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The use of oxygen in high pressure acid leaching of nickel laterites
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Minerals Engineering 21 (2008) 533–538 This article is also available online at: www.elsevier.com/locate/mineng
The use of oxygen in high pressure acid leaching of nickel laterites B.K. Loveday * School of Chemical Engineering, University of KwaZulu-Natal, Durban 4000, South Africa Received 21 August 2007; accepted 7 November 2007 Available online 21 February 2008
Abstract Preliminary leaching tests on a sample of limonitic laterite at 220 °C, gave satisfactory recoveries of nickel. It was noted that a significant proportion of the iron remained in solution (about 11 g/l) due to its presence as divalent iron. This was equivalent to about 55 kg per ton of ore or 14% of the iron. The use of oxygen in the leach was investigated as a means of precipitating this iron as hematite and a few tests with oxygen were performed. Reaction with manganese dioxide was used subsequently, as a means of quantifying oxygen consumption and the ferrous content of the ore. A series of tests at various acid additions was used to test the benefit of oxidizing conditions. The concentration of iron in solution was reduced to about 1 g/l when oxidizing conditions were used, resulting in a calculated saving in acid consumption of about 88 kg/t. The data confirmed that the saving in acid consumption was of this order. The oxygen consumption of the ore, based on conversion of manganese dioxide, was very variable, with an average value of about 15 kg/t. This is about twice the calculated value (of 7 kg/t) based on reduction of iron in solution. This could imply that a similar proportion of ferrous iron was present in the residue. The use of a high temperature (270–280 °C), for precipitation of ferrous compounds, may not necessarily reduce acid consumption, if these compounds contain sulphate or if they re-dissolve on cooling, as filtration at high pressure and temperature is required. The use of oxygen in the leach offers the potential of reducing acid consumption and eliminating unstable precipitates associated with ferrous compounds. Ó 2007 Elsevier Ltd. All rights reserved. Keywords: Laterites; Nickel; Leaching; Oxygen; Acid
1. Introduction About 60% of current worldwide production of nickel is derived from sulphide ores, but about 70% of the nickel reserves are present in laterite deposits (Dalvi et al., 2004). Hence, there is growing interest in improving the technology for extraction of nickel (and cobalt) from laterites, as the demand for stainless steel continues to grow and the sulphide deposits become depleted. Laterites are iron-rich layers in soils derived from prolonged mechanical and chemical weathering of ultramafic (ultra-basic) rocks, in a tropical climate. Some laterites are classified as bauxites, in view of their gibbsite content, but this paper will be focussed on laterites with a significant *
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concentration of nickel (and associated cobalt). The mechanisms by which the iron and other metals are mobilised and concentrated in lateritic layers is the subject of ongoing debate. It should be noted that ultra-basic rocks contain alkaline species and these rocks react with water to produce an alkaline pH. However, the presence of oxygen and carbon dioxide in rainwater over a prolonged period, makes the picture complicated. Table 1 shows that the solubility of trivalent iron is negligible at a pH value of 7, and thus iron mobility has been linked to the presence of organic compounds and bacteria, which can form complexes with the both ferric and ferrous ions (as in our blood). It should also be noted from Table 1 that nickel and cobalt have relatively high solubility limits at 25 °C and it is generally assumed that these metals were coprecipitated with trivalent iron. Divalent iron has a similarly high solubility and it can also be co-precipitated with
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B.K. Loveday / Minerals Engineering 21 (2008) 533–538
Table 1 Solubility limits of selected metals at 25 °C, for metal hydroxides (solubility data from Jackson, 1986) Solubility limits at pH 7 (metal concentration in g/l) Fe++ Ni++ Co++ Mg++ ++
Mn
4.4 3.7 7.4 >100 >100
FeOH++
1016
Co+++
1022
Al+++ Mn+++
1020 1040
trivalent iron. The solubility limits of magnesium, aluminium and manganese are also shown, as various compounds containing these metals are also present in laterites (for later reference). The form of the precipitate in a typical nickel laterite changes with depth and the zones are usually classified, from top to bottom, as follows: ferricrete, limonite, nontronite, saprolite and altered peridotite (Dalvi et al., 2004). The limonite and saprolite layers usually have higher nickel concentrations, (in climates which are still tropical), and these zones would be targeted. Limonite contains more than 40% iron by mass, which is mainly goethite (a-FeOOH). Saprolite, which is also mainly goethite, contains a significant amount of serpentine (Mg3Si2O5(OH)4. Divalent iron is present in the serpentine as a substitute for magnesium (Rubisov and Papangelakis, 2000) and hence dissolution of the ferrous iron and magnesium in saprolite, results in an increased acid consumption, as these ions are not easily precipitated. The iron content of laterites can be utilised if the ore is smelted directly to form an iron/nickel alloy. However, the fine particle size, the high proportion of slag forming elements and the cost of drying (in a tropical climate) make this option unattractive in most cases. High pressure acid leaching (HPAL), on the other hand, fits in well with a fine, moist feed and it is considered to the best option in many cases for selective dissolution of nickel and cobalt, as the iron is re-precipitated as hematite. The existing users of the HPAL process use a pulp containing 25–35% solids, a residence time of 30–90 min, temperatures ranging from 250 to 270 °C (corresponding to gauge pressures of 38–54 bar), and sulphuric acid consumptions are typically 200– 520 kg/dry ton (Moskalyk and Alfantazi, 2002). A free acid concentration of about 20–30 g/l H2SO4 is used and a significant proportion of the sulphate remains in the residue as hydrated metal sulphates (Rubisov and Papangelakis, 2000). Ferric iron can be precipitated as hematite at temperatures as low as 120 °C, but the use of a much higher temperature is required to dissolve the goethite (and associated nickel and cobalt) at an acceptable rate and to precipitate aluminium (as (H3O)Al3(SO4)2(OH)6 or AlOHSO4) and magnesium (as MgSO4 H2O). The sulphate content of these precipitates should be noted together with the fact that significant re-dissolution occurs when the acidic liquor cools.
The use of temperatures in the range 250–280 °C has also been justified by a saving in acid associated with iron dissolution. However, the data published by Rubisov and Papangelakis (2000) on limonite leaching contradicts this view. As temperature was increased from 230 to 270 °C (with a fixed acid-to-ore ratio), the concentration of iron and aluminium was decreased substantially as expected, but the acid concentration in the reactor also decreased marginally, indicating an increase in acid consumption. This could be due to an increase in the proportion of basic ferric sulphate (FeOHSO4). In view of the remote location of many of the laterite deposits, transport of sulphur, rather sulphuric acid, may be considered. The sulphur could be converted to sulphuric acid locally (by reaction with air) and some of the exothermic heat could be used to pre-heat the leach pulp by heat exchange. An alternative approach is to add the sulphur to the pulp and to inject locally generated oxygen into the leach vessels, to provide in situ heating, in combination with the generation of sulphuric acid. A further option would be to use oxygen and a low grade (high sulphur) nickel sulphide concentrate as a source of sulphur, thereby providing a synergistic means of recovering the nickel in the concentrate. Injection of oxygen into the leach provides a means of converting ferrous compounds into hematite, thereby reducing the amount of sulphate-containing precipitate, the associated acid consumption and environmental problems associated with unstable precipitates. It should be noted that oxidation of ferrous sulphate is used to co-precipitate other metals. It is also possible that the rate of dissolution of the nickel-containing minerals may be improved by oxidising conditions. This paper describes a preliminary investigation of the effect oxidising conditions on pressure leaching of limonite. 2. Experimental Recent published laboratory test work on the HPAL process has been done in batch autoclaves at 230 to 270 °C, with in situ filtration for removal of liquid samples (Rubisov et al., 2000). This method produced results that were technically correct for research purposes, but the overall result may not be achievable in practice, if the precipitates re-dissolve as the pulp cools. Large-scale solid– liquid separation (including washing) at high temperature may not be feasible with the typical combination of a very fine ore (e.g. an average size of 1.7 lm was reported by Georgiou and Papangelakis, 1998) and an acidic solution. The use of a CCD after cooling, would be a practical solution for full-scale application, but re-dissolution of precipitates would occur. Hence, it was decided to allow the pulp to cool before separating the liquid for analysis, to depict the overall results, including re-dissolution of some precipitates. A sample of lateritic ore was provided by Anglo American PLC, which was described as a typical limonite, high
3. Preliminary tests Preliminary tests over a range of acid additions showed that nickel extraction increased with increasing acid addition, up to a maximum of about 90%, based on the head grade of 0.4% Ni. The residue was not analysed. The pH of the solution, (after cooling) decreased marginally with increasing acid addition, as shown in Fig. 1. No magne2.2
100
2
80 70
1.8
60
pH
Ni Recovery (%)
20
250
18
230
16
210
14
190
12
170
10
150
8
130
6
110 Fe Conc
4
Sulphate in Residue
2 0 100
200
300
50 1.6
40 30
1.4
20 Ni Rec
10 pH
200
300
1.2 400
Acid Consumed (kg/t) Fig. 1. Preliminary tests – nickel extraction and final pH.
90 70 50 400
Acid Consumed (kg/t) Fig. 2. Preliminary tests – iron concentration in the leachate and sulphate in residue.
sium was found in solution and the iron and nickel in solution only accounted for only a small portion of the acid consumption. Hence, most of the sulphate remained in the residue, as shown in Fig. 2 (The small amount which may have been associated with aluminium in solution at these pH values was ignored). Fig. 2 also shows that iron concentration increased in sympathy with nickel recovery, but at a much higher concentration (i.e. the maximum concentrations were about 13 g/l Fe vs. 0.7 g/l Ni). Given the pH values shown in Fig. 1 and solution potential values of about 350 mV, it was clear that the iron in solution was predominantly in the divalent form. The relatively large proportion ferrous iron in limonite was not expected. It contributes to acid consumption and higher temperatures are required for its precipitation in the leach. Ferrous iron also contributes to downstream costs in separation and neutralisation. One way of dealing with the problem is to use oxygen in the leach, to convert the iron to the trivalent state, thereby saving acid as indicated in the following equation 4FeSO4 þ O2 þ 4H2 O ¼ 2Fe2 O3 þ 4H2 SO4
90
0 100
535
SO4 in Res. (kg/t feed)
in iron and low in magnesium, with a moderate acid consumption. The metal content was given as approximately: 0.4% Ni, 0.05% Co, 40% Fe, 0.3% Mg and 0.8% Al. A Parr autoclave (2 l) was available for tests up to 220 °C (26 bar gauge), and it was decided to do preliminary tests at this temperature prior to possible upgrading of the equipment. The ‘low0 temperature route may be consistent with economic recovery of nickel from this relatively low grade ore, particularly in view of the temperature effects mentioned above. The limonite sample, which contained lumps of aggregated material, was crushed to pass a 600 lm screen and sub-sampled into 100 g lots. The 100 g of sample, water (500 ml) and acid were added, prior to closing the vessel and adding a heating mantle and lagging. A hot plate was used to boost heating capacity during pre-heating. The air was displaced by steam, prior to closing the vent and bringing the reactor up to 220 °C. The stirrer was started when the temperature reached 220 °C. The temperature was controlled at 220 °C, within a range of 2 °C, using the heating mantle alone. Most of the tests were conducted for an hour (at temperature), after which the lagging was removed and the vessel was allowed to cool overnight. The contents were then removed and filtered. The time of filtration was recorded, as an indicator of the residue’s crystallinity. The pH and Eh of the solution were recorded and the solution samples were analysed for Ni, Fe, Mg and Mn using ICP spectrometry.
Iron Conc. (g/l)
B.K. Loveday / Minerals Engineering 21 (2008) 533–538
ð1Þ
Given the experimental conditions of 0.5 l of water and 100 g of ore, the ferrous iron content of about 12 g/l is equivalent to 6% of the ore, or 15% of the iron. If all the iron was precipitated as hematite, a saving in acid consumption of 105 kg per tonne of ore would accrue. From Eq. (1), about 8.6 kg of oxygen per tonne of ore would be used to precipitate the ferrous iron as hematite. A significant reduction in the amount of potentially problematic ferrous precipitates in the residue could also accrue. A few tests were conducted, in which oxygen was added to the autoclave. The pressure regulator for the supply of oxygen was set at about 1 bar above the calculated vapour pressure of water at 220 °C. These tests demonstrated a substantial reduction in iron concentration and an increase in residual acid. However, it was not possible to measure
B.K. Loveday / Minerals Engineering 21 (2008) 533–538
4. The manganese dioxide method The use of manganese dioxide to oxidise ferrous iron to the ferric state is well known to extractive metallurgists for uranium leaching. It has also been used as a wet analytical method for determination of ferrous iron in solution. Manganese dioxide will only dissolve in acid solution by reduction to Mn++ and hence an excess of MnO2 can be used. The ferric ion acts as a catalyst for oxidation of other species and an excess of MnO2 will maintain a high ferric/ferrous ratio and hence a high solution potential. It can simulate the presence of oxygen under pressure in certain circumstances. The quantity of manganese which goes into solution in laboratory experiments is therefore a direct measure of oxygen consumption. In order to predict the acid consumption which would occur when gaseous oxygen is used, the acid consumption in the laboratory test using manganese dioxide, must be corrected by subtracting the acid associated with manganese dissolution. Eq. (2) illustrates the mass balance for the manganese dioxide method. It may be deduced from Eqs. (1) and (2) that one atom of manganese in solution is equivalent to consumption of half a mole of oxygen. 2FeO þ MnO2 þ H2 SO4 ¼ MnSO4 þ Fe2 O3 þ H2 O
ð2Þ
18 16 14 12 10 8 6
Series 1 Normal Leach Series 2 Normal Leach Series 1 Oxidative Series 2 Oxidative
4 2 0 150
200
250
300
350
400
Acid consumed [kg/tonne ore] Fig. 3. Effect of oxygen (via MnO2) on iron dissolution.
25
Oxygen Consumption (kg/t)
oxygen consumption. The limonite could contain other insoluble species which react with oxygen, and it was important to develop a method of measuring oxygen consumption in the laboratory test.
Fe Conc. (g/l)
536
20
15
10
5
0 100
200
300
400
500
Acid Consumption (kg/t)
5. Results
Fig. 4. Oxygen consumption (via manganese dissolution) vs. acid consumption.
The non-oxidative leaching tests showed that the ore contained about 0.2% soluble manganese (0.4 g/l). Hence, this background level was subtracted from the manganese concentration, prior to calculating oxygen consumption and corrected acid consumption. Experiments were conducted with varying amounts of analytical grade MnO2 and acid. It was found that an addition of 20 g of MnO2 was sufficient to provide an excess, with a maximum of about 11 g going into solution. The data in the following graphs, which refer to ‘oxidative’, were obtained using the manganese dioxide method described above. These tests produced manganese concentrations of between 8 and 14 g/l. The solution potential was increased to about 600 mV. Fig. 3 shows that the iron concentration in the leachate was reduced significantly under oxidizing conditions, as expected. The average reduction in iron concentration was about 10 g/l, which equates to a theoretical oxygen consumption of about 7 kg/t and a saving in acid consumption of about 88 kg/t. Fig. 4 shows a plot of oxygen consumption (based on net manganese dissolution) vs. (corrected) acid consumption. The oxygen consumption of the ore, (based on conversion of manganese dioxide), was very variable, with an average value of about
15 kg/t. This is about twice the value calculated above, based on reduction of iron in solution. This could imply that the residue contained other species in a reduced form, probably compounds containing a similar amount of ferrous iron. It should also be noted that the relatively small amount of soluble manganese in the ore can be oxidized by oxygen (and precipitated as MnO2), but that manganese dioxide method will not will not reflect this. Figs. 5 and 6 confirm that acid consumption (for a given nickel recovery) was reduced when oxidising conditions were used. Two series of tests were performed. The trends were similar to the early work using gaseous oxygen and both of the data sets point to an optimum acid addition, where the nickel recovery reaches a maximum. Nickel recoveries in excess of 100% are shown in Fig. 6 (Series 2). Clearly, the head grade of nickel in the sample used for the second series of tests was higher than the stated 0.4%. However, the trends were similar to those obtained in Series 1. The saving in acid consumption is difficult to assess from this type of plot and it may be better to use the change in iron concentration, as a conservative estimate. No noticeable differences were noted in the filtering times with and without oxidising conditions.
B.K. Loveday / Minerals Engineering 21 (2008) 533–538
Nickel Recovery (%)
120 100
2.
80 60 40 Oxidative Leach
20
Normal Leach
0 100
200
300
400
500
Acid Consumption (kg/t) Fig. 5. Effect of oxygen on nickel recovery vs. acid consumption (Series 1).
Nickel recovery (%)
140 120 100 80 60 Oxidative Leach Normal Leach 0.5h Normal 0.5h Oxidative
40 20 0 100
200
300
3. 400
Acid consumption (kg/t) Fig. 6. Effect of oxygen on nickel recovery vs. acid consumption (Series 2). Additional tests using a half-hour leach.
The reaction time was reduced to half an hour in two experiments (in Fig. 6), in order to test the sensitivity of the leach to kinetics. Nickel recovery was reduced by the reduced reaction time, as expected, and Fig. 6 shows that under these conditions, oxidising conditions improved the rate of leaching at comparable acid concentrations. In view of the scatter in the data in Fig. 6, more tests at reduced reaction time are needed to confirm this interpretation. 6. Conclusions 1. Initial leaching tests at 220 °C showed that the sample of limonite contained about 6% soluble ferrous iron. This contributed significantly to acid consumption and it would result in increased downstream costs. Beverskog and Puigdomenech (1996) have recently published updated Pourbaix diagrams for iron, which show that precipitation of magnetite would not occur at the typical solution potential of about 350 mV, even if the temperature was increased to 300 °C. Oxidation is needed to precipitate the iron as hematite, and hence leaching at about 280 °C may result in the precipitation of some of the divalent iron as a hydrated sulphate, with no reduction in acid consumption. Preliminary leaching tests
4.
5.
537
using gaseous oxygen showed that most of the iron in solution was precipitated, resulting in a significant reduction in acid consumption. The consumption of oxidant was measured by adding manganese dioxide and this method proved to be both simple and effective. The equivalent consumption of oxygen varied significantly, (between 8 and 20 kg/t), with an average of 14.6 kg/t. This is about twice the value 7 kg/t calculated from the reduction in soluble iron. Hence, it may be concluded that there were other reduced species in the residue. These compounds could contain sulphate ions, (for example hydrated ferrous sulphate), and their oxidation would also result in a saving in acid consumption. The total ferrous iron content of the ore could therefore be as high as 12%. There is no apparent trend of oxygen consumption vs. acid consumption in Fig. 4, and this implies that dissolution and oxidation of the reduced species does not require high acid additions. The scatter in the data indicates a significant variability in the ferrous iron content of the ore and the associated acid consumption under non-oxidative leaching conditions. If oxygen was used, oxidation of the small amount of divalent manganese (0.2%) would also be possible. Literature on high pressure acid leaching of nickel laterites has not highlighted the presence of ferrous compounds and how they are precipitated. Injection of oxygen is an effective way of converting divalent iron and manganese ions into stable oxides. Tests at 270– 280 °C with oxygen should confirm the benefit of using oxygen by reducing acid consumption and eliminating the precipitated sulphate compounds which may be associated with these species. Oxygen consumption was shown to be independent of acid consumption in tests at 220 °C. This suggests that the dissolution and precipitation of ferrous compounds occurs relatively quickly. A comparison of single halfhour tests showed that oxidising conditions may improve the rate of leaching of nickel. Further work is required to confirm this finding. The ore (limonite) had a low concentration of magnesium (0.3%) and no magnesium was found in solution. The use of oxygen is not expected to have any effect on precipitation of magnesium during pressure leaching of saprolite, but improved co-precipitation with iron may occur.
References Beverskog, B., Puigdomenech, I., 1996. Revised Pourbaix diagrams for iron at 20–300 °C. Corrosion Science 38 (2), 2121–2135. Dalvi, A.D., Bacon, W.G., Osborn, R.C. 2004. The past and the future of nickel laterites, In PDAC 2004 International Convention, Trade Show and Investors Exchange, March 7–10, 2004. Georgiou, D., Papangelakis, V.G., 1998. Sulphuric Acid Pressure Leaching of a Limonitic Laterite: Chemistry and Kinetics. Hydrometallurgy 49, 23–46.
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Jackson, E., 1986. Hydrometallurgical extraction and reclamation: Chemical precipitation processes. Springer, London, p. 157. Moskalyk, R.R., Alfantazi, A.M., 2002. Nickel laterite processing and electrowinning practice. Minerals Engineering 15, 593–605. Rubisov, D.H., Krowinkel, J.M., Papangelakis, V.G., 2000. Sulphuric acid pressure leaching of laterites – universal kinetics of nickel
dissolution for limonites and limonitic/saprolitic blends. Hydrometallurgy 58, 1–11. Rubisov, D.H., Papangelakis, V.G., 2000. Sulphuric acid pressure leaching of laterites – speciation and prediction of metal solubilities ‘‘at temperature”. Hydrometallurgy 58, 13–26.
Minerals Engineering 21 (2008) 533–538 This article is also available online at: www.elsevier.com/locate/mineng
The use of oxygen in high pressure acid leaching of nickel laterites B.K. Loveday * School of Chemical Engineering, University of KwaZulu-Natal, Durban 4000, South Africa Received 21 August 2007; accepted 7 November 2007 Available online 21 February 2008
Abstract Preliminary leaching tests on a sample of limonitic laterite at 220 °C, gave satisfactory recoveries of nickel. It was noted that a significant proportion of the iron remained in solution (about 11 g/l) due to its presence as divalent iron. This was equivalent to about 55 kg per ton of ore or 14% of the iron. The use of oxygen in the leach was investigated as a means of precipitating this iron as hematite and a few tests with oxygen were performed. Reaction with manganese dioxide was used subsequently, as a means of quantifying oxygen consumption and the ferrous content of the ore. A series of tests at various acid additions was used to test the benefit of oxidizing conditions. The concentration of iron in solution was reduced to about 1 g/l when oxidizing conditions were used, resulting in a calculated saving in acid consumption of about 88 kg/t. The data confirmed that the saving in acid consumption was of this order. The oxygen consumption of the ore, based on conversion of manganese dioxide, was very variable, with an average value of about 15 kg/t. This is about twice the calculated value (of 7 kg/t) based on reduction of iron in solution. This could imply that a similar proportion of ferrous iron was present in the residue. The use of a high temperature (270–280 °C), for precipitation of ferrous compounds, may not necessarily reduce acid consumption, if these compounds contain sulphate or if they re-dissolve on cooling, as filtration at high pressure and temperature is required. The use of oxygen in the leach offers the potential of reducing acid consumption and eliminating unstable precipitates associated with ferrous compounds. Ó 2007 Elsevier Ltd. All rights reserved. Keywords: Laterites; Nickel; Leaching; Oxygen; Acid
1. Introduction About 60% of current worldwide production of nickel is derived from sulphide ores, but about 70% of the nickel reserves are present in laterite deposits (Dalvi et al., 2004). Hence, there is growing interest in improving the technology for extraction of nickel (and cobalt) from laterites, as the demand for stainless steel continues to grow and the sulphide deposits become depleted. Laterites are iron-rich layers in soils derived from prolonged mechanical and chemical weathering of ultramafic (ultra-basic) rocks, in a tropical climate. Some laterites are classified as bauxites, in view of their gibbsite content, but this paper will be focussed on laterites with a significant *
Tel.: +27 31 260 3121; fax: +27 31 260 1118. E-mail address: [email protected]
0892-6875/$ - see front matter Ó 2007 Elsevier Ltd. All rights reserved. doi:10.1016/j.mineng.2007.11.002
concentration of nickel (and associated cobalt). The mechanisms by which the iron and other metals are mobilised and concentrated in lateritic layers is the subject of ongoing debate. It should be noted that ultra-basic rocks contain alkaline species and these rocks react with water to produce an alkaline pH. However, the presence of oxygen and carbon dioxide in rainwater over a prolonged period, makes the picture complicated. Table 1 shows that the solubility of trivalent iron is negligible at a pH value of 7, and thus iron mobility has been linked to the presence of organic compounds and bacteria, which can form complexes with the both ferric and ferrous ions (as in our blood). It should also be noted from Table 1 that nickel and cobalt have relatively high solubility limits at 25 °C and it is generally assumed that these metals were coprecipitated with trivalent iron. Divalent iron has a similarly high solubility and it can also be co-precipitated with
534
B.K. Loveday / Minerals Engineering 21 (2008) 533–538
Table 1 Solubility limits of selected metals at 25 °C, for metal hydroxides (solubility data from Jackson, 1986) Solubility limits at pH 7 (metal concentration in g/l) Fe++ Ni++ Co++ Mg++ ++
Mn
4.4 3.7 7.4 >100 >100
FeOH++
1016
Co+++
1022
Al+++ Mn+++
1020 1040
trivalent iron. The solubility limits of magnesium, aluminium and manganese are also shown, as various compounds containing these metals are also present in laterites (for later reference). The form of the precipitate in a typical nickel laterite changes with depth and the zones are usually classified, from top to bottom, as follows: ferricrete, limonite, nontronite, saprolite and altered peridotite (Dalvi et al., 2004). The limonite and saprolite layers usually have higher nickel concentrations, (in climates which are still tropical), and these zones would be targeted. Limonite contains more than 40% iron by mass, which is mainly goethite (a-FeOOH). Saprolite, which is also mainly goethite, contains a significant amount of serpentine (Mg3Si2O5(OH)4. Divalent iron is present in the serpentine as a substitute for magnesium (Rubisov and Papangelakis, 2000) and hence dissolution of the ferrous iron and magnesium in saprolite, results in an increased acid consumption, as these ions are not easily precipitated. The iron content of laterites can be utilised if the ore is smelted directly to form an iron/nickel alloy. However, the fine particle size, the high proportion of slag forming elements and the cost of drying (in a tropical climate) make this option unattractive in most cases. High pressure acid leaching (HPAL), on the other hand, fits in well with a fine, moist feed and it is considered to the best option in many cases for selective dissolution of nickel and cobalt, as the iron is re-precipitated as hematite. The existing users of the HPAL process use a pulp containing 25–35% solids, a residence time of 30–90 min, temperatures ranging from 250 to 270 °C (corresponding to gauge pressures of 38–54 bar), and sulphuric acid consumptions are typically 200– 520 kg/dry ton (Moskalyk and Alfantazi, 2002). A free acid concentration of about 20–30 g/l H2SO4 is used and a significant proportion of the sulphate remains in the residue as hydrated metal sulphates (Rubisov and Papangelakis, 2000). Ferric iron can be precipitated as hematite at temperatures as low as 120 °C, but the use of a much higher temperature is required to dissolve the goethite (and associated nickel and cobalt) at an acceptable rate and to precipitate aluminium (as (H3O)Al3(SO4)2(OH)6 or AlOHSO4) and magnesium (as MgSO4 H2O). The sulphate content of these precipitates should be noted together with the fact that significant re-dissolution occurs when the acidic liquor cools.
The use of temperatures in the range 250–280 °C has also been justified by a saving in acid associated with iron dissolution. However, the data published by Rubisov and Papangelakis (2000) on limonite leaching contradicts this view. As temperature was increased from 230 to 270 °C (with a fixed acid-to-ore ratio), the concentration of iron and aluminium was decreased substantially as expected, but the acid concentration in the reactor also decreased marginally, indicating an increase in acid consumption. This could be due to an increase in the proportion of basic ferric sulphate (FeOHSO4). In view of the remote location of many of the laterite deposits, transport of sulphur, rather sulphuric acid, may be considered. The sulphur could be converted to sulphuric acid locally (by reaction with air) and some of the exothermic heat could be used to pre-heat the leach pulp by heat exchange. An alternative approach is to add the sulphur to the pulp and to inject locally generated oxygen into the leach vessels, to provide in situ heating, in combination with the generation of sulphuric acid. A further option would be to use oxygen and a low grade (high sulphur) nickel sulphide concentrate as a source of sulphur, thereby providing a synergistic means of recovering the nickel in the concentrate. Injection of oxygen into the leach provides a means of converting ferrous compounds into hematite, thereby reducing the amount of sulphate-containing precipitate, the associated acid consumption and environmental problems associated with unstable precipitates. It should be noted that oxidation of ferrous sulphate is used to co-precipitate other metals. It is also possible that the rate of dissolution of the nickel-containing minerals may be improved by oxidising conditions. This paper describes a preliminary investigation of the effect oxidising conditions on pressure leaching of limonite. 2. Experimental Recent published laboratory test work on the HPAL process has been done in batch autoclaves at 230 to 270 °C, with in situ filtration for removal of liquid samples (Rubisov et al., 2000). This method produced results that were technically correct for research purposes, but the overall result may not be achievable in practice, if the precipitates re-dissolve as the pulp cools. Large-scale solid– liquid separation (including washing) at high temperature may not be feasible with the typical combination of a very fine ore (e.g. an average size of 1.7 lm was reported by Georgiou and Papangelakis, 1998) and an acidic solution. The use of a CCD after cooling, would be a practical solution for full-scale application, but re-dissolution of precipitates would occur. Hence, it was decided to allow the pulp to cool before separating the liquid for analysis, to depict the overall results, including re-dissolution of some precipitates. A sample of lateritic ore was provided by Anglo American PLC, which was described as a typical limonite, high
3. Preliminary tests Preliminary tests over a range of acid additions showed that nickel extraction increased with increasing acid addition, up to a maximum of about 90%, based on the head grade of 0.4% Ni. The residue was not analysed. The pH of the solution, (after cooling) decreased marginally with increasing acid addition, as shown in Fig. 1. No magne2.2
100
2
80 70
1.8
60
pH
Ni Recovery (%)
20
250
18
230
16
210
14
190
12
170
10
150
8
130
6
110 Fe Conc
4
Sulphate in Residue
2 0 100
200
300
50 1.6
40 30
1.4
20 Ni Rec
10 pH
200
300
1.2 400
Acid Consumed (kg/t) Fig. 1. Preliminary tests – nickel extraction and final pH.
90 70 50 400
Acid Consumed (kg/t) Fig. 2. Preliminary tests – iron concentration in the leachate and sulphate in residue.
sium was found in solution and the iron and nickel in solution only accounted for only a small portion of the acid consumption. Hence, most of the sulphate remained in the residue, as shown in Fig. 2 (The small amount which may have been associated with aluminium in solution at these pH values was ignored). Fig. 2 also shows that iron concentration increased in sympathy with nickel recovery, but at a much higher concentration (i.e. the maximum concentrations were about 13 g/l Fe vs. 0.7 g/l Ni). Given the pH values shown in Fig. 1 and solution potential values of about 350 mV, it was clear that the iron in solution was predominantly in the divalent form. The relatively large proportion ferrous iron in limonite was not expected. It contributes to acid consumption and higher temperatures are required for its precipitation in the leach. Ferrous iron also contributes to downstream costs in separation and neutralisation. One way of dealing with the problem is to use oxygen in the leach, to convert the iron to the trivalent state, thereby saving acid as indicated in the following equation 4FeSO4 þ O2 þ 4H2 O ¼ 2Fe2 O3 þ 4H2 SO4
90
0 100
535
SO4 in Res. (kg/t feed)
in iron and low in magnesium, with a moderate acid consumption. The metal content was given as approximately: 0.4% Ni, 0.05% Co, 40% Fe, 0.3% Mg and 0.8% Al. A Parr autoclave (2 l) was available for tests up to 220 °C (26 bar gauge), and it was decided to do preliminary tests at this temperature prior to possible upgrading of the equipment. The ‘low0 temperature route may be consistent with economic recovery of nickel from this relatively low grade ore, particularly in view of the temperature effects mentioned above. The limonite sample, which contained lumps of aggregated material, was crushed to pass a 600 lm screen and sub-sampled into 100 g lots. The 100 g of sample, water (500 ml) and acid were added, prior to closing the vessel and adding a heating mantle and lagging. A hot plate was used to boost heating capacity during pre-heating. The air was displaced by steam, prior to closing the vent and bringing the reactor up to 220 °C. The stirrer was started when the temperature reached 220 °C. The temperature was controlled at 220 °C, within a range of 2 °C, using the heating mantle alone. Most of the tests were conducted for an hour (at temperature), after which the lagging was removed and the vessel was allowed to cool overnight. The contents were then removed and filtered. The time of filtration was recorded, as an indicator of the residue’s crystallinity. The pH and Eh of the solution were recorded and the solution samples were analysed for Ni, Fe, Mg and Mn using ICP spectrometry.
Iron Conc. (g/l)
B.K. Loveday / Minerals Engineering 21 (2008) 533–538
ð1Þ
Given the experimental conditions of 0.5 l of water and 100 g of ore, the ferrous iron content of about 12 g/l is equivalent to 6% of the ore, or 15% of the iron. If all the iron was precipitated as hematite, a saving in acid consumption of 105 kg per tonne of ore would accrue. From Eq. (1), about 8.6 kg of oxygen per tonne of ore would be used to precipitate the ferrous iron as hematite. A significant reduction in the amount of potentially problematic ferrous precipitates in the residue could also accrue. A few tests were conducted, in which oxygen was added to the autoclave. The pressure regulator for the supply of oxygen was set at about 1 bar above the calculated vapour pressure of water at 220 °C. These tests demonstrated a substantial reduction in iron concentration and an increase in residual acid. However, it was not possible to measure
B.K. Loveday / Minerals Engineering 21 (2008) 533–538
4. The manganese dioxide method The use of manganese dioxide to oxidise ferrous iron to the ferric state is well known to extractive metallurgists for uranium leaching. It has also been used as a wet analytical method for determination of ferrous iron in solution. Manganese dioxide will only dissolve in acid solution by reduction to Mn++ and hence an excess of MnO2 can be used. The ferric ion acts as a catalyst for oxidation of other species and an excess of MnO2 will maintain a high ferric/ferrous ratio and hence a high solution potential. It can simulate the presence of oxygen under pressure in certain circumstances. The quantity of manganese which goes into solution in laboratory experiments is therefore a direct measure of oxygen consumption. In order to predict the acid consumption which would occur when gaseous oxygen is used, the acid consumption in the laboratory test using manganese dioxide, must be corrected by subtracting the acid associated with manganese dissolution. Eq. (2) illustrates the mass balance for the manganese dioxide method. It may be deduced from Eqs. (1) and (2) that one atom of manganese in solution is equivalent to consumption of half a mole of oxygen. 2FeO þ MnO2 þ H2 SO4 ¼ MnSO4 þ Fe2 O3 þ H2 O
ð2Þ
18 16 14 12 10 8 6
Series 1 Normal Leach Series 2 Normal Leach Series 1 Oxidative Series 2 Oxidative
4 2 0 150
200
250
300
350
400
Acid consumed [kg/tonne ore] Fig. 3. Effect of oxygen (via MnO2) on iron dissolution.
25
Oxygen Consumption (kg/t)
oxygen consumption. The limonite could contain other insoluble species which react with oxygen, and it was important to develop a method of measuring oxygen consumption in the laboratory test.
Fe Conc. (g/l)
536
20
15
10
5
0 100
200
300
400
500
Acid Consumption (kg/t)
5. Results
Fig. 4. Oxygen consumption (via manganese dissolution) vs. acid consumption.
The non-oxidative leaching tests showed that the ore contained about 0.2% soluble manganese (0.4 g/l). Hence, this background level was subtracted from the manganese concentration, prior to calculating oxygen consumption and corrected acid consumption. Experiments were conducted with varying amounts of analytical grade MnO2 and acid. It was found that an addition of 20 g of MnO2 was sufficient to provide an excess, with a maximum of about 11 g going into solution. The data in the following graphs, which refer to ‘oxidative’, were obtained using the manganese dioxide method described above. These tests produced manganese concentrations of between 8 and 14 g/l. The solution potential was increased to about 600 mV. Fig. 3 shows that the iron concentration in the leachate was reduced significantly under oxidizing conditions, as expected. The average reduction in iron concentration was about 10 g/l, which equates to a theoretical oxygen consumption of about 7 kg/t and a saving in acid consumption of about 88 kg/t. Fig. 4 shows a plot of oxygen consumption (based on net manganese dissolution) vs. (corrected) acid consumption. The oxygen consumption of the ore, (based on conversion of manganese dioxide), was very variable, with an average value of about
15 kg/t. This is about twice the value calculated above, based on reduction of iron in solution. This could imply that the residue contained other species in a reduced form, probably compounds containing a similar amount of ferrous iron. It should also be noted that the relatively small amount of soluble manganese in the ore can be oxidized by oxygen (and precipitated as MnO2), but that manganese dioxide method will not will not reflect this. Figs. 5 and 6 confirm that acid consumption (for a given nickel recovery) was reduced when oxidising conditions were used. Two series of tests were performed. The trends were similar to the early work using gaseous oxygen and both of the data sets point to an optimum acid addition, where the nickel recovery reaches a maximum. Nickel recoveries in excess of 100% are shown in Fig. 6 (Series 2). Clearly, the head grade of nickel in the sample used for the second series of tests was higher than the stated 0.4%. However, the trends were similar to those obtained in Series 1. The saving in acid consumption is difficult to assess from this type of plot and it may be better to use the change in iron concentration, as a conservative estimate. No noticeable differences were noted in the filtering times with and without oxidising conditions.
B.K. Loveday / Minerals Engineering 21 (2008) 533–538
Nickel Recovery (%)
120 100
2.
80 60 40 Oxidative Leach
20
Normal Leach
0 100
200
300
400
500
Acid Consumption (kg/t) Fig. 5. Effect of oxygen on nickel recovery vs. acid consumption (Series 1).
Nickel recovery (%)
140 120 100 80 60 Oxidative Leach Normal Leach 0.5h Normal 0.5h Oxidative
40 20 0 100
200
300
3. 400
Acid consumption (kg/t) Fig. 6. Effect of oxygen on nickel recovery vs. acid consumption (Series 2). Additional tests using a half-hour leach.
The reaction time was reduced to half an hour in two experiments (in Fig. 6), in order to test the sensitivity of the leach to kinetics. Nickel recovery was reduced by the reduced reaction time, as expected, and Fig. 6 shows that under these conditions, oxidising conditions improved the rate of leaching at comparable acid concentrations. In view of the scatter in the data in Fig. 6, more tests at reduced reaction time are needed to confirm this interpretation. 6. Conclusions 1. Initial leaching tests at 220 °C showed that the sample of limonite contained about 6% soluble ferrous iron. This contributed significantly to acid consumption and it would result in increased downstream costs. Beverskog and Puigdomenech (1996) have recently published updated Pourbaix diagrams for iron, which show that precipitation of magnetite would not occur at the typical solution potential of about 350 mV, even if the temperature was increased to 300 °C. Oxidation is needed to precipitate the iron as hematite, and hence leaching at about 280 °C may result in the precipitation of some of the divalent iron as a hydrated sulphate, with no reduction in acid consumption. Preliminary leaching tests
4.
5.
537
using gaseous oxygen showed that most of the iron in solution was precipitated, resulting in a significant reduction in acid consumption. The consumption of oxidant was measured by adding manganese dioxide and this method proved to be both simple and effective. The equivalent consumption of oxygen varied significantly, (between 8 and 20 kg/t), with an average of 14.6 kg/t. This is about twice the value 7 kg/t calculated from the reduction in soluble iron. Hence, it may be concluded that there were other reduced species in the residue. These compounds could contain sulphate ions, (for example hydrated ferrous sulphate), and their oxidation would also result in a saving in acid consumption. The total ferrous iron content of the ore could therefore be as high as 12%. There is no apparent trend of oxygen consumption vs. acid consumption in Fig. 4, and this implies that dissolution and oxidation of the reduced species does not require high acid additions. The scatter in the data indicates a significant variability in the ferrous iron content of the ore and the associated acid consumption under non-oxidative leaching conditions. If oxygen was used, oxidation of the small amount of divalent manganese (0.2%) would also be possible. Literature on high pressure acid leaching of nickel laterites has not highlighted the presence of ferrous compounds and how they are precipitated. Injection of oxygen is an effective way of converting divalent iron and manganese ions into stable oxides. Tests at 270– 280 °C with oxygen should confirm the benefit of using oxygen by reducing acid consumption and eliminating the precipitated sulphate compounds which may be associated with these species. Oxygen consumption was shown to be independent of acid consumption in tests at 220 °C. This suggests that the dissolution and precipitation of ferrous compounds occurs relatively quickly. A comparison of single halfhour tests showed that oxidising conditions may improve the rate of leaching of nickel. Further work is required to confirm this finding. The ore (limonite) had a low concentration of magnesium (0.3%) and no magnesium was found in solution. The use of oxygen is not expected to have any effect on precipitation of magnesium during pressure leaching of saprolite, but improved co-precipitation with iron may occur.
References Beverskog, B., Puigdomenech, I., 1996. Revised Pourbaix diagrams for iron at 20–300 °C. Corrosion Science 38 (2), 2121–2135. Dalvi, A.D., Bacon, W.G., Osborn, R.C. 2004. The past and the future of nickel laterites, In PDAC 2004 International Convention, Trade Show and Investors Exchange, March 7–10, 2004. Georgiou, D., Papangelakis, V.G., 1998. Sulphuric Acid Pressure Leaching of a Limonitic Laterite: Chemistry and Kinetics. Hydrometallurgy 49, 23–46.
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Jackson, E., 1986. Hydrometallurgical extraction and reclamation: Chemical precipitation processes. Springer, London, p. 157. Moskalyk, R.R., Alfantazi, A.M., 2002. Nickel laterite processing and electrowinning practice. Minerals Engineering 15, 593–605. Rubisov, D.H., Krowinkel, J.M., Papangelakis, V.G., 2000. Sulphuric acid pressure leaching of laterites – universal kinetics of nickel
dissolution for limonites and limonitic/saprolitic blends. Hydrometallurgy 58, 1–11. Rubisov, D.H., Papangelakis, V.G., 2000. Sulphuric acid pressure leaching of laterites – speciation and prediction of metal solubilities ‘‘at temperature”. Hydrometallurgy 58, 13–26.