Extraction of Lithium from Primary and Secondary Sources by Pre-Treatment, Leaching and Separation: a Comprehensive Review

Home , Eucryptite, Jadarite, Lithium aluminate, Lithium bicarbonate, Lithium oxalate, Lithium superoxide

Hydrometallurgy 150 (2014) 192–208

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Review Extraction of lithium from primary and secondary sources by pre-treatment, leaching and separation: A comprehensive review

Pratima Meshram a,b, B.D. Pandey a,⁎,T.R.Mankhandb a CSIR—National Metallurgical Laboratory, Jamshedpur 831 007, India b Dept. of Metallurgical Engineering, IIT, BHU, Varanasi 221 005, India article info abstract

Article history: In this comprehensive review resources of lithium and status of different processes/technologies in vogue or Received 27 October 2013 being developed for extraction of lithium and associated metals from both primary and secondary resources Received in revised form 4 October 2014 are summarized. Lithium extraction from primary resources such as ores/minerals (spodumene, petalite and Accepted 12 October 2014 lepidolite) by acid, alkaline and chlorination processes and from brines by adsorption, precipitation and ion Available online 23 October 2014 exchange processes, is critically examined. Problems associated with the exploitation of other resources such

Keywords: as bitterns and seawater are highlighted. As regards the secondary resources, the industrial processes followed Lithium and the newer developments aiming at the recovery of lithium from lithium ion batteries (LIBs) are described Primary ores in detail. In particular pre-treatment of the spent LIBs, leaching of metals from the cathode material in different Brines acids and separation of lithium and other metals from the leach liquors, are discussed. Although spent LIBs are LIBs currently processed to recover cobalt and other base metals but not lithium, there is a good prospect for the Hydrometallurgical recovery recovery of lithium in the coming years. Varying compositions of batteries for different applications require development of a suitable recycling process to recover metals from all types of LIBs. © 2014 Elsevier B.V. All rights reserved.

Contents

1. Introduction...... 193 2. Resourcesoflithium...... 193 2.1. Primary resources — minerals/claysandbrines...... 193 2.2. Secondary resources — lithiumionbatteries...... 194 3. Extractionoflithiumfromprimaryresources...... 195 3.1. Lithiumextractionfromminerals/clays...... 195 3.1.1. Acidprocess...... 196 3.1.2. Alkalineprocess...... 197 3.1.3. Chlorinationprocess...... 197 3.1.4. Otherprocesses...... 197 3.2. Lithiumextractionfrombrines/seawater/bitterns...... 197 3.2.1. Adsorptionprocess...... 198 3.2.2. Precipitationprocess...... 199 3.2.3. Ionexchange/Solventextractionprocess...... 199 3.3. Extraction of lithium from secondary resources — lithiumionbatteries...... 200 3.3.1. Majorindustrialprocesses...... 200 3.3.2. Recentdevelopmentinrecyclingoflithiumionbatteries...... 202 4. Conclusions...... 206 Acknowledgements...... 206 References...... 206

⁎ Corresponding author. E-mail address: [email protected] (B.D. Pandey).

1. Introduction Table 1 Principal commercial lithium minerals with compositiona.

Lithium is the 25th most abundant element (at 20 mg/kg) in the Mineral Formula % Lithium content earth's crust. Lithium ﬁnds an application in rechargeable lithium ion bat- Theoretical Range in commercial teries (LIBs) because of its very high energy density by weight and high minerals electrochemical potential (3.045 V). With a present consumption level Spodumene LiAlSi O or Li O·Al O ·4SiO 3.73 1.9–3.3 of ~22% of the total lithium produced in LIBs, it is expected to reach to 2 6 2 2 3 2 Lepidolite LiKAl2F2Si3O9 or 3.56 1.4–1.9

~40% by 2020 (Wang et al., 2012). Besides batteries, at present it has LiF·KF·Al2O3·3SiO2 – major applications in glass and ceramics (30%), greases (11%), metallurgi- Amblygonite LiAlFPO4 or 2LiF·Al2O3·P2O5 4.74 3.5 4.2 – cal (4%) industries and also in chemicals/pharmaceuticals, rubbers etc. Triphylite LiFePO4 or Li2O·2FeO·P2O5 4.40 2.5 3.8 Petalite LiAlSi O or Li O·Al O ·8SiO 2.27 1.6–2.21 (Garrett, 2004; Holdren, 1971). As per the Madrid Report of July 2010, 4 10 2 2 3 2 Bikitaite LiAlSi2O6·H2O 3.28 1.35–1.7 lithium falls in the border-line of low-to-medium in supply and demand, Eucryptite LiAlSiO4 5.53 2.34–3.3 primarily due to scanty resources (Tedjar, 2013). It is also understood that Montebrasite Li2O·Al2O3·2SiO2 3.93 0.9–1.8 – the demand for lithium is increasing further due to its application in nu- Jadarite LiNaSiB3O7(OH) 3.39 0.096 0.1 – clear and strategic areas. As per recent estimates of lithium reserves, out Zinnwaldite LiKFeAl2F2Si3O10 or 1.7 1.21 1.3 LiF·KF·FeO·Al O ·3SiO of 103 deposits with more than 1,00,000 t of lithium, each deposit has a 2 3 2 Hectorite Na0.3(Mg,Li)3Si4O10(F,OH)2 0.56 0.36 different mineralogical composition and therefore, requires the appropri- Zabuyelite Li2CO3 18.75 – ate technology to process (Gruber et al., 2011). In view of this, lithium is a Industrial Minerals and Rocks (2006), Norton and Schlegel (1955), Schaller (1937),and usually extracted from its mineral that is found in igneous rocks (chiefly Siame and Pascoe (2011). spodumene) and from lithium chloride salts found in brine pools, while ignoring other resources including the low-grade ores. The rising demand for lithium for various applications thus calls for policy about the fate of discarded lithium ion batteries in e-waste that prospecting and processing all viable resources. Lithium extraction is distributed internationally. Lithium batteries also contain potentially from ores/minerals utilizes roasting followed by leaching, while its toxic materials including metals, such as copper, nickel and lead, and extraction from brines includes evaporation, precipitation, adsorption organic chemicals, such as toxic and flammable electrolytes containing and ion exchange (Garrett, 2004). Lithium can be extracted from LIBs LiClO4,LiBF4,andLiPF6. Defunct Li-ion batteries are classified as hazard- by leaching followed by precipitation, ion exchange or solvent extrac- ous due to their lead (Pb) (6.29 mg/kg), cobalt (163,544 mg/kg), copper tion and electrolysis (Shuva and Kurny, 2013). It is estimated that 250 (98,694 mg/kg) and nickel (9525 mg/kg) contents with exceeded limits t of ore (spodumene) or 750 t of brine or 28 t of lithium ion batteries of chromium, lead, arsenic and thallium (Bernardes et al., 2004; Kang of mobile phones and laptops or 256 batteries of electric vehicles et al., 2013). Human and environmental exposures to these chemicals (EVs) are required to produce 1 t of lithium (Tedjar, 2013). Lithium con- are typically regulated during the manufacture of lithium batteries centrate obtained mainly by the flotation of pegmatites (ore), is pulver- through occupational health and safety laws, and potential fire hazards ized and leached in hot acid, and lithium is precipitated as lithium associated with their transportation. These findings support the need carbonate (Tahil, 2010). The processing of pegmatites is expensive as for stronger government policies at the local, national, and compared to that of the brines due to the heating and dissolution international levels to encourage recovery, recycling, and reuse of lithi- steps involved, but the higher metal concentration in pegmatites partly um battery materials. In view of the above, efforts must be made to compensates for the cost. Because of the cost factor in the lithium develop an environmentally benign and economically viable technology extraction from brine compared to the ores, many deposits of spodu- for recycling spent LIBs. mene are not currently being mined/processed. Lithium is also present This review focuses on the primary and secondary resources of in seawater, but the concentration is too low to be economical. As lithium available for exploitation and provides comprehensive details regards lithium metal, it can be produced by both carbothermic reduc- on the conventional/currently practiced lithium extraction methods tion and metallothermic reduction of oxide (at times hydroxide) and vis-a-vis the resource type. The resources that are covered include also by electrolysis of LiCl (Kipouros and Sadoway, 1998). In view ores/minerals/clays and brines/seawater and bitterns, and lithium ion of the scattered literature on the extraction of lithium from primary batteries for the hydrometallurgical recovery of lithium. resources viz., ores, minerals and the brines, it is considered worthwhile to review the details and discuss critically the merits and demerits of 2. Resources of lithium various processes in vogue or being developed. With the increasing use of LIBs in mobile phones, laptops, cam- 2.1. Primary resources — minerals/clays and brines corders, tracking systems, military and medical devices and in large energy storage systems including that of transportation applications Lithium is produced from a variety of natural sources, e.g., minerals (None million EVs expected by 2015), there will be a significant pressure such as spodumene, clays such as hectorite, salt lakes, underground on lithium resources and its supplies (Kuo, 2011). The disposal of spent brine reservoirs etc. Lithium is a minor component of igneous rocks, batteries may involve landfilling, stabilization, incineration or recycling. primarily granite. The most abundant lithium containing rocks/minerals In landfills, heavy metals have the potential to leach slowly into the soil, are pegmatites, spodumene and petalite. Other minerals are lepidolite, groundwater or surface water. amblygonite, zinnwaldite and eucryptite (Ferrell, 1985). Zinnwaldite The methods for recycling spent LIBs are based mainly on pyro-/ is the impure form of lepidolite with higher content of FeO (up to hydro-metallurgical processes (Li et al., 2009a). The disadvantage of all 11.5% Fe as FeO) and MnO (3.2%) (Paukov et al., 2010). Pegmatites con- pyro-recycling processes is that lithium is not recovered. The traditional tain recoverable amounts of lithium, tin, tantalum, niobium, beryllium pyrometallurgical processes can burn off all the organic electrolyte and and other elements. Table 1 lists the principal commercial lithium min- binder, and facilitate the leaching of valuable metals. In the hydrometal- erals found in pegmatites along with their composition. The theoretical lurgical processes, the dismantled electrodes are dissolved in concen- lithium content in these minerals is 3% to 5.53%, but most mineral de- trated acid and the metal rich leach solutions are treated to recover posits have around 0.5%–2% Li and the pegmatite-bearing ores that are the individual metal by the different methods mentioned above. These often exploited have b1% Li (Mohr et al., 2010). Spodumene is the pri- processes may produce wastewater containing fluoride which is difficult mary lithium mineral being mined. to treat, and can pollute the environment due to the incomplete Among the clay minerals, hectorite, a type of smectite is rich in lith- recycling of the organic binder and electrolyte. There is an inconsistent ium and magnesium, and generally contains 0.3 to 0.6% Li. The best 194 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208

Table 2 Geographical availability for various lithium resources and their lithium contents.

Resources Li content % (w/w) Location of the largest amount References

Pegmatites Spodumene 1.9–3.3 Australia Industrial Minerals and Rocks (2006), Legers (2008), Lepidolite 1.53–3.6 Zimbabwe Clarke (2013), AMRA (2013) Zabuyelite 17–18.75 China Petalite 1.6–2.1 Zimbabwe Amblygonite 3.5–4.2% Canada, Brazil, Surinam Eucryptite 2.34 Zimbabwe

Sedimentary rocks Smectite (hectorite) 0.27–0.7 Hector, California. and Nevada, US Industrial Minerals and Rocks (2006) Lacustrine evaporites 1.8% lithium oxide Jadar Valley, Serbia

Brine Continental b2700 mg/L (with K, B) Chile Evans (2008,2009), ~532 mg/L Bolivia Legers (2008), Geothermal b400 mg/L (with Mn and Zn) United States Clarke (2013), Oil ﬁeld b700 mg/L (with Br) United States Mohr et al. (2012) Sea water 0.18 mg/L Chile, Argentina, China and Tibet

Secondary resource Spent LIBs 2–7% – Shuva and Kurny (2013)

known hectorite deposit with 0.7% Li is in Hector, California. Flint clays plant. Lithium carbonate is mostly produced from both ores and brines and other high-alumina clays contain b0.01 to 0.5% Li. Jadarite is a and the production ﬁgures are often expressed as lithium carbonate newly discovered lithium-boron containing mineral found in Serbia equivalent (LCE). Other chemicals such as lithium chloride and hydrox- (Mohr et al., 2012). These minerals are often concentrated to around ide are also produced in varying amounts. 2%–4% Li for use in the ceramics and glass industry (Garrett, 2004). In India small pocket deposits mostly comprising of lepidolite in Seawater contains about 0.1–0.2 mg/L Li (Bach and Wasson, 1981). pegmatites of mica ﬁelds are located. The maximum lithium content

Total amount of metallic lithium in seawater (globally) is estimated to (lepidolite with 2–6% Li2O) is found in Jharkhand followed by that of be ~230 Gt. Brine sources include lithium found in salt water deposits — Chhattisgarh (2.56% Li2O) and Rajasthan (2.25% Li2O as pegmatites in lakes, salars, oilfield brines, and geothermal brines. Oilfield brines are Udaipur, Bhilwara, Jodhpur and Ajmer, and zinnwaldite in Dagana). underground brine reservoirs that are located with oil. Geothermal brines Others include spodumene in Raichur, Karnataka (Banerjee et al., 1994), are underground brines naturally heated, e.g., in the Salton Sea California. amblygonite in granitic rocks of Paddar (Kashmir) and lepidolite at Brines containing lithium make up 66% of the world's lithium resources; Dhir-Bil (Goalpara), Assam. Lithium bearing bauxite has also been identi- pegmatites make up 26% and sedimentary rocks make up 8% (Gruber, fied in the Salal area, Jammu (Brown and Dey, 1955; Krishnaswamy, 2010; Kesler et al., 2012). 1979; Roonwal et al., 2005). Almost 70% of the global lithium deposits are concentrated in South America's ABC (Argentina, Bolivia and Chile) region. Table 2 details the geographical availability of various resources of lithium vis-a-vis lithium 2.2. Secondary resources — lithium ion batteries content and their locations. The lithium concentrations in the salars of Chile, Argentina, and Bolivia are in the range 0.04–0.16%. According to Out of the various secondary resources, spent LIBs are the most prom- Yaksic and Tilton (2009) the resource of lithium is estimated to be inent secondary source of lithium and other metals. To recycle these 64 Mt. Chile has the world's largest resource of brine (7.5 Mt, 1500– materials, it is desired to understand the construction/composition of 2700 mg/L Li) containing lithium, followed by Bolivia (resource: 9.0 Mt the cells in brief and how they transform during their use. with 532 mg/L Li) and Argentina (resource: 2.6 Mt, 400–700 mg/L Li) Lithium ion battery is a term generally used for a battery which has and these three countries account for almost 80% of the world's brine lithium metal, lithium alloy or material adsorbing lithium ions for its reserves (Mohr et al., 2012). Estimates of lithium resources are published negative active material. LIB uses carbon as an anode and lithium ions extensively (Clarke and Harben, 2009; Gruber et al., 2011; Ono, 2009; exist in the carbon material; there is no metallic lithium at any state of Evans, 2010a, 2010b; USGS, 1980, 1986, 2005, 2009, 2010, 2011, 2012, charge during normal usage. Depending on their technical construction 2013). and properties batteries are categorized as either primary or secondary. Lithium rock production began with lithium minerals (1899) in the From the legislative view point batteries are also categorized as portable USA (Garrett, 2004). Since the first lithium production from brines at (household) and vehicle or industrial batteries. Primary cells are con- Searles Lake, USA in 1936, brines are exploited largely in South America structed with metallic lithium. The metallic lithium in a non- and China. The largest producer of lithium in the world is Chile where rechargeable primary lithium battery is a combustible alkali metal that lithium is extracted from salar brine at the Atacama Salt Flat. Lithium self-ignites at 178 °C, and when exposed to water/seawater reacts exo- production from brines is also at salt lakes in Tibet and Qinghai in thermically and releases hydrogen. Primary batteries are single-use as China, besides at Nevada in the United States. Several newer installations irreversible discharge reactions occur in the cells and after use they (by 2013) are on various stages of exploration/operation for brine source are disposed off. Secondary battery cells have a chemistry that allows re- which are: one more in China, six in Argentina, three in Chile and one in versing the discharge reaction and are rechargeable. LIBs are of the re- Bolivia (Clarke, 2013). Currently 8% of lithium is obtained from salt lake chargeable secondary type. The functional parts of LIBs are the cathode, brines and sea by sedimentation. Significant quantities of lithium com- anode, electrolyte and separator, which are housed in a protective pounds and ore concentrates are also produced in Australia, Canada, metal casing. Portugal, Russia and Zimbabwe. Currently, Australia produces lithium The chemical reaction in the cell expressed below shows the forms concentrate from spodumene at the mines in Mt Catlin, Western in which lithium and cobalt can be present in the spent LIBs. Australia. The 160,000 t/annum concentrate produced in Western þ → þ : ð Þ Australia is processed to 16,000 t/annum lithium carbonate in its Chinese LiCoyOz 6C LixC6 Li1−xCoyOz 1 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208 195

Table 3 Various components of LIBs, materials used and their meritsa.

Component Wt.% of the total Material Structure Properties/Merits weight of the battery

Cathode 39.1 ± 1.1 LiCoO2 Layered High structural stability and can be cycled for N500 times with 80–90% capacity retention

LiMn2O4 Spinel Attractive for ecological and economic reasons; discharges ~3 V

LiNiO2 Layered Cheaper & possesses higher energy density (15% higher by volume, 20% higher

by weight), but less stable & less ordered as compared to LiCoO2

LiFePO4 Olivine Suitable for biomedical applications because of higher safety levels and

Li2FePO4F Olivine lower cost

LiCo1/3Ni1/3Mn1/3O Layered/spinel Possesses high capacity with structural and thermal stability, and safe to use

Li(LiaNixMnyCoz)O2 Layered/spinel Anode Carbon Graphite Low cost and availability. It has the ability to reversibly absorb and release large Hard carbon Microspheres quantity of Li (Li:C = 1:6)

Electrolyte Lithium salt like LiPF6, Li[PF3(C2F5)3] Withstands high temperature and possesses high mobility of Li ions

or LiBC4O8 in organic solvents Plastic case 22.9 ± 0.7 Polyethylene terephthalate layers, a polymer layer and a Hermetically sealed battery body which converts chemical energy to electrical polypropylene layer, layers of carbonized plastic energy in order to generate current Outer casing 10.5 ± 1.1 Stainless steel, aluminium Copper foil 8.9 ± 0.3 Copper ~14 μm thick Aluminium foil 6.1 ± 0.6 Aluminium ~20 μm thick Polymer foil & 5.2 ± 0.4 Polyethylene, polypropylene or composite Use of 3–8 μm layers (PP/PE/PP) with 50% porosity electrolyte polyethylene/polypropylene ﬁlms Solvent 4.7 ± 0.2 Ethylene carbonate, dimethyl carbonate Non-aqueous &diethylcarbonate Electrical contact 2.0 ± 0.5 Aluminium and copper Conductive

a Wakihara and Yamamoto (1998), Gaines and Cuenca (2000), Nazri and Pistoia (2004) Paulino et al., (2008) and Fergus (2010).

The active mass (cathode, anode and electrolyte) of LIBs comprises of and Olson (1978) have reviewed methods and techniques for the almost 40% of the weight whereas ~30% (wt) of these components are extraction of lithium from ores, brines and clays. Processes followed carbon (anode). A number of chemical species in the cathode material for the extraction of lithium from different resources have also been within the lithium-ion family have been reported including the type, compiled in detail by Garrett (2004). structure of the materials used and their electrochemical properties In order to process ores/concentrates, acid digestion with H2SO4 may which are listed in Table 3. Generally LIBs have a short life of 2–3years be followed for decomposition of the silicate structure at 250–400 °C whether they are used or not. The spent batteries are a good source of which is suitable for the processing of lepidolite, amblygonite and several metals like Li, Co, Ni, Co, Mn etc. zinnwaldite (Kondás and Jandová, 2006). In the sulfate process lithium minerals such as lepidolite are decomposed at high temperature in the 3. Extraction of lithium from primary resources presence of potassium and/or sodium sulfate. Alkali digestion is suitable for the decomposition of spodumene and lepidolite largely by the treat- 3.1. Lithium extraction from minerals/clays ment of potassium carbonate to produce lithium hydroxide. In the alkaline/gypsum process the mineral is reacted with limestone or a mixture Lithium is extracted from its minerals by two processes — acid and of calcium sulfate with calcium oxide and/or hydroxide by heating to alkaline though chlorination is also attempted in some cases. Averill convert silicate to soluble lithium aluminate from which LiOH or Li2CO3

Table 4 Lithium extraction from its minerals/clays by acid process.

Raw material % Li Experimental conditions %Li Li2CO3 References Extraction purity Pre-treatment Roasting Time Water leaching (%) temp. (°C) (h) L/S ratio Temp. (°C) Time (h)

Lepidolite conc. 2.0 Sulfation roasting 850 0.5 2.5:1 Room temp. 0.5 91.6 – Yan et al. (RT) (2012a) Lepidolite conc. 2.0 Salt roasting with 880 0.5 0.8:1 RT 0.5 ~90 N99.5 Yan et al.

Na2SO4 + CaCl2 (2012b) Lepidolite conc. 2.55 Sulfation roasting 1000 0.5 15:1 85 3 90.4 – Luong et al.

Na2SO4:Li molar (2013) ratio = 2:1 Lepidolite conc. 1.79 Roasting with iron 850 1.5 1:1 RT 1 85 (open – Luong et al. sulfate system) (2014) 93 (closed system) Zinnwaldite conc. 0.96 Roasting with 850 1 10:1 RT 0.5 N90 N90 Siame and sodium sulfate Pascoe (2011) Petalite/Lepidolite 1.9 Calcination 1100 2 7.5/1 50 1 97.3 99 Sitando and

Roasting (H2SO4) 300 1 Crouse (2012) Spodumene 4.21 Calcination 1100 1 Leaching & ––90 – Mcketta (1988),

Roasting (H2SO4) 250 0.167 precipitation Tahil (2010)

Spodumene 2.81 Roasting & H2SO4 1050–1090 0.5 1:4 225 1 96 ~99.6 Chen et al. leaching (2011a, 2011b), Clarke (2013)

Montmorillonite clay 1.2% Li2O Without roasting ––5:1 (H2SO4 leach) 250 1.5 90 Li2SO4 Amer (2008) 196 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208

Table 5 Lithium extraction from its minerals/clays by alkali process.

Raw material % Li Pre-treatment Experimental conditions % Li Li2CO3 References extraction purity (%) Roasting temp. Water leaching (°C), time (h) Time (h) Temp. (°C) (L/S ratio)

Lepidolite 1.4 Deﬂuorination 860, 0.5 h 1 150 4:1; lime: deﬂuor. 98.9 99.9 Yan et al. (2012c) concentrate lepidolite: 1 Zinnwaldite 1.4 Gypsum roast 950, 1 h 0.167 90 10:1 96 99 Jandová et al. concentrate (2009) Zinnwaldite 2.07% Gypsum roast 1050 0.5 85 10:1 84 –

(tailing sample) Li2O Zinnwaldite 0.19 Roasting with 975 1 90 5:1 93 – Kondás and Jandová

concentrate gypsum & Ca(OH)2 (2006)

Zinnwaldite 1.21 Roasting with CaCO3 825, 1 h 1 90–95 5:1 ~90 99.5 Jandová et al. (2010) concentrate

Zinnwaldite 1.29 Roasting with CaCO3 825 4 95 10:1 84 – Vu et al. (2013) concentrate Montmorillonite 0.3–0.6 Roasting with 750 – RT – 80 99 Lien (1985), Crocker

clay–hectorite CaCO3 900 and Lien (1987)

CaCO3–CaSO4

is obtained. Thus most of the alkaline processes involve either the Li2CO3 of ~99% purity was precipitated by the addition of Na2CO3 at heating of lithium minerals with alkali salts or in more advanced hydro- 95–100 °C. Amer (2008) reported the extraction of 90% Li from the thermal processes by decomposition in solutions containing Na2CO3, Egyptian montmorillonite type clay containing lithium in 90 min of NaOH, Na2SO4 and/or other alkali salts at elevated temperature and reaction in sulfuric acid at 250 °C. pressure. Ion-exchange processes are applied for the processing of the In order to process spodumene it is desired to convert α-spodumene leach liquors obtained from spodumene, petalite and partly from to β-phase by roasting at 1070–1100 °C (Chen et al., 2011b; Clarke, zinnwaldite. Tables 4–6 summarize recent work on lithium extraction 2013; Tahil, 2010). Tahil (2010) reported the roasting of spodumene from its minerals and clays by using different approaches. in a kiln at ~1100 °C. The calcine was mixed with sulfuric acid and roasted at 250 °C and then leached in water to yield a solution of lithium

sulfate. Reaction of β‐spodumene with H2SO4 is shown as reaction (2) 3.1.1. Acid process (Mcketta, 1988): Sulfation roasting of lepidolite followed by water leaching was recently reported by Yan et al. (2012a). The lithium extraction effi- ⋅ ⋅ þ → þ ⋅ ciency of 91.6% could be achieved at a mass ratio of lepidolite/Na2SO4/ Li2O Al2O3 4SiO2ðsÞ H2SO4ðconcÞ Li2SO4ðsÞ Al2O3 4SiO2ðsÞ þ : ð Þ K2SO4/CaO of 1:0.5:0.1:0.1 and roasting at 850 °C (Table 4). Roasting at H2OðgÞ 2 880 °C with a mass ratio of lepidolite/Na2SO4/CaCl2 of 1:0.5:0.3 resulted in improved recovery (~95% Li) of lithium (Yan et al., 2012b). Above 99.5% pure lithium carbonate was obtainedbyevaporationandprecipita- Lithium carbonate can be recovered by the addition of sodium carbonate to the solution after pH adjustment, purification and evaporation tion with Na2CO3. Luong et al. (2013) examined the sulfation roasting of a Korean lepidolite ore with sodium sulfate at 1000 °C, while achieving (Reaction (3)). extraction of ~ 90.4% Li. Recently, the roasting of lepidolite with iron sulfate at 850 °C and water leaching were investigated by Luong et al. þ → þ : ð Þ Li2SO4ðaqÞ Na2CO3ðaqÞ Li2CO3ðsÞ Na2SO4ðaqÞ 3 (2014). Leaching of the calcines obtained from the open and closed systems yielded leach liquors containing ~7.9 g/L Li and ~8.7 g/L Li, corresponding to the extraction of ~85% and ~93% Li, respectively. The world's first continuous plant to convert spodumene concentrate

Lithium extraction (N90%) by roasting a low grade zinnwaldite to lithium carbonate by calcination, roasting of calcine with H2SO4 and concentrate (0.96% Li) with sodium sulfate followed by water leaching, water leaching, was commissioned in 2012 by Galaxy Resources in was described by Siame and Pascoe (2011).AstudybySitando and China (Clarke, 2013). Crouse (2012) showed the extraction of 97.3% Li (~5 g/L) by roasting One of the drawbacks of the sulfuric acid method to treat lepidolite, a Zimbabwean petalite ore concentrate (4.1% Li2O) with concentrated petalite and zinnwaldite is the requirement of a high concentration of sulfuric acid at 300 °C followed by water leaching at 50 °C. The leach acid and complex puriﬁcation processes, whereas spodumene needs liquor was evaporated till lithium was concentrated to N11 g/L and to be converted to the more leachable β‐phase at higher temperature.

Table 6 Lithium extraction from its minerals/clays by chlorination process.

Raw material Experimental conditions % Li extraction Product References

Roasting temp. (°C) Time (h) Leaching

L/S ratio Temp. (°C) Time (h)

Lepidolite 935 13 – 80 1 ~100 LiCl Löf and Lewis (1942) Spodumene (3.58% Li) 1000 4 10:1 80 1 58 LiCl Peterson and Gloss (1959)

Lepidolite concentrate (2.0% Li) 880 0.5 2.5:1 60 0.5 92.8 Li2CO3 Yan et al. (2012d)

Spodumene (7.2% Li2O) 1100 2.5 –– – 100 LiCl Barbosa et al. (2014)

Lepidolite ore (3.70% Li2O) With CaCO3 + CaCl2, 950 5 10:1 90 1 80.0 LiCl Vyas et al. (1975) P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208 197

3.1.2. Alkaline process 750 °C, ~98% of the lithium contained in spodumene is converted to its Inthealkalineprocess(Table 5), spodumene or lepidolite ore concen- chloride which can be water leached (Zelikman et al., 1996). Vyas trates are ground and calcined with limestone at 825–1050 °C. The et al. (1975) reported a similar process using CaCO3 and CaCl2 to roast resulting calcine is crushed, milled and treated with water to yield lithi- Indian lepidolite at 950 °C by which 80% Li was recovered as LiCl. The um hydroxide which can be converted to chloride by reaction with hy- chlorination process with calcium chloride and/or sodium chloride drochloric acid. The recovery in this method is approximately 85–90% with lepidolite (M = Li, K, Rb, Cs) may be represented as: (Averill and Olson, 1978). The reaction during calcination of spodumene with limestone can be represented as : þ þ þ → þ ð Þ 2NaCl 6SiO2 M2O Al2O3 2NaAlSi3O8 2MCl 8 ⋅ ⋅ þ → þ ⋅ ⋅ Li2O Al2O3 4SiO2ðsÞ CaCO3 Li2OðsÞ CaO Al2O3 4SiO2ðsÞ þ : ð Þ þ þ → þ ð Þ CO2ðgÞ 4 CaCl2 SiO2 M2O CaSiO3 2MCl 9 Lepidolite was pre-roasted at 860 °C under water steam atmosphere for defluorination followed by pressure leaching of the defluorinated þ þ þ → þ : ð Þ CaCl2 2SiO2 Al2O3 M2O CaAlSi2O8 2MCl 10 mass in a lime–milk autoclave at 150 °C. In this process 98.9% lithium was extracted (Yan et al., 2012c). During the roasting with steam formation of phases such as lithium aluminium silicate (beta-eucryptite) and Chlorination roasting with a mixture of CaCl2 and NaCl gives a better lithium extraction yield because of its lower melting point than either leucite (KAlSi2O6)wasnoticed(Karstetter, 1971)asperreaction(5) of the agents, which increases the fluidity of chloride melt. This allows ⋅ ⋅ ⋅ þ → ⋅ ⋅ þ 2LiF KF Al2O3 3SiO2 4H2O Li2O Al2O3 2SiO2 2KAlSi2O6 diffusion of the chlorinating agent to the surface of the lepidolite facili- þ þ : ð Þ 4H2 2F2 5 tating the lithium extraction selectively and yielding a pure product. Crocker and Lien (1987) also reported a process for selective chlorina- Extraction of lithium from a zinnwaldite containing waste was inves- tion of hectorite (0.3–0.65% Li) in clays with limestone at 750 °C using tigated by Jandová et al. (2009). The concentrate (1.40% Li) obtained 20 wt.% HCl. from dry magnetic separation of the waste (0.21% Li) was treated by the gypsum roast method (with CaSO4 and Ca(OH)2). About 96% Li was leached out from the sinter made at 950 °C. Earlier the concentrate 3.1.4. Other processes prepared from this waste by gravity and dry magnetic separation, was Chlorination roasting of ores requires corrosion-resistant equip- roasted by the gypsum method at 975 °C and water leached to recover ment. To overcome this drawback the autoclave method is used. Chen ~93% Li (Kondás and Jandová, 2006). In another study Jandová et al. et al. (2011b) reported a process to treat β-spodumene obtained in (2010) reported that the roasting of the above concentrate with the calcination of α-spodumene, by sodium carbonate solution in an au- CaCO3·Li2CO3 with 99.5% purity was separated from the leach liquors toclave at a liquid/solid ratio of 4 and Na/Li of 1.25 at 225 °C. During by either converting the alkaline liquor to the carbonated solution by pressure leaching lithium carbonate and analcime slurry are formed CO2 bubbling or by solvent extraction with LIX54 and TOPO as an according to reaction (11). extractant followed by stripping with diluted H2SO4;thefirst method provided a higher yield. Siame and Pascoe (2011) obtained recovery of β− ⋅ ⋅ þ þ → Li2O Al2O3 4SiO2 Na2CO3 nH2O Li2CO3 þ ⋅ ⋅ ⋅ : ð Þ ~84% Li from the roasted zinnwaldite concentrate at 1050 °C with lime- Na2O Al2O3 4SiO2 nH2O 11 stone, gypsum and sodium sulfate. A similar process was reported by Vu The slurry was leached in carbon dioxide to form lithium bicarbonate et al. (2013) from zinnwaldite by sintering with CaCO3 powder at 825 °C, followed by water leaching and precipitation. which on heating produced lithium carbonate of 99.6% purity as per Crocker and Lien (1987) reported the recovery of N85% Li by roasting reactions (12) and (13). montmorillonite type clays (0.3–0.6% Li) with KCl–CaSO4 or CaCO3– CaSO4 followed by water leaching. Lithium silicate in the clay was con- þ þ → ð Þ Li2CO3 CO2 H2O 2LiHCO3 12 verted to Li2SO4 by roasting a pelletized mixture of clay, limestone and gypsum at 900 °C in a direct gas-fired rotary roaster (Lien, 1985). → þ þ : ð Þ Formation of lithium sulfate may be represented by reactions (6) and (7) 2LiHCO3 Li2CO3 CO2 H2O 13

þ → þ þ 1 ð Þ Medina and El-Naggar (1984) developed an alternate method of CaSO4 2H2O SiO2 CaSiO3 SO2 O2 6 2 chlorination to treat spodumene with a recovery of ~87% Li by roasting þ þ 1 → þ : ð Þ at 1150 °C with a mixture of 8:1 (wt.) tachyhydrite:spodumene follow- Li2Si2O5 SO2 O2 Li2SO4 2SiO2 7 2 ed by water leaching. Because the production of lithium carbonate from spodumene is energy intensive and expensive, lithium carbonate is The initial product obtained from the alkaline process is lithium mostly produced from the brines. hydroxide which can be converted to carbonate or chloride salt.

3.1.3. Chlorination process 3.2. Lithium extraction from brines/sea water/bitterns A less common method is the chlorination process in which the ore is roasted in the temperature range 880–1100 °C in the presence of chlorine In order to meet the growing demand of lithium, brines and bitterns gas or HCl depending upon the type of mineral treated (Table 6). For have received increasing attention. Tables 7 and 8 elaborate the summa- instance chlorination of lepidolite with HCl can be achieved at lower tem- ry of work carried out recently for the extraction of lithium from sea perature (935 °C) while giving a high yield (~100%) of lithium during water/brines/bitterns. Production process for brine-water lithium costs leaching as compared to the roasting of spodumene (Löf and Lewis, 30% to 50% less than that of the mined ores (Abe, 2010). As mentioned 1942). Lithium in spodumene can be converted to LiCl almost quantita- earlier, lithium carbonate is produced from brines by an evaporative con- tively at higher temperature (1100 °C) in 2.5 h with Cl2 gas (Barbosa centration and reﬁning method. Firstly, brine is concentrated by solar et al., 2014); the recovery is found to be quite low (58%) at the lower evaporation over a year in a pond to crystallize sodium, potassium and temperature (1000 °C) (Peterson and Gloss, 1959). In the chlorination magnesium chlorides. During the reﬁning process calcium carbonate process, when the ore is sintered with NH4Cl and CaCl2 in a furnace at is roasted and then added to the solution of LiCl for the removal of 198 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208

Table 7 Adsorption process for lithium extraction with product synthesized from seawater, brines and bitterns.

Source/Raw material Adsorbent used Conditions Adsorption/Remarks Product References

Seawater H1.6Mn1.6O4 Adsorbent: 200 mg, sea water: 50 L, Max. uptake: 40 mg Li/g adsorbent – Chitrakar et al. 4 weeks; desorption by 0.5 M HCl, 1 day (2001) Seawater (Li 0.17 mg/L) Spinel type Mn-oxide 30 °C, 15 days b90% Li, LiCl Umeno et al. Li uptake: 10.6 mg/g adsorbent (2002)

Seawater (Li 0.15 mg/L) λ-MnO2 (granulated) 150 days Recovery—264 g LiCl in 791 g dried LiCl Yoshizuka precipitated salt (816 m3 seawater) et al. (2006) Artiﬁcial seawater Ion-sieve type Mn-oxide Adsorption at 0.4 M HCl, 5 days Adsorption: 88% Li by adsorbent I and 89% Li LiCl Chung et al.

(Li 0.2 mg/L, 8.01 pH) spinels: HMg0.5Mn1.5O4(I) by adsorbent II. (2004, 2008)

HZn0.5Mn1.5O4 (II) Equil. sorption: 30.3 mg/g (I) and 33.1 mg/g (II).

Seawater (1 mmol/L Li) HMn2O4 60 °C, 24 h Loading capacity: 1.53 mmol/g sorbent Li salt Wajima et al. (2012) III IV Brine (Salars de Uyuni, LimMgxMn yMn zO4 24 h, pH 6.5 Adsorption capacity: 23–25 mg/g Li2CO3 Chitrakar et al. Bolivia) (0 b x ≤ 0.5) adsorbent (2013) Seawater (Li Surface deposition on 30 °C, 10 days 34% Li – Takeuchi 0.192 mg/L) corrosion product of Al (1980) Salt lake bitterns Hydrated alumina: pH 5.8 Adsorption:0.6–0.9 mg/g sorbent – Dong et al. LiOH = 2 M ratio (2007)

Egyptian bitterns Al(OH)3 30 °C, pH 9 Adsorption capacity: 123 mg/g adsorbent LiAlO2 Hawash et al. (19.5, 5.5, 8.8 mg/L) (2010)

Brine (Salar de Hombre Hydrated alumina LiCl solution— Adsorbed Li eluted by acid and Li2CO3 Clarke (2013) Muerto, Argentina) 1% Li (20 times conc.) by solar evaporation. precipitated by sodium carbonate

+ Mg(OH)2.Ageneralflowsheet of Li2CO3 production from brine water is medium (pH of seawater being ~8) for Li in the presence of alkali and shown in Fig. 1. alkaline earth ions. For instance Kitajou et al. (2003) reported the separa- + + The basic approaches for the separation of mineral products (K, Mg, tion of Li from a large amount of Na by the spinel-type λ-MnO2 Na, Ca, Li) from the seawater comprises of flotation (using anionic col- whereby Li+ was concentrated 400 times leaving most Na+ in the sea- lectors), sorption, ion exchange, solvent extraction etc. as described by water. Extraction/separation of lithium from brines and such resources Koyanaka and Yasuda (1977). The existing evaporation process for lith- is summarized in Table 7. ium recovery from brine lakes is time consuming and suffers from low The manganese oxide (H1.6Mn1.6O4) prepared from the precursor, recovery efficiency. Besides, tremendous burden is posed on the envi- Li1.6Mn1.6O4 by hydrothermal and reflux methods, showed the maxi- ronment due to waste generation and substantial water consumption. mum uptake of 40 mg Li/g of adsorbent from the seawater, the highest among the inorganic adsorbents (Chitrakar et al., 2001). The very fine 3.2.1. Adsorption process size (nano-size range) of the synthesized manganese oxide was found Various types of adsorbents have been used for selective lithium to be responsible for its high adsorption capacity towards lithium as recovery from seawater and brines. In the adsorption method certain compared to other adsorbents. Adsorption of lithium from seawater by + inorganic ion-exchangers such as the spinel-type manganese oxide a spinel type λ-MnO2 produced low purity (~33%) Li ions contaminat- show extremely high selectivity for lithium from seawater (Kitajou ed with Na+ (Yoshizuka et al., 2006). Chung et al. (2004) synthesized et al., 2003; Ooi et al., 1989, 1991; Umeno et al., 2002; Yoshizuka et al., nano-manganese oxide (Li1.33Mn1.67O4) through a gel process. The ion 2002, 2006). Such materials exhibit high adsorption capacities in alkaline sieve-type adsorbent containing magnesium after acid treatment was

Table 8 Precipitation and other processes for lithium extraction from seawater and brines.

Source/Raw material Process Conditions Recovery (%)/Remarks Product References (% purity)

Synthetic solution, Precipitation pH 12.5 for [Al]: 50–1000 mg/L 70% Li – Yoshinaga et al. Geothermal water (1986) (Li = 10 mg/L)

Synthetic solution (2.5 M Precipitation followed by IX using Poly BD® Precipitation with 1.8 M Na2CO3 Good usable volume cap. Li salt Bukowsky et al.

LiCl, 0.3 M CaCl2 and 0.15 M R45HTLO (MC50), Lewatit® (TP 207), at 80 °C. of resin TP207: solution (1991)

MgCl2) Dowex®(Y80) IX—50 °C, 30 min 56.4 g/L

Seawater (0.12–0.16 mg/L Li) Integrated ion-exchange method 1st stage: sorption on λ-MnO2, 56% yield Li2CO3 Nishihama et al. 150 days (N99.9) (2011), Onishi et al. (264 g LiCl in 816 m3 seawater); (2010) 2nd stage: Sep. of Mg(II), Ca(II), Sr(II) and Mn(II) with SK110 resin (pH 9)

Seawater (0.18–0.20 mg/L Li) Two stage precipitation 1st stage pH : 11.5–12.5; Recovery of pure Li2CO3 (99.4) Um and Hirato

2nd stage: Na2CO3 at 100 °C lithium carbonate (2012)

Uyuni Salar brine, Bolivia, Two stage precipitation with lime 1st stage pH: 11.3, Precipitation as Li2CO3 LiCO3 (99.55) An et al. (2012), 15–18 g/L Mg, & Na-oxalate 2nd stage: sodium oxalate, Tran et al. (2013) 0.7–0.9 g/L Li 80–90 °C Brine from Salar de Hombre Precipitation (2.5 g/L LiCl from solar pond) Separation of Mg as hydroxide and LiCl salt feed for Li extr. LiCl Clarke (2013)

Muerto, Argentina with lime and Na2SO4 Ca as sulfate in a chemical plant Brine (high Mg/Li ratio) Electrochemical Electrolyte: 0.5 M NaCl, 10 h ~94% Li; – Zhao et al. (2013) Li sorption: 28.65 mg/g

LiFePO4 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208 199

Brine and the Korea Institute of Geoscience and Mineral Resources, Korea had joined hands to build a pilot plant for the commercial production of Evaporation concentration lithium carbonate from sea water based on the adsorption process (USGS, 2009), the outcome of which remains uncertain presently.

Concentrated brine 3.2.2. Precipitation process Some of the precipitation and other processes used for lithium recov- Lime milk ery from seawaters/brines are summarized in Table 8. Among various co- Magnesium removal precipitating agents used, aluminium salts show the best performance for lithium recovery from geothermal water. The appropriate pH for

lithium recovery is 10–13 and use of NaAlO2 seems better than Magnesium AlCl3.WithahighpurityNaAlO2 solution as co-precipitating agent, Filtration /Washing hydroxide about 98–99% Li recovery was achieved at pH 11.5 from Ca- and SiO2-free geothermal water (Yoshinaga et al., 1986). Despite being the richest lithium resource (10.2 Mt) Uyuni salar Calcium-lithium-salt solution brine containing a high magnesium concentration (Mg/Li mass ratio ~21.2:1) causes difficulties in lithium production. The high magnesium content represents a significant metal value which should be recovered Lithium Calcium removal carbonate with lithium. Consequently, calcium and magnesium had to be removed from the brine by using oxalic acid before the production of lithium. The Mg-oxalate produced was suitable for use as a precursor for the production of MgO by roasting (Tran et al., 2013). An et al. (2012) developed a Calcium carbonate Filtration hydrometallurgical process to recover lithium from Uyuni salar brine containing 15–18 g/L Mg and 0.7–0.9 g/L Li saturated with sodium, chlo- Lithium salt solution ride and sulfate. In a 2-stage precipitation process, magnesium and Roasting sulfate were removed as Mg(OH) and gypsum (CaSO ·2H O) at Evaporation 2 4 2 Na2CO3 pH 11.3 by lime in the first stage. Residual magnesium after lime precip- Precipitation crystallization itation and almost all soluble calcium were then removed by the addi- Hydration tion of sodium oxalate. In the second stage 99.6% lithium carbonate was precipitated at 80–90 °C using sodium carbonate. Residual Li+ Filtration/Washing from the solution was quantitatively extracted in the presence of other alkali metals by a mixture of commercial β-diketones (LIX-51) and TBP (Miyai et al., 1988). A 2-stage lime precipitation process to Lithium carbonate treat seawater was reported by Um and Hirato (2012) to separate lithium from calcium and magnesium, whereas Clarke (2013) men- fl Fig. 1. A owsheet for Li2CO3 production from brine. tioned the use of precipitation method to remove the two metals (magnesium and calcium) from the salar brine of Argentina to produce LiCl then generated which selectively adsorbed lithium (~30.3 mg/g adsor- which was further processed to recover lithium. bent) from seawater. When a polymeric membrane reservoir containing an inorganic ion-exchange adsorbent inside it with zinc was used, lithi- 3.2.3. Ion exchange/Solvent extraction process um recovery from seawater was very effective and kinetically favored For high magnesium and calcium containing bitterns or brines solvent with adsorption of 33.1 mg Li/g sorbent (Chung et al., 2008). This adsor- extraction or ion exchange can be used. After selective stripping/elution bent had excellent lithium adsorption of 89% of 400 mg Li in a day only; lithium can be precipitated out. A combined process consisting of solar the desorption efficiency being 92.88% by dipping in 4 L of 0.5 M HCl so- evaporation and ion exchange for the extraction of lithium was proposed lution in a day. By using Mg-doped manganese oxide, Chitrakar et al. by Steinberg and Dang (1975, 1976). The Dowex resin was used for selec- (2013) also observed very fast adsorption equilibrium (within 24 h) for tive exchange of its H+ with the cations present in sea water in the order: effective recovery of lithium from salar brine (Table 7). K+,Na+,Li+ and Mg2+. Lithium ions were eluted using 0.2–0.5 M HCl

Recently Wajima et al. (2012) prepared HMn2O4 by elution of spinel- and eluted LiCl was transferred into an electrolyser to produce lithium. type lithium di-manganese-tetra-oxide (LiMn2O4) and examined the Strelow et al. (1974) separated lithium from sodium, beryllium and kinetics of lithium adsorption. The intermediate, LiMn2O4, was also syn- many other elements by eluting lithium with 1 M HNO3 in 80% methanol thesized from LiOH·H2OandMn3O4 by acid treatment. Lithium recovery from a column of AG50W-XS, a sulfonated polystyrene cation-exchange from seawater reached ~100% at 60 °C using both products. resin. Samples were loaded onto a 20 cm3 AG50W-XS (200–400 mesh) Aluminium foil immersed in sea water forms a corrosion product on resin column and then eluted with a mixed acid–methanol solution its surface which extracts lithium selectively from sea water at the same (1 M HNO3 and 80% methanol) which ensured an extremely good separa- time, at the optimum temperature of ~30 °C (Takeuchi, 1980). Dong tion of lithium from sodium in a single column pass. et al. (2007) prepared an aluminium salt adsorbent using Al(OH)3 and Earlier studies showed that organic ion-exchange resins exhibited low LiOH at pH 5.8, and a molar ratio of 2, and investigated the recovery of selectivity for lithium ions (Abe and Hayashi, 1984; Alberti and Massucci, lithium from a salt lake bittern by this adsorbent. The adsorbent showed 1970; Ho et al., 1978). However, Bukowsky et al. (1991) demonstrated high adsorption and uptake of 0.6–0.9 mg Li/g rather than the other that precipitation followed by ion exchange can be effectively used for alkali or alkaline metals of bitterns. Use of hydrated alumina for adsorp- separation and recovery of lithium from a synthetic solution of calcium tion of lithium was also reported from Egyptian bitterns (Hawash et al., and magnesium chlorides. Recently Nishihama et al. (2011) applied 2010) and salar brines of Argentina (Clarke, 2013). SK110 resin (sulfonated type) from a concentrated solution to remove The Institute of Ocean Energy at Saga University began operating the divalent metal ions (Mg2+,Ca2+,Sr2+ and Mn2+) due to their world's ﬁrst — but small — lab aiming for practical lithium production higher sorption capacity compared to that of mono-valent ions. from seawater and succeeded in acquiring about 30 g of lithium chloride The separation of Li+ from the resultant solution with Na+ and K+ from 140,000 L of seawater in one month. Similarly in early 2010, POSCO was achieved in a packed column of impregnated resin containing 200 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208

Spent lithium ion batteries 1-phenyl-1,3-tetradecanedione (C11phβDK)/tri-n-octylphophine oxide + (TOPO). Li2CO3 was precipitated from the concentrated Li solution by a (NH ) CO solution with an overall yield of 56% and a purity of 4 2 3 Dry N gas Cut up N99.9%. Earlier this group reported selective adsorption of Li+ in aqueous 2 o chloride media using a novel synergistic solvent impregnated resin (SIR) 50 C Steel β containing both 1-phenyl-1,3-tetradecanedione (C11ph DK) and TOPO Acetonitrile Dissolution of electrolyte (Onishi et al., 2010). About 94% Li+ was adsorbed by SIR containing + 0.66 mmol/g of each extractant at pH 12. Almost 96.9% Li was eluted 50-60 oC with a 1.0 mol/L HCl solution with 99.8% purity at a bed volume (BV) Acetonitrile Solvent evaporation of 2.7. Nelli and Arthur (1970) patented a selective liquid extraction process for lithium from bitterns of high magnesium content. In the presence of Electrolyte stored strong chloride and HCl, lithium was converted into stable lithium tetra chloroferrate (Cl4FeLi2), which can be extracted by a number of solvents. About 90% Li was extracted with a mixture of 20% TBP and 80% diisobutyl 50 oC NMP Dissolution of binder ketone in 7 counter current stages. Any co-extracted magnesium was recycled back to the initial extraction step by washing with water in 4 stages. Finally lithium tetra chloroferrate was stripped in 5 counter current stages. The strip water with ~2 M NaCl was contacted with a mixture of solvent (20% D2EHPA and 30% TBP in benzene) in 6 stages to Filtration Sorting of solids remove FeCl3. The raffinate in this step is the product containing ~0.36% Li, 200 mg/L Mg and 20 mg/L Fe. The second solvent is then 90 oC stripped of its FeCl3 with about 0.3 parts of water in 6 stages. NMP Solvent evaporation Cu Al Plastics Separation of lithium and magnesium is difficult because they exhibit many chemical similarities. However, by utilizing the high charge density 2+ + of Mg which is twice that of Li with almost the same ionic radius and Water + LiOH Electrolysis its easy hydrated properties, Zhao et al. (2013) reported a separation method using LiFePO4/FePO4 as electrodes (Table 8). Lithium exhibits LiOH solution Decantation good reversibility in LiFePO4/FePO4 structures, and the redox peak separation is 0.592 V compared to 1.403 V for Mg2+ indicating its higher polar- ization. At a voltage of 1.0 V in pure lithium solution, the inserted capacity WashingCoO of lithium can reach 41.26 mg/g LiFePO4, which is 93.78% of its theoretical value (44 mg). The subsequent extracted capacity of lithium can be CoO attained to 38.93 mg/g LiFePO4, which is 94.3% of its inserted capacity, while the extraction capacity of Mg2+ from a solution containing magne- Fig. 2. Process flowsheet of AEA technology. sium is only 5.5 mg/g LiFePO4. The lower voltage is found to be beneficial for separating Mg and Li and the method also works for brine to reduce the Mg/Li ratio to 0.45 from 60.

3.3. Extraction of lithium from secondary resources — lithium ion batteries binder-PVDF is dissolved in NMP (N-methyl-2-pyrrolidone) and filtered to separate it from electrode material (Al, Cu, steel and plastics). The NMP Secondary resources like spent batteries have lithium compounds is evaporated from the filtrate and recycled back. Residue is back-washed coupled with valuable metals such as cobalt, nickel, manganese etc. and the suspended particulate (LiCoO2 and carbon) is electrolysed with which can be economically viable to recycle, whereas low value elements LiOH as electrolyte. LiCoO2 is reduced adjacent to the cathode to form such as iron and phosphorous will be a greater challenge to create a prof- CoO while increasing the concentration of LiOH (Fig. 2) which is recov- itable recycling program. World over several companies viz., Toxco, OnTo, ered by decantation. The reaction can be represented as: Sony, Accurec etc. are currently performing and improving battery recycling processes. Though Toxco and Sony were the first to recycle − þ − e þ H O þ LiCoðIIIÞO →CoðIIÞO þ Li þ 2OH : ð14Þ LIBs which they still continue, several others also started processing 2 2 such materials recently (Espinosa et al., 2004; Gaines and Cuenca, 2000; Li et al., 2009a). The Toxco process is designed for all types of lith- 3.3.1.2. Toxco process (Canada). A variety of batteries is processed in cryo- ium containing wastes. The main product is targeted as appropriate genic conditions (in liquid nitrogen) to reduce the reactivity and mechan- (cobalt/other base metals) and the other component being the lithium ical and hydrometallurgical approaches are utilized to recover metal hydroxide. Sony uses incineration which is followed by hydrometallurgi- values (McLaughlin and Adams, 1998). The process flow-diagram is cal techniques. The industrial processes in vogue/developed for recycling shown in Fig. 3. Large batteries are sheared into three pieces in a caustic LIBs are described in brief. bath, which neutralizes any acidic components and dissolves the lithium salts. The salts precipitated and dewatered in filter press, are used to pro-

3.3.1. Major industrial processes duce lithium carbonate. Besides organics, H2 formed by reaction with lithium,burnsoffatthesurfaceoftheprocessbath.SludgefromLIBsissentto 3.3.1.1. AEA technology. A patented process developed in the UK aims at recover cobalt. The large pieces are mechanically treated while the batte- the recovery of all battery materials (Lain, 2002). The casing of the bat- ries are submerged in a process solution of lithium brine, after which tery is first removed in N2 and cells are cut mechanically to remove ferrous and nonferrous metals are recovered. Li-ion fluff (mixture of other components wound into a spiral. The anode, cathode and separa- steel and plastics) can then be separated from the rest of the materials tors are shredded (~1 cm2 size) to leach the electrolyte and the solvent (Cu–Co and slurry). Copper is extracted in the next step and lithium in acetonitrile at 50 °C. The electrolyte and the solvent are recovered by brine is treated with Na2CO3 to obtain Li2CO3. Plastics and paper floating evaporation of the leach liquor and acetonitrile is recycled back. The at the top are recovered for disposal or recycling. The filtered carbon P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208 201

Spent lithium ion batteries Spent lithium ion batteries

Conveyor Mechanical pre-treatment Electronics, plastic casing

Disposal/ steel 250 oC Hammer mill Li-ion fluff recovery Electrolyte, solvent, Vacuum thermal treatment conductive salt Copper -cobalt Water spray Shaker table product Mechanical crushing Slurry

Primary metal Sieving, magnetic, Lithium brine Mixing & Holding production or other Al, steel, Cu, plastics tank applications air separation < 0.2 mm electrode material Cobalt filter Filter press cake Binder Agglomeration

Lithium brine Slag former Reducing meltdown Cobalt manganese alloy Na2CO3 Mixing tank

Residue Spent brine Product filter Acid Leaching to disposal Precipitant Washing Precipitation Filtration

Drying Lithium chloride

Li2CO3(s) Fig. 4. Flow-sheet of Accurec recycling process.

Fig. 3. Toxco's hydrometallurgical recycling process for LIBs. cake from the sludge is un-economical to reuse or even to burn off. Cobalt performed under an inert atmosphere to reduce the reactivity of lithi- in the spent LIBs presents higher economic worth. um. Plastics, steel and copper are then separated by physical treatment. The ﬁne powder obtained by screening is suspended in stirred water for 3.3.1.3. Accurec GmbH (Germany) process. The process involves mechan- subsequent leaching and hydrolysis step. Filtering the hydrolyzed solu- ical treatment to separate the electrode material which is treated by tion produces an alkaline solution of lithium salt and a suspension of pyro-metallurgy to recover the Co–Mn alloy and lithium chloride (Sojka, 1998). Electronics and plastic casings are removed followed by vacuum thermal treatment and pyrolysis to take care of the electrolyte Spent lithium ion batteries and solvent including conductive salts (Fig. 4). The batteries are then crushed and by employing sieving, magnetic as well as air separator, Mechanical treatment CO2 & inert gas mixture aluminium, copper and steel are removed. The remaining electrode material is agglomerated with the addition of a binder and pressed to briquettes, which are smelted (reduction) in a furnace to obtain two Physical treatment Steel, copper, plastics fractions, the metallic Co–Mn alloy and lithium containing slag. By acid leaching of the slag lithium can be extracted as lithium chloride/ pH > 12 carbonate. LiOH(aq) Leaching of fines Li(aq) Precipitation Soluble lithium 3.3.1.4. Sony/Sumitomo (Japan) process. In this process untreated LIBs are calcined at 1000 °C by which the cells open and inﬂammable compo- Mixed oxides + nents such as the plastic casing and organic solvent burn off. The residue Precipitation inserted Li + Carbon Li2CO3(s) consisting of metallic and metal pieces of iron, copper and aluminium can be magnetically separated. The remaining fraction is mainly carbon Lithium phosphate Leaching H2SO4 powder and active cathode material (LiCoO2 and/or LiCoxNi(1−x)O2). The powder is treated at the Sumitomo plant by hydrometallurgical Cu and other Soda methods to recover cobalt; lithium is not targeted. Cobalt oxide of impurities Purification high quality is recovered to use directly in the fabrication of new LIBs and the metallic scraps such as copper and stainless steel can be used as the by-products. Large batteries can also be handled but needs to NaClO Oxidation Li2SO4(aq) be punctured prior to introduction to the furnace (Gaines and Cuenca, 2000; Lupi et al., 2005). Cobalt(III) hydroxide 3.3.1.5. Recupyl process (France). This is also called as the Valibat process for recycling LIBs. Mechanical preparation of the spent batteries is Fig. 5. Recupyl process for LIBs recycling. 202 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208

Spent lithium batteries 3.3.1.7. OnTo Technology (USA). Spent batteries are discharged before dismantling. Batteries are cleaned and then placed in a high pressure fi Protective extraction container lled with liquid CO2 and some additives like alkyl CO2 atmosphere ethers, ammonia etc. The CO2 is transformed into a supercritical fluid by increasing the pressure and temperature in the container. With an increase in pressure, the battery casings are breached by fluid. When Feeder & the desired pressure and temperature are reached the electrolyte extrac- CO2 Shredder tion can either be carried out by circulating CO2 through the system or by soaking the batteries in the supercritical fluid. Batteries without function- Moist air Neutralization al potential are subsequently recycled through pulverization and the various materials separated and recovered based on the physical properties.

Release of protective Gases 3.3.1.8. Umicore process (Belgium and Sweden). Also called as the ValÉas atmosphere process, this is a combined pyro-and-hydro-metallurgical process to separate nickel and cobalt without recovering lithium. The spent batte- Acidified Gas Scrubber Leaching & washing ries are smelted in a furnace. Plastic, solvent and graphite are burnt off aqueous liquid while the metals are reduced and collected in a melt. The process is SO2(aq), Cl2(aq) operated in Sweden while the molten metal consisting of Ni, Co, Cu Leach liquor Metal fractions and Fe is cooled and processed in Belgium by sulfuric acid leaching & plastics and solvent extraction. NiSO4 and CoCl2 are separated from the solution Hydrometallurgical from which cobalt oxide and Ni(OH)2 are obtained. processing Residue A disadvantage of pyrometallurgical recycling processes including (SX/Precipitation) that of Umicore is that lithium recovery is not targeted and for which a combination of pyrometallurgical and hydrometallurgical processing Recovered chemical steps are necessary. The Umicore process has another drawback of substances losing base metals which are slagged off, besides the loss of organic materials as well as carbon. Fig. 6. Batrec Industrie AG process ﬂowsheet. 3.3.2. Recent development in recycling of lithium ion batteries metal oxides and carbon. Lithium is precipitated as Li2CO3 using the Spent LIBs containing lithium and other metals are mostly treated CO2 obtained from the mechanical treatment. The suspension of metal by the hydrometallurgical process which is used at times in combi- oxides is dissolved in H2SO4. Copper is cemented out by the steel nation with pyrometallurgical treatment. This may have integrated shots (Tedjar and Foudraz, 2007). The puriﬁed solution is oxidized by pre-treatment steps like pyrolysis or mechanical processing, i.e. crushing NaClO to precipitate cobalt(III) hydroxide (Fig. 5) and cobalt is separat- and material separation. In order to investigate the extraction of cobalt, ed by electrolysis. The remaining lithium in the solution is precipitated nickel, manganese etc. from LIBs by such processes thermodynamic as- with CO2 gas. pects particularly the stability regions of different phases in aqueous solution under redox conditions may be examined. For this standard Eh–pH

3.3.1.6. Batrec Industrie AG process (Switzerland). The company Batrec diagrams of Li–H2O, Co–H2O, Mn–H2OandNi–H2O systems can be re- mainly runs a mechanical processing plant for LIBs. In this process ferred from literature (Pourbaix, 1966; Schweitzer and Pesterfield, 2010). batteries are crushed in CO2 gas atmosphere and the released lithium Leaching of LIBs are carried out with different acids like HCl is neutralized. With the completion of the neutralization step, the (Contestabile et al., 2001; Zhang et al., 1998), H2SO4 (Aktas et al., 2006; protective environment is released and subsequently treated in a gas Dorella and Mansur, 2007; Kang et al., 2010a,b; Nan et al., 2005; Shin scrubber to reduce the emission during the process. Scrap material is et al., 2005; Swain et al., 2007), HNO3 (Lee and Rhee, 2002, 2003), and a leached and washed in acidified aqueous liquid and leach liquor is few organic acids like DL-malic acid (Li et al., 2010a), citric acid (Li et al., further processed for the recovery of different chemical substances. 2010b) etc. The binder (PVDF) which links the cathode material, LiCoO2 Metal containing a solid fraction can be separated from liquid and treat- with aluminium foils does not dissolve easily in the organic reagents ed to remove impurities. The flowsheet of the process is presented in such as fatty hydrocarbon or alcohol at room temperature making the Fig. 6. leaching reactions more difficult to proceed. The important developments

Table 9 Extraction and recovery of metals from spent LIBs in hydrochloric acid system.

Material Leaching conditions Extraction/Separation—parameters Highlights (with References merits/demerits) Acid concentration Temp. Time Pulp density (M) (°C) (h) (g/L)

Cylindrical- 4 80 1 100 SX with 0.29 M D2EHPA and 0.90 M PC-88A SX: N99.9 Co and 12.6% Li with Zhang et al. shaped LIBs 0.9 M PC-88A. (1998) Higher selectivity with PC 88A.

Spent LIBs 4 80 1 100 Precipitation: pH 6–8with4MNaOH. Co(OH)2 separated easily. Contestabile

Synthesis of precursor material by NixCoyMnz a precursor to cathode et al. (2001) co-precipitation. material

LIBs 4 80 1 20 Precipitation: pH 2 with KMnO4 for Mn(II), Purity (%): 96.9 Li, 98.23 Mn, Wang et al.

pH 9 for Co by DMG and Li2CO3 96.94 Co and 97.43 Ni (2009) LIBs 3 M acid + 3.5% 80 1 50 Precipitation: pH 11–12 with NaOH for Co; Recovery (%): 95 Co and 93 Li in Shuva and

H2O2 Na2CO3 at 100 °C for Li precipitate Kurny (2013) Ash from LIBs 4 90 18 50 Precipitation by NaClO at pH 3 and base Recovery (%): 100 Co and 99.99 Ni. Joulié et al. addition at pH 11 HCl gives higher leaching efﬁciency. (2014) P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208 203

Table 10 Extraction and recovery of metals from spent LIBs in sulfuric acid system.

Material Leaching conditions Extraction/Separation— Highlights (with References parameters merits/demerits) Acid concentration Temp. Time Pulp density (°C) (h) (g/L)

LIBs waste 2 M acid + 15 vol.% 75 0.167 50 – Leaching (%): 95 Co Shin et al.

H2O2 and 100 Li. (2005)

Harmful gas—SOx generated. Square-shaped LIBs 3 M acid 70 4 200 Precipitation & SX using Acorga M5640 Precipitation: 90% Co Nan et al. (2005) and Cyanex 272 SX: ~97% Cu by Acorga M5640, 97% Co by Cyanex 272

LIBs 6% acid + 1% H2O2 65 33.3 Precipitation with NH3 at pH 5. Leaching (%): ~55 Al, 80 Dorella and Co/Li separation by diluted (0.72 M) Co and 95 Li. Mansur (2007) Cyanex 272 Extraction (%):~85 Co by Cyanex 272.

LIB wastes 2 M acid + 5% H2O2 75 0.5 100 SX by 0.5 M Cyanex 272, pH 5.35 Leaching (%): 93 Co Swain et al. and O/A = 1 in 1-stage and 94 Li. (2007)

CoSO4 solution: 99.99% pure.

Spent LIBs 4 M acid + H2O2 80 4 Precipitation using ethanol, ethanol/ Co recovery: 92% as Aktas et al.

solution = 3:1; 15 min, room CoSO4 and (2006)

temperature 8% Co as Co(OH)2 by

adding Li(OH)2 at pH 10. Synthetic solution (10 mol/L ––––SX by Cyanex 272 and DP-8R Separation factor (Co/Li): Swain et al.

Co(II) and 20 mol/L Li2SO4) 497 at pH 5 (2010)

Spent LIBs 2 M acid + 6% H2O2 60 1 100 Precipitation at pH 6.5 for Cu, Fe and Al Leaching (%): N99 Co, Kang et al. SX by 50% saponiﬁed 0.4 M Cyanex 272, SX: 99.9% Co in 2 stages. (2010a, 2010b) pH 6, O/A = 2 Separation factor 750 (Co/Li) and Co/Ni at pH 6. Spent LIBs 1 M acid + 30% 80 2 714 Precipitation by 1 M citric acid, 2 h, Max. Co leaching (%): 88.3. Li and Zeng

H2O2 65 °C; calcined at 450 °C, 4 h Crystalline LiCoO2 (2011) synthesized Spent LIBs 4 M acid + 10% 85 2 100 SX by 25% P507 SX: 98% Co and removal Chen et al.

H2O2 Precipitation by ammonium oxalate, of 97% Ni and Li. (2011a, 2011b) pH 1.5 Spent LIBs 3 M + 0.25 M 90 3 67 – N99% leaching Wang et al.

Na2S2O3 (2012)

on the leaching and separation of metals from the leach solutions are recovery of metals from the leach liquors, was carried out either by sol- summarized in Tables 9–11. vent extraction involving PC-88A, Cyanex 272 etc. to produce the pure

metal salts or the metals that were selectively precipitated as Co(OH)2 3.3.2.1. Metal extraction/recovery from hydrochloric acid leach liquors. In and LiCO3 from the leach solutions. Thus, Zhang et al. (1998) used the most processes reported, optimum extraction of metals from the cath- steps such as leaching of cathode materials of spent LIBs in HCl, separa- ode material of spent LIBs was achieved with a 4 M solution of HCl tion of Co/Li by solvent extraction (PC-88A) and recovery of cobalt as (Table 8). In the presence of a reducing agent such as hydrogen perox- sulfate and lithium as carbonate. Out of sulfurous acid, hydroxyl amine ide, metal leaching was possible even in 3 M HCl (Contestabile et al., hydrochloride and hydrochloric acid, HCl was found to be the most suit- 2001; Joulié et al., 2014; Shuva and Kurny, 2013; Zhang et al, 1998). able lixiviant for economic reasons. The leaching efﬁciency of N99% of

While H2O2 facilitated the dissolution of Co(II) which was reduced Co and Li in a 4 M HCl solution was achieved. The complete extraction from Co(III), the dissolution of lithium was also promoted because of of cobalt from the leach liquor by 0.90 M PC-88A in 1 stage at pH 6.7 the presence of the two metals in the same oxide. The separation and and O/A of 0.85:1, scrubbing of lithium by 30 g/L Co at O/A 10:1 and

Table 11 Extraction and recovery of metals from spent LIBs in other organic/inorganic acid systems.

Material Leaching Leaching conditions Highlights References reagent (with merits/demerits) Time (h) Temp. (°C) Pulp density (g/L)

LIBs 1 M HNO3 + 1.7 vol.% H2O2 1 75 20 Leaching (%): 95 Co and Li. Lee and Rhee (2002, 2003)

Harmful gas—NOx generated.

LIBs 1 M HNO3 + 1.0 vol.% H2O2 1 80 20 Leaching (%): 100 Co and Li. Li et al. (2011)

Harmful gas—NOx generated.

Spent LIBs 1.5 M DL-malic acid + 2% H2O2 0.67 90 20 Leaching (%): 90 Co and 100 Li. Li et al. (2010a) Acid can be recycled and reused.

Spent LIBs 1.25 M citric + 1% H2O2 0.25 90 20 Leaching (%): N90 Co and 100 Li. Li et al. (2010b) Simple and environmental friendly process. Cylindrical spent LIBs 1.25 mol/L ascorbic acid 0.33 70 25 Leaching (%): 94.8 Co and 98.5 Li. Li et al. (2012)

of mobile phone Reducibility of ascorbic acid displaced H2O2. 204 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208

stripping of cobalt from the loaded organic with 2 M H2SO4 at O/A 5:1, Shin et al. (2005) reported almost quantitative leaching of cobalt and lith- produced high purity cobalt sulfate. About 80% lithium was recovered ium with 2 M H2SO4 in the presence of a high amount of H2O2 (15 vol.%) as Li2CO3 by precipitation with saturated soda solution at 100 °C. at 75 °C and 50 g/L pulp density (Table 10). Reductive leaching of the The leaching reaction of the waste LiCoO2 withHClsolutioncanbe mechanically treated LIBs in 6 (v/v)% H2SO4 and 1% H2O2 solution resulted represented as: in a relatively lower recovery of metals (~55% Al, 80% Co and 95% Li) (Dorella and Mansur, 2007). Kang et al. (2010a, 2010b) also reported

þ → þ þ þ : ð Þ the reductive leaching of LIBs with H2SO4. The reactions of LiCoO2 with 4LiCoO2ðsÞ 12HClðaqÞ 4LiClðaqÞ 4CoCl2ðaqÞ 6H2O O2ðgÞ 15 H2SO4 and in the presence of H2O2 are shown below:

A laboratory scale process involving leaching of LiCoO in HCl and pre- 2 4LiCoO þ 6H SO →4CoSO þ 2Li SO þ 6H O þ O ð18Þ cipitation of cobalt hydroxide was developed by Contestabile et al. (2001). 2ðsÞ 2 4 4 2 4 2 2ðgÞ The active material of spent LIBs was dissolved in N-methylpyrrolidone at þ þ → þ þ 100 °C to separate aluminium and copper foil. From the leach liquor, car- 2LiCoO2ðsÞ 3H2SO4 H2O2ðaqÞ 4CoSO4 Li2SO4 4H2O fi þ : ð Þ bon powder was removed by ltration and Co(OH)2 was precipitated out O2ðgÞ 19 from the solution at pH 6–8. In another study over 95% dissolution of Ni, Co and Mn was reported when the battery material of 120 μmsizewas The leaching efficiency of cobalt depends on the reductant concentration. leached at a higher acid concentration (6 M HCl) and lower temperature Among other reducing agents, Na2S2O3 helped in the leaching of N99% (60°C)inthepresenceofH2O2 (Li et al., 2009b). Cobalt from the solution of the metals (Co and Li) in 2 M sulfuric acid at 90 °C in 3 h (Wang was cemented out by iron powder and iron was removed as goethite. et al., 2012). Finally chlorides of Ni, Mn and Co were added in the purified solution Safe dismantling procedures of spent lithium ion batteries have to prepare the precursor of the cathode material (NixCoyMnz) directly often been described in the literature (Nan et al., 2005; Tanii et al., through co-precipitation with ammonium bicarbonate. The leaching of 2003). Zhu et al. (2011) applied a mechanical separation process to metals from the cathode material in 4 M HCl at 80 °C with a metal re- recover copper from these batteries. The anodes from the batteries covery of 99% was also reported by Wang et al. (2009).Manganesein were separated by mechanical treatment, pulverization and sieving. 2+ the leach liquor was precipitated at a molar ratio of Mn to KMnO4 Almost 92% of copper in anode particles was recovered by a gas- of 2 and pH 2 followed by selective precipitation of nickel with fluidized bed separator. dimethylglyoxime (DMG) at pH 9. Cobalt was recovered as hydrox- In the alkali–acid leaching process, the cathode was first treated with ide at pH 11 and Li2CO3 was then precipitated; purity of the metals 10% (w/w) NaOH at 30 °C to dissolve Al, followed by reductive leaching being 96.97% Li, 98.23% Mn, 96.94% Co and 97.43% Ni. A recent study of ~97% Co and 100% Li with H2SO4 and H2O2 (Ferreira et al., 2009; Nan by Shuva and Kurny (2013) demonstrated the reductive dissolution of et al., 2005). Acorga M5640 and Cyanex 272 were used to selectively cathode powder in 3 M HCl in the presence of 3.5% H2O2. Over 95% extract and recover 98% Cu and 97% Co, respectively from the solutions. – cobalt was precipitated as hydroxide at pH 11 12, leaving lithium About 90% Co was recovered as oxalate with b0.5% impurities. LiCoO2 (93%) in the leach solution. positive electrode material with a good electrochemical performance Processing of battery ash obtained from the pyrolysis of spent LIBs by was synthesized by using the recovered compounds. HCl leaching was also attempted. Lin et al. (2003) patented a pyrometal- From the sulfate leach liquor of spent LIBs, 96% copper was recov- lurgical process of waste LIBs combined with hydrometallurgical pro- ered as CuSO4·3H2O with ethanol at a volume ratio of 3:1. Cobalt was cessing. The ash with metal and metal oxides was dissolved in 3–6M recovered in two steps. During the first step, 92% of the cobalt was

HCl containing NaCl. Copper and cobalt were separated out using mem- recovered as CoSO4 by the use of ethanol at a volume ratio of 3:1. The – brane electrolysis. Then Fe(OH)3 and Al(OH)3 were recovered at pH 5 7 remaining Co in the second step was recovered as Co(OH)2 by the addi- followed by the precipitation of lithium carbonate. Recently Joulié et al. tion of Li(OH)2 at pH 10 (Aktas et al., 2006). Lithium, which remained in ﬁ (2014) reported the high leaching ef ciency (~100%) of Li, Ni, Co and the solution, was then recovered to the extent of 90% as Li2SO4 by the Al from the Li–Ni–Co–Al oxide ash of spent LIBs by 4 M HCl at 90 °C addition of ethanol (3:1 volume ratio). Aluminium was recovered as with chloride ions promoting the dissolution. Co(II) in the leach liquor Al(OH)3 with 99% recovery efﬁciency. It was shown that metals could was oxidized to Co(III) with NaClO and recovered as Co2O3·3H2Oby be precipitated/separately by the ethanol/sulfate precipitation tech- selective precipitation at pH 3 (Eqs. (16) and (17)). Nickel hydroxide nique depending on their concentrations present in the solution. was then precipitated at pH 11. The waste cathodic active material generated during the manufactur-

ing of LIBs was also leached in H2SO4 in the presence of H2O2 (Swain et al., 2007). During the separation of Co/Li in a 2-step SX process with 2þ þ − þ þ ⇔ 3þ þ − þ ð Þ 2Co ClO 2H3O 2Co Cl 3H2O 16 1.5 M Cyanex 272 at O/A 1.6, about 85% Co was recovered. The remaining cobalt was extracted in 0.5 M Cyanex 272 at O/A 1 and pH 5.35. The 3þ þ −→ ⋅ : ð Þ 2Co 6OH Co2O3 3H2O 17 purity of cobalt sulfate in solution was found to be 99.99%. Earlier, Swain et al. (2006) reported the highest separation factor (Co/Li) of 62 during extraction with saponified Cyanex 272 from a synthetic solution 3.3.2.2. Metal extraction/recovery from sulfuric acid leach liquors. In most at pH 6.9. leaching processes with sulfuric acid, hydrogen peroxide was used as The mechanism by which a metal ion is extracted from an aqueous a reductant (Castillo et al., 2002; Dorella and Mansur, 2007; Jha et al., solution using a partially saponified cation exchange extractant is as 2013a,b; Kang et al., 2010a, 2010b; Shin et al, 2005). In some cases alkali follows (Ritcey and Ashbrook, 1984): leaching followed by acid leaching was considered to remove aluminium. Metals from the leach solutions were separated and recovered by 2þ þ − þ ð Þ ↔ ⋅ þ þ : ð Þ Maq Aorg 2 HA 2org MA2 3HAorg Haq 20 solvent extraction using PC-88A/P507, Cyanex 272 and Acorga M5640 and precipitation processes very similar to that of the HCl system. The Cobalt is extracted as [CoA2·3HA]org with 65% Na-Cyanex 272. Cobalt can fine sized electrodes were initially contacted with N-methyl pyrrol- be completely stripped from the loaded organic with 0.01 M H2SO4 to idone (NMP) to dissolve the binder and separate active material produce cobalt sulfate of N99% purity.

(LiCoO2) from Al and Cu foils (Castillo et al., 2002). The LiCoO2 powder By acid leaching of waste cathodic material and SX with Cyanex 272, was leached in 4 M H2SO4 at 80 °C to dissolve Co and Li, and Co(OH)2 Swain et al. (2008) produced a pure cobalt sulfate solution (99.99%). In was recovered from the leachate by the addition of sodium hydroxide. another study quantitative separation of Co(II)/Li was reported using P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208 205 the supported liquid membrane (SLM) with a mixed extractant contain- (2011). The cathode material was treated in a vacuum furnace at ing Cyanex 272 and DP-8R as the mobile carrier (Swain et al., 2010). 600 °C for 30 min with the residual gas pressure of 1.0 kPa. Over 99% Very recently, leaching of lithium and cobalt from cathodic material of Co and Li were recovered from peeled Co–Li oxide by leaching with waste mobile phone batteries in sulfuric acid in the presence of H2O2 2MH2SO4 at 80 °C. Paulino et al. (2008) examined two approaches for and separation of metals by Cyanex 272 were reported (Jha et al., recycling spent Li/MnO2 and LIBs. In the ﬁrst process ~90% Li was recov- 2013a, 2013b). Chen et al. (2011a) recovered cobalt oxalate from the ered by calcination at 500 °C followed by leaching with H2SO4 and H2O2 spent LIBs by using an alkali–acid leach process. After roasting the spent at 90–100 °C, and solvent extraction. The second process involved fusion

LIBs at 700–800 °C to burn off carbon and the binder, and leaching with with KHSO4 at 500 °C and water leaching with H2O2 at 90 °C. Cobalt and NaOH to remove Al, and leaching with H2SO4 in the presence of 10 (v)% manganese were precipitated at pH N9 followed by precipitation of LiF H2O2 could recover 95% Co and 96% Li. Iron removal (99.99%) as jarosite by KF solution. with a loss of b1% Co was achieved in the pH range 3–3.5 at 95 °C. The Extraction of Co and Li from a material of a large scale mechanical precipitation reaction is shown below: pre-treatment and recycling plant in Northern Italy is described by Granata et al. (2012). The powder is leached in H SO in the presence ð Þ þ þ → ð Þ ð Þ þ : ð Þ 2 4 Fe2 SO4 3 12H2O Na2SO4 Na2Fe6 SO4 4 OH 12 6H2SO4 21 of 50% excess of a reducing agent, glucose — a waste product of the Complete removal of manganese with ammonium persulfate at pH 4 and food factory. Iron, aluminium and copper are partially precipitated 70 °C is as follows: as hydroxides at pH 5.0. Using solvent extraction high purity cobalt carbonate (47% Co) is obtained by precipitation, whereas without sol- 2þ þð Þ þ → þð Þ þ vent extraction the product containing 36–37% (w/w) Co is obtained. Mn NH4 2S2O8 2H2O MnO2 NH4 2SO4 H2SO4 þ þ 2H : ð22Þ Lithium is recovered by crystallization (yield 80%) with 98% purity. The process with solvent extraction shows economical outputs (gross Copper precipitation (N98.5%) with NaOH was followed by solvent margin and payback time) than the one without solvent extraction. extraction to recover Co(II) while removing 97% Ni and Li (Zhu et al., Suzuki et al. (2012) developed a process as detailed in Fig. 7. Acorga 2012). About 95% Co(II) was extracted selectively from the purified M5640 extracted copper within a pH range of 1.5–2.0 leaving Al, Co and solution with saponified 25 wt.% P507 (2-ethylhexyl phosphonic acid Li in the raffinate. Aluminium is then selectively extracted by PC-88A in mono-2-ethylhexyl ester) at pH 3.5 and stripped with 3 M H2SO4. Cobalt the pH range 2.5–3.0. Cobalt(II) and lithium(I) are separated by PC-88A/ oxalate was produced from the strip liquor with a purity of 99.9%. TOA rather than Acorga M5640 due to its higher stripping efficiency CoC2O4·2H2OandLi2CO3 were also produced by a combination of acid (N98%), although Acorga M5640 provides higher cobalt selectivity. leaching, SX and precipitation from spent mobile phone batteries (Jian Synergistic extraction and separation of Co(II) and Mn(II) with et al., 2012). From the leach solution of the used batteries, Co(OH)2 Li(I) from simulated sulfuric acid leaches of waste cathodic materials was precipitated and was converted to Co3O4 by heating (Yamaji et al., using a mixture of Cyanex 272 and PC-88A in N-heptane have been 2011)asperreactions(23) and (24). Lithium was recovered as carbonate investigated by Zhao et al. (2011). A mixed extractant system was also by adsorption with 2% MnO2 (11.7 mg Li/g). utilized to treat the leach solutions of spent LIBs (Pranolo et al., 2010). In the first stage Fe(III), Al(III) and Cu(II) were extracted using Ionquest þ → ð Þ þ ð Þ 801 and Acorga M5640, and the raffinate containing cobalt, nickel and CoSO4 2NaOH Co OH 2 Na2SO4 23 lithium was treated with 15% (v/v) Cyanex 272 to separate out Co. 1 More than 90% of Co could be separated out at pH 5.5–6.0 and A/O 3CoðÞ OH þ O →Co O þ 3H O: ð24Þ 2 2 2 3 4 2 1:2. An ion-exchange resin such as Dowex M4195 was then used to separate Ni(II)/Li(I). The addition of 2% (v/v) Acorga M5640 to 7% (v/v) The advantage of vacuum pyrolysis to prevent the escape of toxic Ionquest 801 generated a significant pH isotherm shift for Cu which gases and lower the decomposition temperature of organics while resulted in a ΔpH50 of 3.45. protecting oxidization of metals in LIBs, was applied by Sun and Qiu Apart from the hydrometallurgical extraction of valuable metals from spent LIBs, synthesis of LiCoO2 was also reported (Li and Zeng, Sulfuric acid Leaching of spent LIBs 2011). The solvent N-methyl-2-pyrrolidone was used to dissolve the PVDF agglomerate from Co–Li membrane to separate the aluminium foil. About 88% Co was recovered during the leaching with sulfuric Al(III), Cu(II), Co(III), Li(I) acid at pH 0.5 and 80 °C. Then, 1 M citric acid solution was added at

pH 1.5-2.0 65 °C to prepare a gelatinous precursor. LiCoO2 was obtained with the Acorga M5640 Solvent Extraction gel precursor calcined in a crucible at 450 °C for 4 h.

3.3.2.3. Metal extraction/recovery from nitric acid leach liquors. Mechanical, thermal, hydrometallurgical and sol–gel steps were applied to recover Cu(II) Al(III), Co(II), Li(I) Co/Li from spent LIBs and synthesize LiCoO2 as a cathode active material pH 2.5-3.0 (Lee and Rhee, 2002, 2003). Reductive leaching in 1 M HNO3 with 1.7 vol.% H O at 75 °C extracted ~95% Li and 95% Co (Table 11). The PC-88A Solvent Extraction 2 2 molar ratio of lithium to cobalt in the leach liquor was adjusted to 1.1

by adding a fresh LiNO3 solution. Then, 1 M citric acid solution was added to prepare a gelatinous precursor which was calcined at 950 °C Al(III) Co(II), Li(I) for 24 h to produce purely crystalline LiCoO2. Like in other studies, the pH 5.5-6.0 addition of hydrogen peroxide increased the leaching efﬁciency of PC-88A + TOA Solvent Extraction metals (Lee and Rhee, 2003). The reductive leaching of almost 100% Co and Li from the spent batteries in nitric acid and H2O2 was achieved at 80 °C by Li et al. (2011). In the nitric acid leaching system pollution

due to harmful NOx gases is a major problem and needs to be addressed. Co(II) Li(I) As regards the leaching kinetics, lithium dissolution followed a shrinking

core model. The mechanism of the dissolution of LiCoO2 was controlled Fig. 7. SX separation of Al, Cu, Co and Li from sulfate media (Suzuki et al., 2012). by a surface chemical reaction with an apparent activation energy of 206 P. Meshram et al. / Hydrometallurgy 150 (2014) 192–208

52.32 and 47.72 kJ/mol for Co and Li, respectively (Lee and Rhee, 2003). LiCoO2 was separated from aluminium foil with dimethyl acetamide (DMAC). Polyvinylidene ﬂuoride and carbon powders in the active The leaching reaction with HNO3 solution can be represented as: material were then eliminated by roasting at 450 °C for 2 h and 600 °C þ → þ ðÞþ for 5 h, respectively. Finally LiCoO2 was obtained by adding a certain 2LiCoO2sðÞ 3HNO3aqðÞ Li2NO3aqðÞ 2Co NO3 2aqðÞ 3H2O 1 amount of Li2CO3 in the recycled LiCoO2 and calcining it at 850 °C for þ O ðÞ: ð25Þ 12 h. 2 2g

Recycling of Li-ion and polymer batteries while producing 4. Conclusions

LiCoxNi(1− x)O2 as a cathode material was investigated by Lupi and Pasquali (2003). The process consists of cathodic paste leaching, Co–Ni Lithium is one of the rare metals with a variety of applications and separation by SX and recovery of Ni by electrowinning. The separation demand for lithium is expected to increase with the ever increasing of Ni/Co was performed by solvent extraction using saponiﬁed 0.5 M use of electrical and electronic devices/hybrid electric vehicles. A few Cyanex 272. Nickel was electrowon at a current density of 250 A/m2, established technologies are in vogue to produce lithium in the desired

50 °C and pH 3–3.2 with an electrolyte of 50 g/L Ni and 20 g/L H3BO3. form from its primary resources like its minerals and brine, whereas Elaborating further, Lupi et al. (2005) reported the current efficiency limited exploitation of lithium resources from seawater and bitterns and energy consumption of 87% and 2.96 kWh/kg for nickel under the calls for their intensified tapping. Extraction of lithium from its minerals above conditions as compared to the figures of 96% and 2.8 kWh/kg, and clays is fraught with high mining costs and involves high energy, respectively for cobalt at the same current density and temperature, while extraction from brine and bitterns/seawater needs a long time fi but at pH 4–4.2 from a solution containing manganese and (NH4)2SO4. for evaporation. Hence these processes need to be adequately modi ed to yield efficiency and better economic returns. 3.3.2.4. Metal extraction using organic acids/reagents. For the sustainable Extraction processes from secondary resources like batteries depend management of the secondary resource such as LIBs, organic acids on the chemistry of the battery material. Most processes involve dis- such as DL-malic acid, citric acid etc. which have mild acidity, are sug- mantling of LIBs, separation of cathode and anode materials, leaching gested for the leaching of metals (Table 11). Li et al. (2010a) reported of valuable metals like Co, Li, Ni, Mn etc. from the cathode material in that DL-malic acid can dissolve lithium and cobalt of LIBs fairly rapidly different mineral acids, and separation and recovery of metals from under aerobic and anaerobic conditions as compared to the mineral the solutions by solvent extraction/IX/precipitation. At present no lithi- fi acids like HCl, HNO3 and H2SO4, and the waste solutions can be treated um extraction is industrially practiced from LIBs. Therefore, a suf cient easily. Almost 100% Li and N90% Co were leached out with 1.5 M malic scope exists not only to reduce the process steps followed currently but fi acid and 2.0% H2O2 at 90 °C in 40 min. Leaching of lithium and cobalt also to improve the ef ciency of metal extraction and separation, in- in citric acid was also reported after separating the anode and cathode cluding lithium recovery. Points that may be considered include process material by the treatment of NMP (Li et al., 2010b). Nearly 100% Li intensification to save energy and improve the kinetics of leaching, and N90% Co were extracted in 1.25 M citric acid and 1.0% H2O2 at 90 °C. while addressing the problem of selectivity and examining the use of Recently ultrasonic assisted leaching of cobalt and lithium from unexplored/synergistic solvents with an ultra high metal loading capac- spent LIBs in the presence of ascorbic acid was investigated (Li et al., ity to cut down the process steps. There is a need to develop appropriate 2012). Leaching efficiencies of as high as 94.8% for Co and 98.5% for Li technology which can address the limitation of current processes for were achieved with 1.25 M ascorbic acid solution at 70 °C (Table 11). extraction of all valuable metals from its primary as well as secondary Sun and Qiu (2012) used oxalic acid as both leachant and precipitant resources. to separate and recover cobalt and lithium from the spent LIBs. Cathode material consisting of LiCoO2 and CoO from the dismantled batteries, Acknowledgements was peeled off from the aluminium foils after vacuum pyrolysis at 600 °C. Leaching was performed using 1 M oxalate at 80 °C with the The authors are thankful to the Director, CSIR-National Metallurgical fi N reaction ef ciency of 98% of LiCoO2 while separating cobalt and lithium. Laboratory (NML), Jamshedpur, India for giving permission to publish The reaction with oxalate for leaching and precipitation proceeds as: the paper.

þ → þ ð Þ þ þ ð Þ 3H2C2O4 LiCoO2ðsÞ LiHC2O4 Co HC2O4 2 2H2O 2CO2ðgÞ 26 References

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