a reviewofcopperhydrometallurgy - saimm · why hydrometallurgy? itisperhaps difficult toprovide a...

A review of copper hydrometallurgy

SYNOPSIS

by j. C. PAYNTER*, B.Sc. (Ch em. Eng.) (Visitor)

Copper. hydrom~t~lIurgy is discus~ed with special reference to the treatment of chalcopyrite concentrates underSout~ African con~ltlons. The following processes are described and illustrated: Sherritt Gordon's ammonia pressureleac~lng, t~e Arbiter Process, sulphuric acid pressure leaching, roasting processes, the Cymet Process, and theferric chloride process developed at the National Institute for Metallursy. These processes are then compared withrespect to the form of sulphur produced (elemental sulphur or sulphate), the character of the iron product and thehandling of different types of ore.

.

SAMEVATTING

~oper.hidrometallurgie wo~d bespreek met spesiale verwysing na die behandeling van chalkopirietkonsentrate inSul?-Afrl.kaanse. toestande. Die volgende prosesse word beskryf en geillustreer: Sherritt Gordon se ammoniakdruk-logl~g, die. Arblter-pr~ses, sw~wels~urd~ukl?ging, roosterprosesse en die ferrichloriedproses wat by die NasionaleInstltuut vir Metallurgle ontwlkkells. Hlerdle prosesse word dan vergelyk ten opsigte van die vorm van die swawelwat geproduseer word (elementale swawel of sulfaat), die aard van die ysterproduk en die hantering van die ver-skillende soorte ertse.

INTRODUCTION

In giving a review of the hydro-metallurgical processes for copper,I am conscious of its being a subjectthat is widely discussedl-3. Notbeing an expert, I shall not attemptto provide a better review than thosealready available. However, as aresult of discussion with my col-leagues at the National Institute forMetallurgy (NIM) and with metal-lurgists in industry, I have come torealize that, instead of having a clearpicture of the direction in whichcopper processing is moving, par-ticularly processes employing hydro-metallurgical techniques, mostmetallurgists are becoming increas-ingly confused, myself included. It ishoped that this paper, althoughsaying very little that is new, willprovide a suitable basis for dis-cussion so that benefit can be ob-tained from the wealth of experienceand judgment available and so thatthe picture can be clarified just alittle.

I intend to discuss copper hydro-metallurgy with special reference toSouth African conditions because Ibelieve that conclusions drawn forconditions elsewhere, for example inArizona, D.S.A., are unlikely to bethe same as those arrived at for aplant situated in South Africa. Evenwithin Southern Africa, what appliesto a plant located in a remote desertarea does not necessarily apply to aplant located in the Witwatersrand

"National Institute for Metallurgy, Johan-nesburg.

158 NOVEMBER 1973

area. I also hope to highlight thoseareas in which research can probablylead to cheaper and improved pro-cesses.

WHY HYDROMETALLURGY?

It is perhaps difficult to provide aconclusive reason for the upsurge ofinterest in hydrometallurgical pro-cesses during the past five years.I expect that many persons wouldcite the fact that, whereas the 1960sfelt the impact of the computer inalmost every sphere of life, the1970s are destined to experience theimpact of the environmentalist.'Clean environment' campaigns arealready a reality of life, and en-vironmental considerations haveforced the cancellation of construct-ion plans for power stations, theclosing down of copper and zincsmelters, the wresting of power fromthe motor giants of Detroit, andlast, but not least, the clean-up ofthe air in cities like London, whichvisitors now say has air superior tothat we breathe in our own modernmetropolis. Sulphur dioxide pol-lution from smelters has perhapsbeen the strongest driving forcebehind the recent massive expendi-ture on research and development inhydrometallurgy. However, there areprobably other factors that havecontributed towards this trend.

(a) It has been predicted that anincreasing percentage of newcopper production will come fromnon-sulphuric sources and mixed-sulphide ores (e.g., copper-zinc).For a variety of technical and

economic reasons, these ores oftencannot be processed by normalflotation and pyrometallurgicaltechniques. Differential-flotationprocesses for mixed sulphidesseldom give as clean a separationas desired, and in some instancesflotation is quite unpractical.

(b) The successful industrial appli-cation of the copper-selectivesolvent LIX4. 5, even though usedin the treatment of oxide ores,provided the most significantbreakthrough in copper hydro-metallurgy since that achieved bySherritt Gordon in their success-ful development of pressure-leaching techniques.

(c) Many of the basic chemicals re-quired for hydrometallurgicalprocesses are produced in largetonnages at low prices (e.g.,oxygen, ammonia, chlorine, hy-drochloric acid). Of course, ifyour process requires sulphuricacid, a smelter may pay you totake it away by the 1980s!

(d) Suitable materials of construc-tion, e.g., titanium, fibreglass, andplastics, have been developed atreasonable prices.

(e) The ever-increasing cost of trans-port and the tendency for con-centrates to be processed locallypossibly mean that smelters,whose processing costs arefavoured by very large units, areless competitive with hydro-metallurgical plants for the small-er tonnages often applicable tonew discoveries (e.g., 5000 to30000 tonnes of copper

JOURNAL OF THE SOUTH AFRICAN INSTITUTE OF MINING AND METALLURGY

annually). Many local companieshave recently found that thecost of transporting concentrates,custom smelting, and refining hasrisen so much that the need toconsider their own processingfacilities is becoming increasinglypressIng.

(f) As the price of copper has ]isen,the importance of working capitaltied up in copper inventory hasincreased, and it is believed thatvirtually all hydrometallurgicalprocesses will have a lower in-ventory than that of conventionalpyrometallurgical processes.

Despite the enormous effort beingdirected to research in hydrometal-lurgical processes, an equal, if notlarger, effort is being devoted to thefollowing:(i) improvements in existing pyro-

metallurgical processes, withspecial emphasis on the controlof sulphur dioxide pollution, and

(ii) developments in new pyrometal-lurgical processes in which thesmelting and converting stepsare integrated into one con-tinuous operation (or sequenceof operations), and which pro-duce a steady high-quality streamof sulphur dioxide gases, thusmaking the manufacture of sul-phuric acid easier and cheaper(e.g. Mitsubishi, Noranda, andWORCRA processes).

The various new processes beinginvestigated are well summarizedin a recent issue of Engineering &

Mining Jourml6.The cost of bringing sulphur

dioxide pollution under control willvary from region to region, depend-ing on the regulations in force, but,for conditions in the U.S.A., thecost has been estimated2 as anythingfrom 3,5 to 9,0 South African centsper kilogram of copper. These costsare significant when it is realizedthat the custom processing of con-centrates to refined copper in alarge smelter may cost about 15South African cents per kilogram ofcopper, direct operating costs beingas low as 7 to 10 cents.

The philosophy that, becausemany smelters are situated in re-mote areas, particularly in a countrylike South Africa, the abatement ofsulphur dioxide pollution is not veryimportant is unlikely to carry much

support in the future. The erection ofsmelters must involve the develop-ment of a township for personsworking at the smelter, and suchpersons are entitled to breatheclean air when away from work,even though they may have toaccept a certain discomfort while atwork. A partial solution to pollutionby sulphur dioxide does not seem tobe acceptable.

PRODUCTION OF SULPHURICACID

It is interesting to consider whatincrease in sulphuric acid productionfrom smelter gases could be ab-sorbed by other industries duringthe next decade. The 1972 pro-duction of sulphuric acid in SouthAfrica was approximately 1 400 000tonnes of 100 per cent H2SO4' ofwhich 22 per cent was used by themining industry (mainly in uraniumprocessing), 67 per cent in themanufacture of fertilizers, and 11per cent by other consumers.

The quantity of sulphuric acidthat could be recovered from smeltergases currently going into the atmo-sphere is approximately 120 000tonnes per annum. The air-pollutionauthorities are already taking certainsteps about persuading most of theseoperators to recover and use thissulphur dioxide.

It is estimated that, if all newcopper, zinc, lead, and platinummining operations smelt their ownconcentrates and are required torecover the sulphur dioxide fromtheir smelting operations, the pro-duction of sulphuric acid in SouthAfrica would be an extra 1 000000to 1 500 000 tonnes per annum by1985. This is likely to be more thanthe local mining and fertilizer in-dustries would require, unless majornew uses for sulphuric acid arefound. Unfortunately, most of thenew smelter capacity is likely to belocated in the north-western Capeand South West Africa, which arenear neither to the source of phos-phates at Phalaborwa nor to themain farming areas. It is fairlygenerally accepted that it is un-economic to transport acid morethan 100 to 200 miles.

There is a very good opportunityfor the export of fertilizers, especiallyto South America, and the opening


up of Saldanha Bay may make itworth while to site smelters nearSaldanha Bay so that the sulphuricacid produced can be reacted withphosphate from Foskor or from thedeposits at Langebaan, near Sal-danha Bay. It is a well-known factthat apatite (phosphate ore) is al.most as cheap a neutralizing agentfor sulphuric acid as lime is.

These facts highlight the im.portance of co-operation between themining and fertilizer industries inco-ordinating, as far as possible,future expansion in smelter andfertilizer plants. The recently an-nounced plans 7 of Impala and Omniato get together to produce phosphoricacid using sulphuric acid producedfrom Impala's stack gases is anencouraging move in this direction.

The cost of producing sulphuricacid from smelter gases depends onthe size of the plant and the type ofsmelter, but it is likely to lie betweenR20 per tonne for small plants andR5 per tonne for very large plants.In most instances, it is probablethat this byproduct acid will be soldat a negligible profit, and, in anassessment of pyrometallurgicalroutes versus hydrometallurgicalroutes, it seems sensible to neglectany credit from the sale of thesulphuric acid over and above theproduction cost.

HYDROMET ALL URGICALPROCESSES FOR

COPPER

There are many excellent publi-cations reviewing the hydrometal-lllrgy of copper2, 3, 8, 9, and Subra-manian and Jennings3 recently dis-cussed, concisely and clearly, thevarious alternatives that are beinginvestigated for the treatment ofchalcopyrite. Dasher2 recently critic-ally assessed the ability of hydro-metallurgy to compete with pyro-metallurgy. In this paper, I wish todiscuss only those processes thatappear to be attracting majorattention or appear to be promising.Some may regard the omission ofbacterial leaching as serious, but toolittle is known about the engineeringof these processes on a large scale topermit a critical assessment.

Discussion will be mainly limitedto the treatment of chalcopyriteconcentrates since chalcopyrite is

NOVEMBER1973 159

both the most-abundant coppermineral and the most difficult toleach. Any process that is to besuccessful must therefore be capableof treating chalcopyrite. A gooddeal has been written about hydro-metallurgical processes for the treat-ment of oxide ores! °, and these willnot be discussed here. After thevarious processes have been brieflydescribed, they will be comparedwith respect to certain key factors,namely, the degree of oxidation ofthe sulphur (elemental sulphur orsulphate), the character of the ironproduct, and the ability to handlemixed sulphides. The comparisonstrictly applies only to chalcopyrite,and quite different conclusions canbe drawn if, for example, chalco-cite is being considered.

Sherritt Gordon's Ammonia Pressure-Leaching Process

Sherritt Gordon Mines were pio-neers in hydrometallurgy, and muchof the credit for advancing the tech-nology of pressure leaching must begiven to them. They use pressure-leaching processes for treating nickelconcentrates, and they have licensed,among others, Western Mining inAustralia and Impala Platinum touse their technology. In laboratoryand pilot-plant tests on chalcopy-ritell, a 95 per cent extraction ofcopper has been achieved. Most ofthe unleached copper can be floatedfrom the residue and recycled.

Their flowsheet for the treatmentof copper-zinc concentrates is shownin Fig. 1. The basic equations for thedissolution reactions are as follows:

2CuFeS2+8t02+12 NH3+2H20=2 Cu(NH3)4S04+2(NH4)2S04

+Fe203' andZnS+2 °2+4 NH3=Zn(NH3)4S04'It will be noted that all the

sulphide sulphur is converted tosulphate, and, where there is amarket for ammonium sulphate, theprocess has a definite attraction. InSouth Africa, where ammonia usedfor fertilizer purposes is sold at alower price than ammonia used forindustrial purposes, ammonia wouldhave to be purchased at the ferti-lizer price before the ammoniumsulphate could be sold competitively.As it is unlikely that there will be alarge market for ammoniumsulphate, the process economics mustnormally be assessed on the basis

160 NOVEMBER 1973

,LIS SEP.'l.HATI("~~t-

J

RESIDUE

--_J

JFLOTATION

---r--hCRYLATE-

TAILING

COPPEA..lIN~ CONCEN TRATE

JNH3

AIR

PRES~URE LEACHING

COt;CENTRt.TE

HEC'(CLED

PRECIOUS ME:TALS

~COPPER POWDER

..CAKE fOR ZINC

RECOVERY

~ NH3 D'ST~ON l'r=-AIR --J.GfDROLYSI;j

240'C. &00 psiI

RECOVERY][LIS SEPA"~I:J

~-- ~CO21100 PSi> =:G

---,ZINC CARBONATE I

NH3 PR~(,IPITATIO~_J

SOLUTION

~::-J..~H2S---L~S STR~::~

LCRyST::'TlO~

AMMOI~IUM SULPHATE

Fig. I.-Flowsheet for ammonia pressure-leaching process (according toEvans et a/11)

that the ammonia will be regeneratedusing lime and the calcium sulphatewill be disposed of.

Sherritt Gordon Mines found thatpressures of at least 100 Ibfin2 or690 kPa (gauge) were necessary forleaching. Copper was recovered asa metal by precipitation with hydro-gen at high pressures, and zinc wasrecovered by precipitation with car-bon dioxide at high pressures. Allthese high-pressure operations resultin an expensive plant requiringcareful maintenance and operation.For the reasons given above, theseprocesses are unlikely to find muchapplication in copper processing.

The Arbiter Process!The Arbiter Process (Fig. 2), de-

veloped by Anaconda, is essentiallya modification of the Sherritt GordonProcess. It uses ammoniacal pro-cessing but eliminates the use ofhigh pressures. In the leaching step,oxygen, instead of air, is used(tonnage oxygen makes this possiblein certain locations) with specialagitation techniques that result ingood dispersion of the oxygen inthe slurry and with proper controlof the reaction by some undisclosedmeans. The availability of the liquid-liquid-extraction reagent LIX haspermitted the extraction of copperby liquid-liquid extraction and elec-trowinning. Techniques are also nowavailable for the recovery of zincand nickel without resort to the high


pressures used by Sherritt Gordon. Itwill be noted from Fig. 2 that, aswith high-pressure leaching, copperrecovery is not complete and un-reacted copper has to be floatedfrom the residue.

The Arbiter Process was de-veloped by Professor Arbiter ofAnaconda, after a thorough reviewof all the possible hydrometallurgicalprocesses, as the process that re-quired the least development andwas most likely to produce thecheapest copper for Anaconda. Alarge plant having a capacity of100 tonnes of copper per day isbeing erected in Arizona, and Ana-conda have claimed that the capitalcost is only half that of a smelter.

The process still suffers from themajor disadvantage that either am-monium sulphate must be sold orammonia must be recovered byneutralizing with lime. Whereprecious metals are not of anyaccount, Arbiter suggests that thelime should be added to the leachbecause this eliminates one step inthe liquid-solid separation and im-proves the filtration characteristicsof the iron oxide in the leach residue.

Because very few metals formammine complexes, all ammoniacalleaching processes have the ad-vantage of being very selective.This is particularly advantageous inavoiding the extraction of iron,which is almost invariably present inall copper concentrates.

Liquid-liquid extraction usingLIX in ammoniacal systems has theadvantage over acid systems thatthe solvent can take higher loadingsof copper. A disadvantage, however,is that LIX is less selective forcopper over nickel and zinc than inacid systems, but there are tech-niques for overcoming this problem.

Ammoniacal processes are idealwhen the gangue minerals are acid-consuming (e.g., calcareous or dolo-mitic).

The economics of the process areclearly critically dependent on theavailability of cheap oxygen, lime,and ammonia. The losses of am-monia will probably be finally knownonly when a large-scale plant is inoperation but should be less than inthe Sherritt Gordon process. Theseaspects will be discussed in greaterdetail later in the paper.

One further potential advantage ofthe ammoniacal system is thatcopper can exist in the cuprous form,which makes the possibility of directelectrowinning of the copper fromthe leach liquor rather attractive,since the power cost would probablyonly be about 50 per cent of thatfrom cupric solutions and the costof liquid-liquid extraction would beeliminated. This, however, repre-sents a major research project inelectrochemical engineering, and thepurity of the copper might be lowerthan that recovered from the stripliquor of a liquid-liquid-extractionplant.

Sulphuric Acid Pressure-LeachingProcesses

Complete oxidation of sulphideconcentrates can be achieved in therelatively short time of 1 to 2 hoursif temperatures in the range of 140to 200 °C and high pressures of airor oxygen are used. However, theequipment is expensive, and thecosts of heating the slurry and com-pressing the oxidant are considerable.Sulphate is formed under theseconditions, and the costs of oxidantand lime are therefore the same as inthe ammoniacal processes.

A lower-temperature process oper-ating below the melting point ofsulphur in the temperature range100 to 120 °C is to be preferred,since the engineering problems arenot nearly so severe, and the costsof heating and compressing theoxidant are also much less. A furthermajor advantage is that, at thesetemperatures, elemental sulphur,rather than sulphate, is formed. Thismeans that a possible salable by-product is produced instead of thesulphate, which requires lime forneutralization. In addition, the costof oxidizing the concentrate is onlyabout 30 per cent of that for thehigher-temperature process wheresulphate is produced. The differenceis shown by the following two simpleequations:

2 CuFeS2+2,5 O2 +2 H2SO4~2 CuS04+Fe203+4 S+2 H2O,

and 2 CuFeS2+8,5 O2+2 H2O~2 CuS04+Fe203+2 H2SO4'

The leach liquor, after removal ofthe copper by liquid-liquid extract-


ion or electrowinning, is returned tothe process for further leaching.The sulphuric acid shown in theleft-hand side of the first equation istherefore recycle acid, and theprocess has a very low net acidconsumption. In fact, it appearsthat about 5 per cent of the sulphurwill be converted to sulphuric acid,thus providing the necessary make-up of acid12.

These pressure-leaching processeshave the advantage that iron isprecipitated as a ferric hydrousoxide, which reports in the residue.The quantity of iron remaining insolution is a function of the acidityand of the time and temperature ofleaching, but, because the solution isrecycled to the leach after the re-moval of the copper, all the iron willeventually report in the residue. It isalmost certain that the actualoxidation attack on the mineral isby ferric iron, and a small quantityof dissolved iron is therefore neces-sary.

Owing to the relatively low levelof iron in the leach liquor, directelectrowinning of copper from theleach liquor can be considered, thussaving the costs of the liquid-liquid-extraction plant.

The effect of various operatingparameters on the leaching processhave been studied extensively byVizsolyi et al.12, and Fig. 3 is aflowsheet of the process suggestedby them. Under the optimum con-ditions of a temperature of 115 °Cand an oxygen partial pressure of200 to 500 Ibjin2 or 1380 to 3450 kPa(gauge) on a concentrate that hadbeen ground to 99,5 per cent minus325 mesh, the percentage copperextraction was only about 65 percent after 2t hours. After removalof the sulphur, the undissolvedchalcopyrite was floated and re-cycled for further leaching. Althoughmany of the hydrometallurgical pro-cesses being considered appear torequire flotation of the leach residuefor the recovery of undissolvedmineral, such a low percentagedissolution as 65 per cent appears tobe unacceptable, especially when itis considered that very intenseleaching conditions were used.

Vezina13 studied this problem ingreater depth on a chalcopyrite-

NOVEMBER1973 161

OXY9 en

~Cu Concentrate

Ar(1monl a

Direct discard ofleached resldues and9ypscm if Inlernallimefeed IS used.Applicable if concentratefeed contains ne91i9ibleprecious metal values.

Discard of pyriteand 9ypsum,ifinternal lime feedis used.

EXTRACTION

STRIPPING

I

I _\

5tron9

ELECTROWINNING IElectrolyte

--_::JMetallicCopper

Fig, 2-Slmpllfled flowsheet of the Arbiter Process (according to Prlcel)

RETURN

CuFeS, CONC. (-30'/.Cu)

---F "

[;2- _l _l

;]GRINI)iHG

99.S'~'-

-:j;~~~C:SH

250 ,-3;zr~}11 COIK,

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.

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ACID

-100 91>1 H,SO.- 20 ~pl Cu5 ~pl fcr

L--;~;;~;;;-L" ,--r'::'-1-

[~;~-~;)AHATlOr~- LEACH

---~-~ACHT~OL: N

nfSIDUr:

-60 IIpl Cu - 5 9pl fe

I~£,~~'-D_~

(I.

£'~~'!:!~~QQ.f,.?

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-~:I~ JFe(OH.) GANGUE

-

[ S'

.

SE PM) AT 15--> !~I!"!f_~'

IL' l'--

~~?iLu~.r

L--RECYCLE CON,C,

Fig, 3--Example of a pressure-leaching scheme for chalcopyrite (according toVlzsolyl et al. 11)

162 NOVEMBER 1973 JOURNAL OF THE SOUTH AFRICAN INSTITUTE OF MINING AND METALLURGY

pentlandite-pyrrhotite concentrateand showed that the degree ofgrinding is critical to the achieve-ment of high copper extractions.On a product that was ground to93 per cent minus 20 /Lm, the copperextraction was 91 per cent after8 hours at an oxygen pressure of80 Ibjin2 or 550 kPa (gauge). Analysisof the residue showed that theextraction for particles less than20 /Lm exceeded 95 per cent.

The importance of grinding wastaken one step further by Gerlachet al.14. It was found that the sub-jection of chalcopyrite to a millingaction involving a strong impact(vibration mills are commerciallyavailable) resulted in copper ex-tractions of 99 per cent in 1 hour atan oxygen pressure of 2000 kPa anda temperature of 110 °C. The degreeof grinding is not given. Gerlachet al. attribute this surprising resultto the fact that impact milling notonly comminutes the mineral sub-stances but also mechanically dis-turbs the lattice structure of themineral so as to render it moresusceptible to oxidizing chemicalattack. This theory is supported byX-ray-diffraction studies and alsoby electrochemical studies, whichshowed that the redox potential foractivated chalcopyrite was only320 mV, compared with 550 mVfor non-activated chalcopyrite.

The approach being pursued byGerlach et al. clearly represents asignificant advance on the results ofprevious workers, mainly becausecomplete leaching of the copper in

~""',. .Oxi~e ore

t

one step now appears to be a possi-bility. It has been reported thatLurgi are testing the process on apilot-plant scale, and it must cer-tainly be regarded as one of thecontenders for commercial appli-cation.

Other investigators have alsorecognized the importance of acti-vating chalcopyrite to make iteasier to leach. Copper15, sulphur16,nitric acid17, and hydrogen18 pre-treatments, among many others,have been tried. These are all de-scribed and compared by Subra-manian and Jennings3.

Roasting ProcessesRoasting processes are well-

established and have been ex-tensively described in the metal-lurgical literature. The basic patentusing fluidized-bed reactors wasissued in 1957, and successful com-mercial plants are operating inZambia and Zaire. Over 300000tonnes of copper per annum areproduced by this process in Zairealone.

Most processes involve a sul-phation roast, followed by a wateror dilute-acid leach of the copperwith some dissolution of iron. Copperis recovered by electrowinning, andthe sulphate formed is removed byliming. (This process is often re-ferred to as the R-L-E process.)The temperature of roasting iscritical: it is essential to select atemperature that is high enough todecompose Fe2(SO2)4 but is not highenough to result in the formationof insoluble copper ferrite.

E,..:~'.:le :'=eof

"ater

I

:xi1~ :re c'.; ;('P c,..

rACia

1- Gase,-I

't'I1-p.eSio]ue-1

\

II~ S"d

I.o.lectrolyte

t *te31-:1'1e-

11

t S~LruP.IC ACD ?:J..NT

SDen+~~Solution

Co;l

I. S?JN.}E IP.~1f?LI.:r::

IC;l'on~e Iron

CEI.\E,;TA.TICN

tCellOcnt Copper

cR"JSE::;;

~ .

L,B11e:::, <°: '"~;':;.cr::;

. IC;n"~",n;.:e

R:I,S':I;T:

.--Air

1cel:s:",

of... -r--;:ater

lEACHIN} 1,,;;;-',:ASC!IN}I

:.,.eLe.',

Selr:'"

ELECTEIWI!C':2C

C.thOdt CoDper

Fig. 4-Roast-leach-electrowlnnlng process as used by Hecla19


Recla Mining Company19 recentlyconducted an extensive experimentalprogramme to develop the R-L-Eprocess for use at its LakeshoreMine. They operated a pilot plantproducing 1 ton of high-puritycathode copper per day. This pro-cess has its major attraction whenit is required to treat ores of mixedoxides and sulphides, as at Lake.shore, and the flowsheet as de-veloped for Lakeshore is given inFig. 4. The sulphur eventually allreports as sulphuric acid, which canbe used in the leaching of theoxide. It would have been interest-ing if Recla had compared the eco-nomics of liquid-liquid extractionwith those of the sponge-iron pro-cess in the production of copper fromthe oxide leaching circuit.

Recla made extensive studies onthe pilot plant to find the optimumroasting temperature for copper re-covery at a minimum production ofdissolved iron. By operating in thetemperature range 690 to 700 °C,they achieved copper dissolutions of97 per cent and iron extractions ofless than 5 per cent.

Other organizations (e.g.,CSIRO2O, V.S. Smelting & RefiningCompany21, Dowa Mining22) havedeveloped alternative methods forthe recovery of the copper fromsolution after roasting and leaching.Most of these alternative processesare designed for the treatment ofmixed copper-zinc ores.

Recently, the V.S. Bureau ofMines23 proposed a sulphating roastwith lime in the roast to overcomeone of the major disadvantages ofthe conventional R-L-E process,namely the production of a roastergas containing sulphur dioxide.Where there is no market for acid,the sulphate may as well be con.verted to gypsum in the roasterrather than later by neutralization,which is known to be a tricky oper-ation. Roasting is done at about600 DC,and, in work done at NIMon this process, extractions exceed-ing 97 per cent have been achieved.The process appears to have aparticular advantage for flotationconcentrates containing acid-con-suming gangue. It has been shown24that the quantity of lime required tofix the sulphate can be decreased if

NOVEMBER 1973 163

the concentrate contains calcareousmaterial.

The major problem with roastingprocesses is the dissolution of iron.If a strongly acid electrolyte isused for the dissolution of thecopper, iron dissolution may beexcessive, resulting in a neutrali-zation problem and a less purecathode. If water is used to dissolvethe copper, copper extraction maybe decreased, and, because of thelow acid strength, electrowinning isnot so attractive although liquid-liquid extraction now becomes pos-sible. The D.S. Bureau of Mines25have recently shown that copperdissolution in the roasting processcan be improved by the addition ofFe20a to the furnace charge, butthe results need to be confirmed on apilot-plant scale.

The Cymet ProcessThe Cymet Process is one of the

many chloride processes beingstudied. One of the major attractionsof chloride processes is that theleach can be operated at atmo-spheric pressure. It is reported thatDuval and Cominco are also de-veloping chloride-leaching processes,but no details have been published.In contrast, Cyprus MetallurgicalProcesses Corporation have pub-lished the results of their researchfairly extensively26. 27, and are pre-pared to discuss the possibility oflicensing their technology. A pilotplant treating 25 tonnes of concen-trate per day is due to becomeoperational during 1973. The plantcost is about R6 million.

Fig. 5 presents the flowsheet de-veloped and the pilot plant testedfor chalcopyrite. Reactivity is in-creased by an increase in the surfacearea, and the flotation concentrateis ground to 95 per cent minus 200mesh. Leaching with ferric chlorideanolyte solution from the ironelectrolytic cells produces the follow-ing reaction:

6 FeCla+2 CUFeS24-2 Qual+8 FeC12+4 S.

The countercurrent leaching,which takes place at about 75 to80 °C, employs cyclones betweenleaching stages to provide maximumefficiency. This results in a negli-gible concentration of ferric chloridein the leached slurry, which goes tothe anode compartment of the

164 NOVEMBER 1973

electrolytic dissolution cells (dia.phragm-fitted cells with a woven,permeable synthetic material thatseparates anode from cathode com.partments). Anodes are made oftitanium that is coated with con.ducting oxides, which permit highcurrent densities (200 Ajft2) andalso give lower voltage drops. Copperis recovered from the leach andanolyte solutions in the cathodecompartment of the cell. The cath.odes are round copper rods uni-formly spaced and oriented parallelto the anodes in the opposite com-partment. Copper is precipitated asa powder and removed as a slurry.The copper powder goes to electro-refining.

According to Cymet, the equationsgoverning the processes in the cellare as follows:

CuFeS2+3 HCI-3e4-CuCl+FeCI2+2 S+2 H+,

3 CuCl+3 e4-3 Cu °+3 Cl-, and3 H++3 CI-4-3 HCl.

The solids from leaching areautoclaved at about 135 °C for 2hours to form globules of sulphur,which, on being cooled, can bescreened off. The screen underflow issubjected to flotation, the concen-trate being recycled for leachingand the tailings being rejected.

The spent copper catholyte isfurther depleted of copper by ce-mentation with metallic iron. Thesolution is further purified by zinccementation so that residual tracesof copper and other impurities, suchas lead, antimony, bismuth, andarsenic, are removed. Zinc occurringin the feed and added for purifi-cation is removed by liquid-liquidextraction with a tertiary amine.

The purified ferrous chloridesolution is sent to electrolytic ironcells, which plate massive high-purity iron on iron starter sheets,and ferric chloride is regeneratedat the anodes. Because of theimbalance between cathode andanode efficiencies, and the use ofiron to cement the copper, an ad-ditional amount of iron must be re-moved from the circuit. This isdone by hydrolysis. The reactionsoccurring at the iron cells are asfollows:

3 FeC12+6 E4-3 FeO+6 Cl-, and6 FeC12-6 e+6 CI-4-6 FeCla.

The pilot-plant tests indicate thatabout 50 per cent of the copper isdissolved in the ferric chlorideleaching step and another 30 per centin the electrolytic cells, and 20 percent is recycled by flotation. Theoverall extraction exceeds 98 percent.

Costs presented by Cymet for aplant producing 80 000 tonnes ofwirebars per annum indicate thatthe operating and capital costs areslightly less than those for a con-ventional smelter constructed in theD.S.A. If credit is given for theiron and sulphur, the Cymet Pro-cess is overwhelmingly cheaper (seeTable I). A better estimate of thecosts will be made when the largepilot plant has been operated.

The Cymet Process is very in-genious in its use of electricalenergy. No figures for power con-sumption are given by Cymet, butit is estimated that, allowing forinefficiencies, the power consumptionof the two electrolysis steps isprobably in the range of 3,3 to4,5 kWh per kilogram of copper. Inaddition, about 0,3 kWh per kilo-gram of copper is required for theelectrorefining of the copper. Thepower consumption for conventionalelectrowinning from sulphate mediais about 2,2 kWh per kilogram ofcopper. The Oymet Process, inaddition to winning copper, alsooxidizes the leach liquor and pro-duces metallic iron for the ad-ditionall,4 to 2,6 kWh per kilogramof copper, which makes it veryattractive. In fact, if the iron can besold for the price suggested byCymet, the cost of iron recovery willbe less than half the sales value.The production of this salable ironbyproduct appears to be the majoradvantage of the Cymet Process.There are other chloride and sulphateprocesses that produce elementalsulphur, so that the Cymet Process'sproduction of elemental sulphurgives it no particular advantage.

Cymet have taken advantage ofthe fact that, in chloride media,copper can exist in the cuprous


TABLE IOF COPPER PRODUCTION COSTS, CONVENTIONAL SMELTER VERSUS CYMET

PROCESS (U.S. CENTS PER POUND OF RECOVERED COPPER)

Conventionalpyrosmelter +electrorefining

10,1030,781

COMPARISON

Direct operating costs""""""""""""Administrative and indirect costs. . . . . . . . . . . . . .

Total direct and indirect costs. . . . . . . . . . . . . . . .Fixed costs (depreciation)

""""""""""

Totaloperatingcosts ,....Credits.

Gold, silver, lead, and z.inc , . . . . . . . . . . . . .Sulphur and iron ,....

Totalcredits""""""""""""""

10,8842,733

13,617

CymetProcess +

electrorefining9,5681,019

10,5872,022

0,7900,535

12,609

0,7904,361

1,325 5,151

Conversion cost per Ib of recovered copper. . . . . .Conversion cost per Ib of total contained copper. .Capital costs

"""""""""""""""". Credits were estimated on the following values:gold at $42,00 per oz tr., silver at $1,45 per oz tr., lead at $0,06 per Ib, zinc at $0,085per Ib, iron (electrolytic) at $0,05 per Ib, sulphur equivalent at $10,00 per long ton,and power cost at $0,0105 kWh.

12,292 7,45812,939 7,5901,0 compared with 0,83

all the leaching in the primaryferric-leaching step, thus simplifyingthe cell operations.

The Cymet Process certainlymerits consideration where power

form, and the power consumptionfor winning it is therefore onlyabout half of that in sulphate media.A disadvantage is that, because it isnecessary to operate the cells atsuch high current densities and inthe presence of such high concen-trations of iron, copper powder isproduced, which requires furtherthe electrorefining. It is known that,although the power consumption forthe electrorefining of copper is muchlower than for the electrowinning,the capital costs of the plants, ifworking capital for copper is in-cluded, are not very different. Itwould therefore be a great ad-vantage if high-purity copper couldbe produced direct, but the chancesof doing this appear to be small.

The major disadvantage of theprocess concerns the technologicalproblems involved in the operationof the slurry electrolytic cell. Thisslurry operation appears to be un-necessary because, at first sight, theanode reaction in this cell is merelythe oxidation of ferrous to ferric ionand/or cuprous to cupric ion. Thechalcopyrite is then oxidized in situ.It is claimed that the advantage ofthis 'electro-oxidation' step is thatlower leaching temperatures can beused and less sulphate is formedthan in conventional ferric chlorideleaching. Experience with directferric chloride leaching28 does notconfirm this, especially if the ob-jective is to achieve a primary dis-solution of only 80 per cent as inthe Cymet Process. I wonder whetherCymet should not consider doing

Copper concentrates

costs are low and the byproductiron can be sold.

Ferric Chloride Lea£hing and Liquid-liquid Extraction

During the past two years, NIM28has investigated, on a pilot-plantscale, a ferric chloride leaching pro-cess involving copper recovery byliquid-liquid extraction with LIX-64N and electrowinning of thecopper from the sulphate stripliquor. Because of pressure of work'the research is not complete, butcertain basic features of the processhave been proven. A flowsheet of oneversion of the process is Rhown inFig. 6.

Laboratory and pilot-plant ex-perience has shown that, for copperextractions to exceed 90 per cent, itis necessary for the concentrate tobe ground to 95 per cent minus400 mesh, thus confirming the ex-perience of other investigators in

Electro-dissolution

Sulphur agglomeration

Flotation

Tailings

(Solutionl

Cu and Pb cementation

Zn solvent extraction

Iron electrolysis

Iron hydrolysis

(Solution)

--- Copper product

Sulphur

Recycle

concentrates

rCu cement..1 Pb cement

{ZnIOH},

FelOHIJ

1ran cathodes

Iron oxides


Fig. 5-Flowsheet for the Cymet Process 26' .7

NOVEMBER 1973 165

Air oroxygen

Oxidation

Return of HCt or FeCI, solution

the leaching of chalcopyrite.Copper extractions have been

found to be variable, depending onsuch factors as age of the con-centrate, whether it was stored wetor dry, whether it had been pre-leached with hydrochloric acid forthe removal of zinc, etc. Leachingat 95 °C for about 8 hours gavecopper extractions of 95 to 97 percent. Leaching vessels lined withhard rubber were used, and theseappeared to stand up to the leachingconditions without deterioration.Contrary to statements by otherresearchers, the amount of sulphurgoing to sulphate did not exceed5 per cent of the total sulphur.

Copper was extracted from theleach liquors with LIX-64N, theloaded solvent was suitably scrubbedto reduce the iron and chloridelevels, and the copper was finallystripped with spent electrolyte fromthe electrowinning plant. The coppercathodes produced were of a high

166 NOVEMBER 1973

Flotation concentrate

Fine grinding

Solvent extractior.

Cu electrowinning

Copper cathodes

Final copper remo.'al

Iron removal

Fe,O, or Fe

Fig. 6-NIM ferric chloride leaching circuit

quality. Because the loading ofcopper in LIX from chloride mediais lower than that from sulphatemedia owing to the complexingeffect of the chloride ions, it isnecessary to operate with moderatelyhigh levels of copper in the raffinate.This means, then, that copper mustbe removed from the bleed streambefore the iron is recovered. If thebleed stream is first oxidized toconvert ferrous chloride to ferricchloride, the excess hydrochloricacid generated in the liquid-liquid-extraction process is consumed andmost of the remaining copper can beextracted in a secondary liquid-liquid-extraction circuit. Varioustechniques such as iron and zinccementation can be used to removethe last traces of copper.

A special foam aeration columnhas been developed and tested ona pilot-plant scale to oxidize ferrouschloride to ferric chloride. Thisdevice gives good oxygen efficiency,

A'~+ Ag

using air at about 45 Ibjin2 or 300kPa (gauge), has no moving parts,and is of cheap and simple con-struction. Good rates of oxidationare achieved at 95 to 100 °C. Oxi-dation of the leach slurry has alsobeen successfully attempted.

Initial plans for the removal ofdissolved iron from the circuit re-volved round the use of a pyro-hydrolysis technique (well provenfor pickle liquors, e.g., the AmanProcess), in which iron chloridesare converted to iron oxides bywater vapour at high temperatures.This serves to regenerate thechlorides as hydrochloric acid forrecycle to the oxidation step of theprocess. The iron oxide might be asalable byproduct. However, de-tailed studies have indicated that,except when very cheap fuel isavailable, the cost of fuel is ratherhigh. Capital costs represent a fairlysignificant proportion of the totalplant cost.


Tests are being done on an al-teI'Ilative route involving the preci-pitation of FeO. OH by oxidationof the feITous chloride liquor atabout 100 DC using the oxidizeI'previously mentioned. A filterableprecipitate has been produced, butso far it has not been possible toreduce the chloride content of theprecipitate to a level acceptable fordischarging. This process wouldotherwise be much cheaper.

The remaining alteI'Ilative is tomake electrolytic iron and oxidizeferrous to ferric chloride as is donein the Cymet Process. If the iron canbe sold for a reasonable price, this ismost likely to be the best route tofollow.

A detailed comparison of theCymet Process with the NIM feITicchloride process has not been under-taken, but it is believed that theNIM process could well be cheaper,especially as the Cymet Process in-volves the extra step of electro-refining the copper powder. Furtherpilot-plant testwork is required be-fore the process can be regarded asproven on this scale of operations.

Haver and Wong29 of the V.S.Bureau of Mines also studied a feITic-leaching process but suggested theproduction of cement copper, which,in my opinion, is a retrograde step.

DEVELOPMENT IN THEFIELD OF COPPERRECOVERY FROM

SOLUTION

The use of LIX liquid-liquid-ex-traction reagents combined withcopper electrowinning in sulphatemedia is now an industrially provenprocess4, 5. Extremely pure coppercathodes can be produced from theseprocesses, the only problem elementbeing lead, which originates fromthe lead-antimony anodes used. Alead content of less than 10 p.p.m.is highly desirable, and hence agreat deal of attention has recentlybeen devoted to finding a means ofreducing the lead value, which insome plants has been as high as100 p.p.m. It has been found thatthe addition of very small quantitiesof cobalt greatly assists in de-creasing the lead content.

De Nora of Italy and I.C.I. inBritain both have substantial re-search programmes aimed at pro-

ducing titanium anodes coated withprecious metal that have a long lifein copper sulphate electrowinningcircuits. According to present indi-cations, the life of these anodes is notlong enough to justify their veryhigh capital cost, but they do permitthe use of higher current densitiesthan do conventional anodes, andthe voltage drop is slightly less. It ispossible that further developmentwill result in their application tocopper processing.

Recent work by many workers3O-33has demonstrated that it is possibleto operate copper-electrowinningplants at much higher current densi-ties (600 Afm2) than previouslypractised (250 Afm2). This can bedone by periodical interruption ofthe cUITent32 or by specially designedail' nozzles33 to improve the circu-lation of the liquor, especially nearthe cathode surface. Others3O haveimproved the liquor circulation byvarious means.

The production of a purer electro-lyte when LIX is used makes itpossible to operate at higher CUITentdensities than when the liquor isimpure. It may be possible in certainInstances to justify the cost ofiiquid-liquid extraction by the lowercost of the electrowinning plant. Theresults of a cost-optimization study31indicate that the optimum CUITentdensity, considering the operatingand the capital costs, is in the regionof 600 Afm2, which appears to bewithin the bounds of practical appli-cation in the near future. Theproblem of how to control the mistarising from cells operating at thesehigh current densities still requiresto be solved. It is worth thinkingabout the possibility of ducting theoxygen away for use in the leachingstep and thus also overcoming themist problem.

The capital cost of a typicalcircuit for liquid-liquid extractionand electrowinning is proportionedapproximately as follows: liquid-liquid-extraction plant 20 per cent,solvent 20 per cent, and electro-winning plant 60 per cent. Operatingat high current densities will meanthat the cost of the electrowinningplant will be reduced by about one-third, and the cost of the liquid-liquid-extraction plant and itssolvent will then be about equal to


the cost of the electrowinning plant.Research to decrease the cost of theliquid-liquid-extraction plant and itssolvent inventory is therefore urgent.ly required.

An alteI'Ilative solvent, KELEX,is selective for copper and undercertain circumstances promises to bepreferable to LIX, but it does not yethave the advantage of having beenused on a large plant. Many com-mercial firms are actively developingnew solvents for copper.

One of the major disadvantages ofliquid-liquid extraction is that theleach liquors must be clarified afterliquid-solid separation so that thelosses of solvent are minimized. Withthe recent successful development ofa continuous ion-exchange tech-nique34, new horizons open up, andit is possible to consider the possi-bility of ion-exchange processes forthe recovery of copper from adilute-leach sIUITY. This would re-duce the costs of liquid-solid sepa-ration and also eliminate the needfor clarification. The major area ofapplication for ion-exchange pro-cesses would doubtless be the leach-ing of oxide or carbonate ores, wherbthe costs of liquid-solid separationare very high and a dilute coppersolution is produced. The ion-ex-change step could then be used as aconcentrating step, followed by aliquid-liquid-extraction step if neces-sary. For these objectives to berealized, new and improved resinshaving properties similar to or betterthan those of LIX or KELEX mustbe developed.

COMPARISON OFPROCESSES

The only fully realistic and con-vincing basis for a comparisonof processes is, of course,operating and capital cost. How-ever, when so many processes areavailable and such a variety offeed materials can be considered, itseems to be important at least tomake a rough comparison of thevarious major features of the pro-cesses available. A change in the'ground rules' (e.g., CQst of power orlime) can alter a comparison ap-preciably.

In Table n, eight different pro-cesses have been compared, andsome of the features of the pro-

NOVEMBER1973 167

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cesses will now be discussed in motedetail. The numbers used in thediscussion refer to the numbers givenin the Table for the different pro.cesses.

Iron Removal, and Oxidation State ofthe Sulphur Produced

These two features together, morethan any other feature, probablydistinguish the various processesfrom one another, and also have thegreatest impact on the operatingcosts. Those processes that formsulphur (nos. 3, 4, 7, and 8) haveonly about 30 per cent of theoxidation requirement of those pro-cesses that make sulphate (nos. 1,2,5, and 6). The processes in the firstgroup have the additional advantagethat the sulphur byproduct has apotential sales value either do-mestically or on the export market.The processes in the second grouphave the disadvantage that thesulphate must be neutralized withlime to form calcium sulphate, whichmust then be dumped.

In some processes (nos. 1, 2, 3,and 4), the Iron is advantageouslyprecipitated during the leach. Inothers (nos. 5 and 6), the iron isrendered insoluble in the roastingstep. The two chloride processes(nos. 7 and 8) are unique in that ironis dissolved, and they thus requirean extra step for its removal. How-ever, there is a possibility that asalable byproduct can be recovered,which would mean that these pro-cesses have the edge over those Inwhich iron is not recovered.

The cost comparison In Table 11serves to bring out the above pointsmore clearly. The following com.ments can be made.(1) At the price of oxygen and power

assumed, the Arbiter Process haslittle advantage in the operatingcost of the leaching step over theconventional Sherritt Gordonprocess. Its advantages lie Inmuch simpler equipment and alower ammonia loss, togetherwith the introduction of theli qui d -li qui d -extraction/ electro-winning step.

(2) Both ammonia leaching pro.cesses are considerably more ex-pensive than the other processes.In large measure this is due tothe cost of liming, thus demon.strating that these processes have

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their major advantage where(NH4)2S04 can be sold.

(3) The other processes do not differvery much in cost, and the choiceof process must probably be madeon the basis of other factors, in-cluding capital costs, costs ofchemicals, etc.

(4) The sulphuric acid pressure-leaching processes and the chlorideprocesses have the advantage ofproducing elemental sulphur,which, if sold at RIO per tonne,would provide a credit for theseprocesses of 1 cent per kilogramof copper.

(5) The Cymet Process has the ad-ditional advantage that the ironcan be sold. Its value is un-certain but would probably be inthe range of R15 to R50 pertonne, depending on its purityand end use. This would providean additional credit of 1,5 to5 cents per kilogram of copper,which tends to make the CymetProcess very attractive if bothiron and sulphur can be sold.

(6) The ferric chloride process usingelectrolytic recovery of iron (no.8) is also very attractive if bothiron and sulphur can be sold.Its capital cost is likely to beless than that of the CymetProcess.

Copper Recovery from SolutionAt least six of the processes listed

(all except nos. 1 and 7) have aconventional copper-electrowinningplant. The cost of these plantsrepresents anything from 20 to 40per cent of the total capital cost and,therefore, although the cost willvary slightly for the different pro-cesses because of different electro-lyte conditions, in principle all theseprocesses have a common capital andoperating cost. The possible ad-vantages of a very pure electrolytehave already been discussed.

As already indicated, the capitalcost of an electro-refining plantwhen the copper inventory is in-cluded is not much less than thecost of an electrowinning plant, sothat this is a cost common to theCymet Process and the other pro-cesses. The power consumption forelectro-refining is, of course, onlyabout 15 per cent of that for electro-winning (about 1,3 cents per kilo-gram of copper at a power cost of

0,6 cents per unit). The cost in theSherritt Gordon process of precipi-tating copper powder by the use ofhydrogen is not known, but, at theprice of hydrogen in South Africa, itis unlikely to be much less (if any)than the cost of electrowinning. Thecopper powder produced is aspeciality sales product. The factthat Anaconda have chosen liquid-liquid extraction and electrowinning,rather than hydrogen precipitation,suggests that, even in the D.S.A.,cathodes are preferable and cheaper.

The purity of copper produced byhydrometallurgical processes usingliquid-liquid extraction (nos. 2 and7, and possibly 4, 5, and 6 if desired)is superior to that from pyrometal-lurgical processes and from otherhydrometallurgical processes not em-ploying liquid-liquid extraction, andunder difficult marketing conditionsis likely to be a selling advantage.The cost of solvent make-up isestimated to lie in the range 0,3 to0,5 cent per kilogram of copper.

Cost of ChemicalsThe cost of the lixiviant for the

four sulphate-based processes (nos.3 to 6) can be regarded as negligible.

The price of ammonia in SouthAfrica is very high compared withits price in other industrializedcountries, although it is predictedthat, with the increasing shortage ofoil and natural gas and hence therising prices, ammonia producedfrom cheap South African coal willeventually be among the cheapestin the world. No realistic figures areavailable for the consumption ofammonia in the two processes con-sidered, but it is believed that itshould be possible to limit the lossesso that the make-up cost will bein the range 1,0 to 2,0 cents perkilogram of copper. The ArbiterProcess will probably be lower incost than the Sherritt Gordon pro-cess. As mentioned previously, if(NH4)2S04 is to be sold as a fertilizer,ammonia will have to be purchasedat the fertilizer producer's price.

No realistic figures for chloride lossare available, but it is expected thatthe make-up cost would be less thanthat for the ammoniacal processes(say 0,5 to 1,0 cent per kilogram ofcopper), because chlorides are vola-tile like ammonia, and the basicraw-material price (hydrochloric


acid) is somewhat lower. Much ofthe ammonia could be lost with theCaSO4, but no CaSO4 is produced inthe chloride processes.

Type of Concentrate

Where a concentrate containsacid-consuming gangue (e.g., of acalcareous nature), only the am-moniacal and the lime-roast-Ieachprocesses are applicable. The lime-roast-leach process (no. 6) seems tobe particularly suitable for thispurpose, because the presence ofcalcareous material reduces the oper-ating cost.

The ability of the various pro-cesses to treat mixed concentrateE(e.g., copper-zinc, copper-zinc-Iead,and copper-nickel-cobalt) is of im-portance because this is a field inwhich hydrometallurgical processefhave a possible major advantagE,over pyrometallurgical processes. Itis believed that adequate processingknowledge is available to solve theseseparation problems for ammoniacal,sulphate, and chloride processes.Separations can probably be moreeffectively achieved in ammoniacaland chloride systems, but every casehas to be looked at separately. Moreresearch is certainly needed to pro-vide improved and cheaper sepa-ration processes.

From a survey of the experienceof various investigators in the leach-ing of chalcopyrite, it appears thatthe reactivity of chalcopyrite variesconsiderably, and much more at-tention should be given to an under-standing of the reasons for thisdifference. One investigator evenfound that, by dissolving the flo-tation chemicals from a concentratein an organic solvent, the copperdissolution increased from 90 to95 per cent.

Recovery of Precious Metals

Sulphide ores always contain someprecious metals. The degree towhich precious metals will dissolvein the leaching step will dependvery much on how they are presentin the concentrate, e.g., they can bealloyed with base metals. However,it is fairly certain that, the moreintense the leaching conditions andthe finer the grind, the greater willbe the tendency for the preciommetals to dissolve if the leachingconditions permit this. Palladium,

NOVEMBER 1973 169

platinum, and silver, for example,are all soluble in ammoniacalsolutions, and under pressure-leach-ing conditions appreciable quantitiesof palladium, especially, have beenfound to dissolve. Palladium andsilver are soluble in sulphate media,but there is no reported instance ofany investigator having detected thedissolution of these metals insulphate-leaching tests. All theprecious metals are soluble inchloride media, and experience hereseems to vary. Work done at NIMon a local chalcopyrite and copper-nickel matte showed negligible dis-solution of precious metals. However,Cymet report recoveries of silverand gold into the copper of 96 percent and 60 per cent respectively27.How this took place is difficult tounderstand.

Various workers have investigatedthe recovery of gold and silver fromleach residues using the cyanidationprocess, and good success has beenachieved. Thiourea processes mightalso be considered under certaincircumstances.

CONCLUSION

This review has demonstratedthat, within a few years, a varietyof hydrometallurgical processes willbe available that are likely to becompetitive with pyrometallurgicalprocesses when there is no sale forsulphuric acid from the smelter.The selection of the right process willdepend on the character of the feedmaterial, the location and size ofthe plants, and the byproducts thatcan be recovered and sold. Ingeneral, it appears that processesproducing sulphur have an ad-vantage over processes producingsulphate as a reaction product.

REFERENCESI. PRICE, F. C. Copper technology on the

move. Engng Min. J., vo!. 174, no. 4.

170 NOV(:"'!;I~R 19~

Apr. 1973. pp. RR to HHH.2. DAsHER, J. Hydrometallurgy of

copper concentrates? Can. Min.Metall. Bull., vo!. 66. 1973. pp. 80-88.

3. SUBRAMANIAN,K. N., and JENNINGS,P. H. Review of the hydrometallurgyof chalcopyrite concentrates. Can.Metall. Q., vo!. 11, no. 2. 1972.pp. 387.400.

4. AGERs, D. Liquid ion exchange: aneconomic analysis. Paper presented atthe Southern Mineral Conference,Phoenix (U.S.A.), 25th Apr. 1972.

5. JONES, R. L. Operating costs atBagdad Copper Corporation's liquidion exchange-electrowinning plant.Paper presented at the Pacific.Southwest Mineral Industry Con.ference, Phoenix (U.S.A.), 25th-27thApr. 1973.

6. ANoN. Confrontation in copper. EngnJMin. J., vo!. 174, no. 4. Apr. 1973.

7. SWIFT, K. Impala digs in with Omnia.Sunday Times, (Johannesburg), Busi-ness Times Section, 19th Aug., 1973.

8. W ADSWORTH,M. E. Advances in theleaching of sulphide minerals. Miner.Sci. Engng, vo!. 4, no. 4. Oct. 1972.

9. ROMAN, R. J., and BENNER, B. R.The dissolution of copper concen-trates. Miner. Sci. Engng, vo!. 5,no. 1. Jan. 1973.

10. HOLMES, J. A., and FISHER, J. F. C.Development of a process for ex-traction of copper from tailings andlow-grade materials at the ChingolaDivision of N'Changa ConsolidatedCopper Mines, Zambia. Presented atthe Advances in Extraction Metal-lurgy and Refining Symposium, organ-ized by the Institution of Mining andMetallurgy, London, 4th-6th Oct.,1971.

11. EVANS, D. J. 1., ROMANCHUK,S., andMACKIW, V. N. Treatment of copper-zinc concentrates by pressure hydro-metallurgy. Can. Min. Metall. Bull.,vo!. 57. Aug. 1964, pp. 857-866.

12. VIZSOLYI,A., VELTMAN,H., WARREN.1. H., and MACKIW,V. N. Copper andelemental sulphur from chalcopyriteby pressure leaching. J. Metals, N. Y.,Nov. 1967. pp. 52-58.

13. VEZINA, J. A. Further studies on acidpressure leaching a chalcopyrite-pentlandite-pyrrhotite concentrate.Can. Min. Metall. Bull., May 1973.pp. 57-60.

14. GERLACH, J. K., GOCK, E. D., andGHOSH, S. K. Activation and leachingof chalcopyrite concentrates withdilute sulfuric acid. InternationalSymposium on Hydrometallurgy, Chi-cago, 1973. pp. 403-416.

15. BJORLING, G., and LESIDRENSKI, 1. P.Activation of sulphide minerals forhydrometallurgical methods. EighthInternational Congress for MineralDressing, Leningrad, 1968.

16. SUBRAMANIAN,K. N., and KANDUTH,H. Activation and leaching of chal-copyrite concentrate. Can. Min.Metall. Bull., Jun. 1973. pp. 88-91.

17. PRATER, J. P., et al. J. Metals, N.Y.,Dec. 1970. p. 23.SPEDDEN, H. R., et al. Metall. Trans.,Nov. 1971. p. 3115.

18. CONDER, H. Engng Min. J., vo!. 116.p. 943. 1923. V.S. Pat. 1 567916.

19. GRIFFITH, W. A., DAY, H. E.,JORDAN, T. S., and NYMAN, V. C.Development of the roast-Ieach-elec-trowin process for Lakeshore. NewYork, AIME, TMS Paper SelectionPaper no. A 73-64.

20. British Pat. 1131716. 1968. V.S.Pat. 349764. 1970.

21. ULRICH, H. J., and MAcDoNALD,R. D. V.S. Pat. 3523787. 1970.

22. KURUSHIMA, H., and TsuNoDA, S.Trans. AIME, vo!. 203. 1955. p. 634.

23. HAvER, F. P., et al. Min. Engng,N.Y., Jun. 1972. p. 52.

24. Work done at the National Institutefor Metallurgy.

25. U.S. BuREAu OF MINES. Copper re-covery from chalcopyrite by a roast-leach procedure. Washington, theBureau, Technical Progress Report 67.Mar. 1973.

26. ALLEN, E. S., and KRUESI, P. R.Cymet electrometallurgical processesfor treating base metal sulfide con-centrates. Joint Meeting MMIJ-AIME, Tokyo, 24th-27th May, 1972.Print No. T. IV b 4'.

27. KRUESI, P. R., ALLEN, E. S., andLAKE, J. L. Cymet Process-hydro-metallurgical conversion of base-metal sulphides to pure metals.Can. Min. Metall. Bull., vo!. 66,no. 734. Jun. 1973.

28. TUNLEY, T. H., et al. The leaching ofcopper and sulphide concentrations bychloride solutions. Johannesburg, Nat-ional Institute for Metallurgy. To bepublished.

29. HAVER, F. P., and WONG, M. M.U.S. Bur. Mines, RI 7474. 20 pp.

30. BALBERYSZSKI,T., et al. Electrowin-ning of copper at high current densi-ties. Proc. Australas. Inst. Min.Metall., no. 224. Dec. 1972. pp. 11-23.

31. Ibid., pp. 25-33.32. LIEKENS, H. et al. High current den-

sity electrowinning by periodic cellshort circuiting. Wld Min., Apr. 1973.pp. 40-43.

33. ETTEL, V. A., et al. Electrowinning ofcopper at high current densities.Paper presented at 1O2nd AnnualMeeting of AIME, Chicago, Feb. 1973.

34. HAINES, A. K., et al. The recovery ofzinc from pickle liquors by ion ex-change. J. S. Afr. Inst. Min. Metall.,Nov. 1973 (this issue).

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