001 sulphur . . , ~ , . . . . . . 0,02 The reco~ery of copper from the concentrate was 98.4% of the amount of copper fed for smelting, with 0.~9 rel.% passing into reusable silicate slag.
~asic process parameters of E~ample 1 are given in.
Table 1.

~2~8~38 Table 1 Ba~ic Process characte~istics of sulphide copper concentrates treatment . . . _. _ ~os Characteri~tics Ex.l Ex.2 1. Specific proauction rate at the stage of flash smelting of concentrate and fluges, tons per day per sq,meter of area under smelting and oxidation 50 50 2. ~low rate of o~ygen per ton of con-centrate (lOO~o 2) ~ n0l3 380 335 3. Average composition of the copper- metallic white -containing product obtained after copper matte treatment o* initial material, wt.%
copper 97.18 76.65 iron 0. 22 2 .19 zinc 0,001 0.06 lead 0.019 Ø024 sulphur0.11 20.3 4. Composition of slag with regard to nonferrous metals after its treat-ment with carbonaceous material in electric furnace, wt.% -copper 0.21 0.23 zinc 0.09 0.06 5. ~he content of copper and sulphur in the processed product, in percent of tha-t introduced with sulphide material copper 99.33 99.05 sulphur into gases 99~9 85.4 ~LX~8238 _ 6. Production capacity of electric furnace in terms of slag during its treatment wi-th carbonaceous rnaterial, t/m2 per day 10 8 _ E ample 2 The method is carried ou-t as described in Example 1, except that as a result of smelting white matte ~as pro-duced (with the flow rate of oxygen being 335 nm3 per ton of concentrate)~
The test results are given in Table 1.
As proceeds from Examples 1 and 2, it becomes possib-le, during both flash smelting followed by the production of white matte or during smelting followed by the produc-tion of metallic copper, to ensure: .
- a high content of copper in metallic copper and o~ zinc in zinc sublimates at a low degree of reduction o~ iron to metallic iron;
- a low content of copper and zinc in the depleted slag resultant from the treatment o~ melt with a solid carbonaceous material;
- a low degree o~ iron transition into metallic copper, which brings about a fusion temperature close to the fusion temperature o~ pure metallic copper (1083C~.

- 43 ~

~IL2~323~3 Such combination of proces~ characteristics permits effective processing of the sulphide copper- nnd sulphide copper-zinc concentrPtes to be carried out with resultant production of metallic copper, zinc sublimates and of slags poor in copper and zinc after performing oxidizing r~asting and ~elting of the above-mentioned sulphide con-centrates.
Exam~le 3 Initial sulphide copper-zinc concentrate, containing the following components in percent by weight:
copper . . . . . . . . . . . 2200~
zinc . . . . . . . . . . . . ~69 lead . . . ~ . , , . . , . , 1.59 iron . . . . . . . . . . ~ 24.41 sulphur . . . . . . . . . . 32.71 silicon dioxide . . . . . . .1.25 calci~m o~ide . . . . . ~ . . 1.01 magnesium oæide . . . . . ~ . 0.05 aluminium oxide . . . . . . . O.03, was mixed with limestone crushed to a particle si~e of minus 1 mm as calculated in terms of pure CaO 15,3% by weight of the initial sulphide material. The resultant mixture was dried until its humidity was about lwt.%.
The dried charge ~as continuously fed into a miæer to be mixed therein with continuously fed reusable dust ob-tained by cleaning of gases during flash smelting. The charge and dust mixture was suspended in a flow of com-mercial oxygen by means of an oxygen burner so as to be continuously fed into the melting chamber. Specific load in terms of the charge in the melting zone was 50 tons p~r sq. meter per day. Under the influence of high tempe-rature, the charge was ignited in the presence of oxygen, then melted to produce a dispersed mixture of highly basic slag and metallic copper. Directly under the flame in the zone of flash smelting, the dispersed mixture was separated into two layers in accordance with specific weights, of which one was metallic copper and the other molten slag. The resultant metallic copper and slag were continuously fed into an electric furnace, while gases produced during flash smelting were directed to treat-ment to be cleaned of dust. The collected dust was reused for flash smelting of the concentrate. The highly basic slag incoming to the electric furnace, containing about 5.5 Wt~,' copper, 12 wt.% zinc, 2.2 wt.% lead, was treated with coal which was-intermittently-introduced in an amount of 33 kg per ton of concentrate as calculated in terms of pure carbon (in addition to the stoichiometric required for reduction of copper and zinc oxides, it amounted to 12.7 kg per ton of concentrate). The copper oxide from t'ne molten highly basic slag was reduced b~ -means of carbon to metallic copper which was deposited at the bottom of the electric furnace, with metallic lead partially passing thereinto. Zinc oxide was reduced to metal which evaporized and then was removed together with - ~5 -~21~32:~8 the residual amount of lead from the electr:ic furnace in th~ form of vapour-ga~ phase so as to he ther~after col-lected after its oxidation in the form of oxidized zinc sublimates. The yield of zinc (in the form of sublimates) was 48 kg~m .h. Metallic copper and depleted slag were discharged frorn the electric furnace, whereupon the slag was cooled in the air at a rate of 2.5 deg. per min. ~'he slag discharged from the electric furnace contained 0.46%
by weight of copper and 0.75% by weight of zinc. In other words, the waste of these metals with slag amount, respec-tively, to 1.13 and 4.66% of the amount of copper and zinc fed for sme]ting.
The test results are given in Table 2.
Example 4 A method was carried out as described in Example 3, except that silicate flux in the form of quartz sand (fraction minus-0.5 mm) was additionaly introduced into the charge in--the presence~of basic fluxes as calculated for pure silicon dioxide amounting to 0.45% by weight of the initial sulphide material, with non~errous metals being -finally recovered from the depleted slag a~ter spontanèous disintegration thereof by a flotation method.
The test results are given in Table 2.
Example 5 A method was carried out as described in Example 3, except that silicate flux in the form of silicate slag .~

~L2~ 3~3 (with a ~raction o~ minus 10 rnm) was introduced into the highly basic mol-ten sla~ obtairled during flnsh Hmelting of the initi~l sulphide copper-zinc concentrate. 'rhe sili-cate slag resu]tant from refining of metallic copper (36.3wt.% copper, 0,3wt.~o zinc, 9.2wt.~o iron, 0.9wt.% cal-cium oxide, 43.5w-t.~Yo silicon dioæide) wqs taken in an amount of 1.3% by weight o~ the initial sulphide concent-rate~ wi-th no~fe~rous metals being finally recovered from the depleted slag after its spon-taneous disintegration by flota-tion methods.
The test results are given in Table 2, ~ able 2 Test Results on the Study of ~itions Occur-ring During Spontaneous Disintegration of Slags No~ Composition of Depleted Rate of cooling Ef~ect of Ex. Slag, wt.~o at which depleted sponta-Silicon Zinc Cop- slag is cooled to neous dis-dioæide er a temperature o~ integra -P -its solidifica- tion of tion,deg.per min slag (~es _ _ or no~

3 293 C,75 0,46 2,5 No 4 3,1 0,76 0'45 2,5 Ye~
5 4~7 0,77 0943 2,5 Yes 610,3 0,73 '4~ 2,5 Yes 714z5 0~75 0,49 2,5 Yes 16,5 0,77 0~4~ 2,5 Yes ~ ~8~,3 ~

9 1093 . 0,730~47 60 Yes 10 10,3 0,740,48 70 No 11 10,3 0,740,4~ 0,5 Yec 12 20,0 o,~o949 2,5 Ye~

~able 2 (Cont'd) Te~t Results on the Study o~ Coditions Occur-ring During Spontaneous Di~integration of Sla~s ~o~Production Recove- Content Content Losces of Addi-Ex.rate at the ry of of cop- of co~ copper tional stage dur- copper per in per in with slag recovery ing which ~rom flota- ~lota- a~ter flo- in kg per zinc is re- ~elf- tion tion tation, 1000 kg - covered disin- concen tail- ~0 of the o~ con-~rom2slag, tegrat- trate, ings, initial centrate kg/m .hour ed slag wt.~o wt~50 during ~lota-tion9%

3 48,0 not reco- - ~ 1,13 ~ered 4 46,.2 67 B,l 0,15 0,3~
5 46,5 72 9,3 0,12 0,30 copper:
10,8 6 53~7 73 15,2 0,13 0,38 zinc:
. 3~,2

7 48,1 71 14,9 0,14 0,35 copper:
- 1,5 ; 8 42~0 75 15,1 0,12 ,34 9 5398 b6 11~3 0,16 0,47 zinc:
38,2 - 4~ -~:

~L~1823~3 .. . .

10 53,9 Not reco- - - 1,34 2inc:
vered 38,1 11 53,5 77 17,30,110,33Zinc:
38,0 12 40,5 82 11,70,090,26 .. .. ~
_ample 6 A method was carried out as described in Example 3, except that ~inc-bearing silicate slag ~with a fraction minus 10 mm~ was used as the silicate flux introduced into the highly basic molten slag obtained during flash smelt-ing of the initial sulphide concentrate. The silicate slag, containing 20.0 wt.% zinc, 0.64 wt.% copper, 20.8 wt.% iron, 14.7 wt.% calcium oxide and 21.65 wt.%
silicon dioxide, was taken in an amount as calculated for pure silicon dioxide constituting 4.35% by weight of the initial sulphide concentrate, with nonferrous metals being finally-~recovered from the depleted-slag after its spontaneous disintegration by a flotation method.
The test results are given in Table 2.
Example 7 A method was carried out as described in Example 3, except that depleted silicate slag (with a fraction of minus 10 mm) was used as the silicate flux introduced into ., ,.
:

23~3 the deple-ted highly 'ba~:Lc molten slQK. rl'he sllio,ate slag, containing 0.76w-t.~o copper, 39.06wt~% iron, 33.38wt.%
~ilicon dioxide, 8.72wt.~'~ calcium oxide wa~ taken in an amount as calculated in terms o~ pure ~ilicon dioxide constituting 6.6% by weight of the initial sulphide con centrate, v~ith nonferrous metals being finally recovered from the depletéd 31ag after its spontaneous di~i~tegra-tion by a flotation methodO
The test results are given in ~able 2.
Exa~!ple 8 A method was carried out as described in Example 3, e~cept that silicate flux in the form of quartz sand (with a ~raction of minus 0.5 mm) was additionally intro-, duced into the charge in the presence of basic fluxes in an amount as calculated in terms of pure silicon dioxide con3tituting 7.7% by weight of the initial sulphide mate-rial, with nonferrous metals being finally recov,ered from the depleted slag a~ter its spontaneous disintegration by a ~lotation method.
The test results are given in Table 2.
E~ 2 A method was carried out as described in Example 6, except that the depleted slag was cooled to a temperature of its complete solidification at a rate of 60 deg~ per min.
The test results are given in Table 2.

~Z~8~3~

_ample 10 A method was carried out as described in Exarnple 6, except that the depleted slay was cooled to a temperature of its complete solidification at a rate of 70 deg. per min.
Final recovery of nonferrous metals from the depleted slag by flotation was not carried out because of no e~fect of spontaneous disintegration of the depleted slag being observed.
The test results are given in Table 2.
Example 11 A method was carried out as described in Example 6, except that the depleted slag was cooled to a temperature of its complete solidification at a rate o-f 0.5 deg. per min.
The test results are given in Table 2.
Example 12 A method was carried out as described in Example 3, exce~t that silicate flux in the form o* quart2 sand (with a fraction of minus 0.5 mm) was additionally introduced into the charge for flash smelting of sulphide copper-zinc concentrat~ in the presence of basic fluxes in an amount as calculated in terms of pure silicon dioxide constituting 9.6% by weight of the initial sulphide material, with non-ferrous metals being finally recovered from the depleted slag after its spontaneous disintegration by a flotation , ,.

~IL21~ 38 method.
A number of difficulties were encoun-tered at the stage of flash smelting and during treatment of highly basic molten slag with a solid carbonaceous material, which were caused by a high temperature of fusion of the highly basic slag containing 20% by weight of silicon dioxide. Thus, thorough control of the process tempera-ture conditions were to be observed.
The test results are given in Table 2.
As is seen from Table 2, the effect of spontaneous disintegration of slag is observed with the content of silicon dioxide therein being over ~/O by weight. However, any increase in this content above 16% by weight is un-desirable as it causes difficulties in running the pro-cess. Therefore, the higher foundary in the content of silicon dioxide in the depleted slag had been found to be 16% by weight. It proceeds from Table 2 that spon-taneous disintegration of;slag occurs with the rate of its cooling until complete solidification not exceeding 60 deg. per min. However, to perform the cooling of slag at the rate below-0.5 deg. per min~is undesirable since the melt cooling time will exceed 10 hours.
In this case, silicate flux can be introduced either into the charge used for flash smelting of concen-trate together with the basic flux, or into the highly basic molten slag resultant from flash smelting and fed further treatment, or else into the highly basic molten slag from .~ . i ~Z1~3~3 which nonferrous met~l~ have been recovered, ~he ~ilic~te ~lag re~ultant from re~ining of metallic copper can be used as the silicate fluæ; the ~ilicate flux poor in non-~errous me-tal~ can al~o be u~ed along with a ~ilicate flux normally utilized for the purpose.
ExamPl_~a~
Sulphide copper-zinc concentrate with a particle size of minus 74 mcm, containing 8.23wt.% copper, 1.53wt.% lead, 18.69wt.% zinc, 21c41wt.% iron, 34.22wt.% ~ulphur, 6 82wt.%
silicon dioxide, 3.02wt.% calcium o~ide, 0 35wt~% alumi-nium oxide, 0~03wt.% magnesium oxide, was mixed with hyd-rated lime in an amount, as calculated for pure CaO, of 16 5% by weight of the initial sulphide materiQl. The re-sultant charge was then dried to ~ humidity of about 1~ by weight.
~ he dried charge ~as continuously fed into a mixer to be intermixed therein with reusable dust continuously fed into the mixer a~ter being collected in the proces~ o~
cleaning gases during flash smelting. q`he mi~ture of charge and dust was suspended in ~ flow o~ commercial oxygen to be thereby tran~ported b~ means of an oxygen burner into the melting chamber. Under the in~luence of hlgh temperature, the charge was ignited ln the pre~ence of oxygen and then melted to produce a di~persed mi~ture o~ highly ba~ic ~lag and metallic copper. The mixture was then separated in ~ccordance with speci~ic weight~ into ~2i8~:3~

metallic copper and highly b~ic molten slag, ~e ~eparated metalli.c copper was discharged ancl the molten ~lag was poured into the elec-tri.c furnace. The gase~ produced dur-ing flash smelting of the charge were purified to give off dust which was then collected and returned for use in the flaeh smelting process. The pourin~ of slag wa~ terminated when it~ quantity reached 10 tons~ At that moment the 91ag had the ~'ollowing composition, in percent by weight:
copper . . . . . . . . . 2.12 zinc . . ~ . . . . . . .16.92 iron . . . . . . . . O .22069 silicon dioxide . . . ~ 7.22 calcium oxide . . . . ~20.69 aluminium o~ide . . ~ . 0.37 Once the slag pouring process was over, coke was charged in batches o~ 87.8 kg of carbon each (25% of the stoichiometric quantity required for the reduction o~
copper, zinc and lead) onto the sur~ace of the melt in the electric furnace. Each succeeding portion of coke was charged only after the preceding one was completely used up in the reaction lasting for about 30 min. Immediately be~ore charging carbonaceous material, assays of slag and metallic ~opper were made to determine the content of valua~le components thereln, ~he zinc and lead vapours resu]tant frQm treatment of -the highly basic slag were di~charged ~rom the elec~ric ~urnace, whereupon they were .

~Z~L~3;2,3~3 oxidize~ in the air and then col]ected :in the fo~n o~
zinc and lead oxide~. After copper, zinc and lead were recovered from the slag~ it was discharged *rom the elec-tric furnace.
The test results are given in '~able ~.
~ rom the obtained data it follows thnt in c~e o~
stoichiometric con~umption of carbon the waste of non-~errous metals with slag remains rather high, while the excess of carbon above stiochiometric by 20-25% makes it possible to reduce losses of these metals 2-3.5-fold without impairing the quality of metallic copper. It should be observed that in the given Example average ~pecific rate of recovery of zinc from slag until its residual con~ent in the latter w~s 0.55% by weight, amounted to 127 kg of zinc per sq.meter per hour (minimum amount was ~1 and maximum 177 kg/m2.hr). And this was des-pite the f~ct that the reduction of oxides wa~ artifi-cially retarded by reducing the rate o~ charging of the solid c~rbonaceous material o~to the ~urface of the melt.

: ~ .
. .

~ 55 -~ .

82'3~3 ~ able_~
Influence o~ Carbon Con~l~ption on Proce~s Characteristics Obtained During Treatment of Highly Ba~ic Slag .
Nos Carbon con~umption,% of the .qtoichiometric quantity required for reduction of zinc, copper and lead oxides ___ . _ . . ~ . , .~ .. . ._ - ~ . ..... ~__ 4 5 6 Slag composition, wt.%
copper 2.12 1.67 1.29 0.90 zinc 16.92 13~9 9~50 6.25 I lead 1.65 1.03 0~57 0.28 iron 22069 24.19 25.61 26.96 : ~ilicon dioxide7.22 7.82 8.37 8.77 calcium oxide :20.7 22.2 23.3 24.6 : Composition of metallic copper obtai~ed during treatment of hi~hly basic melt, wt.%
copper - 91.9 9102 -~9.0 2 zinc - 0.33 0.84 1.27 lead - 2.49 2.80 4.24 iron _ 0~23 0.23 0.26 .. ~ .. . _ . _ . . _ _ . __ . . _. . . . _ _ : 3 Waste of copper with slag,~o 2.50 1.90 1.39 0.92 4 ~laste of z mc with slag9% 90.5 69.8 43.6 24.8 :
~ .' :
r ~ .
-- :~ D

~`
~: :

12~L823~

'l'able ~ (Cont'd) Inf:Luence of Carbon Con~umption on Proce~
Characteristics Obtained During ~reatment of Highly Ba~ic Slag ~ _ _ . . . . __ _ __ . . . .
No~ Carbon consumption, % oP the stoichiometric quanti~y required Por reduction oP zinc, copper and lead o~i.de~
___ .,_,,, . _ . _ . _ . ~
7 _ _ ~ _ . . 9 . . _ , . . _ . . _ O~ 55 0~24 0,16 3~27 855 0~38 0~09 0~03 -0~01 28 ~ 22 29~ 33 29~ 30 -9~ 12 9 ~ 32 -9~ 32 - 25~6 26~7 26~8 ~5~9 85Q5 86~3.
2 1~18 O 95 OJ69

8~29 7~60 5~62 0~27 1,~;5 2,11 -- -- ' -- ---- _ , -- . . .____ _ __ ~ __ _ __ _____ _ 3 a5S.4 o92~ 6 4 8947 2~28 1~58 - - _. .

~: :
' ~ 57 -~1L2~

~xample 1~
Initial slllphide copper-zinc concentrate, contain-ing 22.08wt.% copper~ 8.69wt.% zinc, 1.59wt.% lead, 24.41wt~% iron, 32~71wto% sulphur~ 1.25wt.% ~ilicon dio~i-de, 1 Olwt.% calcium oxide, 0 05wt % magnesium oxide, 0.03wt.% ~lumi~ium o~ide, W~9 mixed with limeqtone pre-crushed to a particle size of minus 1 mm in ~n amount a~
calculated in terms of pure CaO constituting 15.3% by weight of the initial sulphide material. ~he charge thus prepared was dried until its humidity was about lwt.%
(see Example ~)~
The resultant dry charge was continuously fed into a mi~er to be mi~ed therein with reusable dust continuou~ly admitted thereinto a-fter being collected in the process of cleaning gases during ~lash smelting. The mixture o~
charge and reusable dust wa~ suspended in a flow of com-mercial o~ygen so as to be continuously fed by means of an oxygen burner into the melting chamber. Specific load of the melting zo~e was 61 tons per 1 sq.meter per day.
Under the in~luence of high temperature, the charge was ignited and then melted to produce a dispersed mixture o~
high~y ba~ic ~lag and metallic copper. The dispersed mi~ture was then ~eparated in accordance with specific weights into metallic copper and highly basic molten slag to be continuou~ly fed into the electric ~urnace. ~he gases produced during fl~sh ~melting were purified to give off dust which was the returned for use in the ~2 i ~'3 ~

proce~ of flash qmelting. ~lhe highly basic ~qlag admitted into t`he electric furnace wa~ treated with co~l. The con-sumption of coal per unit of time was altered so QS to enqure permanent high concentration of zinc and lead vapours in the electric furnaceA In order -to determine the concentration of ~inc and lead vapours in the elec-tric furnace, the vapour-gas phase WQ~q gQmma-rayed with two ~luxes of di~ferent energy. The change in the con-cen-tration o~ zinc ~nd lead vapourq was determined in accordance with the chan~e in the intenæity of radiation, whereby the ~eeding rate of the reducing agent was regu-lated. ~he consumption of coal was 30 kg per each ton of concentrate calcul~ted for pure carbon (in addition to stoichiome-tric quantity for reduc-tion of zinc and copper o~ides, it amounted to 9.7 kg per ton of conce~tr~te).
~he copper oxide from the highly baqic molten 61ag was reduced to metallic copper which settled down at the bot-tom of the electric furnace. ~he vapour o* metallic zinc produced during reduction of zinc oxide was dischargea ~rom the electric furnace, whereupon it was oxidized and then collected in the form of oxidized zinc ~ublimate~.
~he furnace production rate in terms of zinc wa~ 62 kg/m.
~hourO Metallic copper and depleted slag were di~charged from the furnace.
The blister copper produced as described above had the following characteristic~:

' ~q ~L2~ 38 Composltlon, wt.% Content of bli~ter copper re-covered f.rom the concentr~te, %

Copper - ~6.39 97.58 I.ead - 6.21 42.48 Zi~c - 1~45 2.99 Iron - l,lO 1.33 The recovery o~ zinc from the concentrate in the form of zinc sublimates in the electric furnace was 88~9~o; the content of lead in zinc sublim~te~ was 55.3~;
the content of copper in met~llic copper was 98.92%.

~ method was carried out a~ in E~ample 14, except that the m0tallic copper produced during flash æmelting -~ g removed from the process after dividing -the dispersed mi~ture of highly basic slag ~nd metallic copper in ac-cordance with apecific weights, whereupon the metallic copper produced during treatment of the highly ba~io molten ~lae with carbonaceous material was separately removed ~'rom the proce~s.
~h~ bl~ster copper produced as described in Example 15 had the followlng characteristics:
Composition o~ blister Yield of blister copper copper, wto% recovered ~rom the con-centrate, %
Metallic copper produced at the st~ge of reasting Copper 94.09 77.67 b ' , ~ 3 ~

IJe~d 3.54 15~42 Zino 0.07 0~11 Iron 0.14 0.12 Metallic copper produced duri.ng depletion of slag Copper 80.30 21.16 ~ead 9.10 12.38 Zinc 1.08 0.52 Iron 2.00 0.54 Average characteri~tics of metallic copper produced at both stages:
~opper 90.80 98,~3 Lead 4~86 27.~0 Zinc 0.31 0~63 Iron 0.58 0.66 The content of z;nc recovered ~rom the concentrate in the form of oxidi~ed zinc sublimates was 94.54%; the content o~ lead in zinc sublimates wa~ 65.2%; the con-tent o~ copper in met~llic copper ~a~ 98,95%.
A~ een from Examples 14 and 15, during ~eparate t~ppin~ of metallic copper (a~ter ~la~h smelting and a~ter depletion o~ the highly basic melt, with almost the ~me amount o~ copper belng recovered from the concentrate in the fo~m o~ metallic copper t9~ o)~ it~ quality is sub-~tantially impro~ed along with an increase in the contents of zinc and lead in oxidized zinc ~ublimate~.

Claims (10)

The embodiments of the invention, in which an exclusive property or privilege is claimed, are defined as follows:-

1. A method of processing sulphide copper- and/or sulphide copper-zinc concentrates, comprising:
a) flash smelting, in the presence of oxygen, of a mixture of said concentrates and fluxes selected from the group consisting of basic fluxes and silicate fluxes;
thereby producing a dispersed mixture of highly basic molten slag containing copper and zinc oxides, metallic copper or white matte; said initial concentrates and fluxes being fed for smelting in such ratios that permit production of highly basic slag;
b) effecting reduction of said copper and zinc oxi-des, contained in said highly basic molten slag, by means of a solid carbonaceous material with the resultant for-mation of a vapour-gas mixture, metallic copper and im-poverished highly basic slag containing not more than 18 percent by weight of silicon dioxide;
c) oxidizing the resultant vapour-gas mixture, con-taining zinc vapour, to zinc oxide;
d) collecting the resultant zinc oxide.

2., A method as claimed in claim 1, wherein the con-tent of silicon dioxide in the depleted highly basic molten slag is maintained within the range of 3 to 16 per-cent by weight with the use of silicate flux; said depleted highly basic slag being cooled down to a temperature of its complete solidification at a rate of 0.5 to 60 deg.

per minute with the resultant production of self-disin-tegrating material from which conferrous metals are finally recovered.

3. A method as claimed in claim 1, wherein the resultant metallic copper is subjected to refining in the presence of silicate fluxes with a view of obtaining refined copper and silicate slag.

4. A method as claimed in claim 1, wherein the metallic copper obtained during flash smelting of the concentrates in the presence of oxygen and the metallic copper obtained during treatment of highly basic molten slag with a carbonaceous material are discharged separately.

5. A method as claimed in claim 1, wherein the silicate slag produced during refining of metallic copper is used as the silicate flux.

6. A method as claimed in claim 2, wherein the silicate flux is fed for use in flash smelting of the concentrate.

7. A method as claimed in claim 2, wherein the silicate flux is introduced into the molten highly basic slag produced during flash smelting of the concentrate.

8. A method as claimed in claim 2, wherein the silicate flux is introduced into the depleted highly basic molten slag poor in nonferrous metals.

9. A method as claimed in claim 8, wherein the depleted silicate slag poor in nonferrous metals is used as the silicate flux.

10. A method as claimed in claim 1, wherein said basic flux has calcium as a major structure-determining component.