Metallurgical processes for scandium recovery from various resources: A review

Hydrometallurgy 108 (2011) 100–108

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Metallurgical processes for scandium recovery from various resources: A review Weiwei Wang, Yoko Pranolo, Chu Yong Cheng ⁎ Parker Centre/CSIRO Process Science and Engineering, CSIRO Minerals Down Under National Research Flagship, 7 Conlon Street, Waterford, WA 6152, Australia

a r t i c l e

i n f o

Article history: Received 19 November 2010 Received in revised form 13 January 2011 Accepted 5 March 2011 Available online 17 March 2011 Keywords: Scandium Leaching Solvent extraction Separation Purification

a b s t r a c t Metallurgical processes for scandium recovery from various resources are reviewed. Scandium is mainly recovered as by-product from residues, tailings and waste liquors in the production of other metals such as rare earths, uranium, titanium, tungsten, aluminium, nickel, tantalum and niobium. Bauxite and nickel laterite ores are proposed as the most promising scandium resources for its production. Currently, the methods combined with hydro- and pyro-metallurgical processes, including ore pre-treatment, leaching, solvent extraction, precipitation and calcination, are commonly used for scandium recovery. New technologies for scandium recovery such as selective leaching and solvent extraction are possible development direction in the future. Crown Copyright © 2011 Published by Elsevier B.V. All rights reserved.

1. Introduction Scandium is the 31st abundant element in the Earth's crust, with an average crustal abundance of 22 ppm (Hedrick, 2010b). Because it readily substitutes the major elements such as iron and aluminium, scandium rarely occurs in concentrated quantities, but distributes sparsely in trace amounts in rocks consisting of ferromagnesian minerals with an abundance of 5–100 ppm. The scandium minerals containing appreciable quantities of scandium such as thortveitite, euxenite, and gadolinite are rare (Kristiansen, 2003; von Knorring and Condliffe, 1987). It also extensively co-exists in low contents in ores of aluminium, cobalt, iron, molybdenum, nickel, phosphate, tantalum, tin, titanium, tungsten, uranium, zinc and zirconium (Hedrick, 2010b). Ores with the scandium content range of 0.002–0.005% can be considered as resources of scandium and deserve exploitation (Xu and Li, 1996). America, Australia, China, Kazakhstan, Madagascar, Norway, Russia and Ukraine are the countries with main scandium resources (Hedrick, 2010a; Munnoch and Worstall, 2003). In America, scandium resources are mainly found in uranium, tantalum, aluminium and zirconium ores; in Australia in nickel laterite ores; in China in iron, tin and tungsten ores; in Kazakhstan in uranium ores; in Madagascar and Norway in pegmatite rocks; and in Russia and Ukraine in iron ores. The absence of reliable and long term production coupled with the high price of scandium has limited the commercial applications of scandium. Scandium is a rare and expensive metal because of its scarce distribution and difficulties associated with its extraction. In 2009, the

⁎ Corresponding author. Tel.: + 61 8 9334 8916; fax: + 61 8 9334 8001. E-mail address: [email protected] (C.Y. Cheng).

prices of scandium oxide (Sc2O3) in 99.0% and 99.9% purity are US$ 900/ kg and US$ 1400/kg, respectively (Hedrick, 2010a). The global scandium production is about 2 tonnes per year in the form of scandium oxide (Deschamps, 2003). Only about 400 kg of that is from primary production while the rest is from the stockpiles of Russia generated during the Cold War. In 2003, the only three existing scandium-mining mines in the world were the uranium and iron mines in Zhovti Vody in Ukraine, the rare earth mines in Bayan Obo, China and the apatite mines in the Kola Peninsula, Russia (Deschamps, 2003). Scandium and its compounds have been found wider applications such as in optical, electronic, aeronautical, automotive and transportation industries (Guo et al., 1988; Hedrick, 2010a; Irvine et al., 2004). The properties of Sc-strengthened alloys and Sc2O3-stabilised ZrO2 materials are particularly promising. The Al–Sc alloys (0.35–0.4% Sc) have a number of superior properties including light weight, high strength, good thermal resistance and long durability (Ahmad, 2003). The principal applications of the Al–Sc alloys were for sporting equipment and military demand (Hedrick, 2010a). For example, scandium alloys were applied in premium bicycle frames and in fighters and missiles. The application of Mg–Sc alloy to aircraft engines could save electricity and energy consumption (Shalomeev et al., 2008). Scandia-stabilised zirconia has extremely high oxygen-ion conductivity for use as a high efficiency electrolyte in solid oxide fuel cells (Ciacchi et al., 1991). The applications of scandium are increasing because of its specific mechanical and chemical properties over other metals, and hence there is a growing market demand. Pyrometallurgical processes are suitable for recovery of scandium from its ores with high contents of scandium. However, the energy

0304-386X/$ – see front matter. Crown Copyright © 2011 Published by Elsevier B.V. All rights reserved. doi:10.1016/j.hydromet.2011.03.001

W. Wang et al. / Hydrometallurgy 108 (2011) 100–108

consumption is intensive. Scandium is often enriched in slags, residues, tailings and waste liquors, and primarily produced as a byproduct during processing of various ores. Precipitation of insoluble scandium compounds from scandium-containing solutions is the easiest method to recover scandium. However, the co-precipitation of other metals makes it unsuitable for recovery from solutions with large amounts of impurity metals. Currently, hydrometallurgical processes, which mainly involve leaching, solvent extraction and precipitation are most commonly used for scandium recovery. The complexities of flowsheets to recover scandium depend on the different types and amounts of impurities. In this paper, the metallurgical processes for scandium recovery from various resources including scandium ores, residues, tailings and waste liquors are reviewed in consideration of selecting processes to recover scandium as a minor element and incorporating the scandium recovery process to the main flowsheet for the production of the main metal. 2. Metallurgical processes for scandium recovery from various resources 2.1. Recovery of scandium from its ores Scandium minerals with high scandium content, such as thortveitite and lolbeckite, are mainly dispersed in the thortveitite-rich pegmatites in Madagascar and Norway (Hedrick, 2010a). One of ores in Norway is a double silicate of scandium and yttrium with the formula of 2SiO2•Y2O3•Sc2O3 (Baptiste, 1959). Another ore in Madagascar appears to be a complex silicate of scandium, zirconium and aluminium with its composition shown in Table 1. The content of scandium in the ores was as high as 42.6%. By fractional sublimation based on the significant difference in the sublimation points of their anhydrous chlorides, scandium can be extracted from the two types of thorveitite ores. The finely milled ore and coal were heated up to 900–1000 °C with a current of chlorine gas passing over them. The chlorides of silicon, titanium, aluminium, iron and zirconium were sublimated as their sublimation temperature points are below 350 °C. The scandium chloride was sublimated at about 967 °C and deposited in a state of high purity in a zone where the temperature decreased to about 400 °C. The yttrium chloride remained in the residue. In the United States, scandium was recovered from thortveitite-rich mine tailings, such as the tailings of the Crystal Mountain fluorite mine near Darby, Montana (Hedrick, 2010a). As mentioned above, minerals containing appreciable quantities of scandium are rare with limited deposits being found which have not been exploited in mass amount (Hedrick, 2010a). Therefore, the recovery of scandium from other ores and waste materials is very important.

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as thorium, iron, calcium, fluoride and phosphorus are co-leached into the leach liquor. The scandium can be separated from impurities with acidic organic-phosphorus extractants such as di-(2-ethylhexyl) phosphoric acid (HDEHP) and 2-ethylhexyl phosphoric acid mono2-ethylhexyl ester (HEHEHP, PC-88A). The scandium extraction with HEHEHP was achieved in the aqueous acidity range of 1–5 M H2SO4 with only about 20% Fe(III) was co-extracted (Li et al., 1980). The extraction efficiency of scandium was over 90% with HDEHP from a solution containing 1 M HCl, leaving most of the yttrium and lanthanides in the raffinate (Ochsenkuhn-Petropulu et al., 1995). However, the applications of these extractants are limited due to stripping with high acidity solutions. The ionic-adsorption rare earth deposit (IARED) in China contains a trace amount of ionic-state scandium (9–11 ppm), which are readily to be leached (Liao et al., 2001a). The rare earth values in the IARED are concentrated together by leaching with sulphate salt solutions and precipitating as oxalate salts as shown in Fig. 1 (Liao et al., 2001b). The REE concentrates are calcined to obtain rare earth oxides. The scandium oxide is dissolved with HCl and the other rare earth oxides are co-dissolved. Liao et al. (2001b) used two solvent extraction circuits to separate scandium from other rare earth elements. An organic solution consisting of naphthenic acid and iso-octanol in sulphated kerosene was used for extraction. The separation factors of scandium over all the other REE (βSc/RE) were higher than 104, indicating very good separation. In the first circuit, the scandium was enriched from 0.02–0.04% to 15–20% Sc2O3 in REO (rare earth oxides) by a cross-current circuit with ten stages of extraction at an organic to aqueous phase ratio (A/O) of 5:1. In the second circuit, the obtained scandium concentrate was further purified to 99.99–99.999% grade by a counter-current circuit with three stages of extraction and three stages of scrubbing. Two solutions containing 0.35 M HCl and 1 M HCl were used for scrubbing and stripping, respectively. China dominates the world rare earth exports for the time being and some scandium is recovered during rare earth production. The separation of scandium from other rare earth elements is difficult due to their very similar chemical properties. Complicated flowsheets are needed to obtain high purity scandium products, which increases the capital and operating costs.

IARED

Leaching

Solid Residue H2C2O4

2.2. Recovery of scandium from rare earth ores

30 g/L (NH4)2SO4

Leach liquor Precipitation Sc, Y, Ln

Scandium is often found in nature together with yttrium (Y) and lanthanides (Ln) (Cotton and Wilkinson, 1988) and they are all called rare earth elements (REE). The scandium content in rare earth minerals such as monazite and bastnasite is in the range of 20– 50 ppm (Liao et al., 2001b). Scandium is recovered from the rare earth ores in Baotou, China. The main mineral in the ore is bastnasite containing scandium, yttrium, lanthanides, iron and thorium and about 50% of the REE mass is cerium. The rare earth elements including scandium can be completely extracted into solutions by roasting the ore in concentrated sulphuric acid at 250–300 °C and then leaching with water (Li et al., 2004). The impurity elements such

Calcination REO concentrate HCl

Re-Leach

RECl3 solution

SX 1 Sc2O3 Product

SX 2

REO 15-20% Sc 2O3

Table 1 Composition of the Madagascar scandium ore (based on Baptiste, 1959).

To REO purification

Component

SiO2

Sc2O3

ZrO2

Al2O3

Fe2O3

Content (%)

43.7

42.6

7.8

3.8

1.8

Fig. 1. A flowsheet to recover scandium from ionic-adsorption rare earth deposit (based on Liao et al., 2001b).

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W. Wang et al. / Hydrometallurgy 108 (2011) 100–108

H2SO4

15% NaOH

Uranium ores

Dissolution &

Leaching

Filtration Organic Recycle PLS

HCl

HCl

Solids

U extraction &

Dissolution &

stripping

Hydrolysis U

Organic bleed (Sc/Th/Ti/Zr/Fe/Si)

ScCl3 solution

NaF solution

H2C2O4 Th/Ti/Zr/Fe/Si

Precipitation

HF

to an oxide product of 99.5% purity. The above process is complicated with multiple precipitation and dissolution steps. Moreover, the application of fluoride acid as precipitation reagent could lead to environmental pollution. Ross and Rosenbaum (1962) reported a solvent extraction process with primary amine such as Primene JMT as extractant for scandium recovery from uranium mill waste solutions. Scandium was almost quantitatively extracted from the waste solution containing 0.6 mg/L Sc in a small scale pilot plant operation with 2.5% Primene JMT in kerosene at an A/O ratio 50:1. Scandium was striped with acidified 2 M NaCl solution (pH ~1) at an A/O ratio of 1:10. The scandium was precipitated from the loaded strip liquor with ammonia solution. However, there is still a potential environmental concern due to the high level of radioactivity caused by uranium.

Sc2(C2O4)3

Precipitation &

2.4. Recovery of scandium from aluminium ores

separation Calcination TiF4 solution Sc2O3 Product

ScF3, ThF4 solids

Fig. 2. A flowsheet of scandium recovery from uranium processing (based on Lash and Ross, 1961).

2.3. Recovery of scandium from uranium ores Trace amounts of scandium are found in most uranium ores such as uraninite. The worldwide production of uranium in 2009 amounted to 50,572 tonnes, of which 27% was mined in Kazakhstan, followed by Canada (20%), Australia (16%), Namibia (9%) and Russia (7.0%) (Anon, 2010a). Therefore, recovery of scandium as a by-product in uranium processing is of significant importance. Lash and Ross (1961) reported that the uranium ores were crushed, ground and leached with sulphuric acid, and the leach liquor contained up to 1 mg/L Sc2O3. The uranium can be completely extracted with dodecyl phosphoric acid (0.1 M) from the sulphuric acid leach solution. Scandium, thorium and titanium were co-extracted but could not be stripped along with uranium using 10 M HCl. The accumulated scandium and thorium were recovered by treating an organic bleed stream with hydrofluoric acid which precipitated both scandium and thorium (Fig. 2). The scandium–thorium fluoride precipitate containing 10% Sc2O3 and 20% ThO2 was filtered from the aqueous phase while soluble titanium fluoride was left in the filtrate. The precipitate was dissolved with a 15% NaOH solution in 75–90 °C for 4 h, resulting in the formation of scandium hydroxide precipitate. After filtration, the scandium hydroxide was leached with hydrochloride acid under certain acidity and temperature to ensure the hydrolysis of the impurities including titanium, zirconium, iron and silicon. The scandium was precipitated with oxalic acid to separate it from the co-dissolved uranium and iron. The obtained scandium oxalate was calcined at 700 °C

Scandium is often associated with aluminium ores in nature. The aluminium phosphate minerals can contain 0.01–0.80% Sc2O3 (Frondel et al., 1968). Bauxite is the most common aluminium ore, containing aluminium oxides and hydroxides often with impurities such as iron oxides. In the world aluminium industry, the Bayer process is the major process used to treat bauxite, in which it is digested at 140–300 °C in a caustic solution to dissolve the aluminium (Solymar et al., 1976; Banvolgyi et al., 1992). Normally, for each tonne of alumina produced, an equal amount of slimy caustic residue known as red mud is generated as by-product/waste. Red mud mainly consists of iron, calcium, aluminium, silicon, titanium and sodium in percent range and also contains small quantities of elements of economical interest such as vanadium, zirconium, niobium, and rare-earth elements including yttrium, scandium and lanthanides (Ochsenkuhn-Petropulu et al., 1995). The composition of red mud changes depending on mining locations of bauxites and ore processing methods (Table 2). Scandium is almost doubly enriched in red mud (OchsenkuhnPetropulu et al., 1994) compared to the original ore. For instance, the scandium level in Jamaican bauxite is 87–113 ppm as Sc2O3 and is accumulated and enriched to as high as 200–390 ppm in the red mud (Wagh and Pinnock, 1987). The red mud generated in Greece has a high and uniform content of about 130 ppm Sc in the dry red mud, corresponding to 0.02% Sc2O3 (Ochsenkuhn-Petropulu et al., 1994), which was considered as a valuable scandium resource. The world production of bauxite in 2008 was 205 Mt in 26 countries (Bray, 2010), resulting in the generation of a huge quantity of red mud. If the average Sc2O3 content in the red mud is 50 ppm (Table 2) and 80% of it is recovered from the 102.5 Mt of red mud, the amount of Sc2O3 would be a very significant amount of 4100 tonnes. Therefore, the Bayer process residue or red mud can be potential large source of scandium. However, it is difficult to directly recover scandium from the red mud because the levels of the main components especially iron, aluminium and titanium are high. For example, the content of iron in the red mud from different sites in Australia is in the range of 28.5–56.9% and that of aluminium 15.6–24.0% (Table 2).

Table 2 Main components and scandium contents of red mud. Country

Australia Brazil China Germany Greece Italy Jamaica Russian Spain USA

Main constituents (%) Fe2O3

Al2O3

CaO

SiO2

TiO2

Na2O

28.5–56.9 45.6 3.0–6.2 44.8 42.5 15.2 42.3 19.7–46.0 37.5 35.5

15.6–24.0 15.1 5.0- 8.6 16.2 15.6 24.7 16.4 11.8–15.4 21.2 18.4

2.3–5.3 1.2 34.0–39.5 5.2 19.7 4.2 9.1 10.6–30.2 5.5 7.73

3.0–30.0 15.6 19.0–20.8 5.4 9.2 18.6 7.8 5.2–8.5 4.4 8.5

3.1–8.0 4.3 2.1–3.6 12.3 5.9 6.2 6.0 6 11.5 6.31

2.2–8.6 7.5 2.1–4.5 4.0 2.4 11.7 4.6 5 3.6 6.1

(ppm) Sc2O3

References

– – 41.2–92.5 – 105–156 – 150–172 80–120 – –

Snars and Gilkes (2009) Snars and Gilkes (2009) Xu (2001) Snars and Gilkes (2009) Ochsenkuhn-Petropulu et al. (1994) Snars and Gilkes (2009) Wagh and Pinnock (1987) Yatsenko and Pyagai (2010) Snars and Gilkes (2009) Snars and Gilkes (2009)

W. Wang et al. / Hydrometallurgy 108 (2011) 100–108

Numerous researchers such as Piga et al. (1993) and Rayzman and Filipovich (1999) studied the recovery of valuable metals from red mud with a combination of pyro- and hydrometallurgy processes. The red mud was mixed with coal, lime and sodium carbonate under reduction-sintering conditions in the temperature range of 800– 1000 °C, followed by alumina re-leaching with hot water at 65 °C to obtain brown mud (Piga et al., 1993). Over 99% of Fe can be recovered from the brown mud as pig iron with only 1% left in the slag (Rayzman and Filipovich, 1999). The slag is a valuable resource for scandium with a high level of scandium (420 ppm) together with TiO2 (19.4%), lanthanide oxides (1470 ppm) and yttrium (180 ppm). After iron separation, the brown mud was leached with H2SO4 (Piga et al., 1993). The leach liquor was hydrolysed at 140 °C to recover titanium, and the scandium enriched in the solution after the hydrolysis of titanium can be further recovered by precipitation or solvent extraction. The mineralogy study with X-ray diffraction showed that the red mud produced by the Bayer Process has the following major constituents: hematite, maghemite, calcite, calcium-titanate, calciumsilicate, rutile, diaspore, aluminium silicate and aluminium sodalite (Ochsenkuhn-Petropulu et al., 1996). The red mud can be totally dissolved in strong mineral acids with or without roasting for the further recovery of scandium with solvent extraction and ion exchange. Zhou et al. (2008) reported that red mud can be completely dissolved with 6 M HCl solution at a volume ratio of 1:4 at 60 °C for 4 h. The concentration of scandium in the leach solution was about 8 mg/L with large amounts of other metals including 9.2 g/L Na, 7.0 g/L Fe, 12.4 g/L Ca, 14.8 g/L Al and 2.8 g/L Ti. Scandium was almost completely recovered by absorption with activated carbon modified by tri-butyl phosphate (TBP). However, the scandium absorption efficiency was decreased by the co-absorption of titanium. Smirnov and Molchanova (1997) investigated a method for treating the Russian red mud similar to that used for treating low-grade uranium ores by direct leaching with sulphuric acids followed by separation of the radioactive and valuable components with the nitrogen–phosphorus-containing ampholite resins. The loading capacity of the resin was reduced by other metals including aluminium and titanium in large amounts. The crude scandium cake obtained by precipitation of the eluate contained 5–7% scandium, 4.5% uranium, 0.9% thorium and 35–40% titanium. Ochsenkuhn-Petropulu et al. (1995) studied the recovery of scandium from the red mud produced in Greece. The red mud with borate/ carbonate was fused at 1100 °C for 20 min and leached with excessive 1.5 M HCl (Fig. 3). The leach liquor was passed through an ion exchange column filled with Dowex 50W-X8 resins. The scandium and the main impurity elements such as Fe, Al, Ca, Si, Ti and Na as well as the minor ones such as Ni, Mn, Cr and V were co-extracted. In the subsequent elution stage, most of the impurities were removed in the elution stage with 1.75 M HCl. The scandium was quantitatively eluted with 6 M HCl and followed by solvent extraction with 0.05 M HDEHP in hexane after neutralising the elute solution to a pH of about 0 with ammonia. Scandium was selectively and nearly quantitatively extracted into the organic phase, leaving yttrium and the lanthanides in the aqueous phase. The loaded scandium was quantitatively stripped in high purity into the aqueous phase as Sc(OH)− 6 anions using 2 M NaOH. With such a combined method using ion exchange and solvent extraction, scandium can also be recovered from nitrate medium (Ochsenkuhn-Petropoulou et al., 2002). Obviously, such treatment of red mud is not economical due to the low content of scandium and the large amounts of impurities in the red mud. In the process, large amounts of impurities were dissolved in the leaching step with acids and co-absorbed by the ion exchange resin, resulting in a decrease in the resin capacity. Moreover, the co-absorbed impurities need to be eluted, which consume large amounts of acids, resulting in the consumption of large amounts of bases for neutralisation, suggesting high operating costs. Recently, methods have been developed to selectively leach minor metals, leaving main impurities such as iron in the red mud un-

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dissolved. In a selective leaching process for rare-earth recovery from red mud, the rare earth elements were leached in a diluted acid medium by introducing gaseous SO2 into the solid–liquid slurry to reduce the aqueous pH to about 1.0, leaving the main impurities in the residue (Fulford et al., 1991). A recent study by Yatsenko and Pyagai (2010) showed that scandium was partially recovered in the process of reducing the alkalinity of red mud by the neutralisation of the red mud pulp with the absorption of acidic gas emissions (CO2, SO2, and NO). Ochsenkuhn-Petropulu et al. (1996) studied the selective leaching of lanthanides, scandium and yttrium from red mud with sulphuric acid, nitric acid and hydrochloric acid. It was found that the diluted HNO3 solution with a concentration of about 0.5 M provided the best recoveries for scandium (80%), yttrium (96%) and ytterbium (70%). The leaching selectivity of scandium over iron was also the highest with only 3% iron being leached. Interestingly, the authors found that pre-treatment such as oxidation, roasting, magnetic separation and sizing prior to leaching was not necessary. Based on the above laboratory-scale results, Ochsenkuhn-Petropoulou et al. (2002) conducted a pilot-plant operation with HNO3-leaching under ambient temperature and pressure. 2.5. Recovery of scandium from titanium and zirconium ores Some titanium minerals such as ilmenite (FeTiO2) and rutile (TiO2) contain significant amounts of scandium. For example, the magnetovana-ilmenite ore (an ilmenite with high content of vanadium) in Panzhihua, China, contains 0.002–0.004% Sc2O3 (Chen, 1990). Australia, Canada, India, Norway and South Africa are the main producers of titanium minerals. The world production of ilmenite is about 4.8 Mt per year (Gambogi, 2000), thus a yield of 96–194 tonnes of Sc2O3 is potentially included. Ilmenite is often further beneficiated to produce synthetic rutile and titaniferous slag. The scandium was enriched in the slag, for example, a significant amount of 128 ppm was included (Mao et al., 1996). The scandium in the tailings was combined with high contents of impurities (40% SiO2, 16% FeO, 15% CaO, 14% MgO, 5.3% TiO2 and 5.8% Al2O3). For the production of titanium dioxide or titanium metal, titanium feed

NaKCO3/Na2B4O7 Red Mud

Roast

1.5 M HCl

Acid Leaching Feed solution

Solid Residue 1 1.75 M HCl

Ion Exchange Dowex 50W-X8 resins

6 M HCl 2

Ammonia

Sc, Y, La

Fe, Al, Ca, Si, Ti, Na

Neutralisation pH = 0

2 M NaOH

SX with HDEHP

Strip liquor [Sc(OH)6]3-

Raffinate Y, La Fig. 3. A flowsheet of scandium recovery from bauxite residue (based on OchsenkuhnPetropulu et al., 1995).

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materials are first fused in the chlorination step to produce a vaporous titanium tetrachloride (Baptiste, 1959). The scandium is concentrated principally in the residue or in the dust as scandium chloride, making the recovery of scandium feasible. It was reported that the dust contains a high level of scandium of at least 132 ppm (Xu and Li, 1996). About 90% of the scandium was leached from the ilmenite slag with a process shown in Fig. 4 (Mao et al., 1996). The process included grinding the slag, roasting with Na2CO3 in the temperature range of 900–1000 °C, and leaching with 30% HCl at 80 °C at a solid to liquid ratio (S/L) of 1:2. Over 94% scandium was extracted from the leach liquor using 30% HDEHP in kerosene at an A/O ratio of 20:1 with only about 2.2% iron coextraction. The extracted iron was removed by scrubbing with 5 M HCl and the scandium was stripped with 2 M NaOH. The obtained Sc(OH)3 precipitate was dissolved with HCl and the scandium was precipitated with oxalic acid. A scandium oxide product of 99% purity was obtained after calcining the oxalate precipitate at 800 °C. Feuling (1991a) patented a process (Fig. 5) for the recovery of the scandium from a titanium chlorination residue with tri-butyl phosphate (TBP) as extractant. The residue was leached by 6 M HCl to produce a scandium-containing aqueous solution, followed by removal of the radium in the leach solution with adsorption on newly formed BaSO4. The filtrated solution was contacted with TBP for separating of scandium from the main elements including sodium, calcium and magnesium and other minor elements such as thorium, yttrium and lanthanides. The scandium in the organic phase was stripped with 0.1 M HCl and the co-extracted iron was left in the organic phase. The scandium was precipitated by ammonia to produce a scandium hydroxide precipitate, which was calcined at 600 °C to produce Sc2O3. The hydrolytic solution from TiO2 production in the classic sulphuric acid method contains 15–20 mg/L scandium and other impurities such as Zr, Ti, Lu, Fe and SiO2 with an acidity of about 2 M H2SO4 (Li and Wang, 1998). The scandium exists in the sulphuric acid solution as Sc(III) ions and therefore can be directly recovered by solvent extraction. The authors studied the extraction of scandium with Cyanex 923 (trialkyl phosphine oxide). The scandium was extracted in high efficiency with co-extraction of titanium, iron and zirconium in low efficiency. The coextracted titanium was removed in the scrubbing stages by acidic H2O2 or diluted H2SO4. A scandium product with 95–96% purity and 94% recovery was obtained using 6–8 stages of extraction, 7–9 stages of scrubbing and 1–2 stages of stripping.

Washed organic solution

Ilmenite slag

Sc ST

Roasting

Raffinate

3 M HCl

Sc(OH)3 Precipitate

30% HCl

Dissolution Leaching

10% H2C2O4 Solid residue

30% HDEHP

Leach solution Sc EX Raffinate 5 M HCl

Loaded organic solution

Fe Removal

6 M HCl Acid Leaching

Solids

Liquid BaCl2+H2SO4 Precipitation & Filtration

RaSO4/Ba2SO4 0.1 M HCl

Feed solution TBP

Extraction

Raffinate (Na/Mg/Ca/Th/Y/REE)

Organic Sc/Fe Organic

Stripping Fe disposal Strip liquor Precipitation &

NH4OH

Filtration NH4OH/NH4Cl Rare earth oxides by-products

Calcination

Sc2O3 product Fig. 5. A flowsheet of scandium recovery from titanium residue (based on Feuling, 1991a).

Zircon is a co-product or by-product of the mining and processing of heavy-mineral sands for the titanium and tin minerals. World reserves of zircon were estimated to be 51 Mt of zirconium oxide content (Gambogi, 2008). Zircon is the principal economic source of zirconium and hafnium. In the process of zirconium production, zircon sands are chlorinated at about 1000 °C, leaving scandium in the residue with a content of 0.34% (Feuling, 1991b). With a procedure similar to scandium recovery from titanium ores (Feuling, 1991a), scandium can be recovered from the zircon chlorination residue. Acidic and neutral organophosphorus extractants are most commonly used for scandium recovery from titanium and zirconium residues. However, the significantly high contents of titanium and zirconium in the residues are readily co-extracted to the organic phases and difficult to be removed by conventional scrubbing methods with diluted acids. It is reported that in some practices, the co-extracted titanium and zirconium can be removed with solutions of acidic H2O2 and fluoride, respectively (Xu and Li, 1996). 2.6. Recovery of scandium from tungsten and tin ores

2 M NaOH

Na2CO3

Titanium-bearing Residue

Sc solution Precipitation & Filtration Precipitate Calcination Sc2O3 product

Fig. 4. A flowsheet for recovering scandium from slag of ilmenite (based on Mao et al., 1996).

Scandium is found in the tungsten minerals such as wolframite (iron–manganese tungstate, FeWO4/MnWO4) and scheelite (calcium tungstate, CaWO4), which are mined and used to produce about 37,400 tonnes of tungsten concentrates per year (Shedd, 2000). Normally, tungsten ores are digested in an alkali medium to form an alkali tungstate solution, leaving calcium, iron and manganese in the precipitate. The residues from processing tungsten ores contain significant amounts of scandium (Gokhale and Bhat, 1967; Vanderpool et al., 1986, 1989; Zhong, 2002). For example, the scandium was enriched to about 0.04–0.06% in the wolframite residue (Table 3). Trace amounts of scandium are also contained in the tin ores such as cassiterite (SnO2). A recent study revealed anomalously high scandium content of up to several thousands of parts per million in wolframite and cassiterite in Sn–W deposits in the eastern Erzgebirge mountain range (Kempe and Wolf, 2006). Much work has been carried out to recover scandium from the tungsten bearing material. Since scandium mainly presents in the form of hydroxide in the residue, it can be converted to soluble scandium salts by leaching with various acids such as sulphuric, hydrochloride and nitric acids. Hydrochloride acid can preferentially

W. Wang et al. / Hydrometallurgy 108 (2011) 100–108

105

Table 3 Main constituents and scandium contents of tungstentic residue. Main constituents (%)

References

Fe2O3

MnO

WO3

SiO2

Al2O3

CaO

PbO

TiO2

ThO2

RE2O3

Sc2O3

21–24 36.2

18–23 24.8

1–2 8.6

7.8

4.4

6.5

4.9

0.26

0.01

0.12

0.06 0.04

leach scandium because of the formation of stable scandium chloride complexes (Guo et al., 1988). A flowsheet (Fig. 6) with a process consisting of HCl leaching and HDEHP solvent extraction was developed for recovery scandium from wolframite residues (Guo et al., 1988). About 95.3% scandium was leached with concentrated HCl at 100 °C. The acidity and scandium concentration in the feed solution were about 2 M HCl and 100 mg/L Sc2O3, respectively. About 90% scandium was extracted with HDEHP in kerosene at an A/O ratio of 1:4. The co-extracted impurities such as iron, calcium, magnesium, aluminium, rare earths and silicon were removed from the loaded organic phase in the scrubbing stage with 3.5 M HCl solution. The scandium was almost completely stripped with 2 M NaOH in two stages. The scandium oxide content in the resultant scandium hydroxide reached 70–78% with total recovery of 76–89% (Guo et al., 1988). Kim and McClintic (1988) reported that nearly 100% scandium was recovered without extracting appreciable amounts of impurities with the HDEHP impregnated resins from an acid leach solutions containing about 60 mg/L Sc, 39 g/L Fe, 19 g/L Mn, 0.40 g/L W and some minor elements. The co-extracted iron can be washed out with 2 M HCl. Wakui et al. (1989) used pressure acid leaching to dissolve a small amount of wolframite with concentrated HCl at 120 °C. After filtering the yellow-white precipitate of tungstic acid, the scandium containing in the filtrate was recovered using ion exchange with PC-88A impregnated resin. With a similar process, the authors also recovered scandium from an HCl-leached solution containing 0.2 mg/L Sc, 0.4 mg/L Y, 68 mg/L Ca, 28 mg/L Fe, 27 mg/L Al and 9.2 mg/L Nb from tin slag after smelting cassiterite (SnO2). The main disadvantages of the HCl-leach method are the evaporation of HCl and the formation of toxic compounds, leading to the increase in operating costs and pollution. The H2SO4-leach method also offers high scandium leaching efficiencies, but without the formation of harmful chlorine gas. Xu and Li (1996) reported that under optimised conditions, about 94.9% scandium can be leached with concentrated H2SO4 at high temperatures from a wolframite ore containing Sc (0.04%), W (69.3%), Fe (11.6%) and Mn (4.8%). Vanderpool et al. (1989) revealed a process in which nearly 100% scandium was leached from a tungsten bearing material containing about 23.7% Fe, 22.5% Mn, 0.06% Sc and 2% W, by digestion with 18 M H2SO4 for about 6 h in the temperature range between 100 and ~ 140 °C. Coal was added in the solution to reduce the manganese and then completely digest scandium, manganese and iron. A typical leach solution contained about 23–24 g/L Fe, 16 g/L Mn, 0.15–0.23 g/L W and about 0.04 g/L Sc. The major portion (84%) of tungsten was left in the residue as tungstic acid. Rourke et al. (1990) reported a process to recover scandium from the waste of a tungsten plant. The wolframite residue was dissolved with 1 M H2SO4 containing 6% H2O2 at an S/L ratio of 1:25 for 2 h. The slurry was filtrated and the resultant leach liquor contained 5.6 g/L Mn, 3 g/L Fe and 14 mg/L Sc. The scandium was completely extracted from the aqueous phase in the pH range of 1.8–2.0 with a chelating extractant athenoyltrifluoroacetone (HTTA) in toluene. After stripping the scandium with 3 M HCl, ammonia or oxalic acid was added to precipitate the scandium as Sc(OH)3 or oxalate. Zhong (2002) reported a process to recover scandium from H2SO4-leach solutions of tungsten slag (Fig. 7). The ferric ions in the solutions were reduced to ferrous ions by addition of iron powder, and the thorium was extracted and separated using 0.2% primary amine N1923 in kerosene at an A/O ratio of 4:1. Over 99% Sc was extracted

Vanderpool et al. (1989) Zhong (2002)

with 4.0% N1923 in kerosene at an A/O ratio of 4:1. The co-extracted REE, Fe and Ti were scrubbed with 3 M H2SO4, 0.5 M H2SO4 and 3% H2O2, respectively. The scandium was stripped with 2 M HCl. A Sc2O3 product with 90% purity and 82% recovery was obtained by precipitation of scandium with oxalic acid and calcination of the scandium oxalate precipitate. The residues from processing tungsten and tin ores normally contain relatively high contents of scandium and can be considered important resources of scandium. Although the residues can be leached with concentrated HCl at high temperatures, the leach process with sulphuric acid is preferred due to its less pollution. Acidic organophosphorus extractants are suitable for the extraction and separation of scandium from the main impurities in the leach solution including iron, manganese and tungsten. Chelating extractants and primary amine extractants are more suitable for scandium extraction from sulphuric acid solutions due to the easier stripping with diluted hydrochloride acids.

2.7. Recovery of scandium from nickel ores The nickel ores in Australia containing relatively high content of scandium are considered as important scandium resources. For example, the nickel and cobalt deposits at Syerston and Lake Innes, New South Wales have an average grade of 76 ppm and in the range of

Wolframite residue

HCl

Leaching

Filter residue

Filtration

Feed solution Organic recycle Extraction Raffinate

HDEHP in kerosene Loaded scrub liquor 3.5 M HCl

Loaded organic Scrubbing Scrubbed organic

2 M NaOH

Stripping & Filtration

Precipitate Sc(OH)3 product Fig. 6. A simplified flowsheet for the recovery of scandium from wolframite residue with hydrochloride acid (based on Guo et al., 1988).

106

W. Wang et al. / Hydrometallurgy 108 (2011) 100–108

Purified feed solution Organic recycle Wolframite slag 4.0% N1923 H2SO4

Table 4 Compositions of H2SO4-leach solution of laterite ores after neutralisation and sulphide precipitation (based on Haslam and Arnall, 1999). Concentration (mg/L)

Sc Extraction

Leaching

H2SO4, H2O2

Raffinate Loaded organic

Al 36.5

Ca 463

Co 0.03

Cr 4.8

Fe 0.27

Mg 14700

Mn 777

Si 49

Zn 1.1

Ni 9.6

Sc 1.12

Filtration Scrubbing Feed solution Fe powder

2.0 M HCl

REE, Fe & Ti Loaded organic

Stripping Re-dissolution & Re-filtration Filtrate solution 0.2% N1923 Th Extraction & Separation

H2C2O4 Strip liquor

Precipitation & Filtration Sc oxalate precipitate Calcination Sc2O3 product

Fig. 7. A flowsheet to recover scandium from tungsten slag by leaching with sulphuric acid (based on Zhong, 2002).

130–370 ppm, respectively (Anon, 1998, 1999). A total amount of 3500–6500 tonnes of scandium was estimated in the two deposits. The scandium can be recovered as a byproduct during nickel and cobalt extraction operations. A typical nickel laterite ore contains Ni (1-2%), Co (0.05-0.10%), Fe (15-50%), Al (2-5%) and trace amounts of Sc (0.005-0.006%) (Haslam and Arnall, 1999). Scandium was readily leached with sulphuric acid from the nickel laterite ores in the high pressure acid leach (HPAL) process with over 94% recovery (Koryakov and Medvedev, 1994). After removal of iron and aluminium by neutralisation in the pH range of 2–4 and recovery of nickel and cobalt as sulphide precipitates, the scandium can be precipitated from the solution by adjusting the pH to higher than 4.0 (Akira et al., 1999). Some impurities were co-precipitated with scandium. The scandium can be alternatively separated from the impurities using SX, after the neutralisation and sulphide precipitation. For example, from a typical solution in the pH range of 1.0–1.5 as shown in Table 4, quantitative scandium extraction was achieved using acidic organic-phosphorus extractants, such as HDEHP, PC-88A and Cyanex 272 (Haslam and Arnall, 1999). The selectivity was high with the extraction order of Sc N Zn N Ca N Al N Cr N Mg N Ni. The industrialised plants for scandium recovery from nickel laterites are under development by several mining companies and are expected to produce scandium oxides in large amount. For example, a hydrometallurgy plant is designed to produce 28,000 kg of Sc2O3 per year from the Nyngan project (Anon, 2010b). The Ni–Co–Sc laterite processing plant with a heated atmospheric acid leach process will be developed for the NORNICO project near Greenvale, Queensland (Anon, 2010c), and the high-purity scandium oxide production is proposed to be available at an annual supply of 10,000–40,000 kg. 2.8. Recovery of scandium from tantalum and niobium ores In the United States, some scandium was produced from mine tailings of the tantalum mining (Hedrick, 2010a). The tantalum sulphate tailings were leached with water and a leach liquor containing scandium and over 20 metals in significant quantities was obtained (Odekirk and Harbuck, 1993; Odekirk, 1996). With the organic systems of phosphoric acid and phosphonic acid such as

HDEHP and PC-88A, the scandium was completely extracted from the solution containing 50–200 g/L H2SO4. The impurities such as niobium, tantalum, yttrium and iron were effectively separated, but significant amounts of zirconium, hafnium, titanium, thorium and uranium were co-extracted. The loaded organic solution was firstly scrubbed by 350 g/L H2SO4 solution to remove thorium, and then scrubbed some of the co-extracted elements by 0.1 M HF solution. HF solutions in the concentration range of 0.5–5 M were used to strip the loaded metals in the following order: Th, Ti N Zr N Hf N Sc N U. As a result, scandium was separated from thorium, titanium, zirconium and hafnium by 3–15 stages of counter-current stripping and from uranium by 6 stages of fractional stripping (Odekirk, 1996). The main disadvantage of this process is the use of HF solutions in large amounts, which leads to environmental risks. It was reported that the niobium deposits in Russian Arctic area contain Sc2O3 in an extremely high grade range of 0.1–0.3% (Dobretsov and Pokhilenko, 2010). Some of the ores contain a mixture of niobium (7% Nb oxide) and rare earth (10% REE oxides). The ores were treated with a combined hydro- and pyro-metallurgical technology by digesting with 45% NaOH to convert the rare-earth minerals (including Sc, Y, and Ln) to hydroxides (Kuz'min et al., 2006). The scandium in the hydroxide cake was dissolved with HCl leaving most of the niobium and titanium in solids. The scandium in the leach liquor was completely extracted and separated from other rare earth elements, aluminium and alkali-earth elements with 80% TBP in a counter-current SX circuit. The co-extracted iron and uranium can be removed by selective stripping with different concentrations of HCl. After precipitation of scandium from the strip liquor with oxalic acid, the resultant scandium oxalate precipitate was calcined to obtain scandium oxide product with a purity of 99.9%. The main modification of this process is that the separation of scandium from niobium and titanium by digestion and dissolution.

2.9. Recovery of scandium from its alloy scraps During the production of Sc-containing alloys such as Fe–Sc, Al–Sc and Mg–Sc, considerable amounts of scandium are lost because scandium has high reactivity with oxygen, chlorine and fluorine, especially under high temperatures (Ditze and Kongolo, 1997). The smelting high-scandium magnesium alloy consists of 86% Mg and 14% Sc. Scandium is contained in the salty wastes or metallic dross in high content from the smelting process. The average composition of the Mg–Sc alloy dross was 64–77% Mg, 12–23% Sc and l–1.6% Fe. Nearly 100% Sc recovery and nearly complete separation from magnesium could be achieved after a single stage of leaching with HCl and a solvent extraction process with HDEHP (Fig. 8). In the extraction stage, 99.9% Mg and 90% Fe were left in the raffinate. Diluted HCl solutions can be used for scrubbing the co-extracted iron from the loaded organic solution. The scandium was stripped with 5 M NaOH to obtain scandium hydroxide precipitate. After calcining, a Sc2O3 product with 64.5% Sc, 0.5% Mg and 0.4% Fe was obtained. This procedure can also be used for recovering scandium from scandiumbearing aluminium and iron alloy scraps. Although the scandium contents in the alloy scraps are high, the gross of the scraps produced in the Sc-alloy production worldwide is not enough as a reliable resource. In addition, no accurate data are available at moment.

W. Wang et al. / Hydrometallurgy 108 (2011) 100–108

3. Summary and recommendations Scandium is commonly found in ores of other metals and recovered from their residues and tailings. Various metallurgical processes of scandium recovery from various resources, such as ores of scandium, rare earths, uranium, aluminium, titanium, tungsten, nickel, tantalum and niobium are reviewed. Considering the scandium content and the availability, bauxite residues (red mud) and nickel laterite ores are the most promising scandium resources. A pilot plant for scandium recovery from bauxite residues was reported and some industrial scale plants for scandium recovery from nickel laterites are under development. Pyrometallurgical processes are not suitable for recovery of scandium from residues and tailings due to intensive energy consumption. Currently, hydrometallurgical processes consisting of ore pretreating, leaching, solvent extraction, precipitation and calcination are the most widely used methods for the production of scandium products. The co-dissolution and co-extraction of metal impurities such as iron, titanium, zirconium, uranium, thorium and rare earths could interfere the scandium extraction. Although much work has been done by many researchers, it is necessary to modify existing processes or develop new processes in the area of leaching and solvent extraction to improve extraction selectivity and efficiency. Leaching with medium to concentrated mineral acids could lead to dissolving large amounts of impurities and is not economical due to the high acid-consumption to dissolve trace amounts of scandium. Therefore, the development of selective leaching of scandium under moderate conditions is necessary to reduce acid consumption and pollution. Both acidic and neutral organophophorus extractants have their drawbacks: the former has the difficulties in stripping and the latter is lack of selectivity. Therefore, it is necessary to develop new extraction systems with high scandium selectivity. Since scandium is mainly produced as a by-product during processing of various ores, residues and other materials, its recovery

Mg-Sc dross 15% Sc, 70% Mg, 1%Fe

Residue

HCl, water

Leaching

Feed solution, pH~0.14

Organic recycle

Raffinate, Mg/Fe Extraction Loaded organic Aqueous solution Scrubbing 6.5 M HCl

NaDEHP

Acidification

5 M NaOH Stripping Stripping liquor

Water

Flocculant

Dilution

Solid-liquid separation

Sc(OH)3 precipitate Sc2O3 product

Drying & Calcination

Waste water

Fig. 8. A simplified flowsheet for the recovery of scandium from Mg–Sc dross (based on Ditze and Kongolo, 1997).

107

should be considered in the development and design of the overall flowsheet. It is suggested that red mud from treating bauxides and nickel laterite ores are proposed as the most promising scandium resources for its production in the future.

Acknowledgements The authors thank Dr Matthew Jeffrey for reviewing the manuscript and providing valuable comments. The support of the Parker CRC for Integrated Hydrometallurgy Solutions is gratefully acknowledged.

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