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 so closely intergrown that it is either difficult to obtain suitable high-grade concentrates at 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 complex8 ores. 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 YurenkoI2 roasted the intermediate products obtained from the Berezovka plant at 800°C in a laboratory fluidised bed roaster with 150-200% more air than was theoretically required, and the calcine was leached using H2S04 acid 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; others were of analytical grade

14 197

4 4

e

?

U

* N

It

m N

8

N m

Q n

c m N e ,

Y , % m e .- c

m

h n

c!

W Q

m P

I

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%.13 Carbonates 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 1 and 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

1 1 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 following equations:

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 shifts to 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 a fluo-solid roaster, such as thorough mixing of the gas phase with the solids, proves an added advantage.

Reaction (2) is of far more importance for sulphate roasting, since the partial pressure of SO3 in the furnace atmosphere, whether higher or lower than the equilibrium partial pressure, decides the

200 L. M.dLour

presence or absence of sulphates in the calcine. In an oxidising atmosphere and at lower tempera- tures, more SO, is formed. At higher temperatures SO2 is more stable; over 700"C, especially in presence of metallic oxides, the reaction rate is higher and more SO3 will decompose to give SO2. Nevertheless some SO3 will always be present and the roaster gases contain almost equal proportions of SO2 and SO3. The relation of equilibrium constant to temperature for the reaction is given by the empirical formula represented by WagnerI4 as:

(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 depends on the partial pressures of SO,. If the SO, partial pressure in the furnace atmosphere is more 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 . Thermodynamic treatment for equilibrium roaster gas compositions Knowing the chemical analysis of the ore being investigated, it is possible to theoretically study the effect of: (1) the varying proportions of 90,95,100,110,120 and 135 rnol of air per mol of Zn to ore in the feed; (2) The enrichment of 80,90,95,100,110,120 and 135 rnol of air per rnol zinc content 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% FeS2 in 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:

1 mol of zinc would require Y2 mol of oxygen to form ZnO 1 Zn+0.5 O,=ZnO 10.0932 Fe+0.75x 10.0915 02=5.0466 Fe203 10.0915 S+10.0915 O2=10.0915 SO2

Table 2. The elemental composition of Umm-Gheig ore 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 0.01 12 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 0.71 17

Total 98.38

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 the feed 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 is known. Thus, for a feed ratio of 135 mol of air per rnol of zinc:

Mol of Oz available in 135 mol air

135 x0.207 = 27.945 rnol Mol of O2 reacted = 18.1601 mol Mol of free O2 = 9.7849 rnol Mol of SOz formed = 10.0915 mol Mol of N2 in 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 SO3 formed, then x moles of SOz would 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)

p s o z =

X p S 0 3 =

(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)}"]

The value of K for any particular temperature can 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 thus obtained, for quantities of air varying between 75 and 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 SOz falls by changing the feed ratio of air to ore from 75 to 135. The changes in SO3 percentage is not much affected; the increased amount of oxygen available

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 Mol of 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) 900 K , (c) loo0 K. (d) I I I X ) K. ( e )

1200 K. 0. normal air; 0. enriched air (oxygen 36%).

means more SO2 is converted into SO,, thus compensating for any solution effect on SO3 percentages due to the increased volume. In the same plot calculated valuesfor the equilibrium mole percentages of various gases are recorded, using enriched air containing 36% oxygen. The advantage of oxygen enrichment is that a higher SO2 content in the roaster gas can be achieved with a smaller volume of air. The use of oxygen enrichment, particularly where the sulphide content is low, may also result in the autogenous roasting of the ore.

Figure 3 indicates the calculated values of the equilibrium gas compositions for roasting between 600 K and 1200 K Umm-Gheig ore at 1 atm, with a ratio of 90.95, 100, 110, 120 and 135 mol of air/mole of zinc. The theoretical equilibrium gas composition obtained on usi. I 80,90,95,100,110,

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

600 It 2

16

12

, 4

2

600 700 800 900 1000 1100 1200 Temperature ( K )

600 700 860 900 1000 1100 1200

lo? (d)\

4 i 2

600 x 700 800 900 1000 1100 1200

Temperature ( K )

Apre 3. Equilibrium gas composition on roastingof Umm-Gheig ore at 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.

120 and 135 moles of air per mol zinc content, with oxygen enriched air, is indicated in Figure 4. It is seen from these figures that SO3 content is higher at lower temperatures, while SO2 content is higher 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 the calcine would start absorbing SO3, forming the various sulphates. 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 decomposition pressures of the various possible sulphates. The values were calculated from the log K values for the various reactions as given by Kellogg."

The decomposition pressures for the two zinc sulphates, normal and basic, and the ferric sulphate, have been plotted by dotted lines as a function of temperature in Figure 5; the points of intersection

2Q4 L. Madkour

4c 0 ,a;

13

600 Temperoture (K)

4 ;i 2

600 700 800 900 1000 I100 I200

Temperoture (K1 Figure 4. Equilibrium gas composition on roasting of Umm-Gheig ore at 1 atm with oxygen enriched air (oxygen 36%). 80, Yo,

95, 100,110. 120 and 135 moles of air per mole of Zn content. Assuming no sulphate formed. 0, enriched air (oxygen 36%).

of these lines with the roaster gas SO3 composition 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

Tabl

e 3. D

ecom

posi

tion p

ress

ure

of v

ario

us su

lpha

tes (

atm

)

3(Pb

S04.

PbO

) %

(PbS

O,.Z

PbO

) Te

mpe

ratu

re

V5 Fe

,(SO

,),

3ZnS

04=Z

n0 M

(ZnO

.ZZn

S0,)

2PbS

04=P

bS04

=2(

PbS0

4.Z

PbO

) H

(PbS

O,.P

bO)

=H(P

bS04

.4Pb

O)

PbS

04.4

Pb0

(K)

=M F

e,O

,+SO

, 2

ZnS

O,+

SO,

WZ

nO)+

SO,

PbO

+SO

, +s

o,

=M(P

bSO

A.4

PbO

) +

so,

=SPb

O+S

O,

-

-

-

-

-

-

-

-

700

2.5

53

~

10-8

0 -

-

1.83

22~ 10

-13

7.%

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10-14

800

1.11

97~

10-~

-

-

900

2.89

1x 10

-2

6.8x

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3.

547x

10-5

2

.92

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10

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3.49

9~10

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3.

112 x

10-

lo

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.33

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loo0

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631X

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3.24

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1.1

43

~

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7.9

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1O

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1100

2.

831

8.73

8x lo

-*

8.51

3x lo

-'

2.5

82

~

1.27

6~

2.10

9~10

-7

-

1.70

6x 10

-8

L. W k w

decomposing at 12% K to its basic sulphate, which is stable up to 1322 K, when ZnO starts forming (Figure 5(a)). Normal zinc sulphate decomposes into its basic sulphate at 1370 K and the basic sulphate into zinc oxide at 1433 K (Figure 5(b)). Thus, between 900 K and 1200 K the decomposi- tion of ferric sulphate takes place, while the zinc and lead sulphates remain 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 practical purposes, it may be assumed that zincore ignites at anywhere between 400-900"C. The rate of combustion increases with increasing temperature and decreases as 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 combustion of the sulphides and the heat losses to the surroundings. As the rate of combustion is reduced, heat input per unit time is also reduced and at a certain point it becomes equal to the heat loss. It is at this point that the spontaneous burning of the sulphides ceases. Too high a roasting temperature may cause the particles to sinter or fuse, which would hamper the inflow of air to the sulphides causing the rate of combustion of the sulphur to drop rapidly.

In roasting, some of the zinc sulphide is oxidised to zinc sulphate, which may be expressed by the following equations:

2S0,+0,=2S03+45 .2 kcal ZnO + SO3= ZnSOI + 55.6 kcal

The impurities contained in the Umm-Gheig ore 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, CaC03 and MgCO3, to form zinc sulphates, aresenates, antimonates, ferritesls or silicates, respectively. Not all of the many possible reactions reach completion here, because either the reacting materials are not present in stoichiometric proportions, or contact between them is upset, or the rate of interaction is too low. The most detrimental secondary reactions in roasting are those producing ferrites of zinc and cadmium and silicates of lead and zinc. The latter, reacting with the sulphuric acid in the subsequent leaching, form colloidal silicic acid which hampers filtration and settling. Zinc ferrite, for its part, reacts with sulphuric acid, 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 increases atabove 650°C.

The suspension roasting process has its origin in the observations on the behaviour of the ore when falling from hearth to hearth in a conventional roaster. In falling, the ore particles come into contact with oxygen-bearing gasps and burn quickly. The rate of combustion is greater than the rate of heat

TaMe4. Percentage dissolution at different roasting temperatures followed by acid and pure water leaching processes

4% H2S0, (by vol) at 60°C) Pure H 2 0 at 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; leaching processes. 1 h.

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 to the surroundings, and the temperature of combustion rises appreciably. As a result, dead- roasting becomes possible without auxiliary heating. Furthermore, a smaller excess of air is required for suspension combustion, as the oxygen is utilised to a fuller extent, and the SOz content 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 successively from 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 g of the ore, divided equally into two boats, were used for leaching in either 4% sulphuric acid (by vol.) at 60°C or 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 used consists in bringing the zinc contained in the ore into solution as zinc sulphate after converting 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 zinc oxide, with the formation and precipitation of ferric hydroxide. This method is commonly 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. To meet these conflicting requirements, the leaching operation is carried out in two stages (double leaching). First, roasted Umm-Gheig ore is treated with a slightly acid solution of ZnSO, containing 100 g dm-, Zn and 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 still carries 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 H2S04 in 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 : l O 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 is hydrolysed to form insoluble basic salts. The underflow, which is the insoluble product of the neutral leach containing by weight 15-20% solids and 80-85% of the neutral solution, is partially filtered and the residue is fed 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.

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 HzS04 and the solid zinc compounds. The grain size of 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 slowly with H2S04. Furthermore, the zinc contained 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 * S O 2 + 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 zinc dissolved increases from 46.8% at 400°C to 91.2% 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 specific temperature, 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°C to 13.2% at 650°C for the same retention time, indicating better sulphate conversion at higher temperature.

Thus, from the results obtained, the optimum condition for controlled roasting of Umm-Gheig ore is at about 650°C for 4 h followed by leaching in 4% sulphuric acid (by vol.) at 60°C for 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 (Tables 4 and 5), but above all, on the free acid content of the lixiviating solution. The yield of extracted zinc increases 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 efficiency are 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 zinc concentration in it is reduced, its acidity increases, and current efficiency decreases, making the complete recovery of zinc from the electrolyte uneconomical. The usual practice is to withdraw the electrolyte from the cell after about half the zinc has been recovered and the equivalent amount of free H2S04 has been regenerated. The depleted electrolyte is then used to leach roasted ore. The least energy consumption can be obtained when an optimum balance between all the factors involved is struck. In the electrolysis of zinc sulphate roasted ore solution, the suitable current density was 40-60 mA Cm-2 at 35°C; the yield of zinc extracted 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 i s 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

210 L. Mdkour

Carbonate 30% Pyrites

Zn-Pb ore

650 'C, 4 h

4% H d G (by vol), 6 O o C

I Sulphate roosting

Acid-leach residue Leach solution PbSO,, 4 0 2 ZnS04, Fe2(S04),

Arnmoniation Coustificotlon

l Fe precipitation

I Arnmbiation Coustificotlon 1 l Fe prec,ipitation / 0'k Na,Zn021each t Zn (NH,)&OH 121

t electrolysis I,' Zinc cathode t product

Figure 6. Sulphate roasting process flowsheet for treatment of Umm-Gheig carbonate Zn-Pb ore.

Zn S04.Fe2(S04)3 directly, 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.x H 2 0 was precipitated and removed by centrifuge. The optimum cathodic current density is 100 mA Cm-2 at 55°C with current efficiency of 9O%, and 97% zinc yield; (3) Elecrolysis of the acid leach solution after ammoniation technique using NH40H. The cathodic current density is 90 mA Cm-2 at 30°C with 88% current efficiency and 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) Geology of 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 Confrucf VI. Eid, A . M.; Abd El-Rehim, M. M. (1963) Mefu//urgicu/ Research on Zn-Pb Oxidised Ore of Umm-Gheig Deposit. Eusfrrn Deserf Geological Survey and Mineral Research Department. Egypt, 22.

Sulptmte roPPtlng of carbonate ore 21 I

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 of a Sulphurion Process for 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. Theorefical Principles 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.