Grind establishment

(1)

4.1 Grind establishment 4. 2 Activox® results

4.3 Scanning Electron Microscopy 4. 4 Moss bauer spectroscopy

4.1 Grind establishment

The results of the grind establishment test are shown in Table 4.1 and also displayed graphically in Figure 4.1. Detailed particle size distribution curves obtained directly from the Malvern sizing are presented in Appendix A.

4.1.1 Establishment of grinding conditions

From the data it is clear that the required milling time to attain a target particle size of 80%- 10 J..Lm was between 20 and 30 minutes. The term P80 represents the particle size

where 80% of the particles were smaller than that size. The power consumption after 30 minutes of milling was 82 kWh/t.

Table 4.1 Grind establishment results

Time Size Pso Power consumption

min _fJ.m KWh/t

0

62.0 0

5

19.0 12

10

13.5 25

20

11.0 55

30

9.5 82

40

9.0 110

50

8.6 137

60

8.5 167

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70 180 60 ₁₅₀ ~

-=

=

50

~

120 ' - '

=

₌

b

₌

.,..; 40 o,..;

-!

... c.

=

90

a

QQ 30

=

=--

~ fl.l

₌

60

=

-~ ~ 00. 20

-~ ~ 10 30

_=--

=

0 0 10 20 30 40 50 60

Figure 4.1 Graph indicating the grind establishment conditions

4.2 Activox® results

The results of the pressure-acid leaching experiments are summarised and represented in Tables 4.2 to 4.8 as well as graphically in Figures 4.2 to 4.17. Detailed data sheets of the results from all the different experiments are given in Appendix B.

Theory suggests (Bredenhann & Van Vuuren, 1999:687) that the reaction is initially controlled by the shrinking particle model and then by diffusion through the product layer model. Results from the present research, however suggest that these two models are competing from the start of the leaching process through to the end, and it is not recommended to fit these well established models separately to different stages of the leaching reaction. This statement is in agreement with a previous investigation by Maurice & Hawk (1999:292).

(3)

Initially a shrinking core model, diffusion through a product layer model as well as a second order model of the form:

ln(a- x) -ln(b- x)

____:. _ ___:;_ _ ___;___...:;_ = k2 t + c

a-b ... (17)

where, x =leached fraction, a and b =parameters, k2 =reaction rate (1/min)

were fitted to the data obtained from the present leaching testwork, but it was observed that the data did not readily fit the expectations of these well established models.

Therefore it was more appropriate to fit the data to a universal first order model (Amer, 1995:230) of the form:

x = 1-e-kt _{... (18)}

where, k = reaction rate (1/min), x = leached fraction and t = time (min)

The results of this model were the best of the evaluated models fitted to the relevant data obtained in the present investigation. Note that the reaction rate values, k, of all of the following experimental data are represented in Appendix D.

4.2.1 Effect of agitation rate on the metal extraction

The influence of agitation speed on the metal recovery for the different metals is shown in Table 4.2 and Figure 4.2(a-d).

(4)

Table 4.2 Results showing the influence of agitation rate on the metal extraction

Nickel

Time 400 rpm 600 l]_)_m 800 l]_)_m 1000 rpm

(min) Recovery (%) Recovery (%} Recover_y_(%

l

Recoveru%1

15 16.5 26.6 32.8 36.2 30 23.3 62.2 74.3 84.2 50 34.3 89.0 90.3 92.1 90 58.7 96.7 93.4 94.9 150 90.4 97.7 97.7 98.1 Cobalt Time 400 rpm 600 rpm 800 rpm 1000 rpm

(min) Recovery(%) Recovery_(o/C!} Recovel"Y_(%1 Recovel"Y_(o/C!}_

15 6.4 15.9 21.2 20.1 30 9.6 46.3 63.1 71.3 50 13.5 79.3 82.5 82.0 90 28.3 88.3 89.2 89.5 150 82.4 94.9 94.9 95.1 Copper Time 400 rpm 600 rpm 800 rpm 1000 rpm

(min) Recovery(%) Recovel1'_(%l Recovel"Y_(o/C!}_ Recovel"Y_(o/'!}_

15 10.6 13.3 16.8 16.2 30 13.5 27.9 35.4 39.9 50 16.7 58.7 61.5 64.9 90 31.8 85.2 85.6 86.7 150 86.3 90.9 90.3 90.8 Iron Time 400 rpm 600 rpm 800 rpm 1000 rpm

(min) Recovery(%) Recovery (%1 Recovery_(%}_ Recovery_(o/C!l_

15 7.2 12.2 11.9 13.4

30 8.5 13.4 12.9 10.8

50 10.3 9.0 8.5 7.1

90 12.9 6.0 5.5 4.9

150 10.8 5.8 5.5 5.1

It was observed that the metal extraction was greatly enhanced with increasing agitation rate. From Figure 4.2 it is clear that the metal recovery improved severely when moving from an agitation rate of 400 r/min to 600 r/min, which implies that the initial stage of the reaction is not only controlled by the chemical reaction but also depends on diffusion through to the mineral surface.

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a) b) tOO tOO 90 _.-. 90

-

_~ 80 ~ '$. 80 .._, .._, ₇₀ ~ 70 ~ ₆₀ ;... ;... ~ 60 ~ 50 >

=

> 50

=

40 ~ ~ ~ 40 ~ ;... ;... _JO

=

JO

z

20 u 20 to to 0 0 0 JO 60 90 t20 t50 0 JO 60 90 t20 t50

Time (min) Time (min)

1--.-4oo rpm --a-6oo rpm ---.-8oo rpm -M-tooo rpm

I

1--.-4oo rpm --a--600 rpm ---.-aoo rpm - M - tooo rpm

I

c) tOO d) _t6 90 .-. ₈₀ '$. .._, ₇₀ t4 .-. ~ _t2 ~ .._, ~ 60 ;... ~ > 50

=

~ ₄₀ ~ ;... _JO

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20 u tO ~ to ;... ~ ₈ >

=

~ ₆ ~ ;... 4 ~ ~ ₂ 0 0 0 JO 60 90 t20 150 0 JO 60 90 t20 t50

1---.-400 rpm --a--600 rpm ---.-8oo rpm - M - tooo rpm

I

1--.-4oo rpm --a--600 rpm ---.-8oo rpm - M - tooo rpm

I

(6)

At an agitation rate of 400 r/min about 60% nickel, 32% copper and less than 30% cobalt was extracted after 90 minutes of leaching, as compared to a nickel, copper and cobalt extraction of close to a 100% for nickel and almost 90% for copper and cobalt at 600 r/min after 90 minutes of leaching. The iron hydrolysis and precipitation process increased with an increase in agitation rate. A stirring rate of 600 r/min and more appeared to be adequate to disperse oxygen sufficiently fast to ensure that the reaction rate was not limited by gas-liquid mass transfer.

Results suggest that sufficient agitation is not only needed for adequate gas-liquid mass transfer but also enhances metal extraction by decreasing the thickness of the boundary layer. It can also be noted that the agitator design is of such nature that oxygen from the gas space in the reactor is mixed back into the pulp via the holes in the agitator shaft and the bottom impeller. This means that if the agitation rate is increased, the amount of oxygen fed back to the pulp also increased, which resulted in better gas-liquid mass transfer as well as higher concentrations of oxygen in the pulp. The dependence of the reaction rate on the agitation rate is represented in Figure 4.3. The reaction rate increased significantly with an increase in agitation rate.

0 . 0 5 0 . - - - , 0.045 0.040 0,035

=

] 0030 '""' . ._, 4) 1; ... 0.025

=

Q ;:: ~ 0.020 ~ 0.015 0.010 0.005 -0.000 + - - - . - - - - , - - - . - - - , - - - j 200 400 600 800 1000 1200

Agitation rate (r/min)

~Ni

--llll-Co -tr-cu

(7)

4.2.2 Effect of solid content on the metal extraction

The influence of the solids content on the metal recovery for the different metals is represented in Table 4.3 and Figure 4.4(a-d).

Table 4.3 Results showing the influence of solid content on the metal extraction

Nickel

Time 5% 11% 15% 20%

(min) Recovery (%) Recovery (%) Recovery (%) Recovery (%)

15 34.7 32.8 31.3 26.8 30 73.2 74.3 70.5 56.6 50 94.1 90.3 88.5 85.1 90 95.1 93.4 94.1 92.6 150 97.1 97.7 98.2 97.6 Cobalt Time 5% 11% 15% 20%

(min) Recovery(%) Recovery (%) Recovery (%) Recovery (%)

15 23.1 21.2 17.4 12.0 30 60.3 63.1 60.5 33.1 50 85.8 82.5 83.5 76.0 90 89.6 89.2 89.9 86.4 150 93.5 94.9 94.8 94.8 Copper Time 5% 11% 15% 20%

(min) Recovery(%) Recovery (%) Recovery_(%)_ Recovery_(%)_

15 17.6 16.8 14.5 9.3 30 34.0 35.4 32.4 22.6 50 55.3 61.5 68.5 56.3 90 75.9 85.6 89.8 86.3 150 85.1 90.3 91.7 92.4 Iron Time 5% 11% 15% 20%

(min) Recovery (%) Recovery (%) Recovery (%) Recovery_(%}_

15 9.4 11.9 13.0 12.8

30 11.1 12.9 14.7 16.4

50 7.9 8.5 9.5 10.8

90 4.9 5.5 5.0 5.7

(8)

a) b) 100 100 90 90

-

80 ~ = 70 '-"

-

~ 80 = '-" ₇₀ ~ 60

""

~ 60

""

~ ₅₀

,..

Q 40 ~ ~

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50 Q ~ 40 ~ 30 "" -- ~,~ _____________________________ _ ~

""

30

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_z

20 Q 20 u 10 10 0 0 0 30 60 90 120 150 0 30 60 90 120 150

l~5%--e-ll%-6--15%~20% 1 1 ~5% - e - l l % --6--15% ~20% 1 c) 100 d) 18 90 16

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"$. 80 '-" ₇₀ ~

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60 ~

,..

₅₀ Q ~ ₄₀

-"$. 14 '-" ~ 12

""

10 ~

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Q 8 ~ ~

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30 ~ 6

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=' 20 u 10 ~ 4 ~ 2 0 0 0 30 60 90 120 150 0 30 60 90 120 150

1~5%--G-11%-tr-15%~20% 1 1 ~5% - e - l l % --6--15% ~20% 1

(9)

Metal recoveries of higher than 90% were obtained for 15% solids present in the pulp after 150 minutes of leaching for the metals nickel, copper and cobalt. The extraction of iron increased and the precipitation rate decreased with increasing pulp densities. Although the oxygen flow rate was constant, it was fed in stoichiometric excess for the experiments, to ensure that oxygen did not become the limited reagent when increasing the pulp density.

The dependence of the reaction rate on the leaching on the amount of solids present is represented in Figure 4.5. The reaction rate is clearly sensitive to the pulp density, seeing that the nickel and cobalt recoveries decreased with an increasing pulp density, whilst the copper recoveries appeared to be better at a pulp density of 15% solids, after which the reaction rate decreased for a solid content of above 15%.

0.045 .,.---~ 0.040 - - - -0.035

i

0.030

-

,...; ';' 0.025

e

=

s

0.020 ~ ,... 0.015 0.010 -0.005 -0.000 + - - - . - - - . . . - - - . - - - . - - - 1 0 5 10 15 20 25 solids content (%) ~Ni -m--co -ir-Cu

Figure 4.5 Graph indicating the dependence of the reaction rate, k, on the solid content

(10)

4.2.3 Effect of total pressure on the metal extraction

The influence of the total pressure on the metal recovery for the different metals is shown in Table 4.4 and Figure 4.6(a-d).

Table 4.4 Results showing the influence of total pressure on the metal extraction

Nickel

Time 4 bar 7bar lObar 13 bar

(min) Recovery(%) Recovery (%) Recovery(%) Recovery (%)

15 18.8 25.1 32.8 39.3 30 30.9 51.4 74.3 85.3 50 54.7 85.3 90.3 90.0 90 88.9 94.5 93.4 92.6 150 96.2 97.0 97.7 97.6 Cobalt

(min) Recovery(%) Recovery (%) Recovery (%) Recovery (%)

15 6.9 12.9 21.2 22.4 30 13.9 31.4 63.1 75.1 50 29.0 75.8 82.5 81.1 90 80.9 86.8 89.2 86.4 150 90.9 92.8 94.9 94.5 Copper

Jmin) Recovery_(%)_ Recovery (%) Recovery (%) Recovery (%)

15 10.7 13.1 16.8 16.2 30 14.8 24.2 35.4 34.9 50 25.7 53.1 61.5 56.5 90 73.1 83.2 85.6 82.9 150 90.8 89.6 90.3 89.9 Iron

(min) Recovery(%) Recovery(%) Recovery_(%)_ Recovery_(o/ol

15 8.2 11.3 11.9 13.2

30 10.6 13.4 12.9 10.5

50 12.2 10.2 8.5 6.2

90 10.2 6.3 5.5 4.1

(11)

a) b) 100 ₁₀₀ 90 ₉₀

-

80 ~

-

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'"'

;.-.

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60 ~ 50

...

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50

=

40 y

=

y ₄₀ ~ 30

'"'

~

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30

....

₂₀ z 10

=

20 u ₁₀ 0 0 0 30 60 90 120 150 0 30 60 90 120 150

~~4bar ~7bar ---6-10bar ~13bar ~~4bar - a - 7 b a r ---6-10bar ~13bar

I

c) d) 100 16 90

-~ 80 e ' - ' ₇₀ 14

-~ e ₁₂ ' - ' ;.-.

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60 50

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20 u 10 ;.-. ₁₀

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4 ~ ~ ₂ 0 0 0 30 60 90 120 150 0 30 60 90 120 150

J~4bar - a - 7 b a r ---6-10bar ~13bar ~~4bar - a - 7 b a r ---6-10bar ~13bar

I

(12)

All the valuable metals were leached more than 90% successfully for total pressures in the range of 4 to 13 bar, after a leaching time of 150 minutes. In the initial stage of the reactions, the increase in pressure resulted in significant higher valuable metal extractions, and can be ascribed to the fact that the oxygen solubility increased with an increase in pressure and enhanced oxygen utilisation. Typically, from the results obtained, a 10 bar total pressure seemed to be the optimum pressure for maximum leaching.

The reaction rate for all three valuable metals increased as the total pressure increased from 4 to 13 bar. The influence of oxygen pressure on the reaction rate was more perceptible for nickel and cobalt than for copper. Pressures of above 10 bar did not greatly improve the metal extraction. It is interesting to note that the iron dissolution increased initially and decreased as time progressed, with an increase in total pressure. The dependence of the reaction rate on the total pressure is shown in Figure 4. 7

0 . 0 5 0 . . -0.045 - - - -0.040 0.035

i

_~_0.030 ._,

~

0.025 § := _~ 0.020 ~ 0.015 0.010 - - - -0.005 -0.000 + - - - r - - - , - - - . - - - . - - - , - - - 1 2 4 6 8 10 12 14 Pressure (Bar)

Figure 4. 7 Graph indicating the dependence of the reaction rate, k, on the total pressure

(13)

4.2.4 Effect of temperature on the metal extraction

The influence of the temperature on the metal recovery for the different metals 1s represented in Table 4.5 and Figure 4.8(a-d).

Table 4.5 Results indicating the influence of temperature on the metal extraction

Nickel

Time 90°C 100°C l10°C 120°C

(min) Recovery (%) Recovery_(o/1!)_ Recovery (%) Recovery (%)

15 26.7 31.6 32.8 31.1 30 51.9 64.5 74.3 81.9 50 72.0 86.1 90.3 90.8 90 88.4 92.4 93.4 93.2 150 94.3 97.1 97.7 98.4 Cobalt Time 90°C 100°C l10°C 120°C

(min) Recovery(%) Recovery_(%} Recove.-y_(%}_ Recovery (%)

15 20.4 23.1 21.2 19.7 30 45.0 52.5 63.1 69.1 50 64.5 78.0 82.5 82.0 90 82.8 87.8 89.2 89.9 150 89.8 92.7 94.9 96.7 Copper Time 90°C 100°C l10°C 120°C

15 19.0 19.4 16.8 13.3 30 26.7 31.1 35.4 40.9 50 37.9 49.5 61.5 75.3 90 53.5 71.3 85.6 88.6 150 68.3 85.6 90.3 92.2 Iron Time 90°C 100°C l10°C 120°C

(min) Recovery(%) Recovery (%) Recovery(%) Recovery (%)

15 5.8 8.6 11.9 10.6

30 6.1 11.0 12.9 10.2

50 7.0 8.0 8.5 6.8

90 6.6 5.3 5.5 6.6

(14)

a) b) 100 100 90 ₉₀

-

80 ~

-

~ 80 ' - ' 70 ' - ' ₇₀ ;;.... 60 ""' ;;.... 60 ""' ~

...

₅₀

_...

~ 50 Q 40 u Q u 40 ~ 30 ""' ~ ""' 30

....

₂₀

z

10 Q 20 u 10 0 ₀ 0 30 60 90 120 150 ₀ ₃₀ ₆₀ ₉₀ ₁₂₀ ₁₅₀

Time (min) _{Time (min)}

1--+--90~ ~100~ ~110~ --M--120~ _{j--o--90~ ~100~ ~110~ --M--120~} c) 100 d) 14 90

-

~ 80 = ' - ' ₇₀

-12 ~ = 10 ' - ' ;;.... ""' 60 ~

...

50 Q u ₄₀ ~ ""' 30 ;;.... ""' 8 ~

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Q ₆ u ~ ""' 4

=

20 u 10 ~ ~ ₂ 0 0 0 30 60 90 120 150 0 30 60 90 120 150

j--o--90~ ~100~ ~110~ --M--120~ j--o--90~ ~100~ ~110~ --M--120~

(15)

It has been observed that the temperature had a much greater influence on the dissolution of copper from chalcopyrite than on nickel and cobalt from pentlandite. The dissolution rate of copper increased almost linearly with an increase in temperature, while the other metals (ollowed a parabola curve. At a temperature of 90°C only 70% copper was recovered, whilst at a temperature of 11 0°C, 90% copper was recovered, after 150 minutes of leaching.

Various research projects have been undertaken on the leaching of metal sulphides below and above a temperature of 120°C, the melting point of sulphur, but leaching at exactly 120°C received limited, if any attention. According to Habashi (1999:292) the solid sulphur undergoes a slow transformation at 95.4°C from a rhombic to a monoclinic structure and once in the monoclinic form, elemental sulphur melts at 119.5°C to form a yellow fluid liquid known as A.-sulphur.

It was observed that at 120°C the elemental sulphur underwent a transformation and formed spherical pellets of approximately 1 to 10 mm in diameter (see Figure 4.9). This sulphur transformation however did not impede metal dissolution and metal recoveries of more than 92% were obtained for the nickel, copper and cobalt. These spherical sulphur pellets can be recovered immediately after leaching as a by-product without any further processing, ensuring that adequate facilitation is provided and the possibility of pipe blockage is accounted for.

Figure 4. 9 Photograph of the elemental sulphur pellets obtained from a leach

(16)

Figure 4.10 displays the dependence of the reaction rate of leaching on the temperature.

It is clear from Figure 4.10 that the reaction rate increased significantly for all three the valuable metals with an increase in temperature.

The activation energies for the nickel, copper and cobalt extracted from the sulphide concentrate were determined by fitting the apparent reaction rate constants to the well-known Arrhenius equation:

lnk=c-_g_

RT ... (19)

where, k

=

reaction rate constant (1/min), Q

=

activation energy (J/mol K).

The values of the apparent activation energies of nickel, copper and cobalt were found to be 20.6 (± 4.4) kJ/mol K, 33.6 (± 4.2) kJ/mol K and 17.4 (± 3.5) kJ/mol K, respectively. These values are in agreement with those found by Amer (1995:231).

0 . 0 4 5 - , - - - , 0.040 0.035 ~ 0.030

--e _~_0.025

e

=

:§ 0.020 ~ ~ 0.015 0.010 0.005 -0.000 + - - - . - - - . - - - . - - - r - - - i 80 90 100 110 120 130 Temperature

eq

Figure 4.10 Graph indicating the dependence of the reaction rate, k, on the temperature

-<>-Ni -EI-Co -lir-Cu

(17)

It was observed that the leaching of the iron increased initially with an increase in temperature, which relates to the higher valuable metal recoveries in the early stage of the reaction. The iron dissolution decreased as time progressed due to hydrolysis and precipitation of ferric ion in the form of goethite (FeOOH).

Figures 4.11 and 4.12 represents the concentration of ferrous and ferric iron present in the leach solution, sampled at various time intervals, at the temperatures of 90°C and 11 0°C respectively. These two temperature values were chosen for the specific iron analysis because of its significant influence on the copper recovery from the sulphide concentrate. 4500 4000 - - - -, 3500 3000 }2500 ' - ' + M 2000 4)

...

1500 1000 500 0 0 20 40 60 80 100 120 140 160 Time (min)

Figure 4.11 Graph indicating the amount of ferrous iron present in the leach solution at T=90°C and T=110°C

It is clear from Figure 4.11 that the amount of ferrous iron in solution in the initial stage ofthe reaction is much higher at a temperature of l10°C. After 50 minutes of leaching, where the amount of ferrous iron formed at 11 0°C equalled the amount of ferrous iron formed at 90°C, a third of the copper was recovered at 90°C whilst almost two-thirds of the copper was recovered at 110°C, which can be related to the larger amount of ferrous iron oxidised, in the initial stage of the reaction, at the temperature of 11 0°C.

(18)

At the corresponding 50 minutes of leaching time the amount of ferric iron formed at 110°C is a factor of more or less four times higher, than the amount of ferric iron formed at 90°C. Ferric iron is known for its use as an oxidising agent in a sulphide leaching process. 1750,---, 1500 1250 ~ 1000

s

j; ~ 750 500 -250 0+---~----~----~----~----~----~----~--~ 0 20 40 60 80 100 120 140 160 Time (min) -<>-r=90°C -li!l-T=ll ooc

Figure 4.12 Graph indicating the amount of ferric iron present in the leach solution at T=90° C and T=ll 0° C

Figure 4.12 indicates that a great difference occurs in the concentration of ferric iron in the leach solution at the temperature of 90°C related to the values found at 11 0°C. The high ferric iron concentration in the solution at a temperature of 110°C will enhance the leaching process, which contributed to the much higher reaction rate, revealed in Figure 4.10, as well as the higher metal recoveries found at a temperature of ll0°C, shown in Figure 4.8. After 150 minutes of leaching two-thirds of copper was recovered at a temperature of 90°C compared to more than 90% of copper recovered at a temperature of

(19)

4.2.5 Effect of sulphuric acid concentration on the metal extraction

The influence of the sulphuric acid concentration on the metal recovery for the different metals is shown in Table 4.6 and Figure 4.13(a-d).

Table 4.6 Results indicating the influence of sulphuric acid concentration on the

metal extraction

Nickel

Time 0 kg/ton 30 kg/ton 60 kg/ton 90 kg/ton

15 23.5 25.6 32.8 32.2 30 71.7 66.4 74.3 79.6 50 87.1 90.0 90.3 93.5 90 90.8 93.6 93.4 97.1 150 95.5 97.4 97.7 97.8 Cobalt

15 13.3 15.2 21.2 19.1 30 60.7 55.0 63.1 70.3 50 78.9 79.9 82.5 87.2 90 83.7 86.8 89.2 90.7 150 93.0 94.3 94.9 94.9 Copper

15 6.8 10.5 16.8 17.8 30 32.2 30.9 35.4 38.1 50 60.7 61.0 61.5 66.9 90 81.9 83.5 85.6 88.9 150 90.6 89.2 90.3 91.4 Iron

(min) Recovery (%) Recovery(%) Recovery (%) Recovery (%)_

15 7.3 9.6 11.9 16.9

30 9.7 12.0 12.9 12.1

50 6.5 7.2 8.5 9.7

90 4.2 4.6 5.5 6.3

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a) b) tOO tOO 90 90

-

80 ';;(( =

-

?f?. 80 ._, ₇₀ ._, 70 ;;... 60

..

;;... 60

..

Q;l 50

...

Q;l 50

...

~ ₄₀ y ~ y ₄₀ Q;l JO

..

Q;l

..

JO

....

₂₀ z tO ~ ₂₀ u _to 0 0 0 JO 60 90 t20 t50 0 JO 60 90 t20 t50

I

~o kg/tDn --m--:so kg/tDn ~60 kU/tDn --M---90 kU/tDn

I

c) _tOO d) t8 90 _t6

-

_-

';;(( t4 = ._, t2 ;;...

..

Q;l to

...

₈ ~ y Q;l 6

..

';;(( 80 = ._, 70 ;;...

..

60 Q;l

...

50 ~ y ₄₀ Q;l

..

JO Q;l 4 ~ 2 ;:I ₂₀ u to 0 0 0 JO 60 90 t20 t50 0 JO 60 90 t20 t5o

/--+--0 kU/tDn -m-:so kU/tDn ~60 kg/tDn --M---90 kg/tDn

I

1--+--0

kU/tDn --a-:so kg/tDn ~60 kg/tDn --M---90 kU/tDn

I

(21)

The sulphuric acid was introduced in a batch wise process into the reactor when the experiment commenced. No additional acid was introduced once the experiment started, resulting in the final pH of the leach solution to vary between 1.27 and 1.36 for all the experiments.

Results illustrated that by even no addition of any sulphuric acid, enough sulphuric acid was generated by the leach reaction of the sulphide minerals present, to ensure recoveries of over 90% for the nickel, copper and cobalt. An increase in the acid concentration to above 30 kg/ton did not have a remarkable influence on the metal recovery, signifying that high sulphuric acid concentrations will not improve the extraction of the valuable metals under the specific conditions studied in this project. The formation of elemental sulphur as well as the free acid concentrations increased with an increase in sulphuric acid concentration (see Appendix B). Higher sulphuric acid concentrations resulted in higher iron concentrations in solution as well as lower precipitation rates. It can be noted that the reaction rate thus will not be influenced significantly by the sulphuric acid concentration, as shown in Figure 4.14.

0 . 0 4 5 . , - - - . , 0.040 0.035

:5

0.030 ..e

....

';" 0.025 •

-e

~ 0.020

-!

0.015 -0.010 -0.005 -0.000 +---.---.---,---,---,---,---.--.---,---1 0 10 20 30 40 50 60 70 80 90 100

Sulphuric acid (kg/ton)

Figure 4.14 Graph indicating the dependence of the reaction rate, k, on the sulphuric acid concentration

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4.2.6 Effect of aeration on the metal extraction

The influence of aeration on the metal recovery for the different metals is represented in Table 4.7 and Figure 4.15(a-d).

Table 4. 7 Results indicating the influence of aeration on the metal extraction

Nickel

Time 151/h 40 llh 651/h 90 1/h

(min) Recovery (%) Recovery (%) Recovery(%) Recovery (%)

15 30.3 29.2 32.8 34.9 30 61.0 60.4 74.3 80.9 50 81.4 90.7 90.3 89.6 90 92.4 96.4 93.4 93.1 150 97.9 98.4 97.7 97.8 Cobalt Time 151/h 40 llh 651/h 90 1/h

(min) Recovery (%) Recovery(%) Recovery (%) Recovery (%)

15 16.1 15.0 21.2 18.1 30 42.7 39.4 63.1 72.2 50 69.6 76.8 82.5 82.3 90 84.9 89.1 89.2 88.4 150 94.7 95.4 94.9 94.4 Copper Time 151/h 40 llh 651/h 90 1/h

15 13.1 13.8 16.8 15.3 30 27.7 27.9 35.4 36.9 50 48.6 55.6 61.5 60.6 90 82.0 85.2 85.6 85.1 150 90.9 91.0 90.3 90.8 Iron Time 151/h 40 1/h 651/h 90 1/h

(min) Recovery(%) Recovery (%) Recovery(%) Recovery(%)

15 12.8 12.1 11.9 13.4

30 13.0 13.3 12.9 10.7

50 11.1 10.3 8.5 7.7

90 7.3 6.2 5.5 5.3

(23)

a) b) 100 100 90 90

-

80 ~ "'

-

~ "' 80 ' - ' 70 ' - ' ₇₀ ~ 60

'"'

~ ~

'"'

60 50

...

~

...

50

=

~ 40

=

~ 40 ~ 30

'"'

~

'"'

30

....

₂₀

z

10

=

20 u 10 0 ₀ 0 30 60 90 120 150 0 30 60 90 120 150

1---<>--15 1111 - - a - 4 0 1111 --A--65 1111 ~ 90 1111 1 1---<>--151/h -m--40 1111 ---A--651111 ~90 1111 1 c) 100 d) ₁₆ 90

-

14 ~ 80 "' ' - ' ₇₀

-

-.ft. 12 ' - ' ~

'"'

60 ~ ~ 10

'"'

...

50

=

...

~ 8 ~ 40 ~

=

~ ₆

'"'

30 == 20 u 10 ~

'"'

4 ~ ~ ₂ 0 0 0 30 60 90 120 150 0 30 60 90 120 150

1--o--15 1111 ---s--40 1111 --A--65 1111 ~ 90 1111 1 1---<>--15 1111 -m--40 1111 ---t:--65 1111 ~90 1111 1

(24)

The actual rate of the leach reaction depends on the rate of gas-liquid mass transfer through the mineral surface, or through any boundary layer that may occur. Because in the present research project it is suggested that the kinetics of the leaching ofpentlandite, chalcopyrite and pyrrhotite can be explained in terms of a mixed chemical and diffusion controlled model, it is of importance to ensure that adequate amounts and sufficient transport of reactants to mineral surface must occur.

The oxygen flow rates used in this research, varied between 15 1/h and 90 1/h, seemed to be sufficient to ensure that the reaction was not limited by gas-liquid mass transfer. Figure 4.15 illustrates that an increase in oxygen flow rate increased metal extractions initially but did not greatly enhance the overall metal recovery. A flow rate of 40 1/h 02

resulted in recoveries of higher than 90% for all three the valuable metals nickel, copper and cobalt, after a leaching period of 150 minutes.

0 . 0 4 5 . . , . . . - - - , 0.040 0.035 'd 0.030 ] _... ';' 0.025 -+-Ni

e

-m-co ---k~-~-~-~-~--~-~-~-~---tr--cu

=

~ 0.020

A

,.... 0.015 --- m---..,....--=---=---=---:---0.010 0.005 0.000 +----,---.---,.----,r--~----,----,---,----,---~ 0 10 20 30 40 50 60 70 80 90 100 Aeration (1/h Oz)

Figure 4.16 Graph indicating the dependence of the reaction rate, k, on aeration

Figure 4.16 represents the dependence of the reaction rate on the oxygen flow rate. An increase in oxygen flow rate improved the reaction rate for nickel and cobalt but did not greatly effect the dissolution rate of copper from the sulphide concentrate.

(25)

4.2. 7 Effect of particle size, HN03, AgN03 and FeOOH on the metal extraction

The influence of particle size, HN03, AgN03 and FeOOH on the metal recovery for the

different metals, were conducted to evaluate whether changing the particle size of the concentrate, or adding other oxidising agents will improve the valuable metal recovery.

Optimum conditions, determined in the preceding experiments, were incorporated and were set at a temperature of 11 0°C, a pressure fixed at 10 bar, an oxygen flow rate of 65 1/h and a sulphuric acid concentration of 60 kg/ton.

The particle size of the concentrate, used to determine the influence of particle size on the metal recovery, was 80% smaller than 62 j..Lm. The amount of HN03 and AgN03 added,

as catalysts were 3 g/1 and 0.3 g/1 respectively. The amount of goethite (FeOOH), which was obtained from a previous leaching experiment, added was 30g. In the last mentioned experiments the 80% - 10 j..Lm particle size concentrate was used. The influence of particle size, HN03, AgN03 and FeOOH on the metal recovery for the different metals is

(26)

Table 4.8 Results showing the influence of particle size, HN03, AgN03 and

goethite on the metal extraction

Nickel

Time BeforeUFM _{HN03 =3 g/1} _{AgN03 = 0.3 gil} Goethite = 30e: Baseline (min) Recovery(%) Recovery (%) Recovery (%) Recovery (%) Recovery (%)

15 9.3 41.6 15.8 31.1 32.8 30 15.7 78.4 43.1 78.1 74.3 50 27.2 80.2 54.6 85.9 90.3 90 47.7 86.1 76.6 91.2 93.4 150 65.2 96.9 93.7 97.2 97.7 Cobalt

Time BeforeUFM _{HN03 =3 g:/1} Ae:N03 = 0.3 g:/1 Goethite = 30e: Baseline (min) Recovery(%) Recovery (%) Recovery (%) Recovery (%) Recovery (%)

15 7.4 25.8 4.6 17.7 21.2 30 12.5 63.5 6.5 60.3 63.1 50 23.3 74.1 18.4 74.8 82.5 90 43.0 81.4 62.4 83.7 89.2 150 59.8 93.7 88.1 94.9 94.9 Copper

Time BeforeUFM HN03 = 3 e:/1 Ae:N03 = 0.3 e:/1 Goethite = 30e: Baseline (min) Recovery (%) Recovery (%) Recovery (%) Recovery (%) Recovery (_%)_

15 4.7 22.2 26.0 18.4 16.8 30 5.8 42.1 36.3 47.9 35.4 50 9.9 65.7 58.3 82.4 61.5 90 22.6 84.4 88.1 87.9 85.6 150 37.7 90.1 90.3 90.9 90.3 Iron

Time BeforeUFM _{HN03 =3 g/1} _{AgN03 = 0.3 g/1} Goethite = 30e; Baseline (min) Recovery(%) Recovery (%) Recovery (%) Recovery (%) Recovery (%)

15 10.0 14.1 7.9 14.5 11.9

30 9.9 12.2 8.5 13.1 12.9

50 8.5 7.8 8.0 8.9 8.5

90 6.7 5.2 5.0 6.6 5.5

(27)

b) 100 a) 100 90

-

- - - ---~-~-~---~---90 ::::R 80 =

-

70 ;;....

'"'

60 ~

-

80 ::::R =

-

70 ;;.... 60

'"'

~ 50

=

~ ₄₀ ~ ~ 50 ~

=

~ 40

'"'

30 ~ ₃₀

'"'

₌

20 u

...

_z 20 10 10 0 0 0 30 60 90 120 150 0 30 60 90 120 150

j---+--BeforeUFM ~Baseline j j---+--BeforeUFM -s-BCISEIIne j

c) d) 100 16

-

14 ::::R ₁₂ =

-90

-::::R 80 =

-

70 ;;.... ₁₀

'"'

~ 8 ~

=

~ ₆ ~

'"'

4 ~ ;;....

'"'

60 ~ ~ 50

=

~ 40 ~ 30

'"'

=

20 u ₁₀ ~ ₂ 0 0 0 30 60 90 120 150 0 30 60 90 120 150

j---+--BeforeUFM -e--Basellne j j--o--BeforeUFM ---a--Baseline j

(28)

b) 100 a) 100 90 90

-

~ 80

-

70 I»

""

60 ~

-

80 ~

-

70 I» ₆₀

""

~ ₅₀ c Col ₁₁₀ ~

""

30 ~ ₅₀ ~ c ₁₁₀ Col ~ 30

""

;:I 20 u ... ₂₀

z

10 10 0 0 0 30 60 90 120 150 0 30 60 90 120 150

I~

HN03 =

3

g/1 --a--Baseline

I

~HN03

=3

g/1 ~Baseline

I

c) 100 d) 16 90 ₁₁₁

-~ 80 =

-

~ 12

-

70

-I» 60

""

I»

""

10 ~ 50 ~ ~ ~ 8 c Col 110 ~

""

30 c ₂₀ u ₁₀ c Col ₆ ~

""

II ~ ~ ₂ 0 0 0

30

60 90 120 150 0 30 60 90 120 150

I----4-HN03 =3

g/1 --a--Baseline

I

I-<>-HN03 =3

g/1 --a--Baseline

I

(29)

a) b) 100 100 90 90

-

80 ~ ._, ₇₀ ;;... 60

'"'

-

~ 80 ._, 70 ;;...

'"'

60 Q,) 50

...

Q 40 ~ Q,)

...

50 Q ~ 40 Q,) 30

'"'

Q,)

'"'

30

...

20 z 10 == 20 u ₁₀ 0 0 0 30 60 90 120 150 0 30 60 90 120 150

J--o--A9N03 =0.3 g/1 --m--Basellne

I

J--o--AgN03 =0.3 g/1 -a--Baseline

I

c) d) 100 16 14

-~ 12 ._, 90

-

80 ~ ._, 70 ;;... 10

'"'

;;... 60

'"'

Q,) 8

...

Q,) 50

...

Q ~ ₆ Q,)

'"'

4 Q,) ~ ₂ Q 40 ~ Q,) 30

'"'

Q 20 u 10 0 0 0 30 60 90 120 150 0 30 60 90 120 150

J ~AgN03 =0.3 g/1 - - - a - B - n e

I

l--o--AgN03 =0.3 g/1 -a--Baseline

I

(30)

a) 100 b) 100 90 90

-

80 '#.

-

70

-80 '#.

-

70 ;;... 60 lot ~ 50

..

_~ 40 CJ ~ lot 30 ;;... 60 lot ~

..

50 ~ CJ ₄₀ ~ lot ₃₀

z

20

=

20 u 10 10 0 0 0 30 60 90 120 150 0 30 60 90 120 150

I

~GOeiiiHe=30g -m--Be&elne

I

~Goe111He:30g --m-BIISEIIne

I

c) 100 d) 16 90 ₁₄

-

80 '#.

-

70

-

'#.

-

12 ;;... 60 lot ;;... lot 10 ~ 50

..

~ 8

..

~ CJ 40 ~ lot 30 ~ CJ ₆ ~ lot ~ 20 u 10 ~ 4 ~ 2 0 0 0 30 60 90 120 150 0 30 60 90 120 150

I

~Goe111He=30g --m-BIISEIIne

I

~Goeii!He:30g --m-BIISEIIne

I

(31)

From Figure 4.17 one can conclude that fine grinding plays an important role on the leaching of the sulphide minerals. The metal extraction and leaching rate of the as-received sample decreased remarkably in comparison with the finely ground sample. After 150 minutes of leaching less than two-thirds of the nickel and cobalt were leached and the copper recovery was less than 40% while almost complete leaching of nickel, copper and cobalt was achieved at the same conditions when the fine ground sample was used.

Previous research (Anderson, 2000:8) has been undertaken to determine the possible advantages of nitrogen species to be used as catalysts in the pressure leaching of copper concentrates. No definite advantage of HN03 on the copper extraction has been found

and even a slightly negative influence on the extraction of nickel and cobalt from the other minerals present was suggested.

In the research project done by Miller and Portillo (1979:851) it was proposed that the introduction of silver ions as a catalyst, will react with the chalcopyrite to from a silver sulphide film on the CuFeS2 surface. They suggested that the Ag2S will alter the

morphology of the solid elemental sulphur product layer, making it more porous so that the reaction rate will no longer be controlled by reaction/product diffusion through the elemental sulphur. No exceptional better recoveries were obtained for copper extraction, in the present research project, by adding AgN03 as catalyst. This might be due to the

fact that the concentrate was already pre-treated to increase its reactivity by means of ultra fine milling. The addition of AgN0₃ however had a remarkable effect on the leaching of the nickel and cobalt from pentlandite. The reaction rate for the extraction of nickel and cobalt decreased dramatically with the addition of the silver catalyst and results from this investigation suggested that the addition of silver catalyst is not recommended m copper concentrates containing other valuable minerals such as pentlandite.

(32)

Finally goethite (FeOOH), obtained from a previous leach experiment, was added to investigate if the iron hydrolysis and precipitation process will accelerate due to the presence of the already formed FeOOH, which will in turn improve the leaching kinetics. Figure 4.20 shows that the addition of FeOOH did in fact influence the iron and copper dissolution compared to the previous experiments without the addition ofFeOOH. After 50 minutes of leaching, 62% of the copper was extracted without any further addition of goethite whilst almost 85% copper was extracted when the added goethite was present. This contribution could be due the accelerated hydrolysis process or it might be because of the increase in ferric iron concentration in solution as a result of goethite ionisation, supporting the experiments where the temperature was varied.

4.2.8 Reproducibility of Activox® experiments

Section 4.2. 8 represents the data of repeated baseline tests in order to determine the reliability of the preceding experimental results. The results of the repeated baseline tests are shown in Table 4.9 and Figure 4.21(a-d).

It is clear from Table 4.9 and Figure 4.21(a-d) that the results from the repeated baseline tests are consistent, it is therefore appropriate to approve the experimental data obtained from the present investigation.

(33)

Table 4.9 Results showing the reproducibility of the Activox® tests

Nickel

Time Baseline 1 Baseline 2 Baseline 3 (min) Recovery (%) Recovery (%) Recovery (%)

15 32.8 34.7 35.5 30 74.3 75.7 74.3 50 90.3 91.1 89.6 90 93.4 95.6 93.6 150 97.7 97.9 97.7 Cobalt

15 21.2 23.3 23.5 30 63.1 64.6 65.7 50 82.5 85.3 81.7 90 89.2 89.4 90.7 150 94.9 94.7 94.8 Copper

15 16.8 18.0 16.7 30 35.4 32.7 35.8 50 61.5 58.2 62.6 90 85.6 84.0 87.0 150 90.3 89.4 90.9 Iron

Time Baseline 1 Baseline 2 Baseline 3 _(min) Recovery (%) Recovery(%) Recovery(%)

15 11.9 10.5 12.1

30 12.9 13.9 13.0

50 8.5 9.3 8.8

90 5.5 5.4 6.6

(34)

a) b) 100 100 90 .-.. ₈₀ ?ft. ' - ' 70 90 .-.. ₈₀ ?ft. ' - ' ₇₀ --- -~---~--~---~--~---~--~-->. ₆₀ ;... ~ 50

...

=

u 40 ~ 30 ;... >. 60 ;... ~

...

50

=

u ₄₀ ~ ;... ₃₀

....

₂₀ z _u;::s 20 10 10 0 0 0 30 60 90 120 150 0 30 60 90 120 150

J-o-

Baseline 1 - a -Basellne2 ~Basellne3 !-<>--Baseline 1 -m-Basellne2 ~Basellne3

c) 100 d) 16 90 ₁₄ .-.. ₈₀ ~ 0 '-'• ₇₀ .-.. ~ ₁₂ 0 ' - ' >. 60 ;... >. 10 ;... ~

...

50

=

u 40 ~ ;... ₃₀ ~ 8

...

=

u ₆ ~ ;...

=

20 u 10 ~ 4 ~ 2 0 0 0 30 60 90 120 150 0 30 60 90 120 150

j-o-Basellne 1 -a-BII!iellne2 ~Basellne3

1--<>--

BCISE!IIne 1 -e--BCI5EIIne2 ~ Basellne3

(35)

4.3 Scanning Electron Microscopy (SEM) results

In this section mineralogical characteristics of the minerals present in the solid phase were investigated by means of SEM analysis. The main purpose of these analyses was to determine and observe any kind of mineralogical changes that occurred and reaction products that formed during the leaching process.

4.3.1 SEM micrographs of Activox® leach residues

The samples were prepared by setting small amounts of the specific residue in an Araldite epoxy. After the sample was set, it was polished to ensure a smooth surface, and the samples were then introduced into the microscope.

The samples were further characterised by means of an energy dispersive X-ray spectrometer (EDAX) fitted together with the Scanning Electron Microscope. The quantification spectra of the specific surface areas analysed are presented in Appendix C. In Figure 4.22 the sample, after ultra fine milling and prior to leaching, is shown and it is clear that the required grain size of 80% - 10 ~m was achieved by the ultra fine milling.

(36)

Goethite matrix

Figures 4.23 and 4.24 indicate the difference of the leach residue obtained from a leach experiment after 15 minutes (Figure 4.23) and 150 minutes (Figure 4.24) of leaching. It

is clear that after 15 minutes of leaching, a large amount of metal still remained to be leached, while after 150 minutes of leaching only a goethite (FeOOH) matrix remained.

Figure 4.23 SEM micrograph of leach residue after 15 minutes of leaching

Figure 4.24 SEM micrograph of leach residue after 150 minutes of leaching

Partially

leached metal

(37)

Figure 4.25 and 4.26 represent a pentlandite and chalcopyrite particle respectively after 15 minutes of leaching. After 15 minutes of leaching, the pentlandite particle already underwent a partial transformation, whilst the chalcopyrite particle still remained as an euhedral crystal.

Figure 4.25 SEM micrograph of a pentlandite particle after 15 minutes of leaching

(38)

In Figures 4.27 and 4.28 the leach residue is presented from an experiment performed at temperatures of 90°C and 120°C respectively.

Figure 4.27 SEM micrograph of the leach residue from an experiment performed at a temperature of 90° C

Sulphur drops

Figure 4.28 SEM micrograph of the leach residue from an experiment performed at a temperature of 120° C

(39)

A reasonable amount of chalcopyrite had not been completely leached at a temperature of 90°C, as can be seen in Figure 4.27. Figure 4.28 illustrate that large drops of molten sulphur wetted the mineral surface at a temperature of 120°C, the melting point of sulphur. This phenomenal was expected, but did not negatively influence the reaction kinetics, in fact the reaction kinetics increased at a temperature of 120°C.

In Figures 4.29 and 4.30 leached particles, which were not milled finely to the required particle size of - 1 Oj..Lm, are shown. Figure 4.29 represents a partially leached pentlandite particle and Figure 4.30 represents a partially leached chalcopyrite particle found in the leach residues for an experiment performed at ll0°C.

(40)

Figure 4.30 SEM micrograph of a partially leached chalcopyrite particle

It is clear that the larger particles were not completely leached even after a leaching period of 150 minutes, emphasising the importance of grain sizes when considering the leaching of sulphide concentrates.

leached

(41)

The advantage of fine milling is quite evident from Figures 4.29 and 4.30. Figure 4.31 represents the leach residue from an experiment performed with the as received sample, which did not undergo ultra fine milling, proving evidence that a large amount of metal was partially leached and resulted in low metal recoveries.

4.4 Mossbauer spectroscopy

As mentioned, Mossbauer spectroscopy 1s an attractive tool for studying hydrometallurgical treatment mechanisms of specifically iron-bearing minerals.

A freshly milled ore sample from the Nkomati mine was obtained and a Mossbauer spectrum acquired. From the Mossbauer spectrum, presented in Figure 4.32 and the Mossbauer parameters, presented in Table 4.10, it was clear that the sample consisted of the gangue material (35%) and the massive sulphide ore (65%). The main sulphide minerals identified were pyrrhotite (53%), pentlandite (7%) and pyrite (5%), whilst the gangue minerals consisted mainly of pyroxenes. The Moss bauer parameters, obtained for the gangue minerals (see Table 4.10), are in agreement with those found in literature (Stevens et al., 1998:415).

(42)

..

...

~ I

I

- pyrrhotite

I

- pyrrhotite

I

I I

I

- pyrrhotite -pyroxene

n

-pyrite

n

- pentlandite 99.5

-

.

Figure 4.32 Mossbauer spectrum of the ore sample

Floatation ensured that the gangue minerals were removed, which can clearly be seen in the Mossbauer spectrum of the flotation sample (see Figure 4.33), where the large doublet (IS= 1.06

±

0.02 mms-1, QS = 2.50

±

0.02 mms-1) does not exist in the spectrum

anymore. In addition to the minerals mentioned in the freshly milled sulphide ore, magnetite and chalcopyrite were also identified in the float sample. Due to the low relative abundance ( 4% and 3% respectively) they were not detected in the ore sample where the gangue minerals resulted in a large doublet in the spectrum, shielding the contribution of the minor minerals.

All the minerals, detected in the Moss bauer spectra of the ore and flotation samples, were also described to occur in the massive sulphide ore in an unpublished thesis on the

f

Geology of the area. The pyrite, observed in the ore and float sample was rapidly dissolved and disappeared in the Mossbauer spectra very soon after leaching. The magnetite sextet in the Mossbauer spectrum, found only in the float sample, was also rapidly removed after leaching commenced.

(43)

Table 4.10 The fitted Mossbauer parameters derived from the Mossbauer fits to

the spectra studied at 300K. (The values reported are relative to

a-iron)

Sample IS QS H Relative intensity Assignment

mms-1 mms-1 Tesla (%)

(error±0.02) (error±0.02) (error±2)_

0.63 0.18 28.1 25

0.61 0.15 23.8 18 Pyrrhotite

0.61 0.16 23.5 10'

Ore sample 0.36 0.34

-

7 Pentlandite

0.35 0.52

-

5 Pyrite 1.06 2.50

-

35 Pyroxene 0.65 0.10 28.4 32 0.63 0.14 24.0 25 Pyrrhotite 0.61 0.16 21.5 15 0.37 0.37

-

10 Pentlandite Float sample 0.37 0.53

-

11 Pyrite -0.15 0.70 46.9 4 Magnetite 0.36 0.00 35.5 3 Chalcopyrite 0.64 0.18 29.2 18 0.62 0.15 23.9 11 Pyrrhotite 0.61 0.17 21.2 4

Leach product 0.36 0.36

-

14 Pentlandite

15min

0.36 -0.01 33.5 8 Chalcopyrite

0.39 0.88

-

45 Goethite

0.62 0.18 30.3 6 Pyrrhotite

Leach product 0.35 -0.02 33.5 11 Chalcopyrite

30min 0.32 0.86

-

83 Goethite Leach product 50 min 0.44 0.89

-

100 Goethite Leach product 90min 0.44 0.89 - 100 Goethite Leach product 150 min 0.45 0.89

-

100 Goethite

(44)

...

_)':

...

c: 0

i

_...

!

~ 1-. - - - - 1-. 1-. - - - 1-. - - - - r - - - - r - - - - , -magnetite - chalcopyrite I - pyrrhotite I I I I I I - pyrrhotite - pyrrhotite -pyrite - pentlandite I I I I I I

n

l • • I I 1.01).(1: 1 t t r \1 1" 1

.

99.5 1. •

Figure 4.33 Mossbauer spectrum of the flotation concentrate

Pyrrhotite, the most abundant iron sulphide mineral in the ore sample studied, gave rise to three sextets in the Mossbauer spectra (see Figures 4.32, 4.33 and 4.34). The Mossbauer parameters (see Table 4.10) are in agreement with those reported in literature (Stevens et al., 1998:415). In the resulting 30 minutes leach residue, only the strongest

of the three sextets remained, due to the fact that this mineral had almost been completely dissolved and the bulk of the residue consisted of goethite (IS= 0.44

±

0.02 mms-1, QS =

(45)

..

)C

...

s

...

=

...

I

...

-

.

I I I I

n

- chalcopyrite - pyrrhotite I I - pyrrhotite I I - pyrrhotite -goethite - pentlandite

Figure 4.34 Mossbauer spectrum of the leach product after 15 minutes of leaching

- chalcopyrite - pyrrhotite -goethite

...

_~ "" 98

..

- o. 1 0

(46)

The main nickel and cobalt bearing mineral pentlandite, gave rise to the formation of a doublet (IS = 0.36

±

0.02 mms-I, QS = 0.35

±

0.02 mms-1), with parameters consistent

with data given in The Mossbauer Minerals Handbook (Stevens eta!., 1998:389). The pentlandite was almost completely dissolved after 30 minutes of leaching, as can be seen from the disappearance of this doublet in the spectra (see Figures 4.32, 4.33, 4.34 and 4.35). This is also in agreement with the leaching results, which indicated that under optimum conditions most (7 5%) of the nickel reported in the leach fraction after 30 minutes of leaching.

Chalcopyrite, the copper bearing mineral, was characterised by means of a Mossbauer experiment in a roasting process by Banduopadhyay eta!., (1999:973) and gave rise to a sextet in the Mossbauer spectrum with an isomer shift of between IS= 0.24

±

0.02 mms-1 and IS = 0.30

±

0.02mms-\ but in the present experiment the value was higher and found to be IS = 0.36

±

0.02 mms-I, still within the limits given in the Mossbauer Mineral Handbook (Stevens eta!., 1998:212). The other Mossbauer parameters (QS = 0.00

±

0.02 mm{1 and Hyperfine Magnetic Field= 33.5

±

2.0 Tesla) are in agreement with values found in the literature (Bandyopadhyay et a!., 1999:973). It is interesting to note that the chalcopyrite was the most difficult mineral to leach as it only disappeared from the leach residue after 50 minutes of leaching, whilst the pentlandite, pyrite, magnetite were already dissolved within the first 15 minutes of leaching (see Figures 4.33, 4.34 and 4.35).

Figures 4.36 and 4.37 represent the Mossbauer spectrum of the leach products after 50 and 150 minutes of leaching respectively. It is clear from these spectrums that primarily goethite remained after 50 minutes of leaching. The values of the isomer shift for goethite varied between (IS = 0.32

±

0.02 mms-1 and IS = 0.45

±

0.02 mms-1) and this

can be ascribed to the fact of the presence of impurities, incomplete leaching, or because of very finely formed goethite found in the experiments, resulting in a variation of the isomer shift parameter. Ideally, goethite should be investigated at low temperatures ( 4K) as well, but this falls beyond the scope of this experiment.

(47)

...

!( w

6 ...

Ul 1ft

...

t

~

1-9? -l.O.IJ

n

-goethite rtl I Ill I l i t Ill I I l l I t ' t t l I 1111 t I l t i J l u t • , u t l l I I l t l l l l .. l l l I t'a• I nra 111 11\ll I ••- I I I

uelacJ.tY buot/sl J.O.IJ

Figure 4.36 Mossbauer spectrum of the leach product after 50 minutes of leaching

...

:< w

6 ...

~

...

!

e

1-n

-goethite I Ill Itt•,•'' I 100 ... I t / I I I I , l i t I l j l l l l I l l . : . · ... I I ''" ''''" •' ,•.,• t l I I t I I I I l l I 1 1 11 11 111t ti 111111 1 1 I I /•'

..

98 ' 0 I -1.o.o Uelac:~.ty tttl'l/sl ~, 1 It I at I I

...

I ( U t 1 I I 10.0

(48)

Finally it should be noted that the intermediate leaching product, iron sulphate (FeS04)

with Mossbauer parameters of IS

=

1.30

±

0.01 mms-1 and QS

=

2.90

±

0.01 mms-1 (Fajardo et al., 1999: 133) was not observed in the present investigation. This is most probably due to the fact that transformation occurred at a continuous rate and the sulphate might have already altered to the goethite in the residue and thus being undetectable in the Mossbauer spectra.