CA2055427A1

CA2055427A1 - Process for the treatment of zinc sulphide containing ores and/or concentrates

Info

Publication number: CA2055427A1
Application number: CA002055427A
Authority: CA
Inventors: Murry C. Robinson; Donald R. Spink; Kim D. Nguyen
Original assignee: Materials Concepts Research Ltd; University of Waterloo
Current assignee: Materials Concepts Research Ltd; University of Waterloo
Priority date: 1989-05-03
Filing date: 1990-04-25
Publication date: 1990-11-04
Also published as: AU5546690A; WO1990013679A1; GB8910083D0

Abstract

A process is described for the partial desulphurization roasting of a variety of zinc bearing sulphide ores or concentrates by adjusting the roaster temperature and residence time in the roaster thus providing an oxygen deficient atmosphere so that the required amount of sulphide retention is maintained. The obtained calcine or calcines are subsequently subject to various chemical and/or physical separation process steps to separate the unreacted sulphides which are, according to certain embodiments of the process, then dead roasted and treated for zinc recovery. In another embodiment, the separated unreacted sulphides are again partially desulphurized in an oxygen deficient atmosphere for subsequent treatment to recover zinc, lead and precious metals. The chemical separation process steps include aqueous sulphur dioxide treatment, dilute sulphuric acid solution treatment and treatment with dilute sulphuric acid containing sulphur dioxide in solution, all conducted in the temperature range of 50 to 75·C. Physical separation process steps include one or more of the flotation and magnetic separation techniques although other physical separation steps may also apply in some instances.

Description

: ~ ~ 2o~3~27 WO90/l3679 - 1 - PC~/CA90/00130 ' -- Title~ ~ Novel Proces~ Por The Treatment Of Zinc Sulphide Containing Ores And/Or Concentrates FIELD OF T~E INVENTION
; 5 Thiq invention deals with racovering zinc from zinc and iron-bearing sulphides which flre either in the rorm of conventional zinc sulphide concentrate~, low-grade zinc ~ulphlde aonc~ntr~tos, or in the orm o~ bulk zinc ~ulphldo concentrnto~, the latter which consiq~ o~ zinc;
iron, and non-f0rrou~ m0tal aulphide minerals including lead sulphLdes and precious metals ln complex form.
Additionally, the accompanying valuable non-ferrous metals ~,:1 are recovered, which include lead, copper, cadmium and ~S~t sil~er ~' 15 ~h~oughout the world, convontional proce~ing of oraD cont~ining zlnc 0ulphide ~a tho moat valuable con~ti~uont but includln~ ~ub~tanti~l Mmounto e lroll Ll~
w~ 10 1~!l0er ~mourlt~ o~ oth~s~ b~lo~ mot~l ~ul~hlcle~
ln~olvou tho ~ona~ntr~tion o~ thQ ~ln~ uulphl~ aomperl0nt by ~lo~tlon tai~hnl~uo~ thu~ ~ooultln~ .Ln tho ~o~matlon o~
zlna conc~ntr~t~ aont~ining ~ome 4S~ to 650 zinc a~ zinc 0ulphido wlth tha chle~ Lmpurity baing iron in the form of iron sulphlde~. Such con~entional zinc concentrates are generally treated ln a daad ro~ting step wherein the re~ulting cnlcine contalns zinc oxide as a ma~or ~' constituent~ however, a ~ubstantial portion of the zinc will generally be in the form of zinc ferrites.
! This calcine is then leached in what can be referred to as a neutral sulphuric acid leach wherein the zinc oxide content i9 readily dissolved but without di~olving the zinc present as zinc ferrite. The separated zinc solution is then purified and electrolyzed to produce zinc metal with the ma~or portion of the spent electrolyte recycled to the neutral leaching steps. The undissolved le~ch residue containing substantial amounts of zinc as zinc ferrite was in earlier days disposed of, usually to 4~,, wo 90/l~79 - 2 - PCT/CA90/00l30 a s~ockpile.
In more reeent years, the leach residueq have been directed to a hot conc~trated sulphuric acid 301ution to dissolv~ both the zinc and iron pr~sent in th~
S zin~ f~rrite. Procesaes have be~n d~veloped and are now used in practice to treat the ~o-di~solv2d zinc and iron ~ulphate3 in order to precipitate the iron rather than the ~inc. Two such proce~e3 are the ~arosite and goethlte process~s, both o~ which provide means to recover the zinc ~g ~ zlnc ~ulphate ~olution ~hich then c~n be tre~tod to produce m0talllc zinc. Un~ortunately the iron-contalning re~idu~s nro qulto voluminou~ and c~n contaln toxlc ~ùbst~ncaa which rosult in envixonmental as well as sconomlcal problem~.
In certain circumstances, zinc sulphide concentrates may be too poor in grade to warr~nt conventional roMstlng followed by conventional zinc rqflning steps. ~n th~so in~t~nce~, tho zlnc concontr~tion migh~ b~ too low, aar in thc r~nye o~ 30~-493 zinc or tho 20 iron concontr~tlon might be too high, unr ln thff r~ngo o~
103-200 iron ~nd ln m3ny in~tanc~ both o~ th~o ~cto~
may ~pply. 8Om~ cxampl~ Oe tha ~oa~lbLllty o~ protucln~
~uah t~p~ o~ ~lna con~ntx~tQ~ ~ollo~ n on~ c~00, nn or~body m4y hA~ b~n minnd An~ mlllad to tho ~x~ont thAt 25 only ~ lo~r ~r~d~ un~conomical zlnc concentrate could be pxoducad. In ~nother l~stance, ~n orebody mlght be ~uch thnt only a lo~or grnde 21nc concentrnte could be produced in ~ub~tantinl quantitieo. In a thlrd situaton, thu ~lotation proce~ln~ might be ~uch thnt n mlddling~
product could be produeed that would be of an unsatisfaetory grade so that ~ccordlng to present teehniques, it would hsve to be di~carded, thus resulting in zinc lo~se On the other hand, our novel teehnology ean reduce the formation of zinc ferrite in the produeed ; ~ caleina to 3ueh a small fraetion that the need for such ~ ~ expensive, troubles30me and environmentally unde~irable 205a~27 WO90/13679 PCT/CA90/00130 ~-,: ~
~ 3 treatment steps as in conventional processing can be eliminated; other beneficial results are also obtained in the overall zinc refining proce~sing.
The novel technology also has a great deal of 5 flexibility in that it can be used beneficially for treating a wide variety of zinc concentrates.
Por example, optlon~lly, our novel technology cAn be applied to up~rado low-grnde zinc concentrates so '- th~t ~ zlnc conaentr~to aould bo produced o~ equal or 10 better gr~de to that now belng eco~omically processed by conventional means. Alternatively, a lower grade zinc sulphide flotation concentrate might be produced at a much higher yield; pretreatment of such a concentrate using our " novel technology would readlly upgrade th0 concentrate to 1 15 convention~l level9 with a con9equ0ntl~ gr~ter productlon ; of zinc.
In ~tlll othar clrcum~t~na~ lnc ~ulphLdQ
aont~ining er~ bodlaD c~n ba ao aompla~ ln phynlG~l natu~
~h~t ~h~00 ar~ hnr 1na~pablo o~ b~ln~ a~cl by 20 conv~n~lonnl ma~n~ to p~o~ua~ la~blo p~oduG~ or ~ o~
such a n~turo th~t ~ ~lne sulphide eont~ining orffbody is tro~ted in A mAnnor that results in the production o~ a bulk concentrate that is low in zinc, ~ay in the 30~-40~
ran~e, high in iron, ~ay in the 10~20~ range, hlgh in 25 lead, say in the 10~-70~ range ~nd sulphur in the range o~
30~-36~. Many of such bulk concentrates are shipped to Imperial Smelting Furnaces resulting in low revenues for recovery of the contained zinc, lead, other base metals and precious metal values contained therein. Depending 30 upon the economical climate at any given time, such bulk -~
concentrates might be rejected as being impractical to treat. -Our novel process technology has the potential to circumvent the difficulties described in the foregoing paragraph~ and will be described later in thiq disclosure.
Our process involves a two stage roast whereln the first caleine can contain part or nearly all of the 2 0 S ~ 4 -iron ln an easily acid soluble iron oxide form leaving most of the zinc and other ba~e metalq pre~ent as sulphide3. Such a ~irst stage roa8t also produce~ a higher purity ~ulphur dioxide-containing ga~ than that conventionally produ~d which may bs useable a~ 3uch or more 9a8ily treated to produ~e 8ulphurl~ acid. The iron oxide aomponent in the first calcine can readily be dis-~olved along with any base metalg co-oxidized in a warm ~queous ~ulphur dioxide solution or ~ warm dllute sulphuric ~cid solution or a combination o~ ~ warm dilute sulphuric acid solutlon ~nd ~gueous ~ulphur diaxlde, thus leavlng ~ le~ch r~idue th~t 1~ hl~her ln zinc content and muah low~r ln lron content. ~h~ le~chlng te~p~rature would norm~lly be ln the r~nge o~ 50C and 75C And prefer~nti~lly between 60C ~nd 70C. The agueous sulphur dioxide ~olutlon would be preferentially At or close to ~atur~tion ~nd, 1~ sulphuric acid solution were used lt ; would be in the range o~ 2-5 wt~ H2S0~ ~nd pre~er~bly 3-4 wt3 H2S0~. Where ~ ~ulphuric ncld ~nd ~queoun S02 nolution ~re combined, tho ~ulphurlo ~cid ~olutlon ~ould ~tlll b~
bet~e~n 2-S ~tb H2S0~ and the aquaou~ S0~ di~nolv~d th0rQln ~ould r~n~o Srom ~ mlnor ~ddltlon to claao to s~tuxation.
Sp~nt ~loatxolyta mny ~a~vo ~a n ~ub~tltuto ~or ~ul~huric ~ld ln ~holo o~ ln pnPt ln ~om~ ln~t~noa~. ~n ~omo 1 2g ln~t~naa~ th~ la~ching pulp d~n01ty ~ ould be bet~r~qn 60 .~ ~nd 120 ~ nd the l~flchlng period would be three hours : or 108~. However, ~t~ged le~chlng could result in broader pulp densitie~ nnd longer lo~ahing times.
~ n ~ome in~tanoes~ p~rticularly when tho zlnc sulphide i~ pr~dominantly ~phaleritlc rathor th~n marmatitic in nature, optionally physical separation techniques such as flotation and/or magnetic methods might be employed either alone or in combination with the chffmical di~solutlon methods already described above to separate the oxides from the unreacted sulphides and thus provide a useful separation technique. The~e phy~ical 3eparation tech~iques might be applied to the partially . ' , ` ,., ,: ~' . . .

~ 2B~

WO90/l3679 PCT/CA90/00l30 _ 5 _ desulphurized calcines and/or to upgrade the leach residue produced.
In any event, the leach residue or physical resldue remaining after the iron oxide ~raction has been separated, provide-~ the feed to the second stage roasting ~tep.
In one èm~odiment usin~ conventional zinc ~ulphide concentr~to~ or low gr4de zinc sulphide concentrates, the leach re0idue or physlcal residue containing the unreacted sulphides pre~ent in the first calcine would be sub~ected to a second stage conventional dead roast to produce a calcine which would contain less zinc ferrite than that which would bo convontionally produced. ~he ~lr~t ~t~qe roa~t would b~ conduc~ed in the prouence o~ ~n oxyg~n baaring g~o wh~oin ~ do~roi~ o~
~ulphlde ~ulphur rntention ln tho ~alclna la m~int~ln~cl ~y controllln~ tho oxy~on elow r~ta ~nd/or th~ r~0id~nce ~lm~
o~ ~hQ ~ad m~torl~l ln ~h~ ~on~t~ thu~ ultin~ In an oxyg~n du~ nt ~tmo,~ph~re. 'rh~ p~ra~ntago Oe ~ulphur removed wlll be ~ ~unatlon oP the iron content in the concentrate to bia tre~ted and the degree of iron removal denired uaing the partiAl deaulphurizatlon roAst and loaching and/or ph~0ic~1 ~oparatlon ~tep~. The partially de~ulphurized roast will norm~lly contain between 15~ and 27~ sulphur ~nd preferentially between 20~ and 25~ sulphur by controlling the retention time of the ore or concentrate in the oxygen deficient atmosphere. The first stage roast would be conducted in the temperature range of 650C to below the sintering temperature and preferably between 700C and 1050C and more preferably between 850C
and 100~C. Under these specific conditions, zinc ferrite formation is reduced in the dead roasting step to the degree desirable for any given application. In thi~
context, the words '~partially de~,ulphurizedl' re~ers to partial oxidation of the contained metal sulphides.
As an alternative embodiment, in the case o~
bulk concentrate treatment, two oxy~en dePicient roa~s WO90/l3679 PCT/CA90/00130 Q 9~~ight be conduct~d prior to the ~inal ~ulphide conversion operation. In thi~ in~tance, the fir~t stage roast would be conducted under ths condition~ described in the pr~vious paragraphs to produce a calcine wherein mo~t of 5 the iron sulphidas ars converted to an ea~ily solubl2 iron o~ide form ~uch that all the ~oluble oxide~ can be lsached u~ing ons of the leaching and/or physical separation technique~ herelnbefore describsd leaving the bulk o~ the zinc ~ulphide~, lead sulphides, other base ~etal~
10 sulphides and contained precious metal~ in the separated sulph~de contalnlng residue.
~hi~ sop~rated uulphlde-cont~ining resldue is then conceived to be roastod ln a ~econd stage ro~ster ~l~o u~lng dn oxygnn de~icient atmosphere to the extent 15 wherein more than 80i and pre~erably more than 90~ oi' the contained zlnc sulphide is oxidized to form zinc oxide le~ving the lead sulphldes e9sentlally unreacted. This second calclne would then be leachod u~ing a neutral leach or one o~ the le~chin~ t~chniqueo prevlou~ly descrlbed 20 which a~tor li~uld-solld aop4r~tlon ~ould l~avo a loach realdu~ that i~ prim~rily le~d aulphldo but rich in preciou~ motal~ whlch c~uld bq e~d dlr~c-tly to conv~n~ion~ d ~m~lt~r. Ylotation tochnlquen ml$ht ~lt~rn~lv01y b~ u~d to ffap~rat~ o ~ol~tivaly hlgh g~AdO
25 ~in~ oxlda ~rom the la~d ~ulphid~ ~r~c~ion. ~n thl~
ln~t~nca th~ prc~arr0d ro~ating tqmporature rango would be 6so4c to 850C but pre~erably between 675C to 750C for each o~ th0 two ~t~g~ ro~tlng oper~tions. ~rior to fe~ding thq le~d sulphldo concentrat~ to the-lead ~m~lter, 30 ch~mlcal or physlcal 0epar~tlon technlque~ mlght be employed to ~eparate the preclous metals ~rom the lead sulphide. These techniques will be described later in this disclosure.
By the usa of the techniques described above, it 35 should be possLble to use mining and milling techniques on complex metal ~ulphide depo~itq to recoYer a much higher - ~
percentage of metal sulphides such as zinc, copper, lead, : ~ , f~. 2~33~27 cadmium, silvar and gold from the mineral deposits than is presently practiced and could, in addition, make tha treatmen~ of some of the now dormant orebodies of such a ~ature into prof$table operations. This ls because lower grade hulk concentr~tes could be upgrad~d.
Some of the advantages dlscussed abov~ can be ~chleved by ù01ng the proce~ for the recovery of zinc bearin~ ore~ or conc0ntr~te~ aompri01ng the steps of:
,~a) roastlng the ore or concentrate in the presence of an oxygen-containing ga9 wherein sulphur retention is maintained at a level of lS~ or higher by controlling the residence time of the ore or concentrate ln an oxygen deficient atmosphere ln ~ ~o~ter u~ing n t~mp~ture ln tho ranqe Oe 6S0C to lOgOC ln ord~r to pra~or~ntl~lly oxidlzo the ixon ~th~ th~n th~
zlnc An~ othor b~ mot~l conot.Ltu~ne~ ancl ~hu~
roduco o~ ollmln~tq ~inc eo~rl~o ~orm~tlon ~b) lanchln.~ tho parti~lly ~ou.lphurl~od roaot~d or~ or aonc~ntrata, u~lng ~ lixi~l~nt ~elected ~rom thc group conslsting of aqueou~ sulphur dioxide 001utlon, a 2-5 wt~ ~ulphurlc acid .~olution, and a 2-5 wt~ sulphurlc acid solution containing 001ubllized ~ulphur dioxlde, at a temperature between 50C and 75C in a pH range of 1.0 to 2.5 for one to three hours in order to preferentially dissolve the iron oxide thus formed as well as base metals that have co~
oxidized to obtain a leachate and a leach residue (c) subjecting the leach slurry to liquid-solid . .
separation ~d) treating the leachate to separate the .. . .
dissolved iron and other dissolved oxides of base metal~ from the remalning sulphides, thus leaving a leach resldue which i9 richer in 2inc ~ulphide content and depleted in iron content ~0 3 3 ~ - 8 - pcT/cAso/ool3o which dsad roasted as in conventional practice with the rssultant reduction or effective elimination of zinc ferrite formation or ~ii) if the original concentrate iq rich ln lead, 3uch as in certain bulk concentrataa, 3ub~ecting sa1d leach residue to a ~econd _tage partial de~ulphu~lz~tion ro~t u~lng controlled retention time in the t~mper~ture xange o~ 650C to 850C
~uch thnt the rotained sulphur content ls 8% or higher depending on the percent of lead present, to produce a second stage calcine whlch has a very high zinc oxide content ~nd nubstnntinlly all of the le~d in tho ~ulphide ~o~m, 0ub~ecting s41d saCOnd cAlcin~ to phynic~l ~p4rntlon tachn1qu~ o~ ~o n~utr~l lo~ch or ~o ~n ~qu~ou~
~ulphu~ dloxld~ l~ach undo~ 01~
~ondltlono ~ a~rll~r do~cr1bod~ in o~dar to nop~r~to th~ zinc oxide ~raction f~om the lead ~ulphide fraction, the sulphide-containing re~idue i8 then ~ed directly into n le~d ~elter or ia ~urther trea~ed to ~eparate the lead sulphide from the precious metals.
According to the above deqcription, the process ~ay be applied to low grade zinc 3ulphide concentrates or bulk zinc _ulphide concentrateq or to conventional zinc sulphide concentrates. In the ca-qe of conventional zlnc concentrates these would contain ln the range of 45 to 65%
zinc, 3 to 15% iron and le~3er ~mounts of COppQr, cadmium , ~,j ~
Wo9U~13679 2 0 ~ ~ ~ 2 7 PCT/CA90/00130 and lead, all predominantly in their sulphide form with a variety of other minor i~puritie~ present. In the case of low grade zinc concentrates, the~e would oontain in the range of 30 to 45~ zlnc, lO to 20~ i~on and smaller S amount~ of copper, cadmlum and lead all predomlnantly in th0ir ~ulphide form wlth a vnriety of other mlnor lmpurltle~ present. In the ca~e of bulk zlnc concentrates, th0~0 would cont~ln ln th~ r~ng~ o~ 25 to 40~ zinc, lO to 254 iron~and lO to 250 lead and sm~ller amounts of copper and cadmlum all predomlnantly ln thelr sulphide form wlth a variety of other minor impurities present.
The ~ollowing paragraphs describe detailed technique~ ~or treating the le~chate, ref~rred to in item (d) ~bov~, ob~ainud betwoon th~ flr~t ~nd ~cond ~tag0 o~
roa3ting. Al~o to bo d~crlbad aro lanching ~tep~ that could be ~mployod in tho G~ wh~ro la~ahln~ 1~ d~ bla ~ft~r a ~Q~ond ~ n oxy~n ~a~lclont ~o~u~, ~uah a~ ln ~h~ GA00 0~ ona op~lon ~or tho trOMtmont o~ ~lnc uulphlcl~
bulk con~untrat~. Al80 a m~tho~ o~ loAchln~ a l~ad ~0 nulphldc conc~ntr~to pru~uct ~or ~eluctiv~ly separ~tLng preciou~ met~l~ from the predominant le~d sulphide species will bo de~cribed.
Xn the first inst~nce, we will describe one optLon for trenting the c~lcine or, i~ you will, a partially desulphurized concentrate, aPter a conventional zinc concentrate has been sub~ected to the first stage of a two stage roast. In this in8tance~ we will concern ourselves with an ore or concentrate wherein the zinc sulphide is primarily present in the marmatitic form. In instances where the zinc sulphide is present primarily in the sphaleritic form, it i~ to be expected that the conver~ion of iron 3ulphides to readily soluble iron - oxides will-be accomplished with less conversion of zinc 3ulphide to zinc oxide.
In order to get the iron level oP a zlnc sulphlde concentrate down from 8~ to ll~ lron to 0~ to 3~
iron prlor to de~d roastlng, a quantlty of zinc oxide wlll WO 90/13679 PCI'/CA90/00130 2~3 ~ o-al~o be formed during the partial desulphurLzation roast ~tep ~ pre~iou~ly de~cribed, ~esulting ln zinc dis~olution lnto the leachate of about 5% to 25~ of the total zinc dependent upon thé type of zinc concentrate treated.
An embodimnnt of the inventlon i~ shown in flowsheet form by combining Figure I with Fiqure IVA for tre~t~ng ~ conventional z~nc concentr8te- In one ln~t~nce a conventional zlnc conccntr~te contsinln~ ~9.0 ~t~ Zn, 9.10 ~t~ Po, 0.70 ~t~ Cu, 0.24 ~t~ Cd and 32.4 ~t~ S wa9 giv~n a partlal d~3ulphurlzatlon ro~t to the extent th~t a pnrtlslly de8ulphurlzed concentrate was produced analy~ing 53.4 wt~ Zn, 9-87 wt~ Fe, 1.06 wt~ Cu, 0.27 wt~
Cd, and 25.70 ut3 S. ~See Bxample No. 1) Accordlng to this p~rticulsr ex~mple sn squeous ~0~ lesch ~a0 employed ~t a tomper~turo o~ 65 ~ 5C and A
pH o~ 1.8 to ~ or two hourD st sn lnltl~l pulp denalty o~ about 0~ 91~1.
A~t~r llquld-8011d ~ap~tion~ tha ro~ult~nt l~ch ro~ldu~ ~a8 xaportad to an~l~z~ ~9.3~ tina, 1.163 lron, 1.30~ op~0r, 0.31~ c~dmlum and 30.8~ aulphu~ ~nd ~ould p~o~ld~ an axaqll~nt ~acd m~tarlal ~o ~ do~d ro~
~tap dua to lt~ lo~ i~on contant.
~ho 104ch~to ~n~ly81~ aho~nd that 90.83 of tho lron, 14.~ o~ the zlnc, 2.52~ o~ the copper snd 5.474 o~
th~ cad01um h~d b~n dl~nolv~d ~om the p~rti~
de~ulphurlz~d ro~t~d product.
In thi~ example, the troatmant o~ the leachate involvod thermal d~composltion to dri~e off S02-containing 30 gaA ~or recycle thu~ precipltating a ~olid consi~tlng chiefly of iron sulphite and zinc sulphite.
The solid mixture wa8 treated by an ammonia leach uh~reby most of the zinc dissolved and all of the iron ~a9 left as a re idue. The iron re~idue wa~ separated by liquid-solid 3eparation. In one in~tance wherein the start~ng material wa~ a much lower grade of concentrate analyzing 34.5% Zn, 15.7~ Fe, 1.40~ Cu, 0.~3~ Cd, 32.6~ S, $UBSTmJTE ~I IEr ` ~ ~

,., ~ ",, ,, `. ~ "~. , ", ~

. , ,' " '' ' ' """ . ' ., ' ' ', , .; ' '` , ,,' " ' ~ ~. ' ' ' " ' ` '' ' `' ~ ' ' ' ' .- 20~5~27 WO90/l3679 PCT/CA90/00130 the iron re~idue produced was reported to analyze 61.1%
Fe, 4.31% zn, 0.03~ Cu, 0.07~ Cd and 1.30~ S.
The leachate was steam stripped to remove ammonia for recycle thus precipitating basic zinc sulphite which ~fter liquid-iYolid ~eparation would be treat~d with spent olQctrolyt~ to produce zlnc sulphat~ solutlon for Eoedlng to ~ zlnc r~ln~ry and S0~ g~s for recycle to the aqueou~ SO2 leachLng ~t~p. ~n the oxperiment conducted, starting with a much lower gr~de of zinc concentrate, the recovery of zinc from the SO2 leachate was reported as 89.4~. After calcination of the basic zinc sulphite in thi~ experiment, ~ zinc oxide product wa8 p~oduced which wa~ ~port~d to cont~ln 77.6~ 8n, 0.005~ F~, 0.03~ Cu, 0.34~ Cd ~nd 0.61~ S. ~ha impurity lovol o~ th~ produced zlnc oxlde wa~ romark4bl~ con~ldorin~ thnt no purl~la~tlen ~t~ps 0uah na ~ln~ du~t c~men~tlon wQra aArrl~d oUt.
~ n thi~ ambodlm~nt thQ b~olo ~xlna ~lllphlt~ wo~
b~ dahydr~tu~ 1~ n~o~0~ry, and th~n tr~t~d wlth ~pQnt ~l~ctrolyta llquor to dlia~olv~ lt nnd ~oad thQ resul~nnt zinc ~ulphate product directly to a conventlonal electrolytic zinc re~inery. ~he reaction between zinc ~ulphlte ~nd ~12SO~ will give o~f pur~ S2 ~or recycle when nnd ai~ need~d.
~he spent llquor from the thermal decomposltion step where ba~ic zinc sulphite i9 produced wou}d be treated with lime or calcium hydroxide in order to free and recycle its ammonia content in liquid form and to recover its contained zinc in the precipitate thus formed.
After liquid-solid separation, a sulphuric acid treatment step on ~he precipitate would be required to dissolve the base metal compounds for recovery from the insoluble calcium compounds. However, the circuitry would be quite small and therefore quite inexpensive.
Other embodiments exist for treating the leachate from the aqueous sulphur dLoxide solution leaching step.
Thes0 include ~olvent extractlon o~ the leaaha~0 .. : : . :.: : , . . , , . .: . . .. .. : j.. , :, : .
.: :, .: :- : : :. .... : : .... : .. . :: :,. : .: . .

90~ 9 PCT/C~90~00130 to separate the zinc ~raction from the lron fraction by method3 si~ilar to tho~Q di~clo~ed by Clitheroe.~U.S.
PatQnt 4,053,552) wlth modlflcatlons (Sse Figure IV-8) Another method is to use hydrogen sulphidQ at a pH range of 3 to 6 to convert the 2inc pre~ent ~9 bl~ulphite in the leach liquor to insoluble~zlnc sulphid~ which i9 then ~ed to either the first ~tage roaster or the s~cond ~tage roa~ter dependant upon its iron content. With proper control of pH during the zinc eulphidq pracipltation ~tep, virtually complete r~covery o~ the zinc con~tituents ohould be ~chioved wlth little or no lron co-preclpi~atQd.
~9ee Flgurn VII).
Anothor ~mbodlmont ~or t~o~tlng the partl~lly do~ulphurlz~d zinc concontratu or, i~ you will, c~lclne produced a~ter the ~lrst stage partial desulphurization roast of a convention~l zinc concentrate is to le~ch thls calcine in warm dilute ~ulphurlc acld ~olutlon containing ~gueou3 Dulphur dioxlde ~ollo~qd by llquid-~olid aaparation t~chnique~ ao pr~viou~ly do~c~ibod, ~heraby tho lo~ch raoi,duo i~ ~ed to ~ conventlonAl doad ~0~4t ~n~ th~
lonchnto ia ~ub~act~d to n ~ol~ont e~t~Action tochni~uo to ~elactlv~ly ~p~r~t0 th~ dl~oolvod ~lnc ~nd lron ond thu~
~roduc~ ~I zlnc ~ul~h~t0 nolu~ion~ Th0 ~lnc ~ulph~to ~olution in th~n ~ad to ~n ~ op~i~t- pl~oa ln ~
2~ conv~ntlono.l ~lnc re~lninq alxcuit ~hlle~ tho r~ in~te i3 tre~tad by one of the th~aa optlonu shown ln Figure III.
Another embodi~ent for troating the leachate produc~d ~tar the ~irst ~t~e d~ulphurlzatlon roas~ i3 de~crlbad h~xo~lth. Th~ lesch~t~ 1~ tr0atod ~lth oxyg~n (alr) to oxldlze the dl~olved ~arrous i~on to ~erric and by hydrolytlc action to preclpitate tho errlc iron as goethite. To enable iron pr~clpitation to go to complotion, lime is required to neutralize the acid released by the hydrolytic re~ction ~hile holding the pH
in the range of 3-6. After ligyiid-solid separation, the resulting solution i~ fed into the zinc refining circuit : evol~ing 52 for recycle, ~hile the iron-containing solid ..

2 ~ 2 7 .~

WO90/13679 PCT/CA90/00~30 is sant to disposal.
Other methods of treatlng the partially desulphurized concentrate or calcine ln a two stage roasting process when a conventional zinc concentrate is to be treatsd in order to retard zinc ~errite formation m~y be found in the public domain or be obviou~ to one normally skilled in the art.
Re~orrlng b~ck to the le~ch r~i~idue produced a~tor th~ con~entlon~l neutr~l le~ch o~ the dead roasted product, the lea~h reslduu would be of very much smaller volume th~n th~t conventionally produced because of its very low zlnc ferrite content and thus would be enriched in lead and precious metals content. Thls might be treated by flotation technique~ to ~eparat~ the lead as well a~
thc pr~clou~ metal3 ~rom other ganguo mAterial nnd ~130 to ~opnr~te the ~llv~r from the lead compon~nt. ~n altorn~tive m~thod would ba to u~ oodium cyanld~ or thlour~n to lonch ~nd than oop~r~t0 tho ollvor ~ulphld~
~om the otho~ lQ~oh roul~ue m~torl~ 0 Fl~lro V~ ln thl~ l~to~ ln~tana~) Anoth~r ~mbodlmant la ln tho c~e o~ th~
tre~tmont of ~ low gr~de zin~ aulphlde c~ncentrate not sultable ~or conventional roa~tlng, where the low grade zinc concentrato would be given n ~lr~t ~t~ge p~irtlal de~ulphurization roa~t to the extent th~t i~ ~ould bring it up to a g~ade equivalent or better than a normal concentrate suitable for dead roasting after the intermediat~ leaching step. In this case, the preferred leachant might be a mixture of dilute sulphuric acid ~olution containing sulphur dioxide as previously described. (See Figure III which shows an aqueous SO
leachant variant, al~o Figure V).
In one instance, a low grade zinc concentrate reported to contain 34.524i Zn, 15.95% Fe, 1.15% Cu, 0.23%
Cd and 32.68~i S was given a partial desulphurization roast to the extent that the calcine analy~ls was reported as 43.864 Zn, 16.230 Fe, 1.570 Cu, 0.280 Cd, 0.0820 Pb and ' ~ ' . , ,: ~',.,', :, ' : ' : :' ' : ', WO ~0/~3679 PCI/CA90/00130 24.05~ s.
~ hi~ partially roasted concentrate wa~ th~n leached at an initial pulp density of ao 9l 1 in a 3~ HzSo~
301ution co~tain$ng di~solved S02 at a temperature of 65C
to 69C for approxi~ately three hours. After leaching and liquid-solid sep~ration, ~he leach residus was reported to contain 50.8~ Zn, 7.114 Fe, l.a6% Cu, 0.29~ Cd, 0.062~ Pb and 30.12~ S, wlth only 3.724 o~ the zlnc extracted into the leachate. A8 can be 3een a non-useable zinc concentr3te had thu~ been converted to an uquiva}ent or superior gr~de o~ conventlonal zinc concontrate with only ullght loa~ o~ zinc, wh~ch would otherwi~e be lo~ in any event.
It i~ expected th~t the leachate being high in ixon content and very low in other dis~olved base metals could be treated wLth llme and dlspo~ed to a tailing pond or be oxidized to preclpltata goethLte ~nd then limed for di~po~
Othcr method~ o~ dlapoaal or recov~ry o~ the iron ml~ht ~l~o b~ d~viued depondlng on clrcumflt~nce~ and the ~urroundln~ onvironmant.
~ ha ~bov~ d~c~lb~d ~x~m~l~o lllu~ra~Q ~ho ~loxlblllt~ o~ tha proaqo~ to upgr~d~ low~r4d~ zlnc con~antrat~ or aonv~ntlon~l ~lnc conc~ntrato~ ~o 2~ aonc~ntr~tau h~lng ~ny d~8ir~bla 1~VQ1 a~ iron contont.
In tha ~mbodlment ~or the tre~tment o~ zinc ~ulphide bulX concentrates produced ~om complex mas0ive ~ulphido mineral depo~it~, ~evo~al op~ion0 exi~t. Por ex~mple, a conv~ntion~lly produc~d zlnc ~ulphLde bulk concentrate could be upgraded to produce a superior grade of bulk concentrate having the attractive ~eatures such as a higher level of zinc and lead and a lower level of sulphur and a much lower level of iron thus reducing both tran~portatlon and treatment char~es, the latter substantlally. One example of this type of processing is sho~n in Fi~ure II.
With convantional methods of mineral processing ~ 20~27 of such massive deposits, zinc ~ulphide bulk concentrates of relatively low value are produced. Example No. 3 shows a method of increasing the value of a zinc sulphide bulk concentrate. Also with our novel process it is conceivable that hlgher recoveries of all valuable metals could re~ult. This i9 because lower grades of bulk concentrates could be upgraded to acceptable levels by us1ng more of the orebody and discarding lo~ mlne~al proce~sing t~lling~
Al~o our proc~s~ ia adaptable to orebodies that contain les~ lead but are not 9uLtable for conventional processing (See Example 4).
A flow~heet is provided for one method of treatment of zlnc sulphide bul~ concentrate~ produced or to be produced ~rom complex m~sslve base m~tal sulphlde concantrates cont~inln~ ~ub~tantl~l level~ o~ zlnc, l~ad ~nd lron ~ulphlde~. ThL0 flowDhaat 1~ ~r~ontad in Fl~ura II.
Examplo No. 3 providoo lAbo~A~ory rooul~ ~o~
ono mothod o~ trQ~-tln~ 3 zlna oulphido bulk ~oncon~r~to.
(5~ o ~ ur~ ~ aomblnod wlth ~l~uro ~VA).
~ he ~olLowlng d~crLptlon glvo~ one embodlment o~ a m~thod of uaing partial d~aulphurizatlon roasting techniqueo for treatlng bulk cQncentrates to extract high ~5 valu~ o~ zinc, ledd and preciou~ metala and to provlde an envlronmentally ~atlsfactory lron fractlon.
A zinc ~ulphlde concentrate consisting prlncipally of z1nc sulphldes, iron sulphides and lead sulphides with lesser amounts of other metallic sulphides, usually in the form of complex sulphide compounds or solid solutions thereof ia treated in a roaster using an oxygen containing feed gas, presumably but not necessarily ordinary air, in a manner that results in an oxyyen deficient atmosphere at all times by controlling the retention time o~ the Rolid feed material at temperatures bet~een 650C and 1050C in order to selectlvely convert its iron-containing constituents into readily soluble iron .... .. .. ~ .. . . . . . . . .

WO90/1367s PCT/CA90/00130 .~ 4~ oxide, leaving unreacted the major portion of all the ~9~ remainlng sulphid~s resulting in a calcine or, if you will, a partially desulphurized conc~ntrate.
Thi9 calcine i~ then treated with a medium temperature (50C~75C) dilute 5ulphuric acid solution containing dissolved ~ulphur dioxide or a medium temperatur~ ~50C-~5C) aqueou8 sulphur dioxid~ solution a~
previou~ly described, to leach ~ny qoluble oxides, which includes the ma~or portion of the total iron ~nd a minor portion of the converted ba9e metal oxide9. The resulting ~lurry i9 oub~ected to liquid-oolid ~epar~tion to sopnr~te the Joluble oxldes portion from the inaoluble rem~inlng ~ulphlde~.
~he iron containlng solution ls thon treated in one or more o~ the ~ethodo previously de~crlbod in order to dlspo~e of tho iron fraction in an environmentally sati~factory manner.
Tho remainin~ ~ulphides which contains chiefly zLnc oulphide ~nd le~d nulphlde but a1JO other mot~llic ~ulphld~ 18 ~ubJacted to n necond at~e pnrti~l da~ulphuri~atian ~oa~tlng opar~tlon u~ing ~n oxyg~n aant~lning g~ ~cr tha convar~ion o~ tho bulk Oe tha cont~lna~ ~lnc ~ulphld~0 to ~lnQ oxlda undo~ oxyg~n do~iclont ~o~tlng Gonditiono ~ ~roviou01y do~crlbad and pr~ar~bly in tha tompo~ntu~a ranga o~ 650C to 8~0C.
~ h~ r~uultlng oecond otMge p~rtial danulphurization ro~tlng operation would be designed to pxoduce a calcin~ ~hlah i~ chie~ly compo9ed of zlnc oxido n~d un~o~cted oulphlde9~ chl~ly l~d ~ulphide plua a concentrat~d amount o~ preclou~ metals. ThiD ~econd ~tage c~lclne i8 either leached and sub~cted to liquid-~olid ~eparatlon techniques or treated by flotation or other physical separation techniques in order to selectively separate the zinc oxide and other base metal oxide~
fraction, containing the bulk of the zinc, from the remaining ~ulphide~ which would be chiefly composed of lead aulphide but would also contain almost all of the 2~ 27 precious metalq.
Th2 separated oxlde fraction which contains the bulk of the zinc as zinc oxide would then be sent to a zlnc refinery operation for produclng zinc m~tal or might be qold as a zinc oxide product. If the aqueous SO2 leaching roast were used the re9ulting laachate containing ahie~ly ~lnc ln thè bl~ulphit~ ~orm would be treated wlth ~pent elactrolyte ln order to convert the zinc to it~
~oluble sulphate ~orm ~or feed to a zinc refinery, thus regenerating sulphur dioxide for recycle.
The remaining sulphide fraction which would contain chiefly lead sulphide but would also contaln a concentrated amount of precious motals and perhap~ somq gangue material could be ~ent directly to ~ lead smelter.
Alte~n~tlv~ly~ the prociou~ mot~ ractlon would be aepar~ted ~rom the ram~inin~ ~ulphida ~r~ction by ~lot~tion tQchnigue~ or by l~hln~ wlth A CyAnlde or thlou~en oolutlon bo~ore th~ le~d uulphl~o Cxnc~ion lo ~ont ~o ~ lo~d amaltar ~or tho productlon o~ lo~d.
Som~ bulk ~ulphlda concontr~t~o mlght contnin signL~lc~nt portion0 o~ aruenic and perh4p~ other elements not speci~ic2illy mentioned ln the descrlption th~1s far.
l~he~e type~ of bulk ¢oncontr~tes m~y be ad4pted to the techniques hereinbe~oxe descrlbed but Inay complic~te th~
roaster exit g~ tre~tment ~ystem which would norm~lly contain a high purity SO2-laden gas stream which might be used as a source of sulphur dioxide gas for make-up p~rposes where aqueous sulphur dioxide leaching in employed but would ~ainly be u~ed for conversion to 30 sulphuric acid. - -Other modifications to the circuitry may also be neces~ary to take into account the presence of amounts of other extraneous impurities.
n fluid bed dead roasting of conventional ~inc sulphide concentrates, ~rom about 50~ to as much as 90~ o~
the product may be collected from the ga~ str~am leaving the top of the roaster. Under any such condltion~, it m~y woso/13679 pcT/cAso~ool3o A,9~ be difficult to control the residence time in the roaster and thu~ of the 8ulphur content of the product. secau~e thia i~ our goal, i.e. to control the l~al of 3ulphur of the partial desulphurization ro~8t product through control of the residence time or oxYgen availabllity in the roa~ter, it m~y be necessary to ngglomerate the concentrate ~uch that a maximum of ~5~ of the product i9 elutriated from the roa3ter bed. Preliminary studies ~howed that this goal wa~ achleved when the concentrate was agglomerated to -40 to +60 mesh size r~nge. This also led to very smooth operatlcn o~ the ~luid bed roaster.
1~ tho aoncentrate to be treated i~ ground to tha -40 to ~60 mosh ~i2e prior to ~lot~tion, this could a~ect the yield Oe zlnc (and other ba3e and precious metal~). In practice, concentrate analyses as a function of grind ~mesh size) would be determlned to see whether this approach was desirable. I~ the yield would be gre~ter but the tenor of th0 concentr~te lower, a partlal de~ulphuxlaation roast would be conduct~d wlthout the need ~or an agglomex~tlon ~top whllo tho lo~ch re~lduo a~ke~
th~ p~rtlal do~ulphurl~ation ro~t could b~ ~qu~l to o~
~uperior to mo~t ~ino sulphld~ conc~ntrnto0 avAll~blo tod~ ~nd aould concqi~bly b~ loo~ th~n lO iron.
X~ ~ ~luld b~d ~oaot~r ~oro uo0d ~or th~ p~rtl~l ~5 deaulphurla~tion ro~ting, control o~ the partlcle si~e dl~tributlon to the ~luld bed roa~ter may be deslr~ble to achieve a do~ired residence tl~e in the ro4ster ~nd thu~
controlled degrea o~ ~ulphlde Julphur removal, in order to provlde ~n improved operatlon.
~ 30 Fi~ures I through IX by means of flowsheets, show a variety of embodiments for the treatment of zlnc sulphide concentrates including conventional zinc concentrates, low grade zinc concentrate and bulk zinc concentrates using partlal desulphurization roasting techniques. Some of the embodiments that have been de~cribed have not been ~hown in flowsheet form.
Flgures I through VII provide a variety o~
:
:

WO90/13679 2 ~CT/CA90/00130 tr2atment~ of zinc sulphide concentrates including conventional concentrates, lo~ grade zinc concentrates and bulk concentrates. Additional methods of treatment are included i~ the text. These all depend on two stage roa~tlng wherein at least the fir~t 5tage Ls conducted in an oxygen de~lcient atmosph~r~. Other methodn o treatment u~ing ~uch two st~ge ro~tlng may become apparent to those normally ~klll~d ln thu nrt because o~ the wide range of floxlbllity.
EXAMPLE NO. 1 A conventional zinc sulphide concentrate of the following analysis was partially de~ulphurized ln aLr at a temperature of 850C in a fluidlzed bed roa~ter to produce a partially deaulphu~lz~d ~lclno o~ the ~n~ly~i~
~iven b~low C~ Cala~ 19 ~n~9.0 Wt~ 53.40 Wt~
P~~.10 W~0 ~ ~7 Cu0.70 Wt~ 1.0~ W~
Cd0.24 Wt~ 0.27 Wt~
S32.40 Wt~ 25.70 Wt~
Tha SO2 lnden off-ga~ was reported to contnln 19 volg SO2 ~nd les~ than 0.1 vol~ oxygen.
Forty gr~m~ (40.0 g) o~ the resul~ing p~rtl~lly de~ulphurized calcine was leached ln 503 ml of an aqueous solution of sulphur dloxide. ~he leach was conducted at a temperature of 65 t 5C and a pH of 1.8 to 2.1 for two (2) hours. The leach pulp was filtered from the leachate.
Analysis of the leachate showed a 14.20~ zinc extraction, 90.77~ removal of iron, 2.52% removal of copper, and 5.47%
removal of cadmium. The reRulting leach residue analyzed 59.30~ zinc, 1.16% iron, 1.30~ copper, 0.31~ cadmium, and 30.80% sulphur.
Further treatment of the leachate to recover its 14.20~ zinc content cnn be accompli~hed ~y various route~.
EXAMP~E NO. ~
A low-gr~de zinc ~ulphide cancentrate was . , , . : ;. ;

4 ~ ~ partially de~ulphurized at a temperature of 850C in a fluidized bed roaster to produce a partially desulphurized calcine. The concentratq and calcine analysi-~ were reported to be a9 followss Concentrate Analysis Calcine Analysis Zn34.5 ~t~ 43.86 Wt~
Fe16.5 Wt~ 16.23 wt%
Cu 1.15 W~ 1.57 Wt~
Cd0.23 Wt~ 0.23 Wt~
10 S 32.7 Wt~ 24.05 Wt~
The off-gas was reported to contain 19 vol% SO2 and less than 0.1 volO oxygen.
Porty gr~m~ (40.0 g) o~ the resulting p~rtially desulphurlzed calcine w~ leaahed in 500 ml o~ aqueous 3~
15 H2SO~ plU9 S02 ln ~olutlon. The leach wa~ conduct~d at a temp~rature o~ 6SC i 5C and a pH of 1.1 for three (3) hours. The leachnte was then flltered ~rom the leach pulp.
Analysl~ of the leachate showed an extraction of 3.72~ Zn, 63.54~ Fe, 0.35~ Cu, and 5.32~ Cd. The resulting leach re~ldue analyz~d 50.80~ zn, 7.113 Pe, 1.85~ Cu, 0.28~ Cd, ~nd 30.12~ 5.
EXAMP~E NO. 3 ._ A complox Nffw ~unowlck ~inc ~ulphido bulk concantr~t~ o~ th0 ~n~ly~lu ~hewn b~low WA~ ~iv~n p~rtial d0~ulphuri~tion roA~t ln Mlx ~t 750~C in M
~luldl~od b0d re~t~r to p~o~uc~ rti~lly d~ulphurlz~d concontr~t~. Figur~ X provldea ~ temper~tur~ pro~lle during th~ continuou~ p~rtlal de~ulphurlz~tlon roast along with 80z and 2 o~ ao concentration3. Products r~moved during conot~nt op~ratlng condition~ aro ~hown ao Pl~ P2 ~nd P3. A lower temperature ~A~ employed on the roa~t becau~e of the high lead content o~ the complex concentrate. The partially desulphurized concentrate (sample P3) was given a warm SO2 leach as described in Example 1. After flltration and washing, the lsach residue had the analysiis i~ihown below:
Concentrate AnalYsis Leach Residue ;: : : . : . :: ,: : ;: .: .: , , . , , , ......... . , : . , , . : , : : .,, .. :: : .; , .: : , 20aa~27 Zn 34.6 Wt% 40.8 Wt~
Fe 12.6 Wt~ 1.87 ~t~ ~
Cu 0.9 Wt~ 1.23 Wt~ .
Cd 0.1 ~t~ 0.1 Wt~
Pb 16.4 Wt% 28.9 Wt~
S 33 Wt~ 24.2 Wt~
Two 2ddltlonal pnrtially desulphurized conc~ntrates were produced ~rom ~h~ samo New 3run~wick bulk zinc concentrate. In the firot t0ample Pl), the ~ S was reduced ~rom 33 wt% to 24.9 wt~ and ln the second ~sample P2), the & S was reduced to 18.8 wt~. ~he partially desulphurized concentrates from each roast were given identical aqueous S02 leaches. The extr~ction of zinc and lron ~or both leacho~ wa~ dotermin~d ~a a fun~tlon o~ la~chlng tlmo ~nd the reDult~ plotted ln Flgure XI~l. Fox the xo~nt tha~ had le~ sulphur removed ~Pl), les~ thnn 2~ o~ the z.Lnc di~401vad 410ng with ~bout fiO0 o~ tho lr~n~ ~or ~ho roA~t that hn~ ~ muah ~r~ter ~mount O~ oulphur ~omov~cl ~ P2 ), ~omo 200 oP tho z.~na w~ oxtr~ctod dlQn~ wlth up to 900 o~
th~ iron .
~ hlo provldes two di~tlnct possibilitiess an upgradlng roast to remove part of the iron along wlth a ve~y small amount of zlnc or a roast carri~ out to permit di~solution o~ ne~rly all of the iron and a minor portlon of the zinc while leaving the more stable sulphide~
untouched.
EXA~PLE NO. 4 A second New Brunswick zinc sulphide concentrate of the analysis shown below was given a parti~l`
desulphurization roast as profiled in Figure XII. The ; temperature at the end of the roast was raised ~o above 900~C to determine whether the 1.2~ lead content would re~ul~ in a defluidized bed. No operational problems were noted.
Concentrate Analysis Zn 51.1 Wt3 Pa 10.2 Wtt ~ .

, wo9o/l367s PCT/CA90/00l30 u0.24 wt%
Cdo.lo wt~ `-Pb1.20 Wt4 S34.4 Wt~
Four ~amplas of the partially roasted concentrate were removed during the roast (i.e. Pl, P2, P3 and P4). Slmilar S02 leache3 to that in Example No. 1 ~ere conducted on the flr3t three sample3 which were taken when the roaster was operating at ~bout 750C. Slight changes in the feed rate to the ron3ter and thus the residence time of roastlng re3ulted in ~lightly different 3ulphur content3 ln e~ch pnrtldlly de~ulphurlzed roM~ted product.
The le~chlng pro~ll0~ for each s~mpl~ are presented in Flgure X~t. ~hn~e re~ult~ lllustr~te the ablllty to control the roasting oper~tlon by residence time alone to dchleve ~ust about any level of iron removal desired. The amount of zinc co-leached wlth the iron is also shown.
There ~re a number of approaches th~t can be used to recover the zlna, ~nd racycl0 tho 92 and dlspose of the ao lron.

: ' ~ ; ~ ; i ', , ' . , , ' ' ,, ; , ' ~

~'' ' ~` ' ' ' ' .. ' ` '' ` ' ' ' .. : ' :.. ' ' ''' ' 'I' ' ' .

Claims

THE EMBODIMENTS OF THE INVENTION IN WHICH AN EXCLUSIVE
PROPERTY OR PRIVILEGE IS CLAIMED ARE DEFINED AS FOLLOWS:

1. A process for the recovery of zinc from iron-containing zinc sulphide ores or concentrates comprising the steps of:
(a) roasting the ore or concentrate in the presence of an oxygen-containing gas wherein sulphur retention is maintained at a level of 15% or higher by controlling the residence time of the ore or concentrate and/or the flow of oxygen in the roaster to provide an oxygen deficient atmosphere in the roaster using a temperature in the range of 700°C to 1050°C in order to preferentially oxidize the iron rather than the zinc and other base metal constituents and thus reduce or eliminate zinc ferrite formation, (b) leaching the partially desulphurized roasted ore or concentrate using a lixiviant selected from the group consisting of aqueous sulphur dioxide, a 2-5 wt% sulphuric acid solution, and n 2-5 wt% sulphuric acid solution containing solubilized sulpher dioxide, at a temperature between 50°C and 75°C in a pH range of 1.0 to 2.5 for one to three hours in order to preferentially dissolve the iron oxide thus formed as well as base metals that have co-oxidized thus obtaining a leach slurry (c) subjecting the leach slurry to liquid-solid separation techniques (d) treating the leachate to separate the dissolved iron from the dissolved base metals, thus leaving a leach residue which is richer in zinc sulphide content and depleted in iron content which (i) is dead roasted as in conventional practice with the resultant reduction or effective elimination of zinc ferrite formation or (ii) if the original concentrate is rich in lead, such as in a zinc sulphide bulk concentrate, subjecting said leach residue to a second stage partial desulphurization roast using a controlled retention time in the temperature range of 650°C to 850°C
such that the retained sulphur content is 8% or higher which relates to the amount of lead present to produce a second stage calcine which has a very high zinc oxide content with substantially all of the lead remaining in the sulphide form, subjecting said second calcine to physical separation techniques or to an aqueous sulphur dioxide leach under similar conditions as earlier described in order to separate the zinc oxide fraction from the lead sulphide fraction, the sulphide-containing residue is then fed directly into a lead smelter or is further treated to separate the lead sulphide from the precious metals.

2. A process according to Claim 1 wherein the sulphur retention is between 15 and 27%.

3. A process according to Claim 1 wherein the sulphur retention is between 20 and 25%.

4. A process according to Claims 1, 2 or 3 wherein the first stage roasting temperature is between 850°C and 1000°C.

5. A process as described in Claims 1, 2, 3 or 4 wherein a 3-4 wt% sulphuric acid solution is used as the leachant.

6. A process as described in Claims 1, 2, 3, 4 or 5 wherein the leaching temperature range is between 60°C
and 70°C and the pulp density is between 60 and 120 gpl and the leaching period is two hours or less.

7. A process as described in Claims 1, 2, 3, 4 or wherein the feed material is a conventional zinc concentrate and there is a first stage oxygen deficient roast followed by a leach step and a dead roast of the leach residue.

8. A process as described in Claim 7 wherein the feed material is a conventional zinc concentrate or a low-grade zinc concentrate containing less than 2% by weight of lead sulphide.

9. A process as described in Claims 1, 2, 3, 4 or 5 wherein the leachate produced as a result of the leach step is thermally decomposed to drive off sulphur dioxide for recycle and to produce a precipitate consisting principally of a mixture of zinc sulphite and iron sulphite and a depleted liquor solution. The solid sulphite mixture is then treated by an aqueous ammonia leach to selectively dissolve the base metal components and leave effectively all of the iron content in the residue. The iron-containing residue after liquid-solid separation is then subjected to thermal and physical treatment steps to produce a high purity iron oxide product. The solution portion is then thermally decomposed to drive off ammonia for recycle producing a basic zinc sulphite precipitate and a partially depleted ammonia liquor. The basic zinc sulphite precipitate, after liquid-solid separation, is dehydrated to produce a zinc sulphite product suitable for treatment with a portion of the spent electrolyte stream being recycled to a conventional zinc refinery neutral leach circuit thus to convert the zinc sulphite to soluble zinc sulphate and thus drive off sulphur dioxide for recycle and the partially depleted ammonia liquor is then treated with calcium oxide or calcium hydroxide to precipitate its sulphite content as calcium sulphite and after liquid-solid separation, the dissolved ammonium hydroxide is recycled to the ammonia leaching circuitry. The calcium sulphite precipitate is treated with a sulphuric acid solution to dissolve any base metal components of the precipitate and fed to a liquid-solid separation step to separate the insoluble calcium sulphate thus formed and permit the base metal constituents to be recovered as well as the generated sulphur dioxide.

10. A process as described in Claims 1, 2, 3, 4 or 5 wherein the aqueous sulphur dioxide leachate is subjected to solvent extraction by a method similar to Clitheroe's U.S. Patent No. 4,053,552 to recover the contained zinc as zinc sulphate whereas the iron bisulphite raffinate is further treated before disposal.

11. A process as described in Claims 1, 2, 3, 4 or 5 wherein the aqueous sulphur dioxide leachate is treated with hydrogen sulphide at a pH range of 3 to 6 to precipitate the zinc as zinc sulphide so that it can be separated from the iron component remaining in its soluble form.

12. A process according to Claims 1, 2, 3, 4 or 5 wherein the feed material is a low-zinc high-iron content zinc concentrate wherein the first stage partial desulphurization roast leads to the formation of a conventional or superior zinc concentrate after leaching and a solution containing a relatively high iron level with only a relatively small amount of dissolved zinc, wherein the solution is treated with calcium hydroxide or calcium oxide and the resultant iron-containing residue is fed to a tailings pond.

13. A process according to Claim 1 wherein a zinc sulphide bulk concentrate containing substantial quantities of zinc, lead and iron sulphides as well as some precious metals is treated so that the product of the second stage partial desulphurization roast contains little iron but large quantities of zinc oxide, lead sulphide and perhaps some precious metals wherein the zinc oxide is separated from the lead sulphide and precious metals by described leaching techniques followed by liquid-solid separation techniques leaving an insoluble residue containing essentially lead sulphides, enriched precious metals and perhaps some gangue material which is then treated with a thiourea or cyanide solution by a leaching process to selectively separate the precious metals from the lead sulphides. The lead sulphide concentrate is then fed to a lead smelter.

14. A process wherein the bulk concentrate is treated by a partial desulphurization roasting technique by controlling the residence time in the roaster to the extent that the sulphide sulphur retention is between 20 and 25% sulphur and the roasting temperature is in the range of 650°C to 850°C in order to convert the contained iron into its oxide form in preference to the contained base metals and precious metals and after separating the iron oxides from the remaining sulphides by leaching and liquid-solid separation as described in Claim 1 by the use of leachants consisting of aqueous SO2 solution, sulphuric acid or a mixture thereof, the remaining sulphide fraction is treated by a second partial desulphurization roast in the temperature range of 650°C to 850°C to reduce the sulphur content in the second stage roast to a level where 80% of the contained zinc or more is converted to its oxide form, and preferably 90% of zinc or more is thus converted, and the resulting calcine is then treated by selective flotation techniques to separate the remaining sulphides which will consist primarily of lead sulphides which are enriched with precious metals, so that the lead sulphide fraction can be fed to a lead smelter either before or after separating the precious metals.

15. A process wherein the bulk concentrate is treated by a partial desulphurization roasting technique by controlling the residence time in the roaster to the extent that the sulphide sulphur retention is above 8% and the roasting temperature is in the range of 650°C and the 850°C in order to convert the contained iron and contained zinc in preference to the contained lead and precious metals into their oxide form but short of converting lead sulphides or precious metals sulphide to their oxide form.
The oxidized components are selectively dissolved with aqueous sulphur dioxide solution under the conditions described in Claim 1 and after liquid-solid separation, the leachate is treated by solvent extraction to separate the zinc from the iron components as in Claim 9. The leach residue is treated by thiourea or cyanide leaching to separate the precious metals from the lead sulphides, after which the silver is recovered and the lead concentrate is fed to a lead smelter.

16. A process as described in Claim 15 wherein a dilute sulphuric acid solution is used to dissolve the oxidized components and after liquid-solid separation the leachate is treated with air and lime to a pH of 5 to 6 to separate the zinc and iron components.

17. A process as described in Claim 16 wherein the original feed material is either a conventional zinc concentrate or a low grade zinc concentrate rather than a bulk concentrate.

18. A process according to Claim 1 wherein the off-gas from the partial desulphurization roasting step is used as sulphur dioxide make-up in the aqueous sulphur dioxide leaching process and the balance of high quality sulphur dioxide-containing gas is converted into sulphuric acid.

19. A process according to Claim 1 wherein the concentrate feed material contains substantial amounts of arsenic wherein the arsenic is oxidized and is emitted with the off-gas and is then separated form the off-gas stream prior to feeding the SO2-laden gas to the sulphuric acid plant.

20. A process according to each claim where one or more of the roasters are fluid bed roasters.

21. A process wherein the size distribution of concentrate particles is maintained at >100 and preferably <20 mesh (>200µm; <900 µm) where it is desired to control the residence time of the solids fed to a fluid bed roaster, and thus to reduce the elutriation of small particles from the bed.

22. A process as in Claim 21 wherein the preferred size distribution is approximately <60 mesh; >40 mesh (>400 µm; <700 µm).

23. A process for recovering zinc from zinc bearing ores or concentrate substantially as disclosed herein.