1. J. Chem. Tech. Biotechnol. 1985,36, 197-211
Thermodynamic Studies on Sulphate Roasting for
Zinc Electrowinning from Carbonate Ore
Loutfy H. Madkour
Chemistry Department, Faculy of Science, Tanta University, Tanta, Egypt
(Manuscript received I March 1985 and accepted 9 October 1985)
The bulk of the work consists of a theoretical study of the possibility of submitting
Umm-Gheig carbonate ore to sulphate roasting. The use of the admixture with pyrites
is to enable a carbonate ore to be treated in a similar way to a sulphide ore, and by
doing so, to produce a roasted product capable of being treated by orthodox zinc
electrowinning methods using sulphate solutions. Thermodynamic studies have been
made to find the optimum conditions for sulphate roasting, in either normal air or
enriched 36% oxygen air. The results obtained from the experimental work at
different roasting temperatures in a tube furnace indicated that a maximum dissolu-
tion of 91.2% Zn with a 17.9% Fe could be obtained at a roasting temperature of
650°C for 4 h, followed by leaching in 4% H2S04(by vol.) at 60°C. The results of the
electron microscopic investigation confirmed by metal value data given in the ASTM
cards coincide well with results given by chemical analysis.
Keywords: Carbonate ore; sulphate roasting; zinc extraction.
1. Introduction
The polymetal mineralisation of the Red Sea Western coast has been known since the time of the
Pharoahs. Numerous investigators have studied The polymetal deposits of the Red Sea ore belt
(a zone extending NW-SE for a distance of 130 km) represent a complex morphogenetic type of
mineralisation.6 The chief minerals are hydrozincite, zinc blend, smithsonite and cerussite, while
silica and carbonates constitute the bulk of the gangue. The minerals present in this complex ore are
often soclosely intergrown that it iseither difficult to obtain suitable high-grade concentratesat high
recoveries’ by physical methods, or the recovery of metals in the respective concentrates is poor.
Hydrometallurgical methods based on leaching and precipitation rather than smelting played an
important role in meeting the requirements for the treatment of complex8ores. Kellog and others”’
discussed the thermochemistry of complex ore roasting and showed, with theoretical calculations,
that using a fluo-solid roaster, it is possible to control the calcine composition by controlling the
temperature and air-solid ratio. Surnikov and YurenkoI2roasted the intermediate products obtained
from the Berezovka plant at 800°Cin a laboratory fluidisedbed roaster with 150-200% more air than
was theoretically required, and the calcine was leached using H2S04acid at various pH values.
2. Experimental
Mineralised horizon ore (500 kg) was finely powdered to 100% minus 1.0 mm and dried before
roasting or sulphate roasting. The ore was subjected to mineralogical, chemical, spectral,X-ray and
differential thermal analyses.I3
A series of roasting experiments were carried out in a tube furnace at temperatures ranging from
40&9OO0C; the optimum time for the roasting process was found to be 4 h. All chemicals used
sulphuric acid, nitric acid, sodium hydroxide, ammonium hydroxide; otherswere of analytical grade
14 197
3. S.lpb.tcnldllgdurboarttorc 199
and were used without further purification. The cell design, the electrolysis system and general
experimental procedure for electrolysis have been described elsewhere.l3
3. Results and discussion
The mined ore was analysed as: zinc 30.70%; lead 7.99%; iron 5.05%; sulphur 1.14%; silica
6.38%.13Carbonates constitute the bulk of the gangue, whereas galena is the main sulphide
encountered in the Umm-Gheig mine at 15 metres. The X-ray diffraction chart and the powder data
of galena are shown in Figure 1and Table 1, respectively.
Table 1. X-ray powder data of galena from Umm-Gheig
Umm-Gheig GdenaI7 aCerussite17
d(A) Ilk d(A) I/la d(A) Ilk
4.44
4.27
3.60
3.51
3.44
3.07
2.98
2.60
2.52
2.49
2.10
1.933
1.n59
1.852
1.794
1.716
2
2
14
6
54 3.44
2
100 2.98
2
2
4
37 2.10
3
2
1
25
9
4.427
4.255
3.593
3.498
9
10
3.074
2.893
2.644
2.599
2.522
2.487
2.213
2.129
8
2.081
2.009
1.981
1.933
I 359
I ,847
1.796
1.750
17
7
100
43
24
2
2
11
20
32
7
2
27
11
9
19
21
4
2
n
“Data for galena after Berry and Thompson (1962) and for
cerussite after ASTM cards (card no. 5-0417).
3.1. Theoretical considerations
Important reactions that take place when a sulphide ore (MS) is roasted can be represented by the
followingequations:
MS+3/2 Oz%MO+SO, (1)
so2+1/2 02es03 (2)
MO+SO,%MSO, (3)
Reaction (1) is strongly exothermic and for all practical purposes during roasting, the equilibrium
shiftsto the right, with the formation of metallic oxide (MO) and SOz.Normally the heat evolved in
this reaction is enough to sustain the necessary thermal requirements of the roaster. The higher the
temperature,the faster the reaction, and the conditions that are available in afluo-solidroaster, such
as thorough mixing of the gas phase with the solids, proves an added advantage.
Reaction (2) isof far more importance for sulphate roasting, sincethe partial pressure of SO3in the
furnace atmosphere, whether higher or lower than the equilibrium partial pressure, decides the
4. 200 L.M.dLour
presence or absence of sulphates in the calcine. In an oxidisingatmosphere and at lower tempera-
tures, more SO, is formed. At higher temperatures SO2is more stable; over 700"C, especially in
presence of metallic oxides, the reaction rate is higher and more SO3will decompose to give SO2.
Nevertheless someSO3will alwaysbe present and the roaster gasescontain almost equalproportions
of SO2and SO3.The relation of equilibrium constant to temperature for the reaction is given by the
empirical formula represented by WagnerI4as:
(4)
5665.5
-log K=8.8557---1.21572 logtoT
T
The values obtained for K for different temperatures have been utilised in the calculations.
type:
The formation of metallic sulphates depends on the equilibrium constant for the reactions, of the
MO(s)+ SO&) MSO,(s) (3)
Since MO and MS04 are solids, their activities can be taken as unity and thus the values of Kp
dependson the partial pressures of SO,. If the SO, partial pressure in the furnace atmosphere ismore
than the equilibrium pressures of SO, for reaction (3), then more of the oxides formed in the reactor
according to reaction (1) would react to form the sulphates according to reaction (3).
~ 3 .I .I. Thermodynamictreatmentfor equilibriumroaster gas compositions
Knowing the chemical analysisof the ore being investigated, it is possible to theoretically study the
effectof: (1) the varying proportions of 90,95,100,110,120and 135 rnol of air per mol of Zn to ore in
the feed; (2) The enrichment of 80,90,95,100,110,120and 135 rnol of air per rnol zinccontent with
36%oxygen; (3) The roasting temperature from 800 K, 900 K, 1000 K, 1100 K and 1200 K on the
roaster gas composition; and thus arrive at the conditions for selective sulphation.
Table 2 analyses the Umm-Gheig ore in rnol percentages after adding 30% FeS2in the form of
natural pyrite. The last column expresses the various elements present as mole per mole of zinc.
Theoretical requirement of oxygen to convert all the elements into oxide from a quantity of ore
containing 1 mol of Zn can be calculated:
1mol of zinc would require Y2 mol of oxygen to form ZnO
1 Zn+0.5 O,=ZnO
10.0932Fe+0.75x10.0915 02=5.0466Fe203
10.0915S+10.0915 O2=10.0915 SO2
Table 2. The elemental composition of Umm-Gheigore with admixturc of 30%
pyrites. envisaged as the roaster feed
Weight Mol
Mol/mol Zn
_ _ ~ _ _ _ _ ~Component ("/.) ("/.)
ZnS 3.45 0.0531 1.O(MM)
Zn 21.49 0.3306 6.2290
Pb 5.59 0.0270 0.5088
coj~ 21.03 0.3505 6.6036
so:- 1.ox 0.0112 0.2120
S 17.14 0.5356 10.0915
Fe 17.53 0.3130 5.8978
S O , 4.61 0.1646 3.1019
AI,O, 0.58 0.0215 0.4047
MgO 1.81 0.0754 1.4209
CaO 3.39 0.0847 1 ,5967
Moisture 0.68 0.0378 0.7117
Total 98.38
5. Thus, the total stoichiometric requirement of oxygen is 18.1601 mol, which could be obtained from
roughly 87.73 mol of air.
3.1.2. Effect of proportion of air to ore in thefeed
From the stoichiometry of the various reactions shown above, it is possible to arrive at a material
balance for the various gases in the roaster once the proportion of air to ore feed isknown. Thus, for
a feed ratio of 135 mol of air per rnol of zinc:
Mol of Ozavailable in 135 mol air
135x0.207 = 27.945 rnol
Mol of O2reacted = 18.1601mol
Mol of free O2 = 9.7849 rnol
Mol of SOzformed = 10.0915 mol
Mol of N2in air =107.055 mol
No. of mol of moisture = 1.0153 rnol
Total no. of mol after
reaction =127.9467 mol
Now taking into consideration the equilibrium:
so,+fioz so3 (2)
If x is the moles of SO3formed, then x moles of SOzwould have reacted with x/2 mol of oxygen.
Now the number of mol of various gases would be:
O2 =(9.7849-x/2) mol
SOz =(10.0915-x) mol
SO3 =xmol
Total =(127.9467-x/2) mol
Hence the partial pressures of the various gases would be:
(9.7849-~ / 2 )
(127.9467-x/2)
(10.0915-X)
(127.946742)
psoz=
X
pS03=
(127.9467-~/2)
The equilibrium constant is given by:
(5)
~/(127.9467-~/2) __
K=
[{(10.0915-x)/( 127.9467-~/2)}{(9.7849-~/2)/( 127.9467-~/2)}"]
Thevalue of K for any particular temperaturecan be obtained from the Wagner's empirical formula:
5665.5
T
-log K=8.8557---1.21572 10,loT (4)
Substituting this value of Kin equation (5) we can get the value of x for any particular temperature,
and the values of partial pressure of gases or their molar percentage in the roaster gases can then be
calculated. The values thusobtained, for quantities of airvaryingbetween 75and 135 moles per mole
of zinc in the feed, have been plotted in Figure 2 for roasting temperatures of 800 K, 900 K, 1100 K
and 1200 K. The molar percentage of SOzfallsby changing the feed ratio of airto ore from75 to 135.
The changes in SO3percentage is not much affected; the increased amount of oxygen available
6. I.. Mlldkour
c.14
( a )
12
8
6
4
2 ? /de
14
c.
14-
- so2
70 80 90 100 110 120 130 140
- ' 0
70 80 90 100 110 120 130 140
Molof air/mol of Zn
70'80 90 100 110 120 130 140
Mol of air/ml of Zn
Figure 2. Equilibrium gas composition for Umm-Gheig ore roasting at (a) K 0 0 K. (b) 900K, (c) loo0 K.(d) I IIX) K.(e)
1200 K. 0.normal air;0.enrichedair (oxygen 36%).
meansmoreSO2isconverted intoSO,, thuscompensatingforanysolutioneffectonSO3percentages
due to the increasedvolume. Inthe sameplot calculatedvaluesfor the equilibriummole percentages
of various gases are recorded, using enriched air containing36% oxygen. The advantage of oxygen
enrichment is that a higher SO2content in the roaster gas can be achieved with a smaller volume of
air.The use of oxygen enrichment,particularlywhere the sulphidecontentis low,may also result in
the autogenous roasting of the ore.
Figure 3 indicatesthe calculatedvaluesof the equilibriumgas compositionsfor roasting between
600 K and 1200 K Umm-Gheigore at 1 atm, with a ratio of 90.95, 100, 110, 120 and 135 mol of
air/mole of zinc. The theoreticalequilibriumgascompositionobtainedon usi. I 80,90,95,100,110,
7. 9-
7-
5 -
12
=so2/
8
-
Q
a 2
-g 600 700 800 900 lo00 1100 1200
Q
a 2
-g 600 700 800 900 lo00 1100 1200
10
4
600It2
16
12
,4
2
600 700 800 900 1000 1100 1200
Temperature (K)
600 700 860 900 1000 1100 1200
lo? (d)
4
i2
600
x700 800 900 1000 1100 1200
Temperature (K)
Apre 3. Equilibriumgascompositionon roastingof Umm-Gheigoreat 1 atm pressure (a) 90,(b) 95, (c) 100. (d) 110, (e) 120
and (f) 135 mol of air per mole of Zn.0,normal air.
120and 135 molesof air per mol zinccontent, with oxygen enriched air, isindicated in Figure4. It is
seen from these figuresthat SO3content ishigher at lowertemperatures, while SO2content ishigher
at higher temperatures. The values indicated in these figures are, however, valid only at higher
temperatures. If the temperature were lowered, a stage would be reached when the solid oxides
present in thecalcinewould start absorbingSO3,formingthe varioussulphates.The temperatures at
which such reactions would start can be determined by plotting the variation of the decomposition
pressures of the various sulphates, with temperatures in the above Figures. Table 3 gives the
decompositionpressuresof the variouspossiblesulphates.Thevalueswerecalculatedfrom the logK
values for the various reactions as given by Kellogg."
Thedecompositionpressuresforthe two zincsulphates,normal and basic, and the ferricsulphate,
have been plotted by dotted linesas a functionof temperature in Figure5; the pointsof intersection
8. 2Q4 L.Madkour
4c 0 ,a;
13
600
Temperoture (K)
4
;i2
600 700 800 900 1000 I100 I200
Temperoture (K1
Figure4. Equilibriumgascompositionon roastingof Umm-Gheigore at 1 atmwithoxygenenrichedair(oxygen 36%).80,Yo,
95, 100,110. 120 and 135 moles of air per mole of Zn content. Assumingno sulphate formed.0,enrichedair (oxygen 36%).
of these lines with the roaster gas SO3composition line, represent the temperature up to which the
various sulphates indicated are stable in the roaster atmosphere. Decomposition pressures for the
various lead sulphates are much lower than for the other sulphates, and these have not been plotted.
It is observed from Figure 5, that at 937 K the ferric sulphate starts decomposing to form ferric
oxides, while all other sulphates are,quite stable at this temperature. Normal zinc sulphate starts
11. decomposingat 12% K to itsbasicsulphate, which isstableup to 1322 K, when ZnOstarts forming
(Figure 5(a)). Normal zinc sulphate decomposes into its basic sulphate at 1370 K and the basic
sulphateinto zinc oxide at 1433 K (Figure5(b)). Thus, between 900 K and 1200 K the decomposi-
tion of ferric sulphate takes place, while the zinc and lead sulphatesremain stable.
3.2. Results of the preliminary experiments
3.2.1. Roasting zinc ore
The principle reaction of ore roasting is:
2ZnS+30,- 2Zn0+2S02+223.6 kcal
ZnC03+ ZnO+CO,
For practicalpurposes,it may be assumedthat zincore ignitesat anywherebetween400-900"C. The
rate of combustion increaseswith increasingtemperature and decreasesas more sulphur is burned
out, because the oxide film which forms on the surface of each grain shuts out oxygen. The heat
balance of the roasting operation is made up of the heat input from the combustionof the sulphides
and the heat losses to the surroundings. As the rate of combustion is reduced, heat input per unit
time is alsoreduced and at a certainpoint it becomesequalto the heat loss. It isat thispoint that the
spontaneous burning of the sulphides ceases. Too high a roasting temperature may cause the
particlesto sinter or fuse, which would hamper the inflowof air to the sulphidescausingthe rate of
combustion of the sulphur to drop rapidly.
In roasting, some of the zinc sulphideis oxidisedto zinc sulphate,which may be expressed by the
followingequations:
2S0,+0,=2S03+45 .2 kcal
ZnO+SO3=ZnSOI+55.6 kcal
The impuritiescontained in the Umm-Gheigore are oxidised in roasting to form FeZ03,CuO and
CdO. The acid oxides SO3,As20S,Sb2OS,Fe203, SiO,, etc., react with the basic oxides and
carbonates, i.e. CdO, FeO, CuO, PbO, CaC03and MgCO3, to form zinc sulphates, aresenates,
antimonates, ferritesls or silicates, respectively. Not all of the many possible reactions reach
completion here, becauseeitherthe reactingmaterialsare not present in stoichiometricproportions,
or contact between them is upset, or the rate of interaction is too low. The most detrimental
secondaryreactionsin roastingare those producingferritesof zincandcadmiumandsilicatesof lead
andzinc.Thelatter, reactingwith the sulphuricacid in the subsequent leaching,formcolloidalsilicic
acid which hampers filtration and settling. Zinc ferrite, for its part, reacts with sulphuricacid, but
slowly, and the zinc fixed in it does not readily pass into solution. At low temperatures, the rate of
zinc ferrite formation is insignificant,but it rapidly increasesatabove 650°C.
Thesuspensionroastingprocesshasitsoriginin theobservationson the behaviourof the orewhen
fallingfrom hearth to hearth in aconventionalroaster. In falling,the oreparticlescomeintocontact
with oxygen-bearinggasps and burn quickly. The rate of combustionis greater than the rate of heat
TaMe4. Percentagedissolutionatdifferentroastingtemperaturesfollowedbyacidandpurewaterleachingprocesses
4% H2S0, (by vol) at 60°C) Pure H 2 0at 60'C
Experiment Temperature Zn2+ Fe'+ SOj- Zn2+ Fe3+ Sot-
number ("C) (%) (%) (%) (%) (%I (%)
1 400 46.8 35.8 6.9 40.1 30.7 5.2
2 500 70.4 29.7 8.7 61.8 22.3 6.1
3 600 87.5 24.5 10.3 79.2 18.5 7.0
4 650 91.2 17.9 13.2 83.7 11.6 7.6
76.5 5.7 -5 700 85.7 10.5 -
70.2 3.8 -6 800 79.6 7.8 -
7 900 67.8 4.6 - 58.3 3.8 -
Roasting time, 4 h; leachingprocesses. 1 h.
12. 208 L. Msdkour
Table 5. Percentage dissolution of metals at 650°C after acid and pure water leaching processes
4% H,SO, (by vol) at 60°C) Pure H,O at 60°C
Experiment Roasting Znz+ Fe3+ S q - Znz+ Felt SOi-
number time (%) (%) (%) (%) (%) (%)
1 I h 80.9 10.5 9.7 75.3 8.1 6.3
2 3 h 87.8 13.9 11.1 79.2 9.4 7.0
3 4 h 91.2 17.9 13.2 83.7 11.6 7.6
4 6 h 85.8 12.0 10.8 78.9 10.5 6.7
transfer tothe surroundings, and the temperatureof combustion risesappreciably. As a result, dead-
roasting becomes possiblewithout auxiliary heating. Furthermore, a smallerexcessof air isrequired
for suspension combustion, as the oxygen is utilised to a fuller extent, and the SOzcontent of the
roasting gases increases.
A series of roasting experiments were carried out in a tube furnace at temperatures ranging from
4OCk900"C;the time for roasting was increased successivelyfrom 1-6 h, as given in Tables 4 and 5.
The Umm-Gheig roasted product was then subjected to leaching processes.
3.2.2. Batch leaching of roasted zinc ore
In each roasting run, 20 gof the ore, divided equally into two boats, were used for leaching in either
4% sulphuric acid (byvol.) at 60°Cor pure water at 6O"C,after the roasting process. The slurry was
filtered and the leached liquor in both cases was analysed for zinc, iron and sulphate. The process
usedconsistsin bringing the zinccontained in the ore intosolution aszinc sulphate afterconverting it
into the oxide, or directly into the sulphate by roasting or sulphate roasting.
ZnO+H2S04=ZnSO,+ HzO
Many of the impurities can be reduced or eliminated by neutralising the zinc sulphate solution with
zincoxide, with the formation and precipitation of ferrichydroxide. This method iscommonly called
'iron purification' and is usually carried out simultaneously with leaching. Any ferrous iron present is
first oxidised to the ferric state by hydrolysing the ferric sulphate.
2FeSO4+MnOZ+2HzSO4=Fez(S04),+MnS04+2H20
Fez(SO,),+ 2Hz0=2Fe(OH) SO4+H2S04
The solution should be neutral towards the end of the leaching operation if the iron is to be
withdrawn successfully.Tomeet these conflictingrequirements, the leachingoperationiscarriedout
in two stages (double leaching). First, roasted Umm-Gheig ore istreated with a slightlyacid solution
of ZnSO, containing 100 g dm-, Znand 2 g dm-) HzS04.The acid present will not leach out all the
zinc, but only some of it will pass into the solution which will be neutral and therefore clean of iron
(neutral leach). The insoluble residue of the neutral stage stillcarries a lot of zinc, and it is re-treated
by depleted electrolyte containing 100 g dm-) H2S04 (acid leach). Towards the end of the second
stage, the concentration of H2S04in the solution drops to -3 g dm-3 and it is used for neutral
leaching.
At the beginning of leaching, the solid to liquid ratio is about 1:lO by weight. No auxiliary heating
is required, as the temperature of the pulp is upwards of about 50"C,due to the heat from the added
calcine, exothermic reactions and the heat of hydration. The pulp remains in the neutral leach for
about 1 h.
Towards the end of the neutral leach, the ferric sulphate in the mother liquor ishydrolysed to form
insoluble basicsalts. The underflow, whichisthe insoluble product of the neutral leach containing by
weight 15-20% solidsand 80-85% of the neutral solution, ispartially filtered and the residue isfed to
the acid leach step.
The rate of leaching depends on the concentration of HzS04.As it is higher in the acid than in the
neutral leach, the bulk of the zinc passes into solution during the second stage. The other factors
affecting the rate of leaching are temperature, grain size of the roasted ore, and agitation.
13. The rate of leaching increases as the temperature rises, due to an increase in the rate of diffusion
and the rate of chemical reactions between the HzS04and thesolidzinccompounds. The grain sizeof
the ore affects the rate of leaching above all because the coarse and fine particles differ in chemical
composition. The coarse particles are mainly sintered zinc sulphides, ferrites and silicates which
react slowlywith H2S04.Furthermore, the zinccontained in the larger particles passes into solution
more slowly than from the fine particles. As the grain size decreases, the surface area of solids per
unit weight increases, and the rate of solution is directly proportional to the surface area of the
particles. Thus coarse-grained material should preferably be reground prior to leaching.
The agitation of the pulp, consisting of solid particles and solvent, speeds up diffusion. The solid
particles should be always held in suspension for better contact between their surface and the
solvent.
At the end of the leaching processes the solid to liquid ratio was increased to 1:20, due to the
dissolution of some zinc.
The acid leach step destroys the zinc silicates to form colloidal silicic acid:
Zn 0* SO2+H2S04+(n-1) H20=ZnS04+Si02.nHzO
The acid leach residue carries -0.1% of the zinc in the original ore and all of the lead.
The percentage of leaching after different temperature roasting for 4 h is shown in Table 4. It is
observed that the percentage of zincdissolvedincreasesfrom 46.8%at 400°Cto91.2%at 650"C,and
thereafter decreases. The iron dissolved during leaching, however, decreases continuously from
W90O"C.
The percentage dissolution of these metals at different temperatures in case of leaching in 4%
HzS04(by vol.) is higher than in pure water (Table 4). Roasting time also has an effect on the
percentage dissolution at a specifictemperature, as given in Table 5.
For complete conversion of lead and zinc to their respective normal sulphates, the calcine should
theoretically contain 17.14%sulphur. The sulphate, sulphur in the calcine, increases from 6.9% at
400°Cto 13.2%at 650°Cfor the same retention time, indicating better sulphateconversion at higher
temperature.
Thus,from the resultsobtained,the optimum condition forcontrolled roastingof Umm-Gheigore
is at about 650°Cfor 4 h followed by leaching in 4%sulphuric acid (by vol.) at 60°Cfor 1 h (Tables 4
and 5).
3.2.3. Electrolytic production of zinc
The amount of zinc and impurities which pass into solution depends on the composition of the
starting mineral; its granulation, iron content, temperature and length of roasting (Tables4 and 5),
but above all, on the free acidcontent of the lixiviatingsolution. The yieldof extracted zincincreases
with the concentration of free acid in the solution used for treating the roasted ore, but the quality of
impurities dissolved also increases.Ih
The factors affecting current efficiencyare the opposite of those governing applied voltage (they
call for increased current density, reduced temperature, and reduced acidity of the electrolyte).
As electrolysis progresses, the zincconcentration in it is reduced, its acidity increases, and current
efficiencydecreases, making the complete recovery of zinc from the electrolyte uneconomical. The
usual practice isto withdraw the electrolyte from the cell after about half the zinchas been recovered
and the equivalent amountof free H2S04has been regenerated. Thedepletedelectrolyte is then used
to leach roasted ore. The least energy consumption can be obtained when an optimum balance
between all the factorsinvolved isstruck. In the electrolysis of zincsulphate roasted ore solution, the
suitable current density was40-60 mA Cm-2at 35°C;the yield of zincextracted varies between 80%
and 93%. It is not possible to extract all the zinc present in the original ore both because a certain
proportion remains unattacked. particularly if the iron content is relatively high (17.53%), and
because another part remains trapped in the solid residue of the lixiviation; this is gelatinous in
nature due to the presence of silicic acid and Fe20,.x HzO.
It is felt that zinc in Umm-Gheig ore can be recovered through three alternative pathways as
illustrated in the flowsheet of Figure 6, as follows: (1) Electrolysis of the sulphate acid leach solution
14. 210 L.Mdkour
Carbonate 30% Pyrites
Zn-Pb ore
650 'C, 4 h
4% H d G (by vol), 6OoC
ISulphate roosting
Acid-leach residue Leach solution
PbSO,, 4 0 2 ZnS04, Fe2(S04),
Arnmoniation Coustificotlon
lFe precipitation
I Arnmbiation Coustificotlon
1 lFe prec,ipitation /
0'k Na,Zn021eachtZn(NH,)&OH 121
telectrolysis
I,'Zinc cathodet product
Figure6. Sulphate roasting process flowsheet for treatment of Umm-Gheig carbonate Zn-Pb ore.
Zn S04.Fe2(S04)3directly, in the presence of 10 g d m 3 concentrated H2S0, at 50 mA Cm-?
cathodic current density and 35°C. The cathodic current efficiency is 65%, with 90% zinc recovery;
(2) Electrolysis of the acid leach solution after the caustification process is applied using 3 mol dm-'
sodium hydroxide in excess, whereas Fe203.xH 2 0was precipitated and removed by centrifuge. The
optimum cathodic current density is 100 mA Cm-2at 55°Cwith current efficiency of 9O%, and 97%
zincyield; (3) Elecrolysisof the acid leach solution after ammoniation technique using NH40H.The
cathodic current density is90 mA Cm-2at 30°Cwith 88%current efficiencyand 95% zinc recovery.
This flowsheet (as illustrated in Figure 6) has many advantages: Umm-Gheig ore is used directly,
without any application of concentration techniques, and large savings in reagents and chemical
processes should occur. So it is more economical in both the treatment of carbonate ores and
sulphide ores, and it is the technique generally adopted in industry.
References
1.
2.
3.
4.
5.
6.
7.
Beadnell. H. L. (1924) Report on the Geology of the Red Sea Coast Between Ooseir and Wadi Ranga. Prfrol. Bull.
Government Press. Cairo. 13.
El-Shazly. E. M.; Mansour. A. 0.;Afia. M. S.: Ghobrial. M. G. (1957)Miocene Lead and Zinc Deposits in Egypt. lrrrern.
Geol. Congress. XX Session, Section XII. Mexico, 119-134.
El-Shzaly, E. M. (1959) Controls of Tertiary Ore Deposition in Egypt. Chroniyue des Mines d'oufro Mer ef de [es
Rechrrches Miniers.
El-Akkad. S . ; Dardir. A. A. (1966)Geologyof the Red Sea Coast Between Ras Shagra and Mersa Alam with Short Note of
Results of Exploratory Work at Gebel El Rusas Lead-Zinc Deposit. Geological Survey of Egypr Cairo. 35
Sahet, A. H.; El Kholy. S.; Selim. E. T. (1973) Geochronology of Some Leud Mineralization in Egypf 7th Arah Science
Congress. Cairo.
Sabet. A. 11.;Tsogocv. V. B.;Bordonosov. V. P.; Beloshitsky. V.A,; Kuznetsov. D. N.; El-Hakim. H. A. (1980, 1976)
Annuls of the Geo/ogicu/Survey of Egypf According fo ConfrucfVI.
Eid, A. M.; Abd El-Rehim, M. M. (1963) Mefu//urgicu/Research on Zn-Pb Oxidised Ore of Umm-GheigDeposit. Eusfrrn
Deserf Geological Survey and Mineral Research Department. Egypt, 22.
15. SulptmteroPPtlngof carbonateore 21I
8. Viswanathan. P. V.; Yedavalli. B.V. S.; Srinivasan. S. R.: Bhatnagar, P. P. Symposium on Recent Development in Non-
Ferrous Metals Technology. Vol. II,Copper. 4 December. 1968.
9. Kellogg, H. H. (IW)A CriricalReviewofSulpharion Equilibria Trans. Metallurgical Society of A.I.M.E. December 230,
10. Smithson Jr. G. R.; Hanway Jr. J.E. (1962) Bench Scale Developmenr ofa Sulphurion Processfor Complex Sulphide Ores
Tram. Metallurgical Society of A.I.M.E. 224, 827.
11. Tdha. F.;Afifi. S. E.; Madkour. L. H. J. (1982) Tabbin Institute for Metallurgical Studies (T.I.M.S.) 47.
12. Snernikar, A. P.; Yurenko. V. M. (196.5) Laboratory Studies on the Hydrometallurgical Treatment of Cu-Pb-Zn
Intermediate Products. Russian Journal of Non-Ferrous Merals 11. 77.
13. Madkour, L. H. (19x5) Recommended Flowsheets for the Electrolytic Extraction of Lead and Zinc from Red Sea
Polymetal Ore. J. Chem. Tech. Bioferhnol. 35A, 106114.
14. Kellogg, H. H. (1964) A Crirical Review of Sulphurion Equilibrium Metallurgical Society of A.I.M.E. 230,1662-1661.
15. Sevryukov. N.; Kuzmin. B.;Chelishchcv. Y. (1969) General Metallurgy Mir Publishers, Moscow, 2nd edn.
16. Milaz.zo. G. (1963) Eleclrochemisfry. TheoreficalPrinciples and Pracfical Applicarions Elsevier Publishing Co..Amster-
dam. 467.
17. Berry. L. G.;Thompson. R. M. (1962) X-ray Powder Data for Ore Minerals, the Peacock Atlas. Geol. Soc. Amer. Mem.
85.
1622- 1634.
