Hydrometallurgy 97 (2009) 53–60

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Zinc recovery from the water-jacket furnace flue dusts by leaching andelectrowinning in a SEC-CCS cell

Tshikele Mukongo a, Kasonde Maweja b,⁎, Bilali wa Ngalu c, Ilunga Mutombo c, Kabamba Tshilombo b

a Columbus Stainless, Middleburg, Mpumalanga, Republic of South Africab Council for Scientific and Industrial Research, CSIR/Metals and Metals Processes, Meiring Naudé road, Pretoria 0001, Republic of South Africac Department of Metallurgical engineering, University of Lubumbashi, Democratic Republic of the Congo

⁎ Corresponding author. Tel.: +27 83 365 0952.E-mail address: mawejak@yahoo.fr (K. Maweja).

0304-386X/$ – see front matter © 2009 Elsevier B.V. Adoi:10.1016/j.hydromet.2009.01.001

a b s t r a c t

a r t i c l e i n f o

Article history:

Zinc containing flue dusts g Received 1 April 2008Received in revised form 3 January 2009Accepted 11 January 2009Available online 22 January 2009

Keywords:Flue dustsZinc recoveryLeachingSelective purificationElectrowinning

enerated during a copper matte smelting process in a water-jacket furnace wasleached in a zinc electrowinning return solution at 55 °C. Ninety percent of themetal was dissolved after 2 h oftreatment. The leaching residues contained near 43 wt.% PbSO4.Iron was removed from leaching solutions by precipitation using a saturated lime solution at pH 3–4. Theprecipitate formed mainly consisted of gypsum CaSO4.2H2O. Iron and magnesium co-precipitated into achloro-hydroxide compound. No zinc compound had co-precipitated during iron removal. Copper andgermanium contents lower than 10 mg/l were achieved by solvent extraction using 10% LIX 64N andcementation. Cobalt contents lower than 1mg/l in the solutionwere achieved bycementation on zinc granulesfor 2 h at 72 to 82 °C in presence of antimony. Cadmium contents of solutions were reduced to 10 mg/l bycementation on zinc granules at room temperature.Electrons transfer regime was identified as a limiting step during zinc electrolysis in a symmetric electrolysiscurrent–continuous circulating system, SEC-CCS. Electrolysis current efficiency higher than 94% and 3.5 kWh/kg of specific energy consumptionwas achieved under 500–600 A/m2 at 35 to 40 °C in the presence of gelatine.

© 2009 Elsevier B.V. All rights reserved.

1. Introduction

The copper matte smelting inwater-jacket furnaces in Lubumbashiplants, south-east of the Democratic Republic of the Congo, yieldedthree separated phases consisting of a copper matte, a slag and thefume dusts. The dusts naturally collected all the volatile compounds atthe operating temperature of the furnaces, at about 1250 °C. They alsocontained particles resulting from the wear of the agglomeratedsulphide concentrates and other additives fed at the top at the mouthof the furnace and drawn out of the furnace by the ascending aircurrent. The mechanically generated dusts had particle sizes rangedfrom 10 to 500 μmwhereas the particle sizes of chemically generateddusts formed by the volatilisation and oxidation of low-meltingtemperature metals i.e. Zn, Pb, Cd, Ge etc. ranged from 0.5 to 2 μm(Usines de Lubumbashi, 1993). Chemical dusts formed in the water-jacket furnace contained 17 wt.% Zn and 18 wt.% Pb. Such high levelmetal containing fine dusts may serve as raw materials for zinc andlead productions, which would subsequently reduce their toxicitywhen stored as landfill in open areas exposed to air and rainfalls.Mango and Mashala studied the leaching of the water-jacket furnacesfume dusts to optimise the solubilisation of germanium (Mango,1993), and the simultaneous dissolution of zinc, copper and cadmium

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in sulphuric acid solution (Mashala, 1993). They have analysed theeffects of leaching parameters i.e. dust to liquid mass ratio, acidcontent of the lixiviant, temperature and leaching time on the processefficiency and contamination of the solution by iron, magnesium andlead. It resorted from their works that the optimum conditions are alixiviant containing 50–100 g/l H2SO4 (~0.5 to 1 M), a ratio 1 g:7 ml offume dusts mass to volume of lixiviant, and a temperature of 55 °C.Leaching of some furnace dusts of similar mineralogy was studiedby others.

Zeydabadi et al. (1997) leached selectively the valuable elementsof a blast furnace flue dust which principally consisted of iron, withsome zinc and other elements oxides dust by sulphuric acid at lowacid concentration and room temperature, giving about 80% recoveryof zinc. Previous works also demonstrated that it was possible tobeneficiate the dusts by leaching the non-magnetic fraction of EAFdusts with sulphuric acid at low acid concentration and roomtemperature, giving high recoveries of Zn (≈80%) (Cruells et al.,1992; Caravaca et al., 1994). Langová et al. and Havlík et al. leached EAFdusts at ambient pressure. Zinc extraction reached almost 100% andiron extraction exceeded 90% in 3 M H2SO4 at 80 °C and S/L ratio of 1/5 after 6 h. A good selectivity with regard to zinc altogether with therelatively high zinc extraction was achieved in 0.1–0.3 M H2SO4 at80 °C. In that range the zinc extraction was about 30% and the Zn/Feratio was about 9 (Langová et al., 2007; Havlík et al., 2006). Othersreported that about 81% of zinc could be removed from the coarse size

Table 1Chemical composition of the dusts from the water-jacket furnace.

Compound Zn Cu Pb Ge MgO SiO2 S CaO Cd Co FeO

wt.% 17 10 18 0.14 3.85 3.2 7 6.8 0.89 0.24 2

54 T. Mukongo et al. / Hydrometallurgy 97 (2009) 53–60

fraction (i.e., N38 μm) of the BOF sludge sample by sulphuric acidleaching at a pH of about 2. The corresponding amount of iron co-extracted by acid leaching was about 18% (Kelebek et al., 2004).

Diverse alkaline leaching processes of zinc from various EAF dustshave also been foreseen previously. It was suggested that optimumconditions would be: 95 °C, 1/7 solid/liquid ratio, 10 M NaOH and 2 hleaching time (Orhan, 2005). Under these conditions, 85% Zn and 90%Pb were recovered. Dutra et al. (2006) found that the highest zincrecovery from the EAF dust, containing about 12% of zinc, was about74%. This was achieved after 4 h of leaching in a temperature of 90 °Cand with a sodium hydroxide concentration of 6 M of the leachingagent.

Purification of the leaching liquor is a critical step of thehydrometallurgy of zinc as the electrolysis is rendered inefficient byeven low content of some impurities in the electrolyte. Elimination ofelectropositive ions (Cu2+, Fe2+, Cd2+, Co2+ and Ge2+) from theleaching solutions is imperative before electrolysis. Most hydrometal-lurgical zinc plants remove iron from leaching liquors before electro-winning by precipitating as goethite or hematite at a controlled pH andtemperature (Elgersma et al., 1993a,b). Gouvea and Morais (2007)optimized the parameters for the cadmium cementation on zinc as 2 hreaction time, temperature 50 °C, pH 1.5 and 20% excess metallic zinc. Itis well known that zinc electrolysis is made possible by the highoverpotential of hydrogen formation on the layer of zinc covering thealuminium cathode. To achieve this overpotential condition thetemperature of the electrolyte should not be high, and the electrolytemust bequasi pure. It iswell established that additions of inhibitors suchas colloids are effective in inhibiting the detrimental influences of theimpurities in Zn electrolysis, and the current density must be high,N400 A/m2 (Mureşan et al., 1996). The high electrolysis current densityhowever has a detrimental effect on the quality of the deposition as itleads to the formationoffine loose zincparticles.Diverse electrowinningtechniques are continuously being investigated from ammoniacal,alkaline, nitrate, chloride, sulphate, acid solution to improve the currentefficiency and the structure of the cathode deposit (Rerolle and Wiart,1995;Yoshida et al., 2004;GürmenandEmre, 2003;Huajunet al., 2008).

The effectiveness of additives in improving the current efficiency toabove 94%, the quality of the deposited zinc metal and in developpingpreferred crystal orientation was also established by Saba andElsherief (2000), Gomes and Pereira (2006), Recéndiz et al. (2007)and Fosnacht and O'Keefe (1983). These studies also demonstratedthat addition of gelatine, in the range of 50 mg/l, improved the qualityof electrodeposited zinc.

Table 2Composition of the lixiviant (return solution from the zinc electrowinning plant).

Element Zn2+ Cu2+ Cd2+ Fe2+ Fe3+ Co2+ Cl− Mn2+ Ni2+

Concentration [mg/l] 35000 0.13 0.72 25 25 0.57 17 3.3 0.01

2. Experimental procedures

Thewater-jacket furnace flue dusts whose chemical composition ispresented in Table 1, were leached in acid solution containing Fe2(SO4)3 as oxidant agent.

The sulphur content of these dusts is low enough to foresee ahydrometallurgy route as a possible process for the zinc extraction.The iron oxide andmagnesium oxide were also present in proportionslower than 2 wt.% and 3.9 wt.% that would lead to low acidconsumption and less contamination of the leaching solution by thecorresponding ions.

The lixiviant used in this work consisted of a return acid solutionfrom the zinc electrowinning cells at the Zinc Plants in Kolwezi (Southof the Democratic Republic of the Congo). The chemical composition

of the electrowinning return solution is given in Table 2. The acidcontent of the return solution averaged 100 g/l.

The reactor was filled with 700 ml of lixiviant (the zincelectrowinning return solution) and heated up to 55 °C. A mass of7 g Ferric sulphatewas then added and dissolved by vigorously stirringthe solution at 900min−1 previous to the introduction of the dusts intothe reactor. 100 g of dusts was then introduced into the reactor andleached at 55 °C for 2 h. The mixture of dusts and leaching liquor wascontinuously stirred at 60 min−1.

Afterward, the leaching residues were separated from the leachingsolution by decantation in a 1 l glass container, and filtered on papersunder atmospheric pressure.

Samples of the initial flue dusts dried for 24 h at 80 °C and those ofthe leaching residues were prepared for XRD analysis using a backloading preparation method. They were analysed with a PANalyticalX'Pert Pro powder diffractometer with X'Celerator detector andvariable divergence- and receiving slits with Fe filtered Co-Kαradiation. The phases were identified using X'Pert Highscore plussoftware. The Reference Intensity Ratio method (RIR) was used forsemi-quantitative analysis as not all crystal structures needed for theRietveld method were available. Amorphous phases, if present, werenot taken into account in the quantification.

The samples of the same dusts and leaching residues were alsoprepared for scanning electron microscopy (SEM) analysis by hotmounting in resin, grinding on successive grit papers and polishing on3 and on 1 µmdiamond pasts. Cross sections of the respective powderswere analysed in backscattered electron mode and energy dispersiveanalysis were performed.

Particle size analysis was carried out on a representative sample ofthe dust and of the leaching residues using Saturn Digisizer 5200 V1.12which is a laser particle size analyser.

The purification of the filtrate solution obtained after leaching wasconducted in four steps to reduce the contents of ions noble than zincin the electrolyte.

i. Iron was removed from the filtrate obtained after leaching byprecipitation using a saturated lime solution at pH 3–4. Theprecipitate formed at this stage was separated by decantation andfiltration, dried at 80 °C for 24 h, and analysed by XRD. The filtratewas taken to the next step of copper and germanium removal.ii. Ions Cu2+ and Ge2+ were removed from the filtrate by solventextraction using 10% volume LIX 64N in solution in ESCAID 100.This mixture constituted the extracting organic phase (OP). Theorganic phase was regenerated by stripping the extracted ions bythe aqueous solution containing 100 g/l H2SO4. The volume ratio of(organic phase OP)/(aqueous phase AP) used is 1:1. That meansthe filtrate obtained after precipitation (AP) was mixed to theidentical volume of the organic phase, and stirred for 5 min. Themixture was left for 20min of decantation and separation betweenAP and OP. The operation was repeated and the [Cu2+] and [Ge2+]of the aqueous phase were determined by atomic absorptionspectroscopy (AAS) after 6, 12, 18 and 21 extraction–strippingcycles. The final aqueous solution was submitted to a two stepcementation on zinc granules.iii. Cobalt was first removed from the electrolyte by cementationon zinc granules at temperatures comprising between 72 and82 °C, in the presence of antimony. Antimony was added to thesolution in three different concentrations, i.e. 15 mg/l, 20 mg/l and

Fig. 1. Schematic representation of the experimental procedure.

55T. Mukongo et al. / Hydrometallurgy 97 (2009) 53–60

25 mg/l. Three samples of solution were taken at 40 min intervalup to 120 min in each case to determine the remaining [Co2+] byatomic absorption AAS method. Part of the cadmium contained inthe solution was also removed during this step.

iv. Cadmium was finally removed from the electrolyte bycementation on zinc at room temperature. Samples of the solutionwere taken at 40 min interval during cementation, filtrated andtheir cadmium content determined by AAS method.

Fig. 2. Particle size distribution of the initial fume dusts and the residues obtained afterleaching for 2 h at 55 °C.

A schematic representation of the above procedure is shown inFig. 1.

The solution hence purified constituted the electrolyte for the zincelectrowinning.

The symmetric electrolysis current–continuous circulating system,SEC-CCS cell used for electrolysis is a 280 ml prism made of glass intowhich the electrolyte is introduced via a series of perforated channelsthat lie parallel to the base of the prism. A peristaltic pump maintainsa forced circulation by injecting and sucking up the electrolyte inparallel flow lines to the electrodes from the bottom of the cell. Hencethe electrolyte sweeps across the electrodes. The aluminium cathodeis placed in between two anodes in Pb-1% Ag. This arrangement allowsthe formation of symmetric electrolysis current lines in the electro-lyte. The distance between each anode and the cathode is 3 cm. Thecircuit of the fluid starts in a 2 l container heated up to the requiredtemperature by standing it in simmering water.

The aluminium cathode surfaces were readably polished mechani-cally and stripped in 65% HNO3, dried and weighed before theelectrolysis. Stripping in nitric solution helped writing off thepolishing marks from the surface of the cathodes.

The effects of three variables on the electrolysis process wereinvestigated. The temperature of the electrolyte was varied from 30 to50 °C, the electrolysis current density (J) was varied from 300 to 900 A/m2, and the electrolyte flow rate was varied from 60 to 120 l/h.

The electrolysis timewas considered as a constant 2 h in this study.One variable (electrolyte temperature, electrolysis current density,

electrolyte flow rate) was changed at a time when the other two werekept constant.

The cathode polarisation curves ηc= f(J) were obtained bysystematically imposing a constant current density, J, and measuringthe subsequent quasi static cathode overpotential ηc using a referencecalomel saturated electrode.

3. Results and discussion

3.1. Leaching

The analysis byX-raydiffractionof thedried dusts (startingmaterial)showed that they contained mainly lead sulphate, becherite Zn7Cu(OH)13((SiO(OH)3)0.7(O4H7)0.3SO4), diopside ferroan (Mg0.982Fe0.018)(Ca0.976Mg0.020Fe0.004)Si2O6, corrensite Mg(Si, Al)8O20(OH)10.2H2O andmuch lower proportion of cobaltian sphalerite Zn3CoS4. The chemistryof these dusts rich in sulphate of the low melting temperature metals(Pb and Zn) indicates they were chemically generated throughvolatilization and oxidation of lead and zinc sulphides contained inthe copper concentrates treated in the water-jacket furnace. Theircomposition revealed that the dusts have interacted with the rainfallwater and humidity during their long storage in landfill area.

It is also inferred from the particle size distribution in Fig. 2, thatthe proportion of the dust particles finer than 500 μm is higher than78 vol.%. Mechanically generated dust particles, coarser than 500 μm,are therefore present in proportion less than 20 vol.% in these fluedusts. This might be the reason of the observed lower sulphur contentof the dusts as sulphur in the dusts would be indicative of mechanical

Table 3Chemical composition of the solution after leaching for 2 h at 55 °C and filtration.

Element Zn2+ Cu2+ Cd2+ Co2+ Fe (total) Ge2+

[mg/l] 55,700 8890 1140 260 1650 100

56 T. Mukongo et al. / Hydrometallurgy 97 (2009) 53–60

erosion of the sintered sulphide concentrates treated in the water-jacket furnace. Indeed the mechanically generated dust particles areproduced by erosion by the air flowing through the sinteredchalcopyrite–chalcosine–bornite concentrate, containing less than5 wt.% ZnS and PbS, used in copper matte smelting in the waterjacket furnace.

Leaching these dusts using the return solution from the zincelectrowinning with the composition in Table 2, for 2 h at 55 °Cproduced the solution whose chemical composition is presented inTable 3.

The amount of zinc hence dissolved is equal to:m=(55.7−35) g/l×700 ml=14.49 g.

The total amount of zinc introduced in the leaching reactor isM=17%×100 g=17 g.

The zinc leaching efficiency in the above condition isηleaching = 14:49

17 = 90k:

The phases present in the washed and dried residues of lixiviationwere identified by diffraction of X-rays. They contained 43 wt.% PbSO4

and 32 wt.% KAl3Si3O10(OH)2. The high lead content makes theseresidues not appropriate for landfill. However, their low iron content(b3% Fe), and the absence of zinc and arsenic make them suitable to apyrometallurgical extraction of lead and the potential by-products richin KAl3Si3O10(OH)2 as raw materials in ceramics or cement industries.

No zinc containing phase was detected by X-ray diffraction of theleaching residues. This suggested that zinc was practically fullyrecovered from the water-jacket furnace flue dusts by leaching inthe above-mentioned conditions.

The phase distribution in the dust particle cross sections beforeleaching could not be distinguished in backscattered SEM. Thehomogeneity hence observed may be a confirmation of the chemicalorigin of the major proportion of these dusts particles formed by avolatilisation–oxidation and condensation mechanism during coppermatte smelting of the sintered complex sulphide concentrates.

The zinc peak distinctively observed at about 7.5 keV in the X-rayenergy dispersive spectrum of the initial dusts completely disap-peared after leaching as it was not present in the EDS spectrum of theresidues. This observation confirmed the early XRD results thatshowed no zinc compounds present above the detection limit in theleaching residues.

The SEM of the cross sections of the dusts particles showed a lot ofporosity (small dark spots) that might have contributed to the highzinc leaching efficiency from these dusts.

The particle size distributions of the dusts and the leachingresidues are compared in Fig. 2. It is inferred from this figure that thesizes of the coarser dust particles decreased more significantly uponleaching comparatively to the finer ones. The porosity of the dusts andthe stirring of the reacting mixture might also contribute to thedecrease of the particle size upon leaching by enhancing the breakingoff of the initial particles as the reaction proceeded. However, particlessmaller than 8 μm tend to agglomerate on to the bigger ones. This factmay explain the relative trends of the curves of particle sizes below8 μm. It is also possible that the particles became so fine below thedetection limit of the Digisizer upon leaching and were not accountedfor.

Fig. 3. Evolution of [Cu2+] and [Ge2+] in mg/l in the purified solution during solventextraction by 10 vol.% LIX 64N in solution in ESCAID 100.

3.2. Purification

Iron was removed from the leaching solution in a one-stepprecipitation by a saturated lime solution at room temperature, at

pH value comprised between 3 and 4, within the published limits ofpH 3 to 6 by Elgersma et al. (1993a) and Mureşan et al. (1996). Thereaction was completed quasi-instantaneously.

The XRD pattern revealed that the precipitate hence formed wasmainly constituted of gypsum CaSO4.2H2O and bassanite CaSO4.0.5-H2O, the last being formed by partial dehydration of gypsum duringdrying. The analysis also revealed that iron and magnesium containedin the solution were simultaneously removed by precipitation into achloro-hydroxide compound of the two metals (iowaite). However atthe selected pH range no zinc compound had co-precipitated duringiron removal.

It was observed that the separation of liquid from the precipitate bydecantation and filtration on paper was slow. This might be due to thehydration of gypsum, which is well known for its poor decantabilityand filterability.

The total iron content of the solution dropped from 1650 mg/l, inthe leaching liquor, to 21 mg/l upon precipitation in the abovementioned conditions. This iron content was too low to affect the zincelectrolysis efficiency; precipitation was therefore assumed complete.

The filtrate obtained from the iron precipitation was subsequentlysubjected to a solvent extraction treatment to remove copper andgermanium. The results of this process using LIX 64N in solution inESCAID 100 are presented Fig. 3.

The copper content of the solutionwas reduced to 500mg/l after 12solvent extraction cycles, whereas 15 cycles were needed to decreasethe germanium content to the same level.

However it was not necessary to completely remove copper fromthe solution as the presence of the more electropositive copper helpsin solubilising antimony powder added during the cementation ofcobalt on the zinc powder in the next purification stage as shown byMango (1993), Tripathy et al. (2003), Bøckman and Østvold (2000)and by James et al. (2000).

Cobalt and cadmium were removed from the solution in two-stepcementation on zinc pure powder. The first step of cementation wasconducted at temperatures comprised between 72 and 82 °C toremove cobalt and part of the cadmium. Bøckman and Østvold (2000)and James et al. (2000) observed that removal of Co2+ duringcementation on zinc, in the presence of antimony, is due to tworeactions. One reaction is the formation of a Co alloy or Co in solidsolutions. The second reaction is the formation of basic cobalt salts,and is a function of local high pH at the zinc surface and theconcentration of Zn2+. The high local pH close to the zinc surface iscreated by evolution of hydrogen. This hydrogen evolution stops whenthe core of metallic zinc in the cemented particle is consumed, andbasic cobalt salts in the particle may redissolve. Basic cobalt salts

Fig. 4. Evolution of the cobalt content [mg/l] and the residual antimony content [mg/l]in the purified solutions as functions of the cementation time and of the initialantimony powder mass added to the solution before cementation.

Table 4Residual cadmium content in the solution after cementation on zinc granules at roomtemperature.

Reaction time [min] 0 40 80

[Cd2+] mg/l 744 11 0.81

Table 5Composition of purified solution before electrolysis.

Element Zn2+ Fe (total) Ge2+ Co2+ Cd2+ Sb3+ H2SO4

[M] mg/l 91000 21 0.47 0.91 0.81 0.47 17000

57T. Mukongo et al. / Hydrometallurgy 97 (2009) 53–60

redissolvemore easily than Co alloys and is themajor reason for cobaltredissolution during the technical purification of the zinc sulphateelectrolyte before zinc electrolysis (Bøckman and Østvold, 2000).

The residual cobalt and antimony contents of the solution are givenin Fig. 4.

The cobalt content of the cemented solution decreased from 260 to1 mg/l after 120 min when 25 mg of antimony powder was added to1 l of the solution to be purified. The residual Sb3+ content was about0.5 mg/l in these conditions. It was noticeable that the [Cu2+] in thecemented solution simultaneously decreased to less than 10mg/l thatis lower enough to have no significant effect on the zinc electrolysisprocess. It is also seen from Fig. 4 that the kinetics of cobaltcementation is strongly dependent on the initial amount of antimonypowder added rather on the reaction time.

Fig. 5. Tafel's lines for the electrolyte heated up to 35 °C, 40 °C and 45 °C und

Part of the cadmium contained in the leaching solution was alsoeliminatedduring this step of cobalt cementation but its residual level inthe solutionwas still very high (744mg/l). This rendered necessary thesecond step of cementation at room temperature to decrease [Cd2+] tobelow12mg/l. Thiswasachievedbycementationof thefiltrate obtainedfrom the precedent step on zinc granules at room temperature. Noantimony was needed in this last cementation step. The evolution ofcadmium content during cementation at room temperature of thesolution is given inTable 4. Cadmiumcontents lower than11mg/l in thesolution were hence achieved after 40 min cementation at roomtemperature.

The chemical composition of the purified solution hence obtainedis shown in Table 5. This constituted the electrolyte used for zincextraction in the SEC-CCS cell in this study. It is inferred from thechemistry in Table 5 that the sulphuric acid content of the solutiondecreased to 17 g/l, mainly due to a neutralisation by the lime duringiron removal by precipitation at pH 3–4 and the precipitation ofhydrated calcium sulphates and hydroxysulphates. The hydrating ofthe calcium sulphates (gypsum and bassanite) and the complexchloro-hydroxide of iron and magnesium (iowaite) consumed asignificant proportion of the water present leading to the observedincrease in zinc content of the purified solution in Table 5, to 91 g/l.

3.3. Electrolysis

Drawing of cathode polarisation curves enables the determinationof working regime and the characterization of the kinetics of theelectrolysis phenomenon.

The polarisation curves at 35, 40 and 45 °C for three different flowrates, i.e. 60, 80 and 100 l/h are presented in Fig. 5.

er three different flow rates (60 l/h, 80 l/h, 100 l/h) in the SEC-CCS cell.

Table 6Slopes and J0 of the Tafel's lines for the electrolyte at 35, 40 and 45 °C.

Electrolyte temperature 35 °C 40 °C 45 °C

Slope 0.0656 0.0683 0.0845J0 A/m2 31 37 77

Fig. 7. Effect of the electrolyte flow rate on the overall cell potential UA/C, the cathodeoverpotential ηc, the specific energy consumption w and the % current efficiency Rc.

58 T. Mukongo et al. / Hydrometallurgy 97 (2009) 53–60

The absence of the diffusion plateaus in this figure suggests thatthe electrolysis takes place under a electrons transfer regime. This wasalso shown by the insensibility of the cathode overpotential tochanges of the electrolyte flow rate at the three temperatures.

The slopes of the Tafel's lines and the current densities J0corresponding to zero volt overpotential at the cathode for the threeelectrolyte temperatures are shown in Table 6.

It is inferred from Fig. 5 that increasing the temperature of theelectrolyte lowers the cathode overpotential ηc but increases the slopeof the straight line part of the polarisation curve, hence the increaserate of the cathode overpotential, as the current density J increases.The extrapolated current densities at zero volt overpotential J0 alsosuggest that increased electrolyte temperature increased the currentdensity limit for the transition from an ionic to an electronicconducting mode in the electrolyte. The limit is lower than 77 A/m2

in these cases, which is lower than the 300–900 A/m2 density rangeused in the experiments.

The effects of the electrolyte temperature on the overall electrolysispotential, the cathode overpotential, the specific energy consumptionandon the current efficiencywerefirst investigated at500A/m2,120 l/hand 100 mg/l gelatine. The results are presented in Fig. 6.

The electrolysis current efficiency dropped drastically at tempera-tures higher than 40 °C due to a decrease of the hydrogen evolutionoverpotential and an increase of the conductivity of the electrolyte thatleads to a higher current and higher power loss along the electric wiresout of the electrolysis cell. The slight decrease in the overall electrolysispotential UA/C and the cathode overpotential are dependent on theincreased conductivity of the electrolyte at elevated temperature.

The electrolysis current efficiency decreased with the electrolyteflow rate as shown in Fig. 7. This phenomenon finds its explanation inthe competition between antagonist forces acting on the particulatenear the cathode, i.e. the gravitational interaction and adhesion on thecathode, the drag force and mechanical erosion due to the hydraulicturbulence and the electrostatic force. Increasing the flow rate increasesthe drag force and the cathode erosion. This lead to a decreased amount

Fig. 6. Effect of the electrolyte temperature on the overall cell potentialUA/C, the cathodeoverpotential ηc, the specific energy consumption w and the current efficiency Rc.

of metal collected on the cathode after electrolysis. This loss is purely ofmechanical origin. Particulates ofmetal could be collected in the bottomof the cell and in the channels. Increasing the electrolysis time wouldlead to complete dissolution of these zinc particulates by the sulphuricacid contained in the electrolyte.

It is also observed from Fig. 7, that the electrolyte flow rate, hencethe agitation does not affect ηc and has negligible effect on the UA/C

and on the energy consumption, w. These observations corroboratewith the earlier notice that there was no diffusion plateaus in thecathode polarisation curves and the electrolysis kinetics was con-trolled by electrons transfer phenomenon.

It ressorted from this investigation that the optimum electrolyteflow rate in a SEC-CCS type cell is somwhere between 80 and 100 l/h.

Gelatine was added to the electrolyte to improve the structure andcompacity of the cathode metal by controlling the current density onthe cathode surface. Experiments conducted at 35 °C, 80 l/h and500 A/m2 have shown that introducing gelatine into the electrolyteimproved the electrolysis current efficiency. This is due to the fact thatgelatine favours the formation of dense and compact metal, hence lessmechanical losses of zinc particulates from the cathode. Results areillustrated in Fig. 8. The structure of the metal formed on the cathodewas dendritic for gelatine contents of 0 and 100 mg/l. Dendritesdissapeared at gelatine content of 200 mg/l and 400 mg/l. Gelatinehelps also leveling the cathode metal surface by sticking at the tips ofthe dendrites. This increases the local electric resistivity in front of thedendrite and favours the growth of metal on the slack areas of thecathode. Current efficiency higher than 94% could be achieved in thepresence of gelatine against 90% without gelatine.

The effect of the current density J on the overall cell potential, thecathode overpotential, the specific energy consumption and on thecurrent efficiency was investigated in the electrolyte maintained at35 °C, 200mg/l gelatine and100 l/hflow. The results are plotted in Fig. 9.

Increasing the electrolysis current density up to 600 A/dm2 resultsin increased current efficiency Rc. This observation may be explainedby the easy formation of stable nuclei due to a large number of ionsdischarged in the cathode. The hydrogen evolution overpotential onthe cathode also is high when the electrolysis current density is high,favouring the discharge of the zinc ions.

Above J=600 A/m2 the current efficiency dropped drastically dueto the Joule's effect and the conversion of the electric energy into heatin the electrolyte proportionally to the square of the electrolysis

Fig. 8. Effect of the gelatine on the overall cell potential UA/C, the cathode overpotentialηc, the specific energy consumption w and the % current efficiency Rc.

Fig. 9. Effect of current density on the overall cell potential UA/C, the cathodeoverpotential ηc, specific energy consumption w and the % current efficiency Rc.

59T. Mukongo et al. / Hydrometallurgy 97 (2009) 53–60

current. The Joule's effect in the electrolyte also leads to a temperaturerise that reduces the hydrogen evolution overpotential on the cathode.This effect enhances the undesired formation of hydrogen to thedetriment of the zinc reduction in the cathode.

The overall electrolysis potential UA/C increased when J increaseddue to the rise in electrode overpotentials and the resistivity of theelectrolyte. The specific energy consumption is therefore the con-sequence of the above mentioned phenomenon.

4. Conclusion

The major proportion of water-jacket furnace flue dusts formedduring copper matte smelting are chemically generated through avaporisation–oxidation–condensation cycle. Zinc and lead sulphidescontained in the chalcocite–chalcopyrite–bornite concentrate aresublimated at the furnace working temperature, 1250 °C. Theoxidation of these fumed sulphides forms a mixture of sulphatesand hydroxi-sulphates that solidify out of the furnace to form the dustparticles with finely dispersed phases. Indeed the main constituent ofthe dusts, zinc and lead are present as sulphates and complex hydroxi-sulphates formed during the matte smelting of a complex copper,cobalt, zinc sulphide concentrate. Further hydration of these com-pounds might have occurred upon interaction with air and rain waterin the landfill area. The scanning electron microscopy analysisrevealed that the phases present in the dusts were finely distributedbeyond the resolution limit of the SEM equipment. This observationsupports the thesis of chemically generated dusts.

These dusts are leachable in aqueous acid solution, i.e. the returnsolution from the zinc electrolysis containing100 g/lH2SO4, 35 g/l Zn2+,in presence of 10 g/l Fe2(SO4)3. In these conditions, 90% of the zinccontent of the dusts was dissolved after 2 h at 55 °C.

The leaching residues contained 43 wt.% PbSO4 making them notappropriate for landfill. However, their low iron content, and theabsence of zinc and arsenic make them suitable to a pyrometallurgicalextraction of lead.

Separation of the residues from the leaching solution by decanta-tion was fast and there was no need of using any flocculant.

Purification of the leaching solution was possible by selectiveelimination of iron by precipitation using a saturated lime solution atpH3–4. Iron andmagnesiumwere precipitated in a complex hydroxideMg4Fe(OH)10Cl(H2O), simultaneously with hydrated gypsum andbassanite. No zinc compound was detected, by XRD or EDX analysis,

in the precipitate hence formed. However, solid/liquid separation ofthe precipitate hence formed from the leaching liquor by decantationand filtration on papers was slow due to the hygroscopicity of thecalcium sulphates, i.e. gypsum and bassanite that formed fine massiveprecipitates. A significant amount of water contained in the leachingsolution was consumed in forming the hydrated calcium sulphates.This resulted in a decrease of the volume of solution and an increase ofthe zinc content in the solution from55 g/l after leaching to 91 g/l aftertreatment by the lime milk. Copper and germanium were simulta-neously removed from the electrolyte by solvent extraction using 10%LIX 64N in ESCAID 100. It was not necessary to completely removecopper fromthe electrolyte at this stage as thiswouldnecessitatemanyextraction–stripping cycles. It was rather observed that the coppercontent of the aqueous solution was efficiently reduced to less than10 mg/l in the stage of cementation for 2 h at 72 to 82 °C of cobalt onzinc granules in presence of antimony.

Finally cadmium levels were reduced to less than 10 mg/l beforeundertaking the zinc electrolysis by cementation on zinc granules atroom temperature for 40 min.

The electrolyte solution obtained after purification contained 91 g/l zinc and 17 g/l H2SO4.

The investigation into the electrolysis process of the purifiedsolution in a symmetric electrolysis current–continuous circulatingsystem, SEC-CCS, the cell revealed an electrons transfer regime as thelimiting step. This observationwas confirmed by the constant cathodeoverpotential and the negligible effect of the electrolyte flow rate onthe overall electrolysis cell potential.

Addition of gelatine to the electrolyte improves the quality of thecathode deposited metal and the electrolysis current efficiency.

Optimum conditions for the zinc electrolysis in a SEC-CCS type cellwere determined as:

– Current density 500 to 600 A/m2

– Electrolyte temperature comprised between 35 and 40 °C

– Electrolyte flow rate 80 to 100 l/h

– Gelatine level comprised between 100 and 200 mg/l.

Zinc electrolysis current efficiency higher than 94% and the specificenergy consumption about 3.5 kWh/kg were achieved in the aboveconditions.

60 T. Mukongo et al. / Hydrometallurgy 97 (2009) 53–60

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