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The chemistry of uranium. Part XXVIII. The development of a combined gold and uranium leach of randfontein ore
Hydrometallurgy, 6 (1981) 203-218 Elsevier Scientific Publishing Company, Amsterdam - Printed in The Netherlands
203
THE CHEMISTRY OF URANIUM. PART XXVIII. THE DEVELOPMENT OF A COMBINED GOLD AND URANIUM LEACH OF RANDFONTEIN ORE
J.G.H. DU PREEZ*, Uranium Chemistry Africa)
D.C. MORRIS, C.P.J. VAN VUUREN Research
Unit, University
of Port Elizabeth,
Port Elizabeth
(South
P. HENDRIKS and M. OERTELL Johannesburg
Consolidated
Znvestment
Co., Ltd., Johannesburg
(South Africa)
(Received July 18th, 1979; accepted in revised form July lOth, 1980)
ABSTRACT Du Preez, J.G.H., Morris, D.C., Van Vuuren, C.P.J., Hendriks, P. and Oertell, M., 1981. The chemistry of uranium. Part XXVIII. The development of a combined gold and uranium leach of Randfontein ore. Hydrometallurgy, 6 : 203-218. The extraction of uranium from Randfontein ore has been studied at a temperature of 110°C and at different oxygen pressures, using sodium carbonate and sodium bicarbonate. The pressure leach was followed by cyanidation of the ore to extract the gold. The results show that 90% of the uranium and 98% of the gold can be extracted by this technique but the consumption of sodium carbonate, at 65 kg ton-‘, is too high for the process to be economically competitive. The use of ammonium carbonate was also investigated and gave similar extraction values with an apparently lower reagent consumption. In an attempt to reduce the carbonate consumption, due to the reaction with pyrite, a flotation step was included in the process, the flotation concentrate being roasted to remove the sulphur. The roasted concentrate was then subjected to pressure leaching with both sodium carbonate/bicarbonate and ammonium carbonate solutions followed by cyanidation of the residue. This technique reduced the consumption of carbonate but the extractions of gold and uranium were also considerably reduced. INTRODUCTION
It has previously been shown by Du Preez et al. [l] that a combined leach of uranium dioxide and gold is possible in a solution containing 4% Na&03, 1.5% NaHC03 and 0.025% KCN. These experiments were carried out on pure chemicals and in order to investigate the process further it was necessary to apply it to an actual ore. This process has previously been applied to Witwatersrand ores, at atmospheric pressure by Laxen et al. [2] who found that the extraction of uranium was low in many cases and that the consumption of sodium carbonate was *To whom correspondence should be addressed.
0304-386X/81/0000-0000/$02.50
0 1981 Elsevier Scientific Publishing Company
204 too high for the process to be economically competitive. They managed to reduce the carbonate consumption by flotation but the construction of a pilot plant was thought to be too expensive and so the project was discontinued. It was therefore decided that future studies should be carried out at increased pressures. The present work involves the pressure leaching of Randfontein ore with sodium carbonate and bicarbonate solutions followed by the addition of cyanide to extract the gold. The possibility of using ammonium carbonate in place of sodium carbonate and bicarbonate has also been investigated because of cost and availability. An attempt was made to reduce the consumption of carbonate, due to reaction with pyrite, by flotation followed by subsequent roasting of the concentrate prior to leaching. The sodium carbonate process involves the oxidation of uranium(IV) to uranium(VI) using oxygen. Uranium(VI) is readily soluble in carbonate solutions as the tricarbonato species [3]. The presence of bicarbonate is neceso sary to prevent the pH of the system from rising to so high a value that precipitation of sodium diuranate occurs. The overall reaction may be presented as follows: U 3 0 8 -I- 1/~02 4- 3CO32- 4- 6HCCY3 -+ 3 U 0 2 ( C O 3 ) ~ - 4- 3H20
(1)
The use of pressure has two advantages. (i) It allows the use of elevated temperatures above that of the atmospheric boiling point of water, which increases the rate of dissolution of uranium dioxide [1,4,5] and pitchblende [6] and thus decreases the residence time, and (ii) it enables the concentration of oxygen, in solution, to be increased which, because the rate of dissolution is proportional to the square root of the oxygen partial pressure [4--6] also decreases the residence time. The carbonate leaching of uranium ores has previously been applied industrially many times [7--9]. The temperature chosen for this study was 110°C since it has been shown that maximum extractions are obtained at this temperature [ 10]. The separation of the uranium species from the pregnant solution is effected by the addition of sodium hydroxide to neutralise the excess sodium bicarbonate and hence raise the pH to a value where sodium diuranate will precipitate e.g. NaHCO3 + NaOH -~ Na2CO3 + H20
(2)
2Na4UO2(CO3)3 + 6NaOH -~ Na2U207 + 6Na~CO3 + 3H~O
(3)
Any excess sodium hydroxide can be converted back into sodium carbonate and bicarbonate by contacting the solution with carbon dioxide e.g. 2NaOH + CO2 -+ Na2CO3 + H20
(4)
Na2CO3 + CO~ + H20 -+ 2NaHCO3
(5)
205 The regenerated solution can then be recycled to the leaching step. The chemistry behind the use of a m m o n i u m carbonate is similar to the above except t h a t a m m o n i u m bicarbonate is n o t needed since the ammonium hydroxide produced by reaction (6), 9(NH4)2CO3 + U308
+
1/202 -b 3H20 -~ 3(NH4)4UO2(CO3)3 + 6NH4OH
(6)
will n o t raise the pH sufficiently high to precipitate a m m o n i u m diuranate [11l. The uranium may be recoveredby steam stripping. This drives off ammonia and carbon dioxide, whichmay be collectedin scrubbing towers and precipitates a hydrate of uraniumtrioxide [11]
(NH4)4UO2(CO3)3 A_H 4NH3 + 3CO2 + UO32H20
(7)
It is also possible that ammoniacould be passed through the solution to raise the pH thereby precipitating ammoniumdiuranate in an analogous manner to reaction (3). EXPERIMENTAL
Apparatus and reagents The pressure leaching was carried out in a stainless steel autoclave that was fitted with an oxygen line which had an outlet that could be used for sampling. The leach mixture was agitated by an impeller driven by a magnetic drive. The fittings were enclosed in a set of baffles to prevent the mixture from swirling. The autoclave was heated using a propane gas burner controlled by two thermocouples inserted in a well in the autoclave. The burner was equipped with an infrared detector so t h a t in the event of the flame being extinguished the gas supply would automatically be cut off. At the end of an experiment the autoclave and contents were cooled by passing cold water down the outside of the autoclave. The leach solution for the sodium carbonate experiments consisted of 4% Na2CO3 and 1.5% NaHCO3. Originally the leach mixture also contained 0.025% KCN but after it was discovered t h a t this was oxidised to cyanate the addition of potassium cyanide was discontinued. The solution for the a m m o n i u m carbonate leach contained 150 g dm -3 commercial a m m o n i u m carbonate, NH4COONH2NH4HCO3, which contains varying amounts of a m m o n i u m carbonate and bicarbonate. All the solutions were assayed before use. The solutions for the gold dissolution were the same as described above except that they also contained 0.1% KCN.
Leaching procedu re 2 kg of the ore sample and 4 dm 3 of sodium carbonate leach solution or
206
2.5 kg ore and 2.5 dm 3 of a m m o n i u m carbonate leach solution were placed in the autoclave, the lid bolted and agitation started. Oxygen was charged into the autoclave, after flushing, to the desired pressure and the gas burner lit. About 20 minutes were normally required for the autoclave to reach the desired temperature although the experiments were timed from the m o m e n t the burner was lit. The experiments were carried out at partial pressure of 0.1, 0.2, 0.3, and 0.4 MPa of oxygen although in practice the pressures were approximately 0.15 MPa higher, for the sodium carbonate experiments, owing to the water vapour pressure at l l 0 ° C . In the a m m o n i u m carbonate experiments the total pressures were about 0.8 MPa higher due to the thermal decomposition of a m m o n i u m carbonate. All the experiments were performed in duplicate to check on the reproducibility of results.
Cyanidation experiments The cyanidation of the residue from uranium dissolution was carried out in polythene bottles, open to the atmosphere, placed on a set of rollers. The action of the rollers ensured adequate agitation of the contents. The gold content of the ore was determined from the total of the residue and filtrate values and the percentage extraction determined from this figure*
Flotation 400 kg of Randfontein ore was floated in a Denver cell in 20 kg batches. The pulp (35% solids) was adjusted to pH5 using sulphuric acid and 0.1 kg ton -~ of Promotor 404, 0.01 kg ton -~ of Frother 71 and 0.1 kg ton -~ CuSO4 were added to the pulp and air passed through the mixture. The resulting froth was collected, filtered and dried.
Roasting Initially the roasting was carried out in a muffle furnace at 600°C. 500 g of the concentrate was spread thinly on a silica tray and placed in a furnace, at 600°C, for 6 hours. The concentrate was rabbled frequently to ensure adequate contact with the air. This method, however, proved to be slow and it was difficult to reduce the sulphur content to below 2%. An improved method involved roasting in a rotating tube furnace. The concentrate was fed into the tube using a vibrating feed unit. The tube, at a temperature of 600°C, was rotated at 3 rpm. The end of the tube was linked to a fume cupboard whose fans ensured an adequate supply of air. It was found * T h e h e a d value o f the ore was d e t e r m i n e d by taking t h e average o f several samples. When the h e a d value was used t o d e t e r m i n e p e r c e n t a g e e x t r a c t i o n t h e n in several cases values in excess o f 100 p e r c e n t were o b t a i n e d , due to d i f f e r e n c e s in samples.
207 necessary to pass th e sample through the t ube three times t o achieve p r o p e r combustion. Residence time was appr oxi m at el y 30 minutes. RESULTS Table 1 shows the size analysis for a dry sample o f final mill pulp from R a n d f o n t e i n Estates Gold Mine containing 137 g/ t on U308. TABLE 1 Size analysis of Randfontein ore Size (BSS mesh) -48 -65 -100 -150 -200 -325 -200
+ 48 + 65 + 100 + 150 + 200 + 325
%
%
1.7 4.2 7.9 10.8 17.3 15.7 42.4 58.1
1.7 4.2 7.9 10.8 17.2 15.1 43.1 58.2
Effect of pressure on U308 dissolution The variation o f U308 dissolution with time, calculated from the residue values, at various o x y g e n pressures is shown in Fig. 1. The results obtained show t h a t a b o u t 90% of the UaO8 can be extracted using an alkaline c a r bonat e pressure leach. It can be seen f r om Fig. 1 that the curves do n o t go through the origin. This is p r o b ab ly because zero time was t aken when the gas burner was ignited so that little or no reaction occurred during about the first quarter of an hour, while the autoclave was heating up. Pressure has little e f f e c t on the overall ext ract i on of U308 in the range studied. An increase in pressure, o f course, will increase the initial rate o f dissolution o f U308 b u t this effect does not show up in Fig. 1 except for 0.4 MPa 02.
Sodium carbonate and bicarbonate balances One o f the major drawbacks o f the carbonate leach is t h a t the sodium carb o n a t e will react with any pyrite in the ore according to the equation 2FeS2 + 702 + 8Na2CO3 + 6H 20 -~ 2Fe(OH)2 + 4Na2SO4 + 8NaHCOa
(8)
to p r o d u c e sodium bicarbonate. Typical variations o f sodium bicarbonate p r o d u c t i o n and sodium carbonate c o n s u m p t i o n with time are shown in Fig. 2.
"~OOHUN v 'OO'UN ® "rO gdlAI ~'0 $o oanssoad ~ ~v s~lnsaa IUa!d.~I, -otu!~ ql!,~ oSuuoqxea!q oan!pos t o u o ! ~ a n p o a d pu~ a ~ u o q a ~ a tun!pos t o u o ! $ d m n s u o a ~to uo!~e!xe^ oq,L "7. "~8!el (q) ~LU!i I
I
17
9
I
0
] I I
Q
j
i
o c 3
o~ Q
I
Z
I
Og
~
"~,:IIAI ~"0 • ~dIAI {~'0 ~ 'MIAI ~ ' 0 v edIAI T'O ~ "eua!~ q0,!~ uo!~nioss!p ~O~fl ~o u o ! ~ ! a e ^ oq:{ uo aanssoad $o ~}ao.~$o aq.L "I "g!~ (q) ~
y-, 9
--
!l
~I
I
- -
~o [
/ I
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/
i
i
o~
# m N
i ~09
n --~. ~0
I
gong
209
Pressure had little, if any, effect on these values. The consumption of sodium carbonate is high, being in the region of 65 kg t o n -1. This consumption is an order o f magnitude higher than would be necessary to make the process economically competitive with the present m e t h o d of uranium and gold recovery. The bicarbonate production reaches a maximum and then decreases with time. It is thought that this is due to the thermal decomposition of sodium bicarbonate giving sodium carbonate: 2NaHCO3 ~
Na~CO3 + CO2 + H20
(9)
This idea is partially confirmed when a mass balance is attempted. The values of the sodium carbonate consumptions which are calculated from the loss of sulphur, according to reaction (8), are much higher than the observed values indicating that sodium carbonate is being produced from some reaction. If reaction (9) is included on the mass balance then a better agreement is obtained although this is far from good. See Fig. 3 where the values of carbonate consumption calculated using reaction (9) are designated as "theoretical*". I
I
I
I 2
I 4
I 6
10C
~ 8c
Q.
E
6C
E 0 (,3
o" ij
40
20
Time
(h)
Fig. 3. C o m p a r i s o n o f t h e o r e t i c a l a n d e x p e r i m e n t a l v a r i a t i o n s o f s o d i u m c a r b o n a t e cons u m p t i o n w i t h time. T y p i c a l results a t a pressure o f 0.2 MPa 02. e E x p e r i m e n t a l ~ T h e o retical • T h e o r e t i c a l * .
210 Although the agreement between the theoretical* and the observed values is far from good, the shapes of the two curves are similar giving a plateau in each case in sharp distinction to t h a t obtained for the theoretical values determined from the sulphur consumptions alone.
Variation o f pH with time The variation of pH with time for a typical leach at a pressure of 0.2 MPa 02 is shown in Fig. 4. The decrease in pH with time is due to the production of sodium bicarbonate as described in reaction (8). The variation of pH with time, as shown in Fig. 4, does n o t show the conversion of sodium bicarbonate into sodium carbonate which should result in an increase in pH towards the end of the leach.
I0.5~ I0.0~
9.oL L i
8.5[~ /
C)
~
~
__
I ,4
2
Time
_
_
~ 6
(h)
Fig. 4. The variation of pH with time. Typical results at a pressure of 0,2 MPa 02.
The variation of sodium sulphate production with time One of the products of reaction (8) is sodium sulphate. These values together with the theoretical values calculated from the observed sulphur consumptions are shown in Fig. 5. The agreement between the calculated and the experimental values was found to be good in the majority of cases.
211 I
I
I
i
8o
Q
g g 6o a-
o" " 40 Z
20
0
I 2
I 4 Time
I 6
(h)
Fig. 5. Comparison of the theoretical and experimental variation of sodium sulphate production with time. Typical results at a pressure of 0.2 MPa 02. o Experimental ~ Theoretical.
Gold dissolution The results of the gold leach experiments are given in Table 2. The percentage extraction of gold compares favourably with that found in cyanidation tests carried out at higher pH values but the consumption of potassium cyanide is very high, comparatively speaking -- 0.77 kg ton -~ compared to 0.22 kg ton -~ found in previous tests [14].
TABLE 2 The percentage gold extraction and potassium cyanide consumption obtained from the cyanidation of sodium carbonate pressure leach residues. N.B. Metric tons used Au (Head) (g to~ -1)
Au (Residue) (g ton "1)
KCN (kg ton -~)
Au dissolution (%)
10.42 12.15
0.22 0.25
0.70 0.84
97.9 97.9
212
A m m o n i u m carbonate pressure leach This was carried o ut at a pressure of 0.4 MPa 02 for a period of 6 hours. T h e extraction o f uranium was f ound to be 88.7%. It is difficult to compare the consumptions o f sodium and a m m o n i u m carbonate. The observed cons um pt i on of a m m o n i u m carbonate is 31.7 kg t o n -1 i.e. approximately 50% of t hat of sodium carbonate. Analysis, however, revealed th at no a m m o n i u m bicarbonate was being p r o d u c e d compared to 50 kg ton-~ o f sodium bicarbonate. Since it is reasonable to assume t hat amm o n i u m bicarbonate is being pr oduced by a reaction analogous to reaction (8) then this is probably thermally d e c o m p o s e d to a m m o n i u m carbonate. If this is the case then the true c o n s u m p t i o n of a m m o n i u m carbonate due to (8) must be greater than 31.7 kg t on -I. One i m p o r t a n t difference between the a m m o n i u m and sodium carbonate systems is that the for m er does n o t react with silica in the ore. This was noticeable in the difference between the filtration rates of the two residues. The residue f r o m the a m m o n i u m carbonate leach filtered about five times faster than that from the sodium carbonate leach.
Cyanidation o f the residue from an ammonium carbonate leach The results from t he gold extraction experiments are given in Table 3 and show that the gold extraction is complete after 28 hours and is comparable to t h a t f o u n d in the sodium carbonate system. The consum pt i on of potassium cyanide is greater than that found using sodium carbonate but the pH was lower in these experiments which would be e x p e c t e d t o lead to an increased loss. It is probable th a t a combination of the low pH of the system and the constant agitation in a vessel open to the atmosphere is the cause o f the high cyanide consumption. TABLE3 Results of the cyanidation of the residue from an ammonium carbonate leach Time (h)
Au (head) (g ton -~)
Au (residue) (g ton -~)
KCN (kg ton -~)
Au dissolution (%)
22 28 48
12.31 10.48 12.14
0.54 0.19 0.26
0.88 0.91 0.98
95.6 98.2 97.9
LEACHING OF CONCENTRATES
Flotation The flotation data are given in Table 4.
213 TABLE 4 Flotation data
Wt (kg) Separation (%) S (%)
Concentrate
Tails
Total
20.15 5
380.25 95
400.40 --
Separation (%)
26.07 57.1
1.04 42.9
--
Au (g ton "1) Separation (%)
167.35 79.0
2.51 21.0
11.37 --
U~Os (g ton -~)
550
Separation ( % )
20.2
117
2.30
137
79.8
--
The separation of the pyr i t e was poor, the tails containing over 1% sulphur The aim o f th e flotation had b e e n t o reduce the sulphur in the tails to a b o u t 0.2% and t h e n to reduce t he sulphur c o n t e n t of the c o n c e n t r a t e to a similar value by roasting. Despite the p o o r separation it was decided t hat the c o n c e n t r a t e would suffice f o r th e roasting experiments. A f u r t h e r small scale flotation of the tails showed t hat a c o n c e n t r a t e could be o b tain ed which would be 10% by mass containing 94.5% of the available gold, 91.3% o f the sulphur and 30.7% o f the uranium. The sulphur c o n t e n t o f the tails would be a b o u t 0.2%.
Roasting 9 kg o f calcine was obtained, half f r o m the muffle furnace and half f r o m the rotating tube furnace. T he sulphur grades o f the samples are given in Table 5. T h e samples were used for a m m o n i u m and sodium carbonate pressure leaching followed by cyanidation o f the residues.
TABLE 5 C o m p o s i t i o n and yield of calcine Sample
Wt (kg)
Roasting method
S (%)
1 2
2.25 2.25
Muffle furnace Muffle furnace
2.74 2.25
3
4.5
Rotating tube furnace
0.88
214
Ammonium carbonate leaching The calcines f r om the muffle furnace were subjected to an a m m o n i u m carb o n ate pressureleach at a pressure of 0.4 MPa O~ and a t e m p e r a t u r e of l l 0 ° C . The results are given in Table 6. The extraction of uranium was p o o r in b o t h experiments. The roasting procedure should n o t make the uranium in the ore refractory. If uranium
dioxide is heated in air at 600°C U3Os is usually formed although U O 3 is the thermodynamically stable form ([12], p. 312). It is possible that if uranium VI is formed then this could react with basic constituents in the ore giving uranates which are difficult to dissolve in carbonate solutions ([9], p. 51). The ore from Randfontein, however, contains 2 % sulphur as pyrite which makes it acidic and so the formation of uranates is doubtful in this case. TABLE 6 The results of the ammonium carbonate pressure leach of the calcines from the muffle furnace (pressure 0.4 MPa O2 ; T = ll0°C) Sample 1 2
UsOs (Head)
U30 s (Residue)
U30 s dissolution
(g ton -I)
(g ton-I)
(%)
564 574
148 165
71.8 71.3
Electron micr o pr obe studies have revealed the presence of the refractory uranium silicate in the calcine and brannerite has also been det ect ed whilst the presence o f uraninite has not been confirmed. It is possible that the flotation process has resulted in an upgrading of the refractory uranium minerals in the co n cen tr ate although it is difficult to see why. A microscopic examination of the c o n c e n t r a t e did not confirm t h e above possibility as the uranium was present mainly as uraninite associated with carbon. It has been suggested [13] t hat the reason the uraninite is not detected in the calcine is because it has been finally dispersed by virtue of the burning and coking o f the carbon during the roasting process and is therefore n o t observed even at extremely high magnification. It is difficult to say f r om the available evidence why the extraction of uranium is so low. The co n s u mp tion of a m m o n i u m carbonate and bicarbonate together with pH and time are given in Table 7. The results show t hat the c o n s u m p t i o n o f a m m o n i u m carbonate is very high. It would be e xpe c t e d t h a t the c o n s u m p t i o n would be large since the sulphur contents o f the calcines were high. Also it can be seen t hat a m m o n i u m bicarbonate is being consumed. It is very difficult to give accurate consumptions o f a m m o n i u m carbonate and bicarbonate because of the volatility of solutions of the salt. The solutions, even at r o o m temperature, decompose
215 TABLE 7
Reagent consumption d a t a f o r t h e a m m o n i u m c a r b o n a t e pressure l e a c h o f the calcine from the muffle furnace Sample
Time (h)
pH
1
0 6 0 6
9.20 8.50 9.20 8.60
2
(NH4) ~ CO s (kg t o n -~) -85.8 -100.8
NH4HCO s (kg t o n "~ ) -0.75 -11.8
all the time with evolution of ammonia and carbon dioxide which, in part, could account for the observed high consumptions.
Cyanidation o f the calcine leaching residue The results of the gold leach are given in Table 8. The dissolution of gold was found to be a b o u t 10% lower than is normally found and might be attributed to the presence of iron oxide which could conceivably result in a micro-electrochemical coating of the gold grains with iron oxide leading to an increased dissolution time [ 1 3 ] . This was only an hypothesis however, as no evidence of this type o f coating could be found in electronmicroprobe studies. TABLE 8 D a t a f r o m t h e c y a n i d a t i o n o f t h e residues f r o m t h e a m m o n i u m c a r b o n a t e pressure leach Sample
1 2
Au (Head)
A u (Tails)
Au Dissolution
KCN
(g ton -~)
(g ton -1)
(%)
(kg ton -1)
195.4 179.6
27.2 18.8
86.1 89.1
1.05 0.64
It was thought that the low dissolution of gold could be due to the low pH of the system which would decrease the concentration of " f r e e " cyanide present in the solution. To check on this the experiments were repeated with the addition of sufficient ammonium hydroxide to raise the pH to 10. The dissolution of gold was found to be lower, in this case, however, being : a b o u t 74.5%. One interesting fact observed was that the uranium extractions increased ~during the cyanidation experiments to values of 78.4% for sample 1 and 75.5% for sample 2. It would seem from this that an increase in leaching time would result in an increase in uranium extraction.
216
Sodium carbonate leaching The results of the sodium carbonate pressure leach at a pressure of 0.4 MPa O: and a temperature of l l 0 ° C are shown in Fig. 6. The extraction of uranium is low being approximately the same as found when using a m m o n i u m carbonate. From Fig. 6 it can be seen that the extraction process is still proceeding but it would seem that the time needed to reach 90% extraction would be too long for a pressure leach to be feasible. I
l
T----
I 2
I 4
I 6
? E ©
6@t_ × Lxl
(Z~ 40~-
2o~-
i o
Time(h)
Fig. 6. T h e v a r i a t i o n o f U308 e x t r a c t i o n w i t h time. S o d i u m c a r b o n a t e pressure leach o f t h e calcine.
The consumption of sodium carbonate together with theoretical values are shown in Fig. 7. The consumption of sodium carbonate is about 40% lower than was found in the previous experiments using the complete ore indicating the success of flotation and roasting in this respect. The consumption of sodium carbonate is still too high, however, for the process to be economically competitive but could be decreased if the sulphur content of the calcine was reduced further.
Cyanidation of calcine leach residues The results of the cyanidation experiments are given in Table 9. The extraction of gold was very low, being approximately the same as when using a m m o n i u m carbonate at high pH.
217
4O
#-
v 30 & E 3 020 U
o,~ 10
I 2
I 4
I 6
T i m e (h)
Fig. 7. Comparison of theoretical and experimental variations of sodium carbonate consumption with time. ® Experimental a Theoretical. TABLE 9 Cyanidation data pH
Au (Head) (g ton -~)
Au (Residue) (g ton -~)
KCN (kg ton -~)
Au dissolution (%)
10.30 10.30
228.35 209.2
63.3 45.6
0.35 0.33
72.3 78.2
CONCLUSION T he dissolution o f gold and uranium using a sodium carbonate and bicarb o n a t e solution u n d e r pressure followed by t h e addition o f potassium cyanide was successful with 97.9% o f the gold and a p p r o x i m a t e l y 90% o f the uranium being extracted. The process, however, is n o t economically competitive owing to the high c o n s u m p t i o n o f sodium car bona t e which was in the region o f 65 kg t o n - L T he use o f a m m o n i u m c a r bonat e instead o f sodium carbonate and bicarbonate gave similar extractions with a lower c o n s u m p t i o n of carbonat e although in this case n o a m m o n i u m bicarbonate was p r o d u c e d in contrast to the sodium carbonate system.
218
In both sets of eyanidation experiments the consumption of potassium cyanide was greater than in the conventional process. This is probably due to a combination of the low pH and the agitation of the solutions. The inclusion of a flotation step followed by roasting of the flotation concentrate into the process decreased the consumption of sodium carbonate to a b o u t 35 kg t o n -~ b u t at the same time the extractions of gold and uranium were considerably reduced. ACKNOWLEDGEMENTS
The authors wish to thank Johannesburg Consolidated Investment Company Limited for funds and also for the use of equipment and services at their Minerals Processing Research Laboratory. REFERENCES 1 D u Preez, J.G.H., Van Vuuren, C.P.J. and Morris, D.C. The chemistry of uranium part XXVI: alkaline dissolution of A u and/or U O 2 powders, Hydrometall. 6 (I+2) (1980) 147--158. 2 Laxen, P.A., Pammenter, R.V., Bayne, D.L.G., Litvin, S.S. and Vorster, J.H. South African Atomic Energy Board, Extractive Metallurgy Division Reports, Project No. C30/65 A, 1967. 3 Forward, F.A., Halpern, J. and Peters, E., Studies in the carbonate leaching of uranium ores, Trans. Can. Inst. Min. Metall. LVI (1953) 354--360. 4 Pearson, R.L. and Wadsworth, M.C., A kinetic study of the dissolution of UO 2 in carbonate solution, Trans. Metall. Soc. A.I.M.E. 212 (1958) 294--300. 5 Schortmann, W.E. and De Sesa, M.A., Kinetics of the dissolution of uranium dioxide in carbonate--bicarbonate solutions. Proceedings of the 2nd International Conference on the Peaceful uses of A t o m i c Energy. Geneva--United Nations, 3 (1958) 333--341. 6 Peters, E. and Halpern, J., Studies in the carbonate leaching o f uranium ores. II. Kinetics of the dissolution of pitchblende, Trans. Can. Inst. Min. Metall., LVI (1953) 360--364. 7 Hannay, R.L., Eldorado Beaverlodge Operation. 5: Metallurgical plant, Can. Min. J., 81 (1960) 113--136. 8 Jones, J.Q. and De Jong, L.C., Sapin uranium ore processing, in Arbiter, N. (Ed.), Milling Methods in the Americas. Gordon and Breach Science Publishers, New York, 1964, pp. 283--312. 9 Merrit, R.C., The Extractive Metallurgy o f Uranium, Colorado School of Mines Research Institute, Denver, 1971, pp. 83--102 and chs. 15, 16. 10 Beverley, R.G., Griffith, A.W. and Millsap, W.A., Atmospheric vs. pressure leaching of uranium ores, J. Met., 9 (1957) 745--751. 11 Langston, B.G., MacDonald~ R.D. and Stephens Jnr., F.M., Ammonium carbonate pressure leaching of uranium ores, J. Met., 9 (1957) 752--756. 12 Katz, J.J. and Rabinowitch, E., The Chemistry of Uranium, Dover Publications Inc., New York, 1961, p. 609. 13 Kinloch, E.D., Mineralogical examination of Randfontein bench scale test products, Report M]77/137., Johannesburg Consolidated Investment Company Limited, Johannesburg, 1977, pp. 4. 14 Loo, P.T. and Gussmann, H.W., Pressure leach of final mill pulp from Randfontein Estates, Report No. 15/75, Johannesburg Consolidated Investment Company L i m i t e d Johannesburg, 1975.
203
THE CHEMISTRY OF URANIUM. PART XXVIII. THE DEVELOPMENT OF A COMBINED GOLD AND URANIUM LEACH OF RANDFONTEIN ORE
J.G.H. DU PREEZ*, Uranium Chemistry Africa)
D.C. MORRIS, C.P.J. VAN VUUREN Research
Unit, University
of Port Elizabeth,
Port Elizabeth
(South
P. HENDRIKS and M. OERTELL Johannesburg
Consolidated
Znvestment
Co., Ltd., Johannesburg
(South Africa)
(Received July 18th, 1979; accepted in revised form July lOth, 1980)
ABSTRACT Du Preez, J.G.H., Morris, D.C., Van Vuuren, C.P.J., Hendriks, P. and Oertell, M., 1981. The chemistry of uranium. Part XXVIII. The development of a combined gold and uranium leach of Randfontein ore. Hydrometallurgy, 6 : 203-218. The extraction of uranium from Randfontein ore has been studied at a temperature of 110°C and at different oxygen pressures, using sodium carbonate and sodium bicarbonate. The pressure leach was followed by cyanidation of the ore to extract the gold. The results show that 90% of the uranium and 98% of the gold can be extracted by this technique but the consumption of sodium carbonate, at 65 kg ton-‘, is too high for the process to be economically competitive. The use of ammonium carbonate was also investigated and gave similar extraction values with an apparently lower reagent consumption. In an attempt to reduce the carbonate consumption, due to the reaction with pyrite, a flotation step was included in the process, the flotation concentrate being roasted to remove the sulphur. The roasted concentrate was then subjected to pressure leaching with both sodium carbonate/bicarbonate and ammonium carbonate solutions followed by cyanidation of the residue. This technique reduced the consumption of carbonate but the extractions of gold and uranium were also considerably reduced. INTRODUCTION
It has previously been shown by Du Preez et al. [l] that a combined leach of uranium dioxide and gold is possible in a solution containing 4% Na&03, 1.5% NaHC03 and 0.025% KCN. These experiments were carried out on pure chemicals and in order to investigate the process further it was necessary to apply it to an actual ore. This process has previously been applied to Witwatersrand ores, at atmospheric pressure by Laxen et al. [2] who found that the extraction of uranium was low in many cases and that the consumption of sodium carbonate was *To whom correspondence should be addressed.
0304-386X/81/0000-0000/$02.50
0 1981 Elsevier Scientific Publishing Company
204 too high for the process to be economically competitive. They managed to reduce the carbonate consumption by flotation but the construction of a pilot plant was thought to be too expensive and so the project was discontinued. It was therefore decided that future studies should be carried out at increased pressures. The present work involves the pressure leaching of Randfontein ore with sodium carbonate and bicarbonate solutions followed by the addition of cyanide to extract the gold. The possibility of using ammonium carbonate in place of sodium carbonate and bicarbonate has also been investigated because of cost and availability. An attempt was made to reduce the consumption of carbonate, due to reaction with pyrite, by flotation followed by subsequent roasting of the concentrate prior to leaching. The sodium carbonate process involves the oxidation of uranium(IV) to uranium(VI) using oxygen. Uranium(VI) is readily soluble in carbonate solutions as the tricarbonato species [3]. The presence of bicarbonate is neceso sary to prevent the pH of the system from rising to so high a value that precipitation of sodium diuranate occurs. The overall reaction may be presented as follows: U 3 0 8 -I- 1/~02 4- 3CO32- 4- 6HCCY3 -+ 3 U 0 2 ( C O 3 ) ~ - 4- 3H20
(1)
The use of pressure has two advantages. (i) It allows the use of elevated temperatures above that of the atmospheric boiling point of water, which increases the rate of dissolution of uranium dioxide [1,4,5] and pitchblende [6] and thus decreases the residence time, and (ii) it enables the concentration of oxygen, in solution, to be increased which, because the rate of dissolution is proportional to the square root of the oxygen partial pressure [4--6] also decreases the residence time. The carbonate leaching of uranium ores has previously been applied industrially many times [7--9]. The temperature chosen for this study was 110°C since it has been shown that maximum extractions are obtained at this temperature [ 10]. The separation of the uranium species from the pregnant solution is effected by the addition of sodium hydroxide to neutralise the excess sodium bicarbonate and hence raise the pH to a value where sodium diuranate will precipitate e.g. NaHCO3 + NaOH -~ Na2CO3 + H20
(2)
2Na4UO2(CO3)3 + 6NaOH -~ Na2U207 + 6Na~CO3 + 3H~O
(3)
Any excess sodium hydroxide can be converted back into sodium carbonate and bicarbonate by contacting the solution with carbon dioxide e.g. 2NaOH + CO2 -+ Na2CO3 + H20
(4)
Na2CO3 + CO~ + H20 -+ 2NaHCO3
(5)
205 The regenerated solution can then be recycled to the leaching step. The chemistry behind the use of a m m o n i u m carbonate is similar to the above except t h a t a m m o n i u m bicarbonate is n o t needed since the ammonium hydroxide produced by reaction (6), 9(NH4)2CO3 + U308
+
1/202 -b 3H20 -~ 3(NH4)4UO2(CO3)3 + 6NH4OH
(6)
will n o t raise the pH sufficiently high to precipitate a m m o n i u m diuranate [11l. The uranium may be recoveredby steam stripping. This drives off ammonia and carbon dioxide, whichmay be collectedin scrubbing towers and precipitates a hydrate of uraniumtrioxide [11]
(NH4)4UO2(CO3)3 A_H 4NH3 + 3CO2 + UO32H20
(7)
It is also possible that ammoniacould be passed through the solution to raise the pH thereby precipitating ammoniumdiuranate in an analogous manner to reaction (3). EXPERIMENTAL
Apparatus and reagents The pressure leaching was carried out in a stainless steel autoclave that was fitted with an oxygen line which had an outlet that could be used for sampling. The leach mixture was agitated by an impeller driven by a magnetic drive. The fittings were enclosed in a set of baffles to prevent the mixture from swirling. The autoclave was heated using a propane gas burner controlled by two thermocouples inserted in a well in the autoclave. The burner was equipped with an infrared detector so t h a t in the event of the flame being extinguished the gas supply would automatically be cut off. At the end of an experiment the autoclave and contents were cooled by passing cold water down the outside of the autoclave. The leach solution for the sodium carbonate experiments consisted of 4% Na2CO3 and 1.5% NaHCO3. Originally the leach mixture also contained 0.025% KCN but after it was discovered t h a t this was oxidised to cyanate the addition of potassium cyanide was discontinued. The solution for the a m m o n i u m carbonate leach contained 150 g dm -3 commercial a m m o n i u m carbonate, NH4COONH2NH4HCO3, which contains varying amounts of a m m o n i u m carbonate and bicarbonate. All the solutions were assayed before use. The solutions for the gold dissolution were the same as described above except that they also contained 0.1% KCN.
Leaching procedu re 2 kg of the ore sample and 4 dm 3 of sodium carbonate leach solution or
206
2.5 kg ore and 2.5 dm 3 of a m m o n i u m carbonate leach solution were placed in the autoclave, the lid bolted and agitation started. Oxygen was charged into the autoclave, after flushing, to the desired pressure and the gas burner lit. About 20 minutes were normally required for the autoclave to reach the desired temperature although the experiments were timed from the m o m e n t the burner was lit. The experiments were carried out at partial pressure of 0.1, 0.2, 0.3, and 0.4 MPa of oxygen although in practice the pressures were approximately 0.15 MPa higher, for the sodium carbonate experiments, owing to the water vapour pressure at l l 0 ° C . In the a m m o n i u m carbonate experiments the total pressures were about 0.8 MPa higher due to the thermal decomposition of a m m o n i u m carbonate. All the experiments were performed in duplicate to check on the reproducibility of results.
Cyanidation experiments The cyanidation of the residue from uranium dissolution was carried out in polythene bottles, open to the atmosphere, placed on a set of rollers. The action of the rollers ensured adequate agitation of the contents. The gold content of the ore was determined from the total of the residue and filtrate values and the percentage extraction determined from this figure*
Flotation 400 kg of Randfontein ore was floated in a Denver cell in 20 kg batches. The pulp (35% solids) was adjusted to pH5 using sulphuric acid and 0.1 kg ton -~ of Promotor 404, 0.01 kg ton -~ of Frother 71 and 0.1 kg ton -~ CuSO4 were added to the pulp and air passed through the mixture. The resulting froth was collected, filtered and dried.
Roasting Initially the roasting was carried out in a muffle furnace at 600°C. 500 g of the concentrate was spread thinly on a silica tray and placed in a furnace, at 600°C, for 6 hours. The concentrate was rabbled frequently to ensure adequate contact with the air. This method, however, proved to be slow and it was difficult to reduce the sulphur content to below 2%. An improved method involved roasting in a rotating tube furnace. The concentrate was fed into the tube using a vibrating feed unit. The tube, at a temperature of 600°C, was rotated at 3 rpm. The end of the tube was linked to a fume cupboard whose fans ensured an adequate supply of air. It was found * T h e h e a d value o f the ore was d e t e r m i n e d by taking t h e average o f several samples. When the h e a d value was used t o d e t e r m i n e p e r c e n t a g e e x t r a c t i o n t h e n in several cases values in excess o f 100 p e r c e n t were o b t a i n e d , due to d i f f e r e n c e s in samples.
207 necessary to pass th e sample through the t ube three times t o achieve p r o p e r combustion. Residence time was appr oxi m at el y 30 minutes. RESULTS Table 1 shows the size analysis for a dry sample o f final mill pulp from R a n d f o n t e i n Estates Gold Mine containing 137 g/ t on U308. TABLE 1 Size analysis of Randfontein ore Size (BSS mesh) -48 -65 -100 -150 -200 -325 -200
+ 48 + 65 + 100 + 150 + 200 + 325
%
%
1.7 4.2 7.9 10.8 17.3 15.7 42.4 58.1
1.7 4.2 7.9 10.8 17.2 15.1 43.1 58.2
Effect of pressure on U308 dissolution The variation o f U308 dissolution with time, calculated from the residue values, at various o x y g e n pressures is shown in Fig. 1. The results obtained show t h a t a b o u t 90% of the UaO8 can be extracted using an alkaline c a r bonat e pressure leach. It can be seen f r om Fig. 1 that the curves do n o t go through the origin. This is p r o b ab ly because zero time was t aken when the gas burner was ignited so that little or no reaction occurred during about the first quarter of an hour, while the autoclave was heating up. Pressure has little e f f e c t on the overall ext ract i on of U308 in the range studied. An increase in pressure, o f course, will increase the initial rate o f dissolution o f U308 b u t this effect does not show up in Fig. 1 except for 0.4 MPa 02.
Sodium carbonate and bicarbonate balances One o f the major drawbacks o f the carbonate leach is t h a t the sodium carb o n a t e will react with any pyrite in the ore according to the equation 2FeS2 + 702 + 8Na2CO3 + 6H 20 -~ 2Fe(OH)2 + 4Na2SO4 + 8NaHCOa
(8)
to p r o d u c e sodium bicarbonate. Typical variations o f sodium bicarbonate p r o d u c t i o n and sodium carbonate c o n s u m p t i o n with time are shown in Fig. 2.
"~OOHUN v 'OO'UN ® "rO gdlAI ~'0 $o oanssoad ~ ~v s~lnsaa IUa!d.~I, -otu!~ ql!,~ oSuuoqxea!q oan!pos t o u o ! ~ a n p o a d pu~ a ~ u o q a ~ a tun!pos t o u o ! $ d m n s u o a ~to uo!~e!xe^ oq,L "7. "~8!el (q) ~LU!i I
I
17
9
I
0
] I I
Q
j
i
o c 3
o~ Q
I
Z
I
Og
~
"~,:IIAI ~"0 • ~dIAI {~'0 ~ 'MIAI ~ ' 0 v edIAI T'O ~ "eua!~ q0,!~ uo!~nioss!p ~O~fl ~o u o ! ~ ! a e ^ oq:{ uo aanssoad $o ~}ao.~$o aq.L "I "g!~ (q) ~
y-, 9
--
!l
~I
I
- -
~o [
/ I
I
/
i
i
o~
# m N
i ~09
n --~. ~0
I
gong
209
Pressure had little, if any, effect on these values. The consumption of sodium carbonate is high, being in the region of 65 kg t o n -1. This consumption is an order o f magnitude higher than would be necessary to make the process economically competitive with the present m e t h o d of uranium and gold recovery. The bicarbonate production reaches a maximum and then decreases with time. It is thought that this is due to the thermal decomposition of sodium bicarbonate giving sodium carbonate: 2NaHCO3 ~
Na~CO3 + CO2 + H20
(9)
This idea is partially confirmed when a mass balance is attempted. The values of the sodium carbonate consumptions which are calculated from the loss of sulphur, according to reaction (8), are much higher than the observed values indicating that sodium carbonate is being produced from some reaction. If reaction (9) is included on the mass balance then a better agreement is obtained although this is far from good. See Fig. 3 where the values of carbonate consumption calculated using reaction (9) are designated as "theoretical*". I
I
I
I 2
I 4
I 6
10C
~ 8c
Q.
E
6C
E 0 (,3
o" ij
40
20
Time
(h)
Fig. 3. C o m p a r i s o n o f t h e o r e t i c a l a n d e x p e r i m e n t a l v a r i a t i o n s o f s o d i u m c a r b o n a t e cons u m p t i o n w i t h time. T y p i c a l results a t a pressure o f 0.2 MPa 02. e E x p e r i m e n t a l ~ T h e o retical • T h e o r e t i c a l * .
210 Although the agreement between the theoretical* and the observed values is far from good, the shapes of the two curves are similar giving a plateau in each case in sharp distinction to t h a t obtained for the theoretical values determined from the sulphur consumptions alone.
Variation o f pH with time The variation of pH with time for a typical leach at a pressure of 0.2 MPa 02 is shown in Fig. 4. The decrease in pH with time is due to the production of sodium bicarbonate as described in reaction (8). The variation of pH with time, as shown in Fig. 4, does n o t show the conversion of sodium bicarbonate into sodium carbonate which should result in an increase in pH towards the end of the leach.
I0.5~ I0.0~
9.oL L i
8.5[~ /
C)
~
~
__
I ,4
2
Time
_
_
~ 6
(h)
Fig. 4. The variation of pH with time. Typical results at a pressure of 0,2 MPa 02.
The variation of sodium sulphate production with time One of the products of reaction (8) is sodium sulphate. These values together with the theoretical values calculated from the observed sulphur consumptions are shown in Fig. 5. The agreement between the calculated and the experimental values was found to be good in the majority of cases.
211 I
I
I
i
8o
Q
g g 6o a-
o" " 40 Z
20
0
I 2
I 4 Time
I 6
(h)
Fig. 5. Comparison of the theoretical and experimental variation of sodium sulphate production with time. Typical results at a pressure of 0.2 MPa 02. o Experimental ~ Theoretical.
Gold dissolution The results of the gold leach experiments are given in Table 2. The percentage extraction of gold compares favourably with that found in cyanidation tests carried out at higher pH values but the consumption of potassium cyanide is very high, comparatively speaking -- 0.77 kg ton -~ compared to 0.22 kg ton -~ found in previous tests [14].
TABLE 2 The percentage gold extraction and potassium cyanide consumption obtained from the cyanidation of sodium carbonate pressure leach residues. N.B. Metric tons used Au (Head) (g to~ -1)
Au (Residue) (g ton "1)
KCN (kg ton -~)
Au dissolution (%)
10.42 12.15
0.22 0.25
0.70 0.84
97.9 97.9
212
A m m o n i u m carbonate pressure leach This was carried o ut at a pressure of 0.4 MPa 02 for a period of 6 hours. T h e extraction o f uranium was f ound to be 88.7%. It is difficult to compare the consumptions o f sodium and a m m o n i u m carbonate. The observed cons um pt i on of a m m o n i u m carbonate is 31.7 kg t o n -1 i.e. approximately 50% of t hat of sodium carbonate. Analysis, however, revealed th at no a m m o n i u m bicarbonate was being p r o d u c e d compared to 50 kg ton-~ o f sodium bicarbonate. Since it is reasonable to assume t hat amm o n i u m bicarbonate is being pr oduced by a reaction analogous to reaction (8) then this is probably thermally d e c o m p o s e d to a m m o n i u m carbonate. If this is the case then the true c o n s u m p t i o n of a m m o n i u m carbonate due to (8) must be greater than 31.7 kg t on -I. One i m p o r t a n t difference between the a m m o n i u m and sodium carbonate systems is that the for m er does n o t react with silica in the ore. This was noticeable in the difference between the filtration rates of the two residues. The residue f r o m the a m m o n i u m carbonate leach filtered about five times faster than that from the sodium carbonate leach.
Cyanidation o f the residue from an ammonium carbonate leach The results from t he gold extraction experiments are given in Table 3 and show that the gold extraction is complete after 28 hours and is comparable to t h a t f o u n d in the sodium carbonate system. The consum pt i on of potassium cyanide is greater than that found using sodium carbonate but the pH was lower in these experiments which would be e x p e c t e d t o lead to an increased loss. It is probable th a t a combination of the low pH of the system and the constant agitation in a vessel open to the atmosphere is the cause o f the high cyanide consumption. TABLE3 Results of the cyanidation of the residue from an ammonium carbonate leach Time (h)
Au (head) (g ton -~)
Au (residue) (g ton -~)
KCN (kg ton -~)
Au dissolution (%)
22 28 48
12.31 10.48 12.14
0.54 0.19 0.26
0.88 0.91 0.98
95.6 98.2 97.9
LEACHING OF CONCENTRATES
Flotation The flotation data are given in Table 4.
213 TABLE 4 Flotation data
Wt (kg) Separation (%) S (%)
Concentrate
Tails
Total
20.15 5
380.25 95
400.40 --
Separation (%)
26.07 57.1
1.04 42.9
--
Au (g ton "1) Separation (%)
167.35 79.0
2.51 21.0
11.37 --
U~Os (g ton -~)
550
Separation ( % )
20.2
117
2.30
137
79.8
--
The separation of the pyr i t e was poor, the tails containing over 1% sulphur The aim o f th e flotation had b e e n t o reduce the sulphur in the tails to a b o u t 0.2% and t h e n to reduce t he sulphur c o n t e n t of the c o n c e n t r a t e to a similar value by roasting. Despite the p o o r separation it was decided t hat the c o n c e n t r a t e would suffice f o r th e roasting experiments. A f u r t h e r small scale flotation of the tails showed t hat a c o n c e n t r a t e could be o b tain ed which would be 10% by mass containing 94.5% of the available gold, 91.3% o f the sulphur and 30.7% o f the uranium. The sulphur c o n t e n t o f the tails would be a b o u t 0.2%.
Roasting 9 kg o f calcine was obtained, half f r o m the muffle furnace and half f r o m the rotating tube furnace. T he sulphur grades o f the samples are given in Table 5. T h e samples were used for a m m o n i u m and sodium carbonate pressure leaching followed by cyanidation o f the residues.
TABLE 5 C o m p o s i t i o n and yield of calcine Sample
Wt (kg)
Roasting method
S (%)
1 2
2.25 2.25
Muffle furnace Muffle furnace
2.74 2.25
3
4.5
Rotating tube furnace
0.88
214
Ammonium carbonate leaching The calcines f r om the muffle furnace were subjected to an a m m o n i u m carb o n ate pressureleach at a pressure of 0.4 MPa O~ and a t e m p e r a t u r e of l l 0 ° C . The results are given in Table 6. The extraction of uranium was p o o r in b o t h experiments. The roasting procedure should n o t make the uranium in the ore refractory. If uranium
dioxide is heated in air at 600°C U3Os is usually formed although U O 3 is the thermodynamically stable form ([12], p. 312). It is possible that if uranium VI is formed then this could react with basic constituents in the ore giving uranates which are difficult to dissolve in carbonate solutions ([9], p. 51). The ore from Randfontein, however, contains 2 % sulphur as pyrite which makes it acidic and so the formation of uranates is doubtful in this case. TABLE 6 The results of the ammonium carbonate pressure leach of the calcines from the muffle furnace (pressure 0.4 MPa O2 ; T = ll0°C) Sample 1 2
UsOs (Head)
U30 s (Residue)
U30 s dissolution
(g ton -I)
(g ton-I)
(%)
564 574
148 165
71.8 71.3
Electron micr o pr obe studies have revealed the presence of the refractory uranium silicate in the calcine and brannerite has also been det ect ed whilst the presence o f uraninite has not been confirmed. It is possible that the flotation process has resulted in an upgrading of the refractory uranium minerals in the co n cen tr ate although it is difficult to see why. A microscopic examination of the c o n c e n t r a t e did not confirm t h e above possibility as the uranium was present mainly as uraninite associated with carbon. It has been suggested [13] t hat the reason the uraninite is not detected in the calcine is because it has been finally dispersed by virtue of the burning and coking o f the carbon during the roasting process and is therefore n o t observed even at extremely high magnification. It is difficult to say f r om the available evidence why the extraction of uranium is so low. The co n s u mp tion of a m m o n i u m carbonate and bicarbonate together with pH and time are given in Table 7. The results show t hat the c o n s u m p t i o n o f a m m o n i u m carbonate is very high. It would be e xpe c t e d t h a t the c o n s u m p t i o n would be large since the sulphur contents o f the calcines were high. Also it can be seen t hat a m m o n i u m bicarbonate is being consumed. It is very difficult to give accurate consumptions o f a m m o n i u m carbonate and bicarbonate because of the volatility of solutions of the salt. The solutions, even at r o o m temperature, decompose
215 TABLE 7
Reagent consumption d a t a f o r t h e a m m o n i u m c a r b o n a t e pressure l e a c h o f the calcine from the muffle furnace Sample
Time (h)
pH
1
0 6 0 6
9.20 8.50 9.20 8.60
2
(NH4) ~ CO s (kg t o n -~) -85.8 -100.8
NH4HCO s (kg t o n "~ ) -0.75 -11.8
all the time with evolution of ammonia and carbon dioxide which, in part, could account for the observed high consumptions.
Cyanidation o f the calcine leaching residue The results of the gold leach are given in Table 8. The dissolution of gold was found to be a b o u t 10% lower than is normally found and might be attributed to the presence of iron oxide which could conceivably result in a micro-electrochemical coating of the gold grains with iron oxide leading to an increased dissolution time [ 1 3 ] . This was only an hypothesis however, as no evidence of this type o f coating could be found in electronmicroprobe studies. TABLE 8 D a t a f r o m t h e c y a n i d a t i o n o f t h e residues f r o m t h e a m m o n i u m c a r b o n a t e pressure leach Sample
1 2
Au (Head)
A u (Tails)
Au Dissolution
KCN
(g ton -~)
(g ton -1)
(%)
(kg ton -1)
195.4 179.6
27.2 18.8
86.1 89.1
1.05 0.64
It was thought that the low dissolution of gold could be due to the low pH of the system which would decrease the concentration of " f r e e " cyanide present in the solution. To check on this the experiments were repeated with the addition of sufficient ammonium hydroxide to raise the pH to 10. The dissolution of gold was found to be lower, in this case, however, being : a b o u t 74.5%. One interesting fact observed was that the uranium extractions increased ~during the cyanidation experiments to values of 78.4% for sample 1 and 75.5% for sample 2. It would seem from this that an increase in leaching time would result in an increase in uranium extraction.
216
Sodium carbonate leaching The results of the sodium carbonate pressure leach at a pressure of 0.4 MPa O: and a temperature of l l 0 ° C are shown in Fig. 6. The extraction of uranium is low being approximately the same as found when using a m m o n i u m carbonate. From Fig. 6 it can be seen that the extraction process is still proceeding but it would seem that the time needed to reach 90% extraction would be too long for a pressure leach to be feasible. I
l
T----
I 2
I 4
I 6
? E ©
6@t_ × Lxl
(Z~ 40~-
2o~-
i o
Time(h)
Fig. 6. T h e v a r i a t i o n o f U308 e x t r a c t i o n w i t h time. S o d i u m c a r b o n a t e pressure leach o f t h e calcine.
The consumption of sodium carbonate together with theoretical values are shown in Fig. 7. The consumption of sodium carbonate is about 40% lower than was found in the previous experiments using the complete ore indicating the success of flotation and roasting in this respect. The consumption of sodium carbonate is still too high, however, for the process to be economically competitive but could be decreased if the sulphur content of the calcine was reduced further.
Cyanidation of calcine leach residues The results of the cyanidation experiments are given in Table 9. The extraction of gold was very low, being approximately the same as when using a m m o n i u m carbonate at high pH.
217
4O
#-
v 30 & E 3 020 U
o,~ 10
I 2
I 4
I 6
T i m e (h)
Fig. 7. Comparison of theoretical and experimental variations of sodium carbonate consumption with time. ® Experimental a Theoretical. TABLE 9 Cyanidation data pH
Au (Head) (g ton -~)
Au (Residue) (g ton -~)
KCN (kg ton -~)
Au dissolution (%)
10.30 10.30
228.35 209.2
63.3 45.6
0.35 0.33
72.3 78.2
CONCLUSION T he dissolution o f gold and uranium using a sodium carbonate and bicarb o n a t e solution u n d e r pressure followed by t h e addition o f potassium cyanide was successful with 97.9% o f the gold and a p p r o x i m a t e l y 90% o f the uranium being extracted. The process, however, is n o t economically competitive owing to the high c o n s u m p t i o n o f sodium car bona t e which was in the region o f 65 kg t o n - L T he use o f a m m o n i u m c a r bonat e instead o f sodium carbonate and bicarbonate gave similar extractions with a lower c o n s u m p t i o n of carbonat e although in this case n o a m m o n i u m bicarbonate was p r o d u c e d in contrast to the sodium carbonate system.
218
In both sets of eyanidation experiments the consumption of potassium cyanide was greater than in the conventional process. This is probably due to a combination of the low pH and the agitation of the solutions. The inclusion of a flotation step followed by roasting of the flotation concentrate into the process decreased the consumption of sodium carbonate to a b o u t 35 kg t o n -~ b u t at the same time the extractions of gold and uranium were considerably reduced. ACKNOWLEDGEMENTS
The authors wish to thank Johannesburg Consolidated Investment Company Limited for funds and also for the use of equipment and services at their Minerals Processing Research Laboratory. REFERENCES 1 D u Preez, J.G.H., Van Vuuren, C.P.J. and Morris, D.C. The chemistry of uranium part XXVI: alkaline dissolution of A u and/or U O 2 powders, Hydrometall. 6 (I+2) (1980) 147--158. 2 Laxen, P.A., Pammenter, R.V., Bayne, D.L.G., Litvin, S.S. and Vorster, J.H. South African Atomic Energy Board, Extractive Metallurgy Division Reports, Project No. C30/65 A, 1967. 3 Forward, F.A., Halpern, J. and Peters, E., Studies in the carbonate leaching of uranium ores, Trans. Can. Inst. Min. Metall. LVI (1953) 354--360. 4 Pearson, R.L. and Wadsworth, M.C., A kinetic study of the dissolution of UO 2 in carbonate solution, Trans. Metall. Soc. A.I.M.E. 212 (1958) 294--300. 5 Schortmann, W.E. and De Sesa, M.A., Kinetics of the dissolution of uranium dioxide in carbonate--bicarbonate solutions. Proceedings of the 2nd International Conference on the Peaceful uses of A t o m i c Energy. Geneva--United Nations, 3 (1958) 333--341. 6 Peters, E. and Halpern, J., Studies in the carbonate leaching o f uranium ores. II. Kinetics of the dissolution of pitchblende, Trans. Can. Inst. Min. Metall., LVI (1953) 360--364. 7 Hannay, R.L., Eldorado Beaverlodge Operation. 5: Metallurgical plant, Can. Min. J., 81 (1960) 113--136. 8 Jones, J.Q. and De Jong, L.C., Sapin uranium ore processing, in Arbiter, N. (Ed.), Milling Methods in the Americas. Gordon and Breach Science Publishers, New York, 1964, pp. 283--312. 9 Merrit, R.C., The Extractive Metallurgy o f Uranium, Colorado School of Mines Research Institute, Denver, 1971, pp. 83--102 and chs. 15, 16. 10 Beverley, R.G., Griffith, A.W. and Millsap, W.A., Atmospheric vs. pressure leaching of uranium ores, J. Met., 9 (1957) 745--751. 11 Langston, B.G., MacDonald~ R.D. and Stephens Jnr., F.M., Ammonium carbonate pressure leaching of uranium ores, J. Met., 9 (1957) 752--756. 12 Katz, J.J. and Rabinowitch, E., The Chemistry of Uranium, Dover Publications Inc., New York, 1961, p. 609. 13 Kinloch, E.D., Mineralogical examination of Randfontein bench scale test products, Report M]77/137., Johannesburg Consolidated Investment Company Limited, Johannesburg, 1977, pp. 4. 14 Loo, P.T. and Gussmann, H.W., Pressure leach of final mill pulp from Randfontein Estates, Report No. 15/75, Johannesburg Consolidated Investment Company L i m i t e d Johannesburg, 1975.