- Email: [email protected]
Reduction-roasting and ferric chloride leaching of copper converter slag for extracting copper, nickel and cobalt values
Hydrometallurgy, 7 (1981) 243--252
243
Elsevier Scientific Publishing Company, Amsterdam
REDUCTION-ROASTING AND FERRIC CHLORIDE LEACHING OF COPPER CONVERTER SLAG FOR EXTRACTING COPPER, NICKEL AND COBALT VALUES S. ANAND, R.P. DAS and P.K. JENA
Regional Research Laboratory, Bhubaneswar- 751 O13 (India) (Received June 1, 1980; accepted January 15, 1981)
ABSTRACT Anand, S., Das, R.P. and Jena, P.K., 1981. Reduction-roasting and ferric chloride leaching of copper converter slag for extracting copper, nickel and cobalt values. Hydrometallurgy, 7 : 2 4 3 - - 2 5 2 In a previous investigation the o p t i m u m conditions for recovering copper, nickel and cobalt from converter slag through ferric chloride leaching have been described. The study of various parameters revealed that nickel and cobalt recovery could not be improved beyond 24 to 26% respectively from converter slag, though more than 90% of the copper could be extracted. Further attempts were made to bring the metal values completely into solution through reduction-roasting followed by ferric chloride'leaching of the slag. The present work comprises a study of various experimental conditions such as concentration of ferric chloride, duration of leaching, duration of reduction-roasting, temperature and nature of reducing agent, to arrive at the optimal recovery of the metal. Under identical experimental conditions a decrease in copper recovery, but an increase in nickel and cobalt recovery has been observed above a roasting temperature of 750 ° C. The decrease in copper recovery has been attributed to copper ferrite formation which has been confirmed both by leaching experiments with synthetic mixtures and by X-ray diffraction studies with both slag samples and synthetic mixtures. Recovery of nickel has also shown little decline when solid reductants were used above 850°C whereas cobalt recovery remains nearly the same even above 850 ° C. Under o p t i m u m conditions 80% copper, 95% nickel and 80% cobalt could be recovered by reducing the slag at 850°C with 10 wt % furnace oil, followed by leaching with ferric chloride.
INTRODUCTION
In a previous study [ 1 ], the o p t i m u m conditions for recovering copper, nickel and cobalt from the converter slag by ferric chloride leaching were determined. These studies on the slag from the Ghatsfla plant of Hindustan Copper Limited, India, had shown that only about 26% of the nickel and cobalt present in slag could be leached by ferric chloride. In another study on the same slag, attempts were made to leach the slag by dilute sulphuric acid [2]. It was observed that although sulphuric acid could leach about 95% of cobalt and nickel, and 90% of copper, its consumption was about ten times the stoichiometricaUy required quantity. The excess sulphuric
0304-386x/81/0000--0000/$02.50 © 1981 Elsevier Scientific Publishing Company
244
acid was consumed by the reaction with iron. Further, the presence of silicic acid in the leach liquor made the solid--liquid separation difficult. Due to these difficulties the present w o r k on reduction-roasting and ferric chloride leaching of the converter slag was carried out. Reduction of molten converter slag by carbon between 1300 and 1500°C had been investigated by several workers [ 3--6 ]. The present study also considers reduction of converter slag by carbonaceous material, b u t at a much lower temperature of 750--950 ° C. Various reducing agents tried in this study were: charcoal, furnace oil, lignite and bituminous coal from the Talcher region of India. Some of the characteristics of these fuels [7] are given in Table 1. In this w o r k the effects of temperature and the time of reduction roasting on the recovery of metal values, and the effect of ferric ion concentration on the leaching efficiency of ferric chloride, have been studied. TABLE 1 Some characteristics of the fuels used (a) Proximate analysis of solid fuels Fuel
Charcoal Bituminous coal from Talcher Lignite
Moisture content
Volatile matter
Ash content
Fixed carbon
(%)
(%)
(%)
(%)
7.62
20.41
4.62
67.35
5.72 16.55
34.75 60.76
24.01 7.01
36.02 15.60
(b) Furnace oil Specific gravity Calorific value, kcal kg -~ Sulphur content, wt. %
0.9654 10,000--10,250 gross 4.5
EXPERIMENTAL
The copper converter slag used for these studies had a typical composition as shown in Table 2. All the chemicals used were of reagent grade. Copper was quantitatively estimated b y the iodometric method, while nickel was gravimetrically estimated as nickel dimethyl glyoxime, and cobalt colorimetrically b y using nitroso-R-Salt as reagent [8]. The interference due to copper, while estimating nickel and cobalt, w a s prevented b y extracting it completely and selectively at pH 2 into 25% LIX 64N in kerosene as solvent. For the reduction-roasting studies, the required a m o u n t of slag was mixed with various types of reducing agents in pre-determined quantities, and then placed in a stainless steel tubular reactor (length, 225 mm; inside diameter
245 TABLE 2 Chemical analysis o f converter slag
Amount (%)
Converter slag
Cu
Ni
Co
Fe
Ca
Mg
A1
SiO 2
4.03
1.98
0.48
38.32
4.01
2.65
0.08
34.32
44 mm; wall thickness, 3.5 m m ) with a flanged end and a flat lid. The system was made airtight by placing an asbestos O-ring between the lid and the flange and tightening the nuts to hold the lid in place. Reduction-roasting of the above mixture was carried o u t by placing the reactor in a muffle furnace which could maintain a specific temperature within + 5 ° C. Synthetic mixtures of copper powder/cupric oxide and magnetite were roasted the same way as slag samples. The purity of the chemicals used for preparing synthetic mixtures was approximately 95%. The roasted products were analysed by X-my and diffraction patterns were compared with those in the ASTM index [ 9 ] . Leaching studies were carried o u t in a 500 ml glass vessel fitted with a Q
I00
E) COPPER • NICKEL A COBALT T- IO0~C, I0 % EOIL RT- 750%, R t - Ih FeCl.~ 1.4TIMES STOICHIOMETRIC
.j I00
BQ
70
•
sot 40
/
7/o L,/I
z o
60
/
o COPPER
/
•
&//
20 tu 20
0 0
NICKEL
RT- 750"C: Rt- Ih
/
,o~. F,OIL
I0 •
o15
i
~.o
[
Ls
TIME, HOUR
~.o
z'.s ~'.o
0
i.o
L
1.2
,:4
,'.6
i
,.s
z'.o 2'.~
z'.4 2:6
FERRIC IRON(g/IOI OF CONVERTER SLAG)
Fig. 1. Effect o f leaching t i m e on extraction o f metal values f r o m reduced converter slag. In all figures T = leaching temperature, t = leaching time, R T = reduction roasting temperature, Rt = reduction roasting time. Fig. 2. Effect o f total iron on extraction o f metal values f r o m reduced converter slag.
246
mechanical stirrer and a thermometer. Heating of the glass vessel was done b y a heating mantle with an energy regulator. The temperature could be adjusted within +2°C. F o r studying the effect of ferric ion concentration on the recovery of metal values, 100 g of slag was reduced with 10 w t % furnace oil at 750°C for one hour. The reduced mass was ground to - 1 0 0 mesh B.S.S. Representative samples each o f 10 g were leached at 100°C with amounts of ferric chloride'varying from the stoichiometric requirement to a 100% excess. Similarly, duration of leaching was varied with the representative samples from once reduced slag. A t w o hour duration was sufficient to leach the reduced metal values using 1.25 times the stoichiometrically requirecl ferricchloride (Figs. 1 and 2) which confirmed earlier observations [1]. The experiments were repeated a number of times to ascertain the reproducibility of the results, which was f o u n d to vary within + 5%. RESULTS AND DISCUSSION
Some attempts were made to identify the forms in which copper, nickel and cobalt were present in the slag. The microscopic (optical) examination and X-ray diffraction patterns did n o t lead to any conclusion due to the low percentage of nickel and cobalt present in the slag. But copper was f o u n d to be in the oxide form. It was observed that ferric chloride could leach a b o u t 26% of nickel and cobalt values while dilute H2SO4 (0.98 M) could leach more than 95% of these metals [1, 2]. Since ferric chloride would leach E) COPPER • NICKEL A COBALT T - IO0°C, t - 2h RT - 750"C s R t - l h PARTICLE SIZE<~ISO MICRONS FeC[ 3 1.4 TIMES STOICHIOMETRIC
Q COPPER • NICKEL COBALT T-IO0°C, t - 2 h I 0 ~ F.OIL, RT-750°C FeC{3 1.4 TIMES STOICHIOMETRIC
_J
lOC
.~ IOC I--
z sc z sc x 4¢ kX uJ
2C
o~
~'.s
~
7.'s
i'o
ds
FURNACE OIL (.Wt % )
I's
4c
2¢
°o
ds
I:o
iis
2'.o 2'.s 3'.0
T I M E OF REDUCTION-ROASTING~HOURS
Fig. 3. E f f e c t of p e r c e n t furnace oil o n m e t a l value r e c o v e r y f r o m c o n v e r t e r slag. Fig. 4. E f f e c t o f reduction-roasting t i m e o n m e t a l values e x t r a c t i o n f r o m c o n v e r t e r slag.
247 nickel and cobalt in metallic/sulphide forms and dilute H2SO4 w o u l d leach these metals in metallic/oxide form, it is inferred that most of the nickel and cobalt in the converter slag are present as oxides. Such a conclusion is supported b y literature evidence [5, 6]. To optimize the recovery of copper, nickel and cobalt b y reduction roasting, the amounts of furnace oil and the times of roasting were varied. Figure 3 shows that the reduction of nickel and cobalt oxides did n o t improve significantly when the furnace oil added was more than 5 wt. % of slag. However, to ensure that a sufficient a m o u n t of reducing agent is available, the experiments were carried o u t with 10 wt. % furnace oil. The duration of roasting was determined from another set of experiments and the results are shown in Fig. 4. It was observed that the recovery of copper after roasting and leaching with ferric chloride remained independent of the duration of roasting. This is due to the fact that most of the copper can be leached b y ferric chloride from unreduced slag [ 1]. However, the recovery of cobalt and nickel attained equilibrium after one hour. An a t t e m p t was made to see if thermodynamic equilibrium was attained in the system by considering the following reactions NiO + C -+ Ni + CO
(1)
CoO + C -~ Co + CO
(2)
Assuming that the activities of carbon, nickel and cobalt are equal to unity, and Pco was the same for these reactions under equilibrium conditions, we have Pco
=
K l a c o o = K2aNio
(3)
where K, and K2 are the equilibrium constants for the reactions (1) and (2) respectively. If we assume that both CoO and NiO follow the ideal Henrian behaviour in the solid slag, the percentage of CoO and NiO left in slag after roasting will have a linear relation. Such behaviour was actually observed as shown in Fig. 5. During reduction-roasting of converter slag with 10 wt. % furnace oil, it was observed that the recovery of cobalt and nickel sharply increased at 850°C, while the recovery of copper decreased to 80% from 95% (Fig. 6). With solid reducing agents like charcoal, lignite and bituminous coal, a similar behaviour as shown in Figs. 7--9 was also observed. While the decrease in copper recovery began at 750°C with each of these solid reducing agents, the sharpest decrease was observed in the case of lignite. Since copper, in the form of sulphide, oxide and metal can be leached by ferric chloride, the decrease in the recovery could be from the combination of copper with the silicates present or b y the formation of copper ferrite. The presence of CuFeO2 was confirmed in the p r o d u c t b y X-ray diffraction studies (Table 4). An a t t e m p t was made to determine whether the conditions for CuFeO2
248 O
COPPER
• NICKEL & COBALT T - IO0~C, t - 2 h 10% F.OIL t R t - t h FeC[ 3 1.4 TIMES STOIGHK)METRIC I00
7c
90 c= I BO
6c z
70 x 0
60
~ 5G
50 u
~
4o
40 30
30
4'o
s'o
6'o
~o
2%0
. 500
7= COBALT AS OXIDE IN RESIDUE
. 600 . .
700
500
' 900
' I000
TEMPERATURE~°C
Fig. 5. Percent nickel o x i d e vs. p e r c e n t cobalt oxide in residue. Fig. 6. E f f e c t of reduction-roasting t e m p e r a t u r e using 10% furnace oil on metal value r e c o v e r y f r o m c o n v e r t e r slag.
Q COPPER • NICKEL A COBALT T - IO0°C, t - 2 h 107o TALCHER COAL~ Rt - I h FeC[~ 1.4 TIMES STOICHIOMETRIC
E) COPPER • NICKEL COBALT T - IO0°C, t - 2 h 107oCHARCOAL, R t - t h FeC[3 1.4 TIMES STOICHIOMETRIC I00
~100
:ESO
so
so
u
6o
~ 4o
=~ 40
w
tu 20
600
7 ~0
~ 0 9 ~00 I000 . 1200. TE MPERAT URE~=C
.
.
.
600 .
700 8 0 0 9 0 0 I000 TEMPERATUREj°C
Fig. 7. E f f e c t o f reduction-roasting t e m p e r a t u r e using 10% charcoal on m e t a l value recovery f r o m c o n v e r t e r slag. Fig. 8. E f f e c t o f reduction-roasting t e m p e r a t u r e using 10% Talcher coal on metal value r e c o v e r y f r o m c o n v e r t e r slag.
249 o COPPER • NICKEL A COBALT T- IOO'C, t - Z h IO7=LIGNITE~ Rt- lh FeC[3 I.ZS TINES 5TOICHIOMETRIC lOG
IE Z
2 4c ,o i
6O0 700 80O 900 I000 TEMPERATUREfC Fig. 9. E f f e c t o f r e d u c t i o n - r o a s t i n g t e m p e r a t u r e using 10% l i g n i t e o n m e t a l v a l u e r e c o v e r y f r o m c o n v e r t e r slag.
formation existed in the system. It was thought that ferrite formation could take place because of one or more of the following (i) the ash fraction from lignite would be reacting with the slag to form ferrite (ii) the existing phases in the slag would be reacting with each other to form ferrite at the roasting temperature (iii) the lignite present in the charge would be reducing the metallic oxides to metals, which in turn, would be reacting among themselves and with oxygen for ferrite formation. To test these three alternatives, synthetic mixtures were prepared with copper powder, cupric oxide and magnetite to represent the slag composition. To these samples 10 wt. % lignite was added and the mixtures were roasted and leached in the same manner as the slag samples. The results shown in Table 3 indicated that the metallic copper or cupric oxide, mixed with magnetite and lignite, produced ferrite (CuFeO2) and hence resulted in poor copper recovery. The studies also indicated that CuFe:O4 would also be formed when the roasting is done in air w i t h o u t lignite. While the conditions for the formation of CuFeO2 and CuFe204 in air are well reported in the literature [11, 1 2 ] , the same under a reducing atmosphere are n o t well known. To explain the mechanism of CuFeO2 formation in the presence of lignite the following may be considered. The X-ray diffraction patterns (Table 4) indicate the presence of metallic copper when either the slag or copper oxide was roasted at 750°C with lignite. Such reduction of cuprous oxide to metallic copper was also observed by Sohn and R y z h o n k o v [ 1 3 ] . It is also known that reduction of magnetite starts above 800°C [ 13] ; the presence of metallic iron in the roasted p r o d u c t was also observed in the present
250 TABLE 3 Results obtained for synthetic mixtures
No.
1 2 3 4 5 6 7 8 9 10 11 12 13 14
Designation
% Cu recovery on leaching with ferric chloride
Copper powder Copper oxide 0.5 g Cu powder + 10% lignite RT* 950°C 0.5 g cupric oxide + 10% lignite R T 950°C 0.5 g Cu powder + 5 g magnetite heated at 950°C for 1 h in air 0.5 g Cu powder** + 5 g magnetite heated at 950°C for 1 h inside the reactor 0.5 g Cupric oxide + 5 g magnetite heated at 950°C for 1 h in air 0.5 g cupric oxide** + 5 g magnetite heated at 950°C for 1 h inside the reactor Converter slag heated at 950°C for 1 h in air 0.5 g Cu powder + 5 g magnetite + 10% lignite, RT 750°C 0.5 g cupric oxide + 5 g magnetite + 10% lignite, RT 750°C 0.5 g Cu powder + 5 g magnetite + 10% lignite, R T 950°C 0.5 g Cupric oxide + 5 g magnetite + 10% lignite, R T 950°C Slag + 10% lignite RT* 950°C
96.07 86.6 86.07 86.6 nil 81.47 nil 83.27 1.98 86.16 85.31 0.8 nil 1.56
* R T = Reduction-roasting temperature. **In these studies the magnetite remained unchanged. It is estimated that Po~, in the reactor would be less than 10 -s, which is the partial pressure of oxygen required for magnetite to hematite conversion at 950 ° C. TABLE 4 Phases identified by X-ray in various slag samples and synthetic mixtures
No.
Designation
Phases identified
Converter slag Converter slag heated at 950°C Converter slag + 10% lignite, RT* 750°C Converter slag + 10% lignite, R T 950 ° C Copper oxide + magnetite heated at 950°C Copper oxide + magnetite + 10% lignite, RT 950°C
Fe2SiO4, CuO, Cu20 , F%O 4 Fe~SiO 4, CuFe204, Fe203, F%O4 Fe2SiO,, F%O 4, metallic Cu Fe2SiO4, CuFeO2, metallic Fe CuFe204, Fe:O3 Major metallic Fe, CuFeO2, F%O 4
*Reduction-roasting temperature.
study. Therefore, it would be reasonable to assume t h a t the initial reactants (CuO, Cu20, Fe304 and C) as well as the products (Cu and Fe) will be participating in the reaction for CuFeO2 formation. Based on the X-ray evidence the following three reactions m a y be considered for CuFeO2 formation: Cu20 + C ~ 2 Cu + CO
(4)
251
Fe304 +4C-~3Fe+4CO
(5)
Cu + Fe + 02 -~ CuFeO2
(6)
The calculated equilibrium oxygen partial pressures for these three reactions are 10 -9"7, 10 -13"ss, and 10 -14"3s respectively [14, 15]. Thus it m a y be concluded that CuFeO2 could be formed in the presence of carbon (or lignite) if the system oxygen partial pressure was around 10 -14 . Since the X-ray evidence showed the presence of metallic copper and iron, it is fair to assume that reactions (4)--(6) m a y be occurring simultaneously in the system. The variation in decrease in the percentage of copper extraction is probably due to the different fixed carbon values (Table 1) and the differing reactive nature of carbon in the various reducing agents as well as the prevailing gaseous atmosphere inside the reactor. Furnace oil, which is the strongest reducing agent, possibly decomposes the ferrite, whereas lignite which has a fixed carbon content of a b o u t 16% is unable to do this. In the case of the reduction of nickel oxide, a m a x i m u m was observed at a reduction-roasting temperature of 850°C with charcoal and Talcher coal as reductants (Figs. 7 and 8). However, there was no significant decrease in nickel recovery at 850°C when lignite or furnace oil was used; R y z h o n k o v et al. [16] have found that the reduction of nickel occurs at a m a x i m u m rate at 775°C and 90% reduction is obtained at 820°C in agreement with our observations (the low percentage recovery of nickel with lignite is probably due to the non-availability of the carbon present). A decrease in the extraction of nickel with solid-reductants m a y be attributed to the recombination of nickel with silicates that form at high roasting temperatures [ 1 7 ] . This could n o t be confirmed b y X-ray diffraction studies due to low percentage of nickel present in the slag. Reduction of cobalt oxide with all the reducing agents taken up for present studies either increased with increasing reduction-roasting temperature (750--950°C) or remained unchanged. CONCLUSION
The studies on reduction-roasting and ferric chloride leaching of converter slag have revealed that only when 10% furnace oil was used as the reducing agent 82% copper, 95% nickel and 80% cobalt could be extracted b y roasting the slag at 850°C followed b y leaching with ferric chloride. In the case of the solid reductant like charcoal, bituminous coal or lignite, nickel and cobalt extractions were low when roasted at 750°C whereas copper recovery decreased at a roasting temperature of 850 ° C. The decrease in copper recovery has been attributed to ferrite formation through the combination of reduced copper and metallic iron with oxygen. Cobalt recovery using solid reductants remained around 60% at 850°C and when the reduction temperature was increased above 850 ° C nickel recovery started declining, probably due to the recombination of nickel with refractory silicates [16].
252 ACKNOWLEDGEMENTS
The authors are thankful to Mr. B.C. Nayak for taking X-ray diffraction patterns of the samples. Thanks are due to Mr. K.N. Jena for helping in nickel analysis. One of the authors (S.A.) is grateful to the Director, Regional Research Laboratory, Bhubaneswar, for providing her research associateship. REFERENCES 1 Anand, S., Kanta Rao, P. and Jena, P.K., Recovery of copper, nickel and cobalt from copper converter slag through ferric chloride leaching. Hydrometallurgy, 5 (4) ( 1980) 355--365. 2 Anand, S., Kanta Rao, P. and Jena, P.K., Leaching behaviour of copper converter slag in sulphuric acid. Trans. IIM, 33 (1) (1980) 77--81. 3 White, R.M., Organs, J.R., Harris, G.B. and Thomas, J.A., Development of a process for the recovery of electrolytic copper and cobalt from Rokhana converter slag. In: M.J. Jones (Ed.), Advan. Extr. Metall. IMM, (1977) 57--68. 4 Banks, C.C. and Harrison, D.A., Recovery of non-ferrous metals from secondary copper smelter discard slags. Can. Metall. Q., 14 (2) (1975) 183--190. 5 Makhmadiyarov, T., Deer, V.I. and Khudyankov, I.I., Kinetics of the reduction of cuprous oxide from molten slags by solid carbon. Fiz. Khim. Issled-Metallurg Protsessov, 2 (1974) 94--96 (in Russian). 6 Zakharov, B.N. and Tikhorov, A.I., Kinetics of the reduction of cobalt oxides from Converter slags. Tsvetn. Met., (7) (1975) 32--33 (in Russian). 7 Panda, S.C., Sukla, L.B., Kanta Rao, P. and Jena, P.K., Extraction of nickel through reduction-roasting and ammonical leaching of lateritic nickel ores. Trans. IIM, 33 (2) (1980) 161--165. 8 Vogel, A.I., Quantitative Inorganic Analysis. Longmans, Green and Co. Ltd., 1962. 9 ASTM index for X-ray diffraction patterns. 10 Habashi, F., Chalcopyrite; its Chemistry and Metallurgy. McGraw-Hill International Book Co., New York, 1978. 11 Rosenqvist, T., Thermodynamics of metal--sulphur--oxygen systems. Presented at International Conference on Advances in Chemical Metallurgy, Bombay, 3--6 Jan. 1979, preprints 2/1--2/18. 12 Shirts, M.B., Bloom, P.A. and Mckinney, W.A.,, Double roast leach electrowinning process for chalcopyrite concentrates. U.S. Bur. Mines, Rep. Invest., 7996 (1975) 33. 13 Sorin, S.B. and Ryzhonkov, D.I., Carbothermic reduction oxides under non-isothermal conditions. Kinet. Zakohomern. Sovmestnogo Vosstanov Okislov Zhelera Drugikhmater, (1973) 91--105 (in Russian). 14 Kubaschewski, O. and Evans, E.I.L., Metallurgical Thermochemistry, 3rd edn. Pergamon Press, Oxford, 1965. 15 Karpet'yants, M.Kh. and Karpet'yants, M.L., Thermodynamic constants of inorganic and organic compounds. Translated by J. Schmork. Ann Arbor, Humphrey Science Publishers, Inc., 1970. 16 Ryzhonkov, D.I., Sazonov, V.V. and Kolchanov, V.A., Thermogravimetric study of the carbon-reduction of ferrous-oxide--nickel monoxide mixtures. Kinet. Zakohomern. S ovmestnogo Vosstanov Okislov Zhelera Drugikhmater, ( 1973) 1 1 0 - - 1 2 8 ( in Russian ). 17 Brooks, P.T., Potter, G.M. and Rosenbaum, J.B., Improvement in reductive-roasting of lateritic nickel ores for ammonia leach processes. Presented at the Annual AIME meeting, Chicago, Feb. 26--March 1 1973.
243
Elsevier Scientific Publishing Company, Amsterdam
REDUCTION-ROASTING AND FERRIC CHLORIDE LEACHING OF COPPER CONVERTER SLAG FOR EXTRACTING COPPER, NICKEL AND COBALT VALUES S. ANAND, R.P. DAS and P.K. JENA
Regional Research Laboratory, Bhubaneswar- 751 O13 (India) (Received June 1, 1980; accepted January 15, 1981)
ABSTRACT Anand, S., Das, R.P. and Jena, P.K., 1981. Reduction-roasting and ferric chloride leaching of copper converter slag for extracting copper, nickel and cobalt values. Hydrometallurgy, 7 : 2 4 3 - - 2 5 2 In a previous investigation the o p t i m u m conditions for recovering copper, nickel and cobalt from converter slag through ferric chloride leaching have been described. The study of various parameters revealed that nickel and cobalt recovery could not be improved beyond 24 to 26% respectively from converter slag, though more than 90% of the copper could be extracted. Further attempts were made to bring the metal values completely into solution through reduction-roasting followed by ferric chloride'leaching of the slag. The present work comprises a study of various experimental conditions such as concentration of ferric chloride, duration of leaching, duration of reduction-roasting, temperature and nature of reducing agent, to arrive at the optimal recovery of the metal. Under identical experimental conditions a decrease in copper recovery, but an increase in nickel and cobalt recovery has been observed above a roasting temperature of 750 ° C. The decrease in copper recovery has been attributed to copper ferrite formation which has been confirmed both by leaching experiments with synthetic mixtures and by X-ray diffraction studies with both slag samples and synthetic mixtures. Recovery of nickel has also shown little decline when solid reductants were used above 850°C whereas cobalt recovery remains nearly the same even above 850 ° C. Under o p t i m u m conditions 80% copper, 95% nickel and 80% cobalt could be recovered by reducing the slag at 850°C with 10 wt % furnace oil, followed by leaching with ferric chloride.
INTRODUCTION
In a previous study [ 1 ], the o p t i m u m conditions for recovering copper, nickel and cobalt from the converter slag by ferric chloride leaching were determined. These studies on the slag from the Ghatsfla plant of Hindustan Copper Limited, India, had shown that only about 26% of the nickel and cobalt present in slag could be leached by ferric chloride. In another study on the same slag, attempts were made to leach the slag by dilute sulphuric acid [2]. It was observed that although sulphuric acid could leach about 95% of cobalt and nickel, and 90% of copper, its consumption was about ten times the stoichiometricaUy required quantity. The excess sulphuric
0304-386x/81/0000--0000/$02.50 © 1981 Elsevier Scientific Publishing Company
244
acid was consumed by the reaction with iron. Further, the presence of silicic acid in the leach liquor made the solid--liquid separation difficult. Due to these difficulties the present w o r k on reduction-roasting and ferric chloride leaching of the converter slag was carried out. Reduction of molten converter slag by carbon between 1300 and 1500°C had been investigated by several workers [ 3--6 ]. The present study also considers reduction of converter slag by carbonaceous material, b u t at a much lower temperature of 750--950 ° C. Various reducing agents tried in this study were: charcoal, furnace oil, lignite and bituminous coal from the Talcher region of India. Some of the characteristics of these fuels [7] are given in Table 1. In this w o r k the effects of temperature and the time of reduction roasting on the recovery of metal values, and the effect of ferric ion concentration on the leaching efficiency of ferric chloride, have been studied. TABLE 1 Some characteristics of the fuels used (a) Proximate analysis of solid fuels Fuel
Charcoal Bituminous coal from Talcher Lignite
Moisture content
Volatile matter
Ash content
Fixed carbon
(%)
(%)
(%)
(%)
7.62
20.41
4.62
67.35
5.72 16.55
34.75 60.76
24.01 7.01
36.02 15.60
(b) Furnace oil Specific gravity Calorific value, kcal kg -~ Sulphur content, wt. %
0.9654 10,000--10,250 gross 4.5
EXPERIMENTAL
The copper converter slag used for these studies had a typical composition as shown in Table 2. All the chemicals used were of reagent grade. Copper was quantitatively estimated b y the iodometric method, while nickel was gravimetrically estimated as nickel dimethyl glyoxime, and cobalt colorimetrically b y using nitroso-R-Salt as reagent [8]. The interference due to copper, while estimating nickel and cobalt, w a s prevented b y extracting it completely and selectively at pH 2 into 25% LIX 64N in kerosene as solvent. For the reduction-roasting studies, the required a m o u n t of slag was mixed with various types of reducing agents in pre-determined quantities, and then placed in a stainless steel tubular reactor (length, 225 mm; inside diameter
245 TABLE 2 Chemical analysis o f converter slag
Amount (%)
Converter slag
Cu
Ni
Co
Fe
Ca
Mg
A1
SiO 2
4.03
1.98
0.48
38.32
4.01
2.65
0.08
34.32
44 mm; wall thickness, 3.5 m m ) with a flanged end and a flat lid. The system was made airtight by placing an asbestos O-ring between the lid and the flange and tightening the nuts to hold the lid in place. Reduction-roasting of the above mixture was carried o u t by placing the reactor in a muffle furnace which could maintain a specific temperature within + 5 ° C. Synthetic mixtures of copper powder/cupric oxide and magnetite were roasted the same way as slag samples. The purity of the chemicals used for preparing synthetic mixtures was approximately 95%. The roasted products were analysed by X-my and diffraction patterns were compared with those in the ASTM index [ 9 ] . Leaching studies were carried o u t in a 500 ml glass vessel fitted with a Q
I00
E) COPPER • NICKEL A COBALT T- IO0~C, I0 % EOIL RT- 750%, R t - Ih FeCl.~ 1.4TIMES STOICHIOMETRIC
.j I00
BQ
70
•
sot 40
/
7/o L,/I
z o
60
/
o COPPER
/
•
&//
20 tu 20
0 0
NICKEL
RT- 750"C: Rt- Ih
/
,o~. F,OIL
I0 •
o15
i
~.o
[
Ls
TIME, HOUR
~.o
z'.s ~'.o
0
i.o
L
1.2
,:4
,'.6
i
,.s
z'.o 2'.~
z'.4 2:6
FERRIC IRON(g/IOI OF CONVERTER SLAG)
Fig. 1. Effect o f leaching t i m e on extraction o f metal values f r o m reduced converter slag. In all figures T = leaching temperature, t = leaching time, R T = reduction roasting temperature, Rt = reduction roasting time. Fig. 2. Effect o f total iron on extraction o f metal values f r o m reduced converter slag.
246
mechanical stirrer and a thermometer. Heating of the glass vessel was done b y a heating mantle with an energy regulator. The temperature could be adjusted within +2°C. F o r studying the effect of ferric ion concentration on the recovery of metal values, 100 g of slag was reduced with 10 w t % furnace oil at 750°C for one hour. The reduced mass was ground to - 1 0 0 mesh B.S.S. Representative samples each o f 10 g were leached at 100°C with amounts of ferric chloride'varying from the stoichiometric requirement to a 100% excess. Similarly, duration of leaching was varied with the representative samples from once reduced slag. A t w o hour duration was sufficient to leach the reduced metal values using 1.25 times the stoichiometrically requirecl ferricchloride (Figs. 1 and 2) which confirmed earlier observations [1]. The experiments were repeated a number of times to ascertain the reproducibility of the results, which was f o u n d to vary within + 5%. RESULTS AND DISCUSSION
Some attempts were made to identify the forms in which copper, nickel and cobalt were present in the slag. The microscopic (optical) examination and X-ray diffraction patterns did n o t lead to any conclusion due to the low percentage of nickel and cobalt present in the slag. But copper was f o u n d to be in the oxide form. It was observed that ferric chloride could leach a b o u t 26% of nickel and cobalt values while dilute H2SO4 (0.98 M) could leach more than 95% of these metals [1, 2]. Since ferric chloride would leach E) COPPER • NICKEL A COBALT T - IO0°C, t - 2h RT - 750"C s R t - l h PARTICLE SIZE<~ISO MICRONS FeC[ 3 1.4 TIMES STOICHIOMETRIC
Q COPPER • NICKEL COBALT T-IO0°C, t - 2 h I 0 ~ F.OIL, RT-750°C FeC{3 1.4 TIMES STOICHIOMETRIC
_J
lOC
.~ IOC I--
z sc z sc x 4¢ kX uJ
2C
o~
~'.s
~
7.'s
i'o
ds
FURNACE OIL (.Wt % )
I's
4c
2¢
°o
ds
I:o
iis
2'.o 2'.s 3'.0
T I M E OF REDUCTION-ROASTING~HOURS
Fig. 3. E f f e c t of p e r c e n t furnace oil o n m e t a l value r e c o v e r y f r o m c o n v e r t e r slag. Fig. 4. E f f e c t o f reduction-roasting t i m e o n m e t a l values e x t r a c t i o n f r o m c o n v e r t e r slag.
247 nickel and cobalt in metallic/sulphide forms and dilute H2SO4 w o u l d leach these metals in metallic/oxide form, it is inferred that most of the nickel and cobalt in the converter slag are present as oxides. Such a conclusion is supported b y literature evidence [5, 6]. To optimize the recovery of copper, nickel and cobalt b y reduction roasting, the amounts of furnace oil and the times of roasting were varied. Figure 3 shows that the reduction of nickel and cobalt oxides did n o t improve significantly when the furnace oil added was more than 5 wt. % of slag. However, to ensure that a sufficient a m o u n t of reducing agent is available, the experiments were carried o u t with 10 wt. % furnace oil. The duration of roasting was determined from another set of experiments and the results are shown in Fig. 4. It was observed that the recovery of copper after roasting and leaching with ferric chloride remained independent of the duration of roasting. This is due to the fact that most of the copper can be leached b y ferric chloride from unreduced slag [ 1]. However, the recovery of cobalt and nickel attained equilibrium after one hour. An a t t e m p t was made to see if thermodynamic equilibrium was attained in the system by considering the following reactions NiO + C -+ Ni + CO
(1)
CoO + C -~ Co + CO
(2)
Assuming that the activities of carbon, nickel and cobalt are equal to unity, and Pco was the same for these reactions under equilibrium conditions, we have Pco
=
K l a c o o = K2aNio
(3)
where K, and K2 are the equilibrium constants for the reactions (1) and (2) respectively. If we assume that both CoO and NiO follow the ideal Henrian behaviour in the solid slag, the percentage of CoO and NiO left in slag after roasting will have a linear relation. Such behaviour was actually observed as shown in Fig. 5. During reduction-roasting of converter slag with 10 wt. % furnace oil, it was observed that the recovery of cobalt and nickel sharply increased at 850°C, while the recovery of copper decreased to 80% from 95% (Fig. 6). With solid reducing agents like charcoal, lignite and bituminous coal, a similar behaviour as shown in Figs. 7--9 was also observed. While the decrease in copper recovery began at 750°C with each of these solid reducing agents, the sharpest decrease was observed in the case of lignite. Since copper, in the form of sulphide, oxide and metal can be leached by ferric chloride, the decrease in the recovery could be from the combination of copper with the silicates present or b y the formation of copper ferrite. The presence of CuFeO2 was confirmed in the p r o d u c t b y X-ray diffraction studies (Table 4). An a t t e m p t was made to determine whether the conditions for CuFeO2
248 O
COPPER
• NICKEL & COBALT T - IO0~C, t - 2 h 10% F.OIL t R t - t h FeC[ 3 1.4 TIMES STOIGHK)METRIC I00
7c
90 c= I BO
6c z
70 x 0
60
~ 5G
50 u
~
4o
40 30
30
4'o
s'o
6'o
~o
2%0
. 500
7= COBALT AS OXIDE IN RESIDUE
. 600 . .
700
500
' 900
' I000
TEMPERATURE~°C
Fig. 5. Percent nickel o x i d e vs. p e r c e n t cobalt oxide in residue. Fig. 6. E f f e c t of reduction-roasting t e m p e r a t u r e using 10% furnace oil on metal value r e c o v e r y f r o m c o n v e r t e r slag.
Q COPPER • NICKEL A COBALT T - IO0°C, t - 2 h 107o TALCHER COAL~ Rt - I h FeC[~ 1.4 TIMES STOICHIOMETRIC
E) COPPER • NICKEL COBALT T - IO0°C, t - 2 h 107oCHARCOAL, R t - t h FeC[3 1.4 TIMES STOICHIOMETRIC I00
~100
:ESO
so
so
u
6o
~ 4o
=~ 40
w
tu 20
600
7 ~0
~ 0 9 ~00 I000 . 1200. TE MPERAT URE~=C
.
.
.
600 .
700 8 0 0 9 0 0 I000 TEMPERATUREj°C
Fig. 7. E f f e c t o f reduction-roasting t e m p e r a t u r e using 10% charcoal on m e t a l value recovery f r o m c o n v e r t e r slag. Fig. 8. E f f e c t o f reduction-roasting t e m p e r a t u r e using 10% Talcher coal on metal value r e c o v e r y f r o m c o n v e r t e r slag.
249 o COPPER • NICKEL A COBALT T- IOO'C, t - Z h IO7=LIGNITE~ Rt- lh FeC[3 I.ZS TINES 5TOICHIOMETRIC lOG
IE Z
2 4c ,o i
6O0 700 80O 900 I000 TEMPERATUREfC Fig. 9. E f f e c t o f r e d u c t i o n - r o a s t i n g t e m p e r a t u r e using 10% l i g n i t e o n m e t a l v a l u e r e c o v e r y f r o m c o n v e r t e r slag.
formation existed in the system. It was thought that ferrite formation could take place because of one or more of the following (i) the ash fraction from lignite would be reacting with the slag to form ferrite (ii) the existing phases in the slag would be reacting with each other to form ferrite at the roasting temperature (iii) the lignite present in the charge would be reducing the metallic oxides to metals, which in turn, would be reacting among themselves and with oxygen for ferrite formation. To test these three alternatives, synthetic mixtures were prepared with copper powder, cupric oxide and magnetite to represent the slag composition. To these samples 10 wt. % lignite was added and the mixtures were roasted and leached in the same manner as the slag samples. The results shown in Table 3 indicated that the metallic copper or cupric oxide, mixed with magnetite and lignite, produced ferrite (CuFeO2) and hence resulted in poor copper recovery. The studies also indicated that CuFe:O4 would also be formed when the roasting is done in air w i t h o u t lignite. While the conditions for the formation of CuFeO2 and CuFe204 in air are well reported in the literature [11, 1 2 ] , the same under a reducing atmosphere are n o t well known. To explain the mechanism of CuFeO2 formation in the presence of lignite the following may be considered. The X-ray diffraction patterns (Table 4) indicate the presence of metallic copper when either the slag or copper oxide was roasted at 750°C with lignite. Such reduction of cuprous oxide to metallic copper was also observed by Sohn and R y z h o n k o v [ 1 3 ] . It is also known that reduction of magnetite starts above 800°C [ 13] ; the presence of metallic iron in the roasted p r o d u c t was also observed in the present
250 TABLE 3 Results obtained for synthetic mixtures
No.
1 2 3 4 5 6 7 8 9 10 11 12 13 14
Designation
% Cu recovery on leaching with ferric chloride
Copper powder Copper oxide 0.5 g Cu powder + 10% lignite RT* 950°C 0.5 g cupric oxide + 10% lignite R T 950°C 0.5 g Cu powder + 5 g magnetite heated at 950°C for 1 h in air 0.5 g Cu powder** + 5 g magnetite heated at 950°C for 1 h inside the reactor 0.5 g Cupric oxide + 5 g magnetite heated at 950°C for 1 h in air 0.5 g cupric oxide** + 5 g magnetite heated at 950°C for 1 h inside the reactor Converter slag heated at 950°C for 1 h in air 0.5 g Cu powder + 5 g magnetite + 10% lignite, RT 750°C 0.5 g cupric oxide + 5 g magnetite + 10% lignite, RT 750°C 0.5 g Cu powder + 5 g magnetite + 10% lignite, R T 950°C 0.5 g Cupric oxide + 5 g magnetite + 10% lignite, R T 950°C Slag + 10% lignite RT* 950°C
96.07 86.6 86.07 86.6 nil 81.47 nil 83.27 1.98 86.16 85.31 0.8 nil 1.56
* R T = Reduction-roasting temperature. **In these studies the magnetite remained unchanged. It is estimated that Po~, in the reactor would be less than 10 -s, which is the partial pressure of oxygen required for magnetite to hematite conversion at 950 ° C. TABLE 4 Phases identified by X-ray in various slag samples and synthetic mixtures
No.
Designation
Phases identified
Converter slag Converter slag heated at 950°C Converter slag + 10% lignite, RT* 750°C Converter slag + 10% lignite, R T 950 ° C Copper oxide + magnetite heated at 950°C Copper oxide + magnetite + 10% lignite, RT 950°C
Fe2SiO4, CuO, Cu20 , F%O 4 Fe~SiO 4, CuFe204, Fe203, F%O4 Fe2SiO,, F%O 4, metallic Cu Fe2SiO4, CuFeO2, metallic Fe CuFe204, Fe:O3 Major metallic Fe, CuFeO2, F%O 4
*Reduction-roasting temperature.
study. Therefore, it would be reasonable to assume t h a t the initial reactants (CuO, Cu20, Fe304 and C) as well as the products (Cu and Fe) will be participating in the reaction for CuFeO2 formation. Based on the X-ray evidence the following three reactions m a y be considered for CuFeO2 formation: Cu20 + C ~ 2 Cu + CO
(4)
251
Fe304 +4C-~3Fe+4CO
(5)
Cu + Fe + 02 -~ CuFeO2
(6)
The calculated equilibrium oxygen partial pressures for these three reactions are 10 -9"7, 10 -13"ss, and 10 -14"3s respectively [14, 15]. Thus it m a y be concluded that CuFeO2 could be formed in the presence of carbon (or lignite) if the system oxygen partial pressure was around 10 -14 . Since the X-ray evidence showed the presence of metallic copper and iron, it is fair to assume that reactions (4)--(6) m a y be occurring simultaneously in the system. The variation in decrease in the percentage of copper extraction is probably due to the different fixed carbon values (Table 1) and the differing reactive nature of carbon in the various reducing agents as well as the prevailing gaseous atmosphere inside the reactor. Furnace oil, which is the strongest reducing agent, possibly decomposes the ferrite, whereas lignite which has a fixed carbon content of a b o u t 16% is unable to do this. In the case of the reduction of nickel oxide, a m a x i m u m was observed at a reduction-roasting temperature of 850°C with charcoal and Talcher coal as reductants (Figs. 7 and 8). However, there was no significant decrease in nickel recovery at 850°C when lignite or furnace oil was used; R y z h o n k o v et al. [16] have found that the reduction of nickel occurs at a m a x i m u m rate at 775°C and 90% reduction is obtained at 820°C in agreement with our observations (the low percentage recovery of nickel with lignite is probably due to the non-availability of the carbon present). A decrease in the extraction of nickel with solid-reductants m a y be attributed to the recombination of nickel with silicates that form at high roasting temperatures [ 1 7 ] . This could n o t be confirmed b y X-ray diffraction studies due to low percentage of nickel present in the slag. Reduction of cobalt oxide with all the reducing agents taken up for present studies either increased with increasing reduction-roasting temperature (750--950°C) or remained unchanged. CONCLUSION
The studies on reduction-roasting and ferric chloride leaching of converter slag have revealed that only when 10% furnace oil was used as the reducing agent 82% copper, 95% nickel and 80% cobalt could be extracted b y roasting the slag at 850°C followed b y leaching with ferric chloride. In the case of the solid reductant like charcoal, bituminous coal or lignite, nickel and cobalt extractions were low when roasted at 750°C whereas copper recovery decreased at a roasting temperature of 850 ° C. The decrease in copper recovery has been attributed to ferrite formation through the combination of reduced copper and metallic iron with oxygen. Cobalt recovery using solid reductants remained around 60% at 850°C and when the reduction temperature was increased above 850 ° C nickel recovery started declining, probably due to the recombination of nickel with refractory silicates [16].
252 ACKNOWLEDGEMENTS
The authors are thankful to Mr. B.C. Nayak for taking X-ray diffraction patterns of the samples. Thanks are due to Mr. K.N. Jena for helping in nickel analysis. One of the authors (S.A.) is grateful to the Director, Regional Research Laboratory, Bhubaneswar, for providing her research associateship. REFERENCES 1 Anand, S., Kanta Rao, P. and Jena, P.K., Recovery of copper, nickel and cobalt from copper converter slag through ferric chloride leaching. Hydrometallurgy, 5 (4) ( 1980) 355--365. 2 Anand, S., Kanta Rao, P. and Jena, P.K., Leaching behaviour of copper converter slag in sulphuric acid. Trans. IIM, 33 (1) (1980) 77--81. 3 White, R.M., Organs, J.R., Harris, G.B. and Thomas, J.A., Development of a process for the recovery of electrolytic copper and cobalt from Rokhana converter slag. In: M.J. Jones (Ed.), Advan. Extr. Metall. IMM, (1977) 57--68. 4 Banks, C.C. and Harrison, D.A., Recovery of non-ferrous metals from secondary copper smelter discard slags. Can. Metall. Q., 14 (2) (1975) 183--190. 5 Makhmadiyarov, T., Deer, V.I. and Khudyankov, I.I., Kinetics of the reduction of cuprous oxide from molten slags by solid carbon. Fiz. Khim. Issled-Metallurg Protsessov, 2 (1974) 94--96 (in Russian). 6 Zakharov, B.N. and Tikhorov, A.I., Kinetics of the reduction of cobalt oxides from Converter slags. Tsvetn. Met., (7) (1975) 32--33 (in Russian). 7 Panda, S.C., Sukla, L.B., Kanta Rao, P. and Jena, P.K., Extraction of nickel through reduction-roasting and ammonical leaching of lateritic nickel ores. Trans. IIM, 33 (2) (1980) 161--165. 8 Vogel, A.I., Quantitative Inorganic Analysis. Longmans, Green and Co. Ltd., 1962. 9 ASTM index for X-ray diffraction patterns. 10 Habashi, F., Chalcopyrite; its Chemistry and Metallurgy. McGraw-Hill International Book Co., New York, 1978. 11 Rosenqvist, T., Thermodynamics of metal--sulphur--oxygen systems. Presented at International Conference on Advances in Chemical Metallurgy, Bombay, 3--6 Jan. 1979, preprints 2/1--2/18. 12 Shirts, M.B., Bloom, P.A. and Mckinney, W.A.,, Double roast leach electrowinning process for chalcopyrite concentrates. U.S. Bur. Mines, Rep. Invest., 7996 (1975) 33. 13 Sorin, S.B. and Ryzhonkov, D.I., Carbothermic reduction oxides under non-isothermal conditions. Kinet. Zakohomern. Sovmestnogo Vosstanov Okislov Zhelera Drugikhmater, (1973) 91--105 (in Russian). 14 Kubaschewski, O. and Evans, E.I.L., Metallurgical Thermochemistry, 3rd edn. Pergamon Press, Oxford, 1965. 15 Karpet'yants, M.Kh. and Karpet'yants, M.L., Thermodynamic constants of inorganic and organic compounds. Translated by J. Schmork. Ann Arbor, Humphrey Science Publishers, Inc., 1970. 16 Ryzhonkov, D.I., Sazonov, V.V. and Kolchanov, V.A., Thermogravimetric study of the carbon-reduction of ferrous-oxide--nickel monoxide mixtures. Kinet. Zakohomern. S ovmestnogo Vosstanov Okislov Zhelera Drugikhmater, ( 1973) 1 1 0 - - 1 2 8 ( in Russian ). 17 Brooks, P.T., Potter, G.M. and Rosenbaum, J.B., Improvement in reductive-roasting of lateritic nickel ores for ammonia leach processes. Presented at the Annual AIME meeting, Chicago, Feb. 26--March 1 1973.