- Email: [email protected]
Dephosphorisation of western australian iron ore by hydrometallurgical process
Minerals Engineering, Vol. 12, No. 9, pp. 1083-1092, 1999
Pergamon 0892-6875(99)00093-X
© 1999 Elsevier Science Ltd All rights reserved 0892--6875/99/$- see front matter
DEPHOSPHORISATION OF WESTERN AUSTRALIAN IRON ORE BY HYDROMETALLURGICAL PROCESS
C.Y. CHENG '~, V.N. MISRA '~, J. CLOUGH t and R. MUN* § CSIRO Div. of Minerals, PO Box 90, Bentley, WA 6982, Australia E-mail: chuyong.cheng @minerals.csiro.au t School of Chemical Engineering, Curtin University of Technology, GPO Box U1987, Perth, WA 6845, Australia
(Received 8 February 1999; accepted 22 May 1999)
ABSTRACT More than 80% of Western Australian iron ore contains an average of O.15% phosphorus, and attracts a penalty due to its high level of phosphorus when it is exported. At the current rate of mining, identified premium grade iron ore with low phosphorus content (<0.05%) will be depleted in 30 years. The development of an economical dephosphorisation process is critical for the future success of the Western Australian iron ore industry. In the current work, effective dephosphorisation of Western Australian iron has been demonstrated. Sulphuric acid was chosen as the leachant on the basis of its availability and low cost. The iron ore sample used in this study typically contained 0.126% phosphorus, was from the Pilbara region of Western Australia. After roasting at 1250°C, lump ore (P8o 5.6 mm), pellet 1 (grinding to 100% -1.5 mm before pelletisation) and pellet 2 (grinding to 100% -0.15 mm before pelletisation) were leached in solutions with different sulphuric acid concentrations. After leaching for 5 hours at 60°C in 0.1 M sulphuric acid solution, 67.2%, 69.0% and 68.7% of the phosphorus was leached from the above three samples, respectively. The phosphorus content was reduced from 0.126% to 0.044%, 0.055% and 0.042% respectively. The dissolution of iron during leaching was negligible. The optimum sulphuric acid concentration was 0.1 M in terms of acid cost and iron loss. The acid consumption cost is as low as $A 0.47/tonne. © 1999 Elsevier Science Ltd. All rights reserved.
Keywords Iron ores; leaching; hydrometallurgy
INTRODUCTION Australia is the largest exporter of iron ore in the world. In 1997, approximately 147 million tonnes of Australian iron ore with a total value of 3 billion Australian dollars were exported. Ninety five percent of Australia's iron ore resources are located in the Pilbara region of Northwest Western Australia. The known Western Australian iron ore resources comprise 34 billion tonnes, which represents six percent of the total identified world iron ore resources. From 1966 to 1990, 1.5 billion tonnes of Western Australian iron ore
1083
1084
C.Y. Chenget al.
have been produced at a total value of A$23.7 billion (O'Leary, 1993). Only 10% of Western Australian iron ore resources are premium grade low phosphorus iron ore. At the current rate of export, premium grade resources will be depleted in 30 years. The remaining high iron grade ore has a high phosphorus content and is unsuitable for iron and steel production without dephosphorisation treatment. Over 90% of Western Australian high grade iron ore has a phosphorus content in excess of 0.05% and the average phosphorus content is 0.15%. At present, the acceptable phosphorus content in iron ore for some foreign companies is 0.075%. For every increase of 0.001% in phosphorus content, it attracts a penalty of $A 0.80/tonne. This means that for the ore containing 0.080% phosphorus, the penalty would be around $A 4.0/tonne. The acceptable phosphorus content in iron ore for some other foreign companies is 0.070%. For every increase of 0.001% in phosphorus content, it attracts a penalty of $A 0.10/tonne. For the iron ore containing over 0.080% phosphorus, every increase of 0.001% in phosphorus content, it attracts a penalty of $A 1.00/tonne. This means that for the ore containing 0.083% phosphorus, the penalty would be around $A 4.0/tonne. These penalties reduce the profit margin substantially and make its export virtually impossible. The successful development of a hydrometallurgical process to reduce the phosphorus content in Western Australian iron ore would extend the export of high grade iron ore with low phosphorus content by over 100 years. Since currently there are no facilities for the dephosphorisation of iron ore in Australia, the development of a successful and economically viable technology is of paramount importance. The dephosphorisation of high phosphorus iron ore has been researched in a number of countries, and a variety of dephosphorisation processes have been developed for different phosphorus compounds. Forssberg and Adolfsson (1981) investigated the economic feasibility of acid leaching of iron ore in which apatite, Ca5(PO4)3(F,C1,OH), was the phosphorus-containing mineral. Nitric and hydrochloric acid leaching agents were used. It was found that the economy of leaching with the aid of hydrochloric acid or alternatively by means of nitric acid depended very much on the extent to which the phosphorus can be recovered as phosphoric acid. The dephosphorisation of iron ore using acid leaching was also investigated by Matsuo e t al. (1980). The iron ore contained phosphorus in the form of apatite, Cas(PO4)3 F. The acids tested were hydrochloric, nitric and sulphuric acid. The optimum acid concentration was determined to be between 2-7%. The optimum leaching time was found to exceed 10 hours. Without the use of heat treatment, a 20% reduction in the phosphorus content was achieved. With heat treatment, a reduction of phosphorus of up to 40% was obtained. Unlike iron ore in other regions of the world, phosphorus in Western Australian iron ore exists within goethite, FeOOH, in the form of solid solution (Graham, 1973). The distribution of phosphorus throughout the goethite crystals prevents the effective use of physical separation techniques, and necessitates the use of chemical separation techniques (Gooden et al., 1979). Phosphorus is less soluble in hematite than in goethite. Goethite can be converted to hematite by roasting, which will cause the ore to be more amenable to leaching. The degree of conversion of goethite to hematite is expected to increase with the roasting temperature and time. The cost of roasting also increases with roasting temperature and time, and it is desirable that the use of a thermal pre-treatment stage be avoided. Peixoto (1991) investigated the effect of heat treatment on the acid leaching of iron ore. When the ore contained phosphorus in solid solution in the goethite phase, it was found that thermal treatment of the ore at 1200°C caused a structural re-arrangement of goethite which converted to hematite and facilitated the dissolution of phosphorus in mineral acids. Without thermal treatment of the ore, hydrochloric acid was not effective. After pelletisation, the leaching with hydrochloric acid achieved a 34% reduction in phosphorus. After sintering, a 23% reduction in phosphorus was obtained. No loss of iron was observed. The dephosphorisation of iron ore was conducted by roasting in the presence of alkaline earth metal halides, particularly calcium chloride, or halides of ammonia, manganese, zirconium, copper and lithium (Feld e t al., 1966). Acid leaching was also conducted. The ore was roasted for 2 hours at temperatures between 500°C and 1200°C. The optimum temperature was determined to be 900°C. The most economical salt was
Dephosphorisation of Western Australian iron ore by hydrometallurgicalprocess
1085
found to be CaCI 2, and the optimum leaching temperature was 30°C. Within 30 minutes the phosphorus content in the ore was reduced by around 23% using sulphuric, nitric and hydrochloric acid. The dephosphorisation of high phosphorus ore from the Pilbara region of Western Australia was investigated by Gooden e t al. (1979). Physical beneficiation methods failed as the phosphorus was in solid solution in goethite. Chemical dephosphorisation methods for iron ore were reviewed and tested on the ore samples from the Pilbara region. Leaching without thermal treatment was found to be quite ineffective. The effect of heat treatment, with and without the addition of salts, was investigated. Experiments demonstrated that in a fairly elevated temperature range, salt was not required for solubilisation, and it actually hindered the dephosphorisation process. The effect of roasting temperature on dephosphorisation was investigated, using a roasting time of 0.5 hours. With heat treatment and leaching with hydrochloric, nitric and sulphuric acid and sodium hydroxide solutions, the phosphorus content in the ore could be reduced from 0.12% to 0.05%. Sulphuric acid was extensively tested as this was the most economical acid. Using sulphuric acid, it was found that the quantity of iron leached was dependent on the solution concentration, and independent of the leaching time. The minimum sulphuric acid concentration required to dissolve phosphorus was found to be 0.2 M. The optimum leaching time was found to be between 0.1 and 0.5 hours. Increasing the leaching temperature above 50°C or 60°C did not improve the leaching of phosphorus. However, leaching at temperatures above 70°C to 90°C was found to cause ferric phosphate to precipitate at much lower iron and phosphate concentrations than at temperatures below 70°C, improving the ease of phosphorus recovery. Sodium hydroxide was found to be equally effective as sulphuric acid at concentrations around 0.2 M. However, the process was not as cost effective as sulphuric acid. A conceptual plant design was provided, using a continuous process with leaching agent regeneration and phosphorus recovery. Sun and Yang (1997) indicated that after leaching of the high phosphorus iron ore with Thiobacillus followed by a chemical treatment using a solution containing tannin, the phosphorus content dropped to less than 0.2%. However, the original phosphorus content was not mentioned in the patent abstract. Another researcher in China performed dephosphorisation work using techniques such as bacteria leaching with chemical treatment and nitric acid leaching (Wang, 1997). In the bacteria leaching with Thiobacillus and chemical treatment using RL-1, the phosphorus dropped from 0.36% to 0.18%. In the nitric acid leaching process, the phosphorus content was reduced to less than 0.05% Based on the previous research, the following are the observations on the dephosphorisation of Western Australian iron ore: Roasting is necessary, Leaching with mineral acid is effective alter roasting, and The use of minimum acid concentration and cheap mineral acid makes the process economically viable.
EXPERIMENTAL Approximately 570 kg of high phosphorus iron ore from the Pitbara region of Western Australia was provided by an Australian mining company. Chemical analysis (Table 1) indicated that the phosphorus content of the ore is high at 0.126%. This ore is considered to be representative of Western Australian high phosphorus iron ore. The entire 570 kg of iron ore sample was dried, mixed, and sampled using a rotary sample divider.
TABLE 1 Chemical composition of the iron ore sample
LOI
Content
(%)
(%)
Fe
P
S
Si02
A1203
MgO
Mn
CaO
K20
TiO2
4.01
62.1
0.126
0.014
4.15
2.68
0.10
0.06
0.06
<0.01
0.11
1086
C.Y.
C h e n g e t al.
Size analysis was performed to determine the size distribution of the lump ore as received (Figure 1). It was found that the lump ore sample was 100% - 2 0 mm with 80% - 5 . 6 mm and 30% - 1 . 5 mm. A b o u t 2 kg lump ore was ground in a Braun pulverising disc mill to 100% - 1 . 5 m m which consisted of the raw material for pellet type 1 and another 2 kg to 100% - 0 . 1 5 m m which consisted of the raw material for pellet type 2.
IO0.O0
J
90.00 80.00 ~"
70.00 6o.oo
~,
50.00 40.00
r)
30.00 20.00 I0.00
•
J 0,00 i 0.053
0.090
0.150
0.250
1.5o0 Size
2.500
5.6oo
8.000
I 1.200
20.000
(ram)
Fig.l Size distribution of the lump ore sample. Pellets were prepared in a 0.45 m diameter motor driven pelletising machine. It was observed that smaller pellets of narrower size range and higher sphericity were obtained for the sample ground to 100% - 1 . 5 m m than for the ore ground to 100% - 0 . 1 5 ram. The pellet size was set as 10 m m and the average pellet size was 9.4 and 10.8 m m for pellets 1 and 2, respectively. The lump ore sample and the two pellet samples were batch roasted separately in an electric furnace for one hour in an open cylindrical container of diameter 200 m m and height 50 mm. The sample was placed in and removed from the furnace when the furnace was at room temperature. The furnace was heated from room temperature to 1250°C in 6 hours. After holding at that temperature for one hour, the furnace was cooled from 1250°C to room temperature in 17 hours. In this way, the cracking of the lumps and pellets by vaporisation of the residual moisture in the sample was minimised. The overall increase in internal porosity after roasting was determined to be 1.1% for the lump ore, 16.7% for pellet 1 and 26.9% for pellet 2. The roasted samples were then riffled to six sub-samples which weighed about 200 g each. The leach solution and ore were heated separately to 60°C and the leaching started when the liquid and solids were poured into a 500 ml conical flask. The slurry contained 40% solids by weight. The conical flask was placed on a Cenco hotplate shaker. The leaching temperature was controlled at 60_+2°C. The conical flask was loosely sealed with a rubber stopper to avoid excessive water loss and allow inverse shaking of the slurry in the flask. The flask was inverted manually every 5 minutes, and in the first 10 minutes of leaching it was shaken every minute. Six tests were conducted for each of the three samples. Three acid concentrations and four leaching times were used and the parameters for the 18 tests are listed in Table 2. No intermediate samples were taken since it was very difficult to take representative samples when the particle size was large and the amount of solids was limited. The four tests leached for 10, 60, 120 and 300 minutes formed one set of tests.
Dephosphorisation of Western Australian iron ore by hydrometallurgical process
1087
TABLE 2 Dephosphorisation leaching test conditions Test
Sample
Acid concentration
Leaching time
No
name
(M)
(min)
1
Lump
0.10
10
2
Lump
0.10
60
3
Lump
0.10
120
4
Lump
0.10
300
5
Lump
0.30
300
6
Lump
0.50
300
7
Pellet 1
0.10
10
8
Pellet 1
0.10
60
9
Pellet 1
0.10
120
10
Pellet 1
0.10
300
11
Pellet 1
0.30
300
12
Pellet 1
0.50
300
13
Pellet 2
0.10
10
14
Pellet 2
0.10
60
15
Pellet 2
0.10
120
16
Pellet 2
0.10
300
17
Pellet 2
0.30
300
18
Pellet 2
0.50
300
At the end of each test, the slurry was vacuum filtered and a solution sample was taken for chemical assay. The acid concentration in the leachant was determined by titration with 0.1 M sodium hydroxide. The solids were washed three times with a total of 700 ml distilled water. The residue was dried at 110°C for 16 hours. The residue was then ground in the Braun pulverising disc mill to 100% - 1 5 0 lam and split to obtain 15 g sub-sample for chemical assay by SGS Australia Limited.
RESULTS AND DISCUSSION
Leaching kinetics Figure 2 shows the phosphorus leaching kinetics over 5 hours at 60°C in 0.1 M sulphuric acid solution. About 45% of the phosphorus was leached in the first 10 minutes from the lump ore, largely due to the presence of fines in the solids. After leaching for 60 minutes, the same amount of phosphorus dissolved from the two pellet samples tested. After leaching for 5 hours, the phosphorus extraction was 67.2%, 59.0% and 68.7% for lump ore, pellet 1 and pellet 2 samples, respectively. The highest phosphorus dissolution in 5 hours for pellet 2 was attributed to the high porosity of the pellets, which increased by 26.9% after roasting. The high phosphorus dissolution from the lump ore was attributed to the amount of fines in the sample.
1088
C.Y. Cheng et al.
I00.00
90.00
80.00
70.00
._--&
60.00
...............
4
50.00
40.00
e~
g,
30.00
j
•
~/.'~,o' ° ,,
j ..
20.00
Lumps
- - 4. - -Pellet 1
,.o
- - - A - - Pellet 2
10.00
0.00 50
150
100
200
250
300
350
Time (rain)
Fig.2 Leaching of roasted samples in 0.1 M sulphuric acid solution. Figure 3 shows the phosphorus reduction during leaching. After 5 hours of leaching, the phosphorus content in the three samples was reduced to 0.044%, 0.055% and 0.042%, for the lump ore, pellet 1 and pellet 2, respectively. Again, the lowest phosphorus content in pellet 2 was attributed to the highest porosity in the sample.
0.16
014
0.12
k
= Lump.]
'..'.,, 0.1 c
ea *d
""..72,.. ,.:Z~....7..?.~..S_ ........ ~ ....
00s
-----p~ml~t21
" 0.06
0.04
0.02
0
'
'
20
40
60
80
100
120
140
160
180
200
220
240
260
280
300
Time (rain)
Fig.3 Phosphorus reduction during leaching in 0.1 M sulphuric acid solution. Compared with the other processes reviewed, the current process was very effective in terms of phosphorus extraction. Over 60-70% of the phosphorus was extracted which was much higher than the other researcher's results. For instance, Peixoto [6] obtained 34% phosphorus reduction after leaching roasted iron ore with goethite mineralisation.
Dephosphorisation of Western Australian iron ore by hydrometallurgical process
1089
Effect of acid concentration Table 3 shows the effect of acid concentration on the phosphorus extraction. As expected, the higher the acid concentration, the higher the phosphorus extraction except for pellet 2 which showed the highest phosphorus extraction in 0.3 M acid concentration. This was most probably caused by experimental error. After leaching for 5 hours at 60°C in 0.5 M sulphuric acid solution, about 80%, 63% and 70% of the phosphorus was extracted from the lump ore, pellet 1 and pellet 2, respectively. Accordingly, the phosphorus in the samples decreased to 0.027%, 0.049% and 0.041%, respectively (Table 4).
TABLE 3 Phosphorus extraction (%) after leaching for 5 hours in different acid concentrations Sample
Acid
concentration
(M)
Name
0. I0
0.30
0.50
Lump ore
67.16
74.63
79.85
Pellet 1
58.96
62.69
63.43
Pellet 2
68.66
75.37
69.40
TABLE 4 Phosphorus remaining (%) after leaching for 5 hours in different acid concentrations Sample
Acid
concentration
(M)
Name
0.10
0.30
0.50
Lump ore
0.044
0.034
0.027
Pellet 1
0.055
0.050
0.049
Pellet 2
0.042
0.033
0.041
Compared with the other processes reviewed, the current process was very effective in terms of acid concentration. Gooden et. al. (1979) pointed out that the minimum acid concentration was 0.2 M. In the current process, 0.1 M sulphuric acid concentration was strong enough to reduce the phosphorus content in the iron ore to the required level of 0.05%.
Iron loss during leaching Figure 4 shows the iron loss during leaching in 0.1 M sulphuric acid solution at 60°C. Due to existence of fines, the iron loss in the lump ore was the highest: about 0.06% iron loss in the first 10 minutes and 0.16% after 5 hours of leaching. The iron loss in the two pellet samples was negligible: only about 0.01% iron dissolved in the 0.1 M acid solution after 5 hours of leaching.
TABLE 5 Iron loss (%) after leaching for 5 hours in different acid concentrations Sample
Acid
concentration
(M)
Name
0.10
0.30
0.50
Lump ore
0.157
0.161
0.199
Pellet 1
0.012
0.017
0.018
Pellet 2
0.007
0.025
0.022
1090
C.Y.
et al.
Cheng
The effect of acid concentration on the iron loss is shown in Table 5. The iron loss for the 3 samples was 0.20%, 0.018% and 0.022%, respectively, indicating that the iron loss in the lump ore was about ten times of that in the pellet samples.
0.20
0.18
0.16
0.14
0.12
/
/ f
/ 0.10
Lumps
• 0.08
- .-A-- • Pellet I 0.06
- - • - - Pellet 2
0.04
0.02 j.
~
...........
20
40
~
..............
.-2.."2.'7..-..'~. . . . . . . . . . . . . . . . . . . . . . .
--2 -"~--'- --'- ~-~-i --2 . . . . .
~. . . . . . . . . . . . . . . . . . . . .
0.00 0
60
80
100
120
140
160 Time
180
200
220
240
260
280
300
320
(rain)
Fig.4 Iron loss during leaching in 0.1 M sulphuric acid solution.
Leaching mechanism Various theories have been proposed to explain the existence of phosphorus in goethite. However, none has been able to explain the dephosphorisation leaching of goethite. Graham (1973) proposed that elemental phosphorus existed in solid solution with goethite. Dukino [10] has disputed this theory on the basis that interstitial solid solution is likely only if the substituting cation is of a similar or smaller size than the lattice bound cation and its charge is the same. The size and valency of phosphorus are not suitable for its existence in solid solution with goethite, and it is considered unlikely that elemental phosphorus exists in solid solution with goethite. A more plausible mechanism has been proposed by Morris (1973) and Barbour (1973) and supported by Dukino (1997), involving surface adsorption. It has been suggested that, prior to the dehydration of ferrihydrate to goethite, a surface hydroxyl group is replaced by a phosphate ligand as shown in the following model:
3 O H q - PO43-
I Fe--O
=
O=Fe +
PO 3-
I
Fe + 11 0
Fe+~O
+ 3OH(1)
This theory is considered to be a likely explanation of the existence of phosphorus in goethite. Furthermore, this theory can be used to explain dephosphorisation. Roasting of the iron ore causes the goethite to dehydrate to hematite, liberating the phosphorus in an acid soluble compound, for example, to aluminium phosphate as shown in equation 2:
Dephosphorisation of Western Australian iron ore by hydrometallurgicalprocess 2(FeO)3PO 4 + A1203 = 2AIPO 4 + 3Fe203
1091 (2)
The chemical analysis of the ore indicated that several acid soluble phosphate compounds are possible, including aluminium phosphate, manganese phosphate and magnesium phosphate. The total mass of ore dissolved during leaching (up to 1.4%) does not correspond to the quantity of any of the components in the ore. Therefore it is likely that more than one type of acid soluble compound is formed.
Acid consumption cost The percentage of acid consumption was the highest for leaching with 0.1 M acid: 40-50% compared to 23-40% for 0.3 M acid and 25-35% for 0.5 M acid. However, the overall acid consumption was the highest for leaching with 0.5 M: triple that of leaching with 0.1 M acid. Acid consumption for leaching with 0.3 M acid was double that of leaching with 0.1 M acid. Acid consumption critically affects the process economics, and therefore the minimum acid concentration necessary to achieve leaching of phosphorus to <0.05% is desired. Therefore using 0.1 M acid was the most economical in terms of acid consumption. The acid consumption for leaching using 0.1 M acid was approximately 0.67 tonnes of 98% acid per 100 tonnes of lump ore, and 0.83 tonnes of acid per 100 tonnes of pellet 1 (grinding to 100% -1.5 m m before pelletisation). Sulphuric acid is available from Coogee Chemicals in WA for $A 70/tonne. The cost of acid consumption for dephosphorisation is $A 0.47/tonne of lump ore and $A 0.58/tonne of pellet 2 (grinding to 100% -1.5 mm before pelletisation). Therefore, the cost of acid consumption is reasonable. To estimate the total capital and operation costs of the whole dephosphorisation process including grinding, roasting, pelletisation, leaching, washing and drying is impossible at this stage. However, it is clear that the dephosphorisation plant has to be integrated with a DRI plant where the use of pellets is preferable. Therefore, the cost of the dephosphorisation process can be absorbed by the DRI plant and become economical viable.
C O N C L U S I O N S AND R E C O M M E N D A T I O N S It has been demonstrated that the current process was very effective for the dephosphorisation of Western Australian iron ore. After roasting of the lump ore and the pellets at 1250°C and leaching in 0.1 M sulphuric acid solution at 60°C for 5 hours, the phosphorus extraction was 67.2%, 59.0% and 68.7% for lump ore, pellet 1 (grinding to 100% -1.5 m m before pelletisation) and pellet 2 (grinding to 100% -0.15 mm before pelletisation) samples, respectively. The phosphorus content in the three samples was successfully reduced from 0.13% to 0.044%, 0.055% and 0.042%, respectively, which met the requirement of high grade iron ore for iron making. The higher the acid concentration of the leach solution, the higher the phosphorus extraction. After leaching for 5 hours at 60°C in 0.5 M sulphuric acid solution, about 80%, 63% and 70% of the phosphorus was extracted from the lump ore, pellet 1 and pellet 2 samples, respectively. Accordingly, the phosphorus in the samples decreased to 0.027%, 0.049% and 0.041%, respectively. Due to the presence of fines, the iron loss in the lump ore was the highest: about 0.06% iron loss in the first 10 minutes and 0.16% after 5 hours of leaching. The iron loss in the two pellet samples was negligible. The iron loss increased with the increasing of acid concentration. From an economical point of view, 0.1M sulphuric acid solution is the best choice to obtain reasonable phosphorus extraction with minimum acid consumption cost. The acid consumption cost was as low as A$0.47/tonne for lump ore and A$0.58/tonne for pellet 2. The surface adsorption theory can be used to explain the existence of phosphorus in goethite and the
1092
C.Y. Chenget al.
dephosphorisation process. Roasting of the iron ore causes the goethite to dehydrate to hematite, liberating the phosphorus in an acid soluble form. The residual sulphate in the leached iron lump ore or pellet may affect the total sulphur content, which should be taken into account in the next stage of research work.
ACKNOWLEDGMENTS The authors would like to thank the staff and management of Mineral Processing Laboratory, Department of Minerals and Energy, Western Australia, for the support of the project. Thanks are extended to the School of Chemical Engineering, Curtin University of Technology, for the provision of funds for the chemical analysis of the iron ore samples.
REFERENCES
Barbour, A.R., Distribution of phosphorus in iron ore deposit of Itabira, Minas Gerais, Brazil, Economic Geology, 1973, 68 (1). Dukino, R., Phosphorus in Hamersley Range iron ore, BHP Internal Report, 1997. Feld, I.L., Franklin, T.W. and Lampkin, W.M., Process for removing phosphorus from iron ores, United States Patent, 1966, No 3,402,041. Forssberg, R. and Adolfsson, G., Dephosphorisation of high-phosphorus iron ores by means of acid leaching, Erzmetal, 1981, 34, 316-322. English translation by BHP Central Research Laboratories, No. CRL/T 13278. Gooden, J.E.A., Walker, W.M. and Allen, R.J., 'AMDEPHOS'--A chemical process for dephosphorisation of iron ore, Proceedings of National Chemical Engineering Conference, Queensland, 1974, 21-33. Graham, J., Phosphorus in iron ore from the Hamersley iron formations, Proceedings of the Australasian Institute of Mining and Metallurgy, 1973, No. 246, 41-42. Matsuo, S., Ikeda, R. and Inaga, S., Method of dephosphorising ore, Japanese Patent 57-67136, English translation by BHP Central Research Laboratories, 1980, No. CRL/T 13278. O'Leary, M.A., Overview of the iron ore industry: Twenty-five years of iron ore developments in Australia, in Australasian Mining and Metallurgy, 2nd edn. The Institute of Australasian Mining and Metallurgy, 1993, 1, 231-237. Peixoto, G., Improvement of the reduction process in P content and other gangues in iron ore and its agglomerates, International patent, 1991, No 93/10271. Sun, D. and Yang Y, Method for microbial-chemical dephosphorisation of iron ore, Chinese patent, 1997, No CN 1107518 A. Wang, W.X., Research background of the Mineral Processing Group at Wuhan Yejin University of Science and Technology, Private communication, 1997. Morris, R.C., A pilot study of phosphorus distribution in parts of the Brockman iron formation, Hamersley Group Internal Report, 1973.
C o r r e s p o n d e n c e on papers published in M i n e r a l s Engineering is invited, preferably by e-mail to b w i l l s @ m i n - e n g . c o m , or by Fax to +44-(0)1326-318352
Pergamon 0892-6875(99)00093-X
© 1999 Elsevier Science Ltd All rights reserved 0892--6875/99/$- see front matter
DEPHOSPHORISATION OF WESTERN AUSTRALIAN IRON ORE BY HYDROMETALLURGICAL PROCESS
C.Y. CHENG '~, V.N. MISRA '~, J. CLOUGH t and R. MUN* § CSIRO Div. of Minerals, PO Box 90, Bentley, WA 6982, Australia E-mail: chuyong.cheng @minerals.csiro.au t School of Chemical Engineering, Curtin University of Technology, GPO Box U1987, Perth, WA 6845, Australia
(Received 8 February 1999; accepted 22 May 1999)
ABSTRACT More than 80% of Western Australian iron ore contains an average of O.15% phosphorus, and attracts a penalty due to its high level of phosphorus when it is exported. At the current rate of mining, identified premium grade iron ore with low phosphorus content (<0.05%) will be depleted in 30 years. The development of an economical dephosphorisation process is critical for the future success of the Western Australian iron ore industry. In the current work, effective dephosphorisation of Western Australian iron has been demonstrated. Sulphuric acid was chosen as the leachant on the basis of its availability and low cost. The iron ore sample used in this study typically contained 0.126% phosphorus, was from the Pilbara region of Western Australia. After roasting at 1250°C, lump ore (P8o 5.6 mm), pellet 1 (grinding to 100% -1.5 mm before pelletisation) and pellet 2 (grinding to 100% -0.15 mm before pelletisation) were leached in solutions with different sulphuric acid concentrations. After leaching for 5 hours at 60°C in 0.1 M sulphuric acid solution, 67.2%, 69.0% and 68.7% of the phosphorus was leached from the above three samples, respectively. The phosphorus content was reduced from 0.126% to 0.044%, 0.055% and 0.042% respectively. The dissolution of iron during leaching was negligible. The optimum sulphuric acid concentration was 0.1 M in terms of acid cost and iron loss. The acid consumption cost is as low as $A 0.47/tonne. © 1999 Elsevier Science Ltd. All rights reserved.
Keywords Iron ores; leaching; hydrometallurgy
INTRODUCTION Australia is the largest exporter of iron ore in the world. In 1997, approximately 147 million tonnes of Australian iron ore with a total value of 3 billion Australian dollars were exported. Ninety five percent of Australia's iron ore resources are located in the Pilbara region of Northwest Western Australia. The known Western Australian iron ore resources comprise 34 billion tonnes, which represents six percent of the total identified world iron ore resources. From 1966 to 1990, 1.5 billion tonnes of Western Australian iron ore
1083
1084
C.Y. Chenget al.
have been produced at a total value of A$23.7 billion (O'Leary, 1993). Only 10% of Western Australian iron ore resources are premium grade low phosphorus iron ore. At the current rate of export, premium grade resources will be depleted in 30 years. The remaining high iron grade ore has a high phosphorus content and is unsuitable for iron and steel production without dephosphorisation treatment. Over 90% of Western Australian high grade iron ore has a phosphorus content in excess of 0.05% and the average phosphorus content is 0.15%. At present, the acceptable phosphorus content in iron ore for some foreign companies is 0.075%. For every increase of 0.001% in phosphorus content, it attracts a penalty of $A 0.80/tonne. This means that for the ore containing 0.080% phosphorus, the penalty would be around $A 4.0/tonne. The acceptable phosphorus content in iron ore for some other foreign companies is 0.070%. For every increase of 0.001% in phosphorus content, it attracts a penalty of $A 0.10/tonne. For the iron ore containing over 0.080% phosphorus, every increase of 0.001% in phosphorus content, it attracts a penalty of $A 1.00/tonne. This means that for the ore containing 0.083% phosphorus, the penalty would be around $A 4.0/tonne. These penalties reduce the profit margin substantially and make its export virtually impossible. The successful development of a hydrometallurgical process to reduce the phosphorus content in Western Australian iron ore would extend the export of high grade iron ore with low phosphorus content by over 100 years. Since currently there are no facilities for the dephosphorisation of iron ore in Australia, the development of a successful and economically viable technology is of paramount importance. The dephosphorisation of high phosphorus iron ore has been researched in a number of countries, and a variety of dephosphorisation processes have been developed for different phosphorus compounds. Forssberg and Adolfsson (1981) investigated the economic feasibility of acid leaching of iron ore in which apatite, Ca5(PO4)3(F,C1,OH), was the phosphorus-containing mineral. Nitric and hydrochloric acid leaching agents were used. It was found that the economy of leaching with the aid of hydrochloric acid or alternatively by means of nitric acid depended very much on the extent to which the phosphorus can be recovered as phosphoric acid. The dephosphorisation of iron ore using acid leaching was also investigated by Matsuo e t al. (1980). The iron ore contained phosphorus in the form of apatite, Cas(PO4)3 F. The acids tested were hydrochloric, nitric and sulphuric acid. The optimum acid concentration was determined to be between 2-7%. The optimum leaching time was found to exceed 10 hours. Without the use of heat treatment, a 20% reduction in the phosphorus content was achieved. With heat treatment, a reduction of phosphorus of up to 40% was obtained. Unlike iron ore in other regions of the world, phosphorus in Western Australian iron ore exists within goethite, FeOOH, in the form of solid solution (Graham, 1973). The distribution of phosphorus throughout the goethite crystals prevents the effective use of physical separation techniques, and necessitates the use of chemical separation techniques (Gooden et al., 1979). Phosphorus is less soluble in hematite than in goethite. Goethite can be converted to hematite by roasting, which will cause the ore to be more amenable to leaching. The degree of conversion of goethite to hematite is expected to increase with the roasting temperature and time. The cost of roasting also increases with roasting temperature and time, and it is desirable that the use of a thermal pre-treatment stage be avoided. Peixoto (1991) investigated the effect of heat treatment on the acid leaching of iron ore. When the ore contained phosphorus in solid solution in the goethite phase, it was found that thermal treatment of the ore at 1200°C caused a structural re-arrangement of goethite which converted to hematite and facilitated the dissolution of phosphorus in mineral acids. Without thermal treatment of the ore, hydrochloric acid was not effective. After pelletisation, the leaching with hydrochloric acid achieved a 34% reduction in phosphorus. After sintering, a 23% reduction in phosphorus was obtained. No loss of iron was observed. The dephosphorisation of iron ore was conducted by roasting in the presence of alkaline earth metal halides, particularly calcium chloride, or halides of ammonia, manganese, zirconium, copper and lithium (Feld e t al., 1966). Acid leaching was also conducted. The ore was roasted for 2 hours at temperatures between 500°C and 1200°C. The optimum temperature was determined to be 900°C. The most economical salt was
Dephosphorisation of Western Australian iron ore by hydrometallurgicalprocess
1085
found to be CaCI 2, and the optimum leaching temperature was 30°C. Within 30 minutes the phosphorus content in the ore was reduced by around 23% using sulphuric, nitric and hydrochloric acid. The dephosphorisation of high phosphorus ore from the Pilbara region of Western Australia was investigated by Gooden e t al. (1979). Physical beneficiation methods failed as the phosphorus was in solid solution in goethite. Chemical dephosphorisation methods for iron ore were reviewed and tested on the ore samples from the Pilbara region. Leaching without thermal treatment was found to be quite ineffective. The effect of heat treatment, with and without the addition of salts, was investigated. Experiments demonstrated that in a fairly elevated temperature range, salt was not required for solubilisation, and it actually hindered the dephosphorisation process. The effect of roasting temperature on dephosphorisation was investigated, using a roasting time of 0.5 hours. With heat treatment and leaching with hydrochloric, nitric and sulphuric acid and sodium hydroxide solutions, the phosphorus content in the ore could be reduced from 0.12% to 0.05%. Sulphuric acid was extensively tested as this was the most economical acid. Using sulphuric acid, it was found that the quantity of iron leached was dependent on the solution concentration, and independent of the leaching time. The minimum sulphuric acid concentration required to dissolve phosphorus was found to be 0.2 M. The optimum leaching time was found to be between 0.1 and 0.5 hours. Increasing the leaching temperature above 50°C or 60°C did not improve the leaching of phosphorus. However, leaching at temperatures above 70°C to 90°C was found to cause ferric phosphate to precipitate at much lower iron and phosphate concentrations than at temperatures below 70°C, improving the ease of phosphorus recovery. Sodium hydroxide was found to be equally effective as sulphuric acid at concentrations around 0.2 M. However, the process was not as cost effective as sulphuric acid. A conceptual plant design was provided, using a continuous process with leaching agent regeneration and phosphorus recovery. Sun and Yang (1997) indicated that after leaching of the high phosphorus iron ore with Thiobacillus followed by a chemical treatment using a solution containing tannin, the phosphorus content dropped to less than 0.2%. However, the original phosphorus content was not mentioned in the patent abstract. Another researcher in China performed dephosphorisation work using techniques such as bacteria leaching with chemical treatment and nitric acid leaching (Wang, 1997). In the bacteria leaching with Thiobacillus and chemical treatment using RL-1, the phosphorus dropped from 0.36% to 0.18%. In the nitric acid leaching process, the phosphorus content was reduced to less than 0.05% Based on the previous research, the following are the observations on the dephosphorisation of Western Australian iron ore: Roasting is necessary, Leaching with mineral acid is effective alter roasting, and The use of minimum acid concentration and cheap mineral acid makes the process economically viable.
EXPERIMENTAL Approximately 570 kg of high phosphorus iron ore from the Pitbara region of Western Australia was provided by an Australian mining company. Chemical analysis (Table 1) indicated that the phosphorus content of the ore is high at 0.126%. This ore is considered to be representative of Western Australian high phosphorus iron ore. The entire 570 kg of iron ore sample was dried, mixed, and sampled using a rotary sample divider.
TABLE 1 Chemical composition of the iron ore sample
LOI
Content
(%)
(%)
Fe
P
S
Si02
A1203
MgO
Mn
CaO
K20
TiO2
4.01
62.1
0.126
0.014
4.15
2.68
0.10
0.06
0.06
<0.01
0.11
1086
C.Y.
C h e n g e t al.
Size analysis was performed to determine the size distribution of the lump ore as received (Figure 1). It was found that the lump ore sample was 100% - 2 0 mm with 80% - 5 . 6 mm and 30% - 1 . 5 mm. A b o u t 2 kg lump ore was ground in a Braun pulverising disc mill to 100% - 1 . 5 m m which consisted of the raw material for pellet type 1 and another 2 kg to 100% - 0 . 1 5 m m which consisted of the raw material for pellet type 2.
IO0.O0
J
90.00 80.00 ~"
70.00 6o.oo
~,
50.00 40.00
r)
30.00 20.00 I0.00
•
J 0,00 i 0.053
0.090
0.150
0.250
1.5o0 Size
2.500
5.6oo
8.000
I 1.200
20.000
(ram)
Fig.l Size distribution of the lump ore sample. Pellets were prepared in a 0.45 m diameter motor driven pelletising machine. It was observed that smaller pellets of narrower size range and higher sphericity were obtained for the sample ground to 100% - 1 . 5 m m than for the ore ground to 100% - 0 . 1 5 ram. The pellet size was set as 10 m m and the average pellet size was 9.4 and 10.8 m m for pellets 1 and 2, respectively. The lump ore sample and the two pellet samples were batch roasted separately in an electric furnace for one hour in an open cylindrical container of diameter 200 m m and height 50 mm. The sample was placed in and removed from the furnace when the furnace was at room temperature. The furnace was heated from room temperature to 1250°C in 6 hours. After holding at that temperature for one hour, the furnace was cooled from 1250°C to room temperature in 17 hours. In this way, the cracking of the lumps and pellets by vaporisation of the residual moisture in the sample was minimised. The overall increase in internal porosity after roasting was determined to be 1.1% for the lump ore, 16.7% for pellet 1 and 26.9% for pellet 2. The roasted samples were then riffled to six sub-samples which weighed about 200 g each. The leach solution and ore were heated separately to 60°C and the leaching started when the liquid and solids were poured into a 500 ml conical flask. The slurry contained 40% solids by weight. The conical flask was placed on a Cenco hotplate shaker. The leaching temperature was controlled at 60_+2°C. The conical flask was loosely sealed with a rubber stopper to avoid excessive water loss and allow inverse shaking of the slurry in the flask. The flask was inverted manually every 5 minutes, and in the first 10 minutes of leaching it was shaken every minute. Six tests were conducted for each of the three samples. Three acid concentrations and four leaching times were used and the parameters for the 18 tests are listed in Table 2. No intermediate samples were taken since it was very difficult to take representative samples when the particle size was large and the amount of solids was limited. The four tests leached for 10, 60, 120 and 300 minutes formed one set of tests.
Dephosphorisation of Western Australian iron ore by hydrometallurgical process
1087
TABLE 2 Dephosphorisation leaching test conditions Test
Sample
Acid concentration
Leaching time
No
name
(M)
(min)
1
Lump
0.10
10
2
Lump
0.10
60
3
Lump
0.10
120
4
Lump
0.10
300
5
Lump
0.30
300
6
Lump
0.50
300
7
Pellet 1
0.10
10
8
Pellet 1
0.10
60
9
Pellet 1
0.10
120
10
Pellet 1
0.10
300
11
Pellet 1
0.30
300
12
Pellet 1
0.50
300
13
Pellet 2
0.10
10
14
Pellet 2
0.10
60
15
Pellet 2
0.10
120
16
Pellet 2
0.10
300
17
Pellet 2
0.30
300
18
Pellet 2
0.50
300
At the end of each test, the slurry was vacuum filtered and a solution sample was taken for chemical assay. The acid concentration in the leachant was determined by titration with 0.1 M sodium hydroxide. The solids were washed three times with a total of 700 ml distilled water. The residue was dried at 110°C for 16 hours. The residue was then ground in the Braun pulverising disc mill to 100% - 1 5 0 lam and split to obtain 15 g sub-sample for chemical assay by SGS Australia Limited.
RESULTS AND DISCUSSION
Leaching kinetics Figure 2 shows the phosphorus leaching kinetics over 5 hours at 60°C in 0.1 M sulphuric acid solution. About 45% of the phosphorus was leached in the first 10 minutes from the lump ore, largely due to the presence of fines in the solids. After leaching for 60 minutes, the same amount of phosphorus dissolved from the two pellet samples tested. After leaching for 5 hours, the phosphorus extraction was 67.2%, 59.0% and 68.7% for lump ore, pellet 1 and pellet 2 samples, respectively. The highest phosphorus dissolution in 5 hours for pellet 2 was attributed to the high porosity of the pellets, which increased by 26.9% after roasting. The high phosphorus dissolution from the lump ore was attributed to the amount of fines in the sample.
1088
C.Y. Cheng et al.
I00.00
90.00
80.00
70.00
._--&
60.00
...............
4
50.00
40.00
e~
g,
30.00
j
•
~/.'~,o' ° ,,
j ..
20.00
Lumps
- - 4. - -Pellet 1
,.o
- - - A - - Pellet 2
10.00
0.00 50
150
100
200
250
300
350
Time (rain)
Fig.2 Leaching of roasted samples in 0.1 M sulphuric acid solution. Figure 3 shows the phosphorus reduction during leaching. After 5 hours of leaching, the phosphorus content in the three samples was reduced to 0.044%, 0.055% and 0.042%, for the lump ore, pellet 1 and pellet 2, respectively. Again, the lowest phosphorus content in pellet 2 was attributed to the highest porosity in the sample.
0.16
014
0.12
k
= Lump.]
'..'.,, 0.1 c
ea *d
""..72,.. ,.:Z~....7..?.~..S_ ........ ~ ....
00s
-----p~ml~t21
" 0.06
0.04
0.02
0
'
'
20
40
60
80
100
120
140
160
180
200
220
240
260
280
300
Time (rain)
Fig.3 Phosphorus reduction during leaching in 0.1 M sulphuric acid solution. Compared with the other processes reviewed, the current process was very effective in terms of phosphorus extraction. Over 60-70% of the phosphorus was extracted which was much higher than the other researcher's results. For instance, Peixoto [6] obtained 34% phosphorus reduction after leaching roasted iron ore with goethite mineralisation.
Dephosphorisation of Western Australian iron ore by hydrometallurgical process
1089
Effect of acid concentration Table 3 shows the effect of acid concentration on the phosphorus extraction. As expected, the higher the acid concentration, the higher the phosphorus extraction except for pellet 2 which showed the highest phosphorus extraction in 0.3 M acid concentration. This was most probably caused by experimental error. After leaching for 5 hours at 60°C in 0.5 M sulphuric acid solution, about 80%, 63% and 70% of the phosphorus was extracted from the lump ore, pellet 1 and pellet 2, respectively. Accordingly, the phosphorus in the samples decreased to 0.027%, 0.049% and 0.041%, respectively (Table 4).
TABLE 3 Phosphorus extraction (%) after leaching for 5 hours in different acid concentrations Sample
Acid
concentration
(M)
Name
0. I0
0.30
0.50
Lump ore
67.16
74.63
79.85
Pellet 1
58.96
62.69
63.43
Pellet 2
68.66
75.37
69.40
TABLE 4 Phosphorus remaining (%) after leaching for 5 hours in different acid concentrations Sample
Acid
concentration
(M)
Name
0.10
0.30
0.50
Lump ore
0.044
0.034
0.027
Pellet 1
0.055
0.050
0.049
Pellet 2
0.042
0.033
0.041
Compared with the other processes reviewed, the current process was very effective in terms of acid concentration. Gooden et. al. (1979) pointed out that the minimum acid concentration was 0.2 M. In the current process, 0.1 M sulphuric acid concentration was strong enough to reduce the phosphorus content in the iron ore to the required level of 0.05%.
Iron loss during leaching Figure 4 shows the iron loss during leaching in 0.1 M sulphuric acid solution at 60°C. Due to existence of fines, the iron loss in the lump ore was the highest: about 0.06% iron loss in the first 10 minutes and 0.16% after 5 hours of leaching. The iron loss in the two pellet samples was negligible: only about 0.01% iron dissolved in the 0.1 M acid solution after 5 hours of leaching.
TABLE 5 Iron loss (%) after leaching for 5 hours in different acid concentrations Sample
Acid
concentration
(M)
Name
0.10
0.30
0.50
Lump ore
0.157
0.161
0.199
Pellet 1
0.012
0.017
0.018
Pellet 2
0.007
0.025
0.022
1090
C.Y.
et al.
Cheng
The effect of acid concentration on the iron loss is shown in Table 5. The iron loss for the 3 samples was 0.20%, 0.018% and 0.022%, respectively, indicating that the iron loss in the lump ore was about ten times of that in the pellet samples.
0.20
0.18
0.16
0.14
0.12
/
/ f
/ 0.10
Lumps
• 0.08
- .-A-- • Pellet I 0.06
- - • - - Pellet 2
0.04
0.02 j.
~
...........
20
40
~
..............
.-2.."2.'7..-..'~. . . . . . . . . . . . . . . . . . . . . . .
--2 -"~--'- --'- ~-~-i --2 . . . . .
~. . . . . . . . . . . . . . . . . . . . .
0.00 0
60
80
100
120
140
160 Time
180
200
220
240
260
280
300
320
(rain)
Fig.4 Iron loss during leaching in 0.1 M sulphuric acid solution.
Leaching mechanism Various theories have been proposed to explain the existence of phosphorus in goethite. However, none has been able to explain the dephosphorisation leaching of goethite. Graham (1973) proposed that elemental phosphorus existed in solid solution with goethite. Dukino [10] has disputed this theory on the basis that interstitial solid solution is likely only if the substituting cation is of a similar or smaller size than the lattice bound cation and its charge is the same. The size and valency of phosphorus are not suitable for its existence in solid solution with goethite, and it is considered unlikely that elemental phosphorus exists in solid solution with goethite. A more plausible mechanism has been proposed by Morris (1973) and Barbour (1973) and supported by Dukino (1997), involving surface adsorption. It has been suggested that, prior to the dehydration of ferrihydrate to goethite, a surface hydroxyl group is replaced by a phosphate ligand as shown in the following model:
3 O H q - PO43-
I Fe--O
=
O=Fe +
PO 3-
I
Fe + 11 0
Fe+~O
+ 3OH(1)
This theory is considered to be a likely explanation of the existence of phosphorus in goethite. Furthermore, this theory can be used to explain dephosphorisation. Roasting of the iron ore causes the goethite to dehydrate to hematite, liberating the phosphorus in an acid soluble compound, for example, to aluminium phosphate as shown in equation 2:
Dephosphorisation of Western Australian iron ore by hydrometallurgicalprocess 2(FeO)3PO 4 + A1203 = 2AIPO 4 + 3Fe203
1091 (2)
The chemical analysis of the ore indicated that several acid soluble phosphate compounds are possible, including aluminium phosphate, manganese phosphate and magnesium phosphate. The total mass of ore dissolved during leaching (up to 1.4%) does not correspond to the quantity of any of the components in the ore. Therefore it is likely that more than one type of acid soluble compound is formed.
Acid consumption cost The percentage of acid consumption was the highest for leaching with 0.1 M acid: 40-50% compared to 23-40% for 0.3 M acid and 25-35% for 0.5 M acid. However, the overall acid consumption was the highest for leaching with 0.5 M: triple that of leaching with 0.1 M acid. Acid consumption for leaching with 0.3 M acid was double that of leaching with 0.1 M acid. Acid consumption critically affects the process economics, and therefore the minimum acid concentration necessary to achieve leaching of phosphorus to <0.05% is desired. Therefore using 0.1 M acid was the most economical in terms of acid consumption. The acid consumption for leaching using 0.1 M acid was approximately 0.67 tonnes of 98% acid per 100 tonnes of lump ore, and 0.83 tonnes of acid per 100 tonnes of pellet 1 (grinding to 100% -1.5 m m before pelletisation). Sulphuric acid is available from Coogee Chemicals in WA for $A 70/tonne. The cost of acid consumption for dephosphorisation is $A 0.47/tonne of lump ore and $A 0.58/tonne of pellet 2 (grinding to 100% -1.5 mm before pelletisation). Therefore, the cost of acid consumption is reasonable. To estimate the total capital and operation costs of the whole dephosphorisation process including grinding, roasting, pelletisation, leaching, washing and drying is impossible at this stage. However, it is clear that the dephosphorisation plant has to be integrated with a DRI plant where the use of pellets is preferable. Therefore, the cost of the dephosphorisation process can be absorbed by the DRI plant and become economical viable.
C O N C L U S I O N S AND R E C O M M E N D A T I O N S It has been demonstrated that the current process was very effective for the dephosphorisation of Western Australian iron ore. After roasting of the lump ore and the pellets at 1250°C and leaching in 0.1 M sulphuric acid solution at 60°C for 5 hours, the phosphorus extraction was 67.2%, 59.0% and 68.7% for lump ore, pellet 1 (grinding to 100% -1.5 m m before pelletisation) and pellet 2 (grinding to 100% -0.15 mm before pelletisation) samples, respectively. The phosphorus content in the three samples was successfully reduced from 0.13% to 0.044%, 0.055% and 0.042%, respectively, which met the requirement of high grade iron ore for iron making. The higher the acid concentration of the leach solution, the higher the phosphorus extraction. After leaching for 5 hours at 60°C in 0.5 M sulphuric acid solution, about 80%, 63% and 70% of the phosphorus was extracted from the lump ore, pellet 1 and pellet 2 samples, respectively. Accordingly, the phosphorus in the samples decreased to 0.027%, 0.049% and 0.041%, respectively. Due to the presence of fines, the iron loss in the lump ore was the highest: about 0.06% iron loss in the first 10 minutes and 0.16% after 5 hours of leaching. The iron loss in the two pellet samples was negligible. The iron loss increased with the increasing of acid concentration. From an economical point of view, 0.1M sulphuric acid solution is the best choice to obtain reasonable phosphorus extraction with minimum acid consumption cost. The acid consumption cost was as low as A$0.47/tonne for lump ore and A$0.58/tonne for pellet 2. The surface adsorption theory can be used to explain the existence of phosphorus in goethite and the
1092
C.Y. Chenget al.
dephosphorisation process. Roasting of the iron ore causes the goethite to dehydrate to hematite, liberating the phosphorus in an acid soluble form. The residual sulphate in the leached iron lump ore or pellet may affect the total sulphur content, which should be taken into account in the next stage of research work.
ACKNOWLEDGMENTS The authors would like to thank the staff and management of Mineral Processing Laboratory, Department of Minerals and Energy, Western Australia, for the support of the project. Thanks are extended to the School of Chemical Engineering, Curtin University of Technology, for the provision of funds for the chemical analysis of the iron ore samples.
REFERENCES
Barbour, A.R., Distribution of phosphorus in iron ore deposit of Itabira, Minas Gerais, Brazil, Economic Geology, 1973, 68 (1). Dukino, R., Phosphorus in Hamersley Range iron ore, BHP Internal Report, 1997. Feld, I.L., Franklin, T.W. and Lampkin, W.M., Process for removing phosphorus from iron ores, United States Patent, 1966, No 3,402,041. Forssberg, R. and Adolfsson, G., Dephosphorisation of high-phosphorus iron ores by means of acid leaching, Erzmetal, 1981, 34, 316-322. English translation by BHP Central Research Laboratories, No. CRL/T 13278. Gooden, J.E.A., Walker, W.M. and Allen, R.J., 'AMDEPHOS'--A chemical process for dephosphorisation of iron ore, Proceedings of National Chemical Engineering Conference, Queensland, 1974, 21-33. Graham, J., Phosphorus in iron ore from the Hamersley iron formations, Proceedings of the Australasian Institute of Mining and Metallurgy, 1973, No. 246, 41-42. Matsuo, S., Ikeda, R. and Inaga, S., Method of dephosphorising ore, Japanese Patent 57-67136, English translation by BHP Central Research Laboratories, 1980, No. CRL/T 13278. O'Leary, M.A., Overview of the iron ore industry: Twenty-five years of iron ore developments in Australia, in Australasian Mining and Metallurgy, 2nd edn. The Institute of Australasian Mining and Metallurgy, 1993, 1, 231-237. Peixoto, G., Improvement of the reduction process in P content and other gangues in iron ore and its agglomerates, International patent, 1991, No 93/10271. Sun, D. and Yang Y, Method for microbial-chemical dephosphorisation of iron ore, Chinese patent, 1997, No CN 1107518 A. Wang, W.X., Research background of the Mineral Processing Group at Wuhan Yejin University of Science and Technology, Private communication, 1997. Morris, R.C., A pilot study of phosphorus distribution in parts of the Brockman iron formation, Hamersley Group Internal Report, 1973.
C o r r e s p o n d e n c e on papers published in M i n e r a l s Engineering is invited, preferably by e-mail to b w i l l s @ m i n - e n g . c o m , or by Fax to +44-(0)1326-318352