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Lithium leaching via calcium chloride roasting from simulated pyrometallurgical slag of spent lithium ion battery
Journal Pre-proofs Lithium leaching via calcium chloride roasting from simulated pyrometallurgical slag of spent lithium ion battery Hui Dang, Na Li, Zhidong Chang, Benfeng Wang, Yifei Zhan, Xue Wu, Wenbo liu, Shujaat Ali, Hongda Li, Jiahui Guo, Wenjun Li, Hualei Zhou, Changyan Sun PII: DOI: Reference:
S1383-5866(19)31783-6 https://doi.org/10.1016/j.seppur.2019.116025 SEPPUR 116025
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Separation and Purification Technology
Received Date: Revised Date: Accepted Date:
4 May 2019 28 July 2019 2 September 2019
Please cite this article as: H. Dang, N. Li, Z. Chang, B. Wang, Y. Zhan, X. Wu, W. liu, S. Ali, H. Li, J. Guo, W. Li, H. Zhou, C. Sun, Lithium leaching via calcium chloride roasting from simulated pyrometallurgical slag of spent lithium ion battery, Separation and Purification Technology (2019), doi: https://doi.org/10.1016/j.seppur. 2019.116025
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Lithium leaching via calcium chloride roasting from simulated pyrometallurgical slag of spent lithium ion battery Hui Danga, Na Lia, Zhidong Changa,, Benfeng Wanga, Yifei Zhana, Xue Wua, Wenbo liua, Shujaat Alib, Hongda Lia, Jiahui Guoa, Wenjun Lia, Hualei Zhoua, Changyan Suna a
Department of Chemistry and Chemical Engineering, School of Chemistry and Biological Engineering,
University of Science and Technology Beijing, Xueyuan Road, Beijing 100083, PR China b
Women University Swabi, KPK, 23430, Pakistan
Abstract In the present case, insoluble lithium in the simulated slag which is obtained by pyrometallurgical processing of spent lithium ion batteries (LiBs) is leached aqueously through transformation of insoluble lithium into soluble lithium via roasting with calcium chloride (CaCl2). The change of both Gibbs free energy and the logarithm of the equilibrium constant, which are taken as functions of temperature for the chlorination reaction of simulated slag with CaCl2, are predicted through the simulation with an enthalpy, entropy and heat capacity (HSC) programme. The simulation indicates that CaCl2 is a favourable and effective chlorine donor in this chlorination reaction to yield lithium chloride (LiCl) when the temperature is not less than 500°C. These predicted results are tested and confirmed experimentally. A maximum of 90.58% lithium recovery can be yielded with the optimal roasting conditions of a temperature of 800°C for 60 min and a molar ratio of Cl/Li of 1.8:1, along with subsequent leaching conditions of 60°C for 30 min with a water/calcines mass ratio of 30:1. X-ray diffraction (XRD) results suggest that most of LiAl(SiO3)2 in the slag disappears under the best reaction conditions and is transformed into LiCl.
Corresopnding author. E-mail address: [email protected]
Keywords: spent lithium ion battery, lithium recovery, pyro-metallurgical slag, aqueous leaching, chlorination roasting; 1. Introduction Lithium ion batteries (LiBs) are considered one of the most promising energy storages due to advantages such as their long cycle life, high capacity, light weight and environmental friendliness [1-5]. These advantages have been used in an increasingly wide range of applications from portable gadgets to heavy-loading vehicles [6,7]. The output of LiBs reaches globally 100.75 GWH year on year with an increase of 39.45% [8], which has resulted in the increasing consumption of lithium [7,9-11]. Generally speaking, lithium is a kind of non-renewable resource and should be used smartly [12,13]. In addition, with the rapidly growing expenditure on these electronic products, a huge quantity of spent LiBs has been generated [14]. Therefore, it is meaningful from the viewpoints of sustainable development and resource conservation to recycle lithium from spent LiBs [15]. Many hydrometallurgical or pyrometallurgical processes to treat spent LiBs have been studied. Nonetheless, nearly all metals in the spent batteries, such as lithium, nickel, cobalt, aluminium and iron and so on, can be recovered effectively using hydrometallurgical treatment [16-26]. However, the process which involves dismantlement or disassembly of cathode scarp from batteries consumes many labour and mechanical powers, which is also applicable to some pyro-metallurgical methods [27-30]. Industrialized processes, such as the Umicore VAL’EASTM smelting technique, which is much simpler and does not even require pretreatment [31, 32], normally do not recycle lithium, which is ultimately transferred and fixed in the smelting slag and used for construction material without further processing [7]. However, the content of lithium in the solid slag is up to 1.4% [10], and it is profitable to isolate this lithium. In our previous work, lithium was evaporated from slag by chlorination roasting via gas-solid separation [33]. Unfortunately, the roasting conditions are comparatively harsh with temperature of higher than 1000°C,
which consumes a large quantity of precious energy, especially exergy [34]. Therefore, it is an urgent challenge to devise a process of extracting lithium energy-efficiently. Yan et.al [35] reported that lithium could be extracted by water from lepidolite after being roasted with either calcium chloride or sodium chloride under 850℃. Subsequently, Lucía [36, 37] proposed that insoluble lithium in β-spodumene could be transformed into soluble lithium through chlorination roasting with a chlorinating agent such as calcium chloride or chlorine gas and then leached aqueously. Because the chlorination roasting and water leaching conditions are comparatively mild and the chemical constitution of spodumene is greatly similar to that of the pyrometallugical slag, lithium could be extracted from simulated slag using the same method. In the present work, the possibility of recovering lithium from simulated pyrometallurgical slag of spent lithium ion battery through chlorination roasting and water leaching was explored. The effect of water leaching conditions and chlorination roasting conditions, such as water/calcines mass ratio, leaching temperature and time, roasting temperature and time, and molar mass of the chlorine donor on the extraction efficiency of lithium, were studied and optimized. 2. Materials and methods 2.1 Materials The main agentia was Li2O, which was purchased from Aladdin Biochemical Technology Co., Ltd., Shanghai, China. CaCl2, CaO, Al2O3 and SiO2 were all bought from Sinopharm Chemical Reagent Co., Ltd., China. Deionized water as produced by Laboratory Center of University of Science and Technology Beijing, China. 2.2 Synthesis of simulated slag Generally, pyrometallurgical recycling of spent LiBs can be divided into three stages: pre-heating, plastic pyrolizing, and smelting along with reducing [38]. In the last stage, metals such as cobalt, nickel,
copper and iron, which can be reduced by carbon, are transformed into liquid metal by reduction roasting, and recovered as an alloy. Lithium and aluminum are removed as impurities through being transferred into refractory slag with the addition of slag-forming agents such as silica and calcium oxide. The chemical composition and structure of slag are similar to those of cement and are generally expressed as xLi2O·yCaO·zAl2O3·nSiO2 [39,40]. In this paper, the possibility of leaching aqueously of lithium from the slag mentioned above is going to be probed through chlorinating the insoluble lithium bound in slag by calcium chloride. In order simplify the research,a simulated slag is adopted in line with our previous work [33] with the following conditions: silica of 20wt% ~ 60wt%, calcium oxide of 20wt% ~ 35wt%, aluminium oxide of 10wt% ~ 30wt% and lithium oxide of 0.5wt% ~ 15wt% [39,40]. The corresponding amount of lithium oxide, silica, aluminium oxide and calcium oxide were mixed, ground, and finally roasted in a muffle furnace (heating rate of 10℃/min) at 1200℃ for 1 hour. Scanning electron microscopy (SEM) was utilized to investigate morphological characterization of the slag after grinding. 2.3 Thermodynamic analysis of reaction process In this study, the thermodynamic calculations were tested within the temperature ranging from 25 to 1000℃ for the chlorination reaction of simulated slag with CaCl2. Equilibrium composition of the system Si–Al–Li–O–Ca–Cl of the reaction, the changes of standard Gibbs free energy (ΔrGm) and the logarithm of the equilibrium constant (LgK) between the simulated slag and CaCl2 were calculated in light of the effect of reaction temperature using the programme HSC chemistry 6.0 [41, 42] to assess the possibility of the reaction and predict the content of products as they change with temperature. HSC chemistry 6.0 software is a thermochemical software used to perform enthalpy (H), entropy (S) and capacity (C) calculations (HSC calculations). Most general thermodynamic parameters can be predicted readily and
cost-effectively using the HSC program, which was developed by Outokumpu Oy Information Center [43-45]. These calculations were accomplished on the basis of system’s free energy minimization method. 2.4 Chlorination roasting and water leaching procedure Lithium was recovered by chlorination roasting and water leaching from simulated slag. The latter was primarily blended with CaCl2 at different molar ratios of chlorine to lithium (1:1-1.8:1). The mixtures were then poured into a muffle furnace for roasting at 800°C for 60 min. Then the mixtures with a suitable molar ratio of Cl/Li were roasted at 500-900℃ for 20-180 min. The resulting calcines were then disposed by removing, cooling down to room temperature and grinding thoroughly. Deionized water was then used to leach ground calcines at 40-80°C for 20-60 min using the mass ratios of water to calcines at 10:1-50:1 to free lithium. Solid and liquid were divided by filtration, and the residues were washed several times thoroughly with deionized water. XRD analysis of the roasted product was performed to explore the proper molar ratio of Cl/Li when the simulated slag reacts with CaCl2 completely. The desiccative leached solution was also analyzed by XRD. The chemical composition of leaching solutions was analyzed with ion chromatography (ICS5000). An atomic absorption spectrophotometer (AAS 6800, Shimadzu, Japan) was employed to measure the concentrations of lithium in the leached solution with an 8 mA current at wavelength 670.8 nm. The extraction efficiency of lithium was calculated according to Eq. (1). cV
Extraction efficiency of lithium = 𝑚
0
(1)
Where V is the volume of the solution, c is the concentration of lithium in the solution measured by AAS and m0 is the mass of the lithium in the sample before roasting. Three parallel experiments were repeated during roasting and water leaching to ensure the repeatability of the results. The maximum relative deviation of the replicate experimental data was ±3%. 3. Results and Discussion
3.1. Characterization of simulated slag
Fig.1. Morphology of the simulated slag particles: (a) overall view; (b) superficial appearance of a particle
The main components in the simulated slag were LiAl(SiO3)2 (JCPDS file no.31-0706) and CaSiO3 (JCPDS file no.31-0300) according to our previous study, which indicated that a new structure could be generated by the sintering of SiO2, CaO, Al2O3 and Li2O. A morphological and structural description of simulated slag particles are given in Fig.1. It was observed that the particle size distribution of the slag was among 5-30 µm after regrinding using planetary ball mill, and its surface was relatively smooth. 3.2. Thermodynamic analysis of reaction process The thermodynamic process for the chlorination reaction of LiAl(SiO3)2 with CaCl2 were predicted with the temperature varying from 25 to 1000℃. The reaction between LiAl(SiO3)2 and CaCl2 molecules during chlorination roasting could be theoretically written as: 2LiAl(SiO3)2 + CaCl2 = 2LiCl + 2SiO2 + CaAl2Si2O8 [36]. The change of both the logarithm of the equilibrium constant (LgK) and Gibbs free energy (ΔrGm) of this reaction as functions of temperature were predicted using HSC thermochemical database software. The results of LgK and ΔrGm are shown in Fig.2. The LgK of this reaction was positive, and the ΔrGm was negative from 500 to 1000℃. The results indicated that LiAl(SiO3)2 could be converted to LiCl spontaneously in a reaction with CaCl2 in this temperature range.
Fig.2. Computed ΔrGm and LgK of the reaction: 2LiAl(SiO3)2 + CaCl2 = 2LiCl + 2SiO2 + CaAl2Si2O8, as a function of temperature
Fig.3. Equilibrium composition of the system Si-Al-Li-O-Ca-Cl as a function of temperature. Abbreviations: CRS = cristobalite; A = andalusite; Q = quartz; s = solid and l = liquid The chlorination reaction of LiAl(SiO3)2 with CaCl2 is favoured in the full temperature region investigated to yield LiCl, SiO2 and CaAl2Si2O8, as shown in Fig.3. LiCl is a liquid or solid rested upon the temperature of chlorination. SiO2 was present at equilibrium with SiO2(Q) and SiO2(CRS) in the entire temperature range researched. CaSiO3 and Al2SiO5 were also present at equilibrium. Accordingly, the pyrolysis of CaAl2Si2O8 might be favoured in the studied temperature range because the amount of CaAl2Si2O8 decreased while the number of moles of CaSiO3 and Al2SiO5 increased in line with the
temperature. It could be confirmed that the reaction could occur at a high temperature range and generate lithium chloride through the above thermodynamic analysis. 3.3. Roasting of simulated slag with CaCl2 to extract lithium 3.3.1. Effects of Cl/Li molar ratios on the amount of chlorine donor The effects of the molar ratios of Cl/Li on the dosage of chlorine donor were studied by roasting the products at 800℃ for 60 min. The peak of LiAl(SiO3)2 decreased gradually when the Cl/Li molar ratios increased from 1:1 to 1.8:1. The XRD patterns of calcines are shown in Fig.4. The main products included LiAl(SiO3)2, CaSiO3, CaAl2Si2O8 and LiCl·H2O when the molar ratio of Cl/Li was less than 1.8:1. However, LiAl(SiO3)2 almost disappeared when the molar ratio of Cl/Li was 1.8:1, which showed that lithium in the simulated slag could be converted to LiCl completely. A peak of LiCl•H2O was found in the spectrum, indicating formation of LiCl during the calcination process. The formation of crystal water might have been caused by the absorbing water of LiCl from the surrounding atmosphere during the pulverization process, which confirmed by Kamali [46]. Therefore, the extraction efficiency of lithium should have been much higher when the molar ratio of Cl/Li was 1.8:1.
Fig.4. XRD patterns of samples roasted at 800°C for 60 min with different molar ratio of Cl/Li; (a) Cl/Li=1 (b) Cl/Li=1.2 (c) Cl/Li=1.4 (d) Cl/Li=1.6 (e) Cl/Li=1.8 3.3.2. Effect of roasting temperature on extraction efficiency of lithium
Roasting temperatures in the range of 500-900°C was tried to identify optimal condition for recovery of lithium. The relationship between roasting temperature and lithium extraction efficiency is shown in Fig.5.When roasting temperature was lower than 800°C, the lithium recovery improved with enhancement of temperature. This might be attributed to insufficient reaction of the simulated slag with calcium chloride, resulting in unreacted lithium still remained in the simulated slag. When the temperature was higher than 800°C, the lithium recovery reduced as the temperature rose. This might be owing to the fact that the evaporation of LiCl is caused by high temperature [33]. Therefore, the roasting temperature determined to be the pivotal factor in managing the recovery efficiency of lithium. The recovery rate of lithium was up to 90.58% at 800°C.
Fig.5. Effect of roasting temperature on extraction efficiency of lithium. Roasting: molar ratio of Cl/Li at 1.8:1, 60 min. Leaching: mass ratio of water/calcines at 30:1, 60℃, 30min 3.3.3. Effect of roasting time on extraction efficiency of lithium The connection between roasting time and recovery efficiency of lithium is displayed in Fig.6. When the roasting time was less than 60min, the recovery efficiency of lithium increased in line with the temperature. This may be due to insufficient reaction time for calcium chloride and the simulated slag, leading to the incomplete conversion of lithium in the simulated slag to LiCl. When the calcination time exceeded 60 min, the recovery rate of lithium sharply reduced accompanied by the increase of roasting
time, which revealed the evaporation of lithium during the roasting procedure. Therefore, the optimal roasting time was 60 min at 800°C.
Fig.6. Effect of roasting time on extraction efficiency of lithium. Roasting: Cl/Li molar ratios of 1.8:1, 800℃. Leaching: water/calcines mass ratio of 30:1.60℃, 30min 3.4. Leaching of lithium from calcines 3.4.1. Effect of water/calcines mass ratio on extraction efficiency of lithium The effect of mass ratios of water to calcines (mass ratio ranging from 10:1 to 50:1) on extraction efficiency was tested by maintaining a leaching temperature of 60°C, a leaching time of 30 min, a molar ratio of Cl/Li at 1.8:1, a chlorination roasting temperature at 800°C and a chlorination roasting time of 60 min Results are shown in Fig.7. Lithium extraction efficiency reached an optimal value which using a mass ratio of water/calcines of 30:1. Although the water/calcines mass ratio has a minor influence on the recovery efficiency of lithium, this value is the best for optimal conditions.
Fig.7. Effect of water/calcines mass ratio on extraction efficiency of lithium 3.4.2. Effect of leaching temperature on extraction efficiency of lithium Maintaining the mass ratio of water/calcines at 30:1, molar ratio of Cl/Li at 1.8:1, the chlorination roasting time at 60 min, chlorination roasting temperature at 800°C and leaching time at 30 min, the effect of leaching temperature on leaching efficiency of lithium was studied and results were shown in Fig.8. The recovery rate was nearly up to 80% at room temperature, and it can be promoted somewhat by the increase of temperature. Fig.8 showed that the lithium extraction efficiency increased slightly from room temperature to 60°C after 30 min. Lithium recovery efficiency almost retained a constant with a further increase in leaching temperature.
Fig.8. Effect of leaching temperature on extraction efficiency of lithium for 30min 3.4.2. Effect of leaching time on extraction efficiency of lithium
Fig.9. Effect of leaching time on extraction efficiency of lithium at 60℃ Maintaining the mass ratio of water/calcines at 30:1, molar ratio of Cl/Li at 1.8:1, the chlorination roasting time at 60 min, chlorination roasting temperature at 800°C and leaching temperature at 60°C, the effect of leaching time on leaching efficiency of lithium was studied and results are shown in Fig.9. It can be noted that the lithium leaching efficiency increased from 2 to 30 min at 60°C and achieved its maximum after 30 min. With further increase in leaching time, lithium leaching efficiency was almost constant. The recovery rate exceeded 80% when the leaching time was 12 min. The leaching time had relatively little effect on the lithium recovery rate. The classical kinetic model can be used to theoretically analyze the water leaching process of the calcines. Models [47,48] adapted for the flooding process include the unreacted shrinking core model and the liquid membrane unsteady diffusion model. The unreacted shrinking core model includes three stages: external diffusion, chemical reaction and internal diffusion. In this study, the stage of external diffusion can be negligible through intensive stirring. In addition, chemical reactions were not involved in the leaching process, which is considered to be controlled by internal diffusion. The leaching kinetics can be expressed as: kt=1+2(1-R)-3(1-R)2/3
(2)
When the leaching process is controlled by the liquid membrane unsteady diffusion model, the
dynamic equation is as follows, which deduced from Fick’s law kt1/2=4R+3[(1-R)4/3-1]
(3)
The above two kinetic models were used to analyse the water leaching process of the calcines and the results are shown in the Fig.10. As can be seen from Fig.10, the leaching process includes two steps. The first step refers to a leaching process with a leaching time of less than 14 min, which is controlled by an internal diffusion step of the unreacted shrinking core model. In this step, the leaching agent diffuses into the interior of the calcines, and the water-soluble chlorine salt dissolves gradually. The second step is the leaching process with a leaching time greater than 14 min. The process is mainly controlled by the liquid membrane unsteady diffusion model. After the leaching agent diffuses into the interior of the calcines for several minutes, the gap between the particles becomes larger. This is accompanied by formation of the liquid membrane, which hinders the internal diffusion process and reduces the diffusion rate of the leaching process.
Fig.10. Fitting curve of (a) internal diffusion model (b) liquid membrane unsteady diffusion model Moreover, the XRD characterization of desiccative leached solution has been added. As can be seen from Fig.11, the main components in the filtrate are LiCl (JCPDS file no.04-0664), CaCl2 (JCPDS file no.49-1092), CaCl2·H2O (JCPDS file no.01-1104), CaCl2·4H2O (JCPDS file no.25-1090) and LiCl·H2O (JCPDS file no.22-1142). The main peak is LiCl instead of LiCl·H2O. It can be proven that the formation of LiCl·H2O is due to the absorbing water of LiCl from the surrounding atmosphere during the
measurement process. Table S1 lists the chemical composition of the leaching solution under optimum condition.
Fig.11. XRD pattern of the desiccative leached solution collected from the reactor Finally, the economic analysis of the recycling process was shown in Table S2. This table suggests that the recovery of slag is economical. 4. Conclusion Lithium was successfully extracted from simulated slag of pyrometallurgical processing of spent lithium ion batteries using chlorination roasting followed by water leaching. In this study, the optimal conditions of lithium roasting were as follows: molar ratio of Cl/Li 1.8:1, roasting time 60 min and roasting temperature 800℃. Following this, the calcines should be leached by deionized water with a mass ratio of 30:1 of water/calcines at 60°C for 30 min. A lithium extraction efficiency of 90.58% can be reached according to the optimized conditions. In the light of the thermodynamic calculations and experimental results, the excellent chlorine donor of CaCl2 contributes to the formation of LiCl. This work was useful for determining a new suitable method to extract lithium, which is, in turn, beneficial for lithium storage and development of lithium ion batteries. Conflicts of interest None.
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Supplementary Materials for
Lithium leaching via calcium chloride roasting from simulated pyrometallurgical slag of spent lithium ion battery by
Hui Danga, Na Lia, Zhidong Changa,, Benfeng Wanga, Yifei Zhana, Xue Wua, Wenbo liua, Shujaat Alib, Hongda Lia, Jiahui Guoa, Wenjun Lia, Hualei Zhoua, Changyan Suna a Department of Chemistry and Chemical Engineering, School of Chemistry and Biological Engineering, University of Science and Technology Beijing, Xueyuan Road, Beijing 100083, PR China b Women University Swabi, KPK, 23430, Pakistan
Corresopnding author. E-mail address: [email protected]
Table.S1 Chemical composition of the leaching solution The concentration of component(μg/ml)
Leaching solution
Li
Ca
Cl
353.67
990.6
3539.3
Table. S2 The analysis of economics for recycling lithium in this study (A)Process
Reagents
Cost of reagent
Base cost
Product
Price of the
consumed in the
of reagents
recovery
recovered
process (in $)
($/t)
(in kg)
product (in $)
roasting
leaching
Slag
188.99
188.99
CaCl2
25.39
872.28
Muffle
14.53
14.53
Deionized
804.58
22.34
water Total cost
LiCl,
1569.4
77.108 1033.49
Total recovery value 1569.4
Profit in recycling
（1569.4-1033.49）=535.91$/t of slag
The cost of slag is same as that of cement because of its use as a building materials [10]. Recovering lithium from slag includes the following processes: roasting and leaching. The average standard reagent price used in this study is presented in Table S1. This table shows that the recovery of slag is economical, with an estimated profit margin of $535.91 per ton of slag.
Highlights 1. Recovering lithium from simulated pyro-slag of spent lithium ion battery. 2. Deionized water was employed as a leaching agent and CaCl2 as a roasting agent. 3. The aqueous leaching rate of lithium is higher than 90%. 4. Thermodynamical parameters provided effective guidance for experimental purpose.
S1383-5866(19)31783-6 https://doi.org/10.1016/j.seppur.2019.116025 SEPPUR 116025
To appear in:
Separation and Purification Technology
Received Date: Revised Date: Accepted Date:
4 May 2019 28 July 2019 2 September 2019
Please cite this article as: H. Dang, N. Li, Z. Chang, B. Wang, Y. Zhan, X. Wu, W. liu, S. Ali, H. Li, J. Guo, W. Li, H. Zhou, C. Sun, Lithium leaching via calcium chloride roasting from simulated pyrometallurgical slag of spent lithium ion battery, Separation and Purification Technology (2019), doi: https://doi.org/10.1016/j.seppur. 2019.116025
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Lithium leaching via calcium chloride roasting from simulated pyrometallurgical slag of spent lithium ion battery Hui Danga, Na Lia, Zhidong Changa,, Benfeng Wanga, Yifei Zhana, Xue Wua, Wenbo liua, Shujaat Alib, Hongda Lia, Jiahui Guoa, Wenjun Lia, Hualei Zhoua, Changyan Suna a
Department of Chemistry and Chemical Engineering, School of Chemistry and Biological Engineering,
University of Science and Technology Beijing, Xueyuan Road, Beijing 100083, PR China b
Women University Swabi, KPK, 23430, Pakistan
Abstract In the present case, insoluble lithium in the simulated slag which is obtained by pyrometallurgical processing of spent lithium ion batteries (LiBs) is leached aqueously through transformation of insoluble lithium into soluble lithium via roasting with calcium chloride (CaCl2). The change of both Gibbs free energy and the logarithm of the equilibrium constant, which are taken as functions of temperature for the chlorination reaction of simulated slag with CaCl2, are predicted through the simulation with an enthalpy, entropy and heat capacity (HSC) programme. The simulation indicates that CaCl2 is a favourable and effective chlorine donor in this chlorination reaction to yield lithium chloride (LiCl) when the temperature is not less than 500°C. These predicted results are tested and confirmed experimentally. A maximum of 90.58% lithium recovery can be yielded with the optimal roasting conditions of a temperature of 800°C for 60 min and a molar ratio of Cl/Li of 1.8:1, along with subsequent leaching conditions of 60°C for 30 min with a water/calcines mass ratio of 30:1. X-ray diffraction (XRD) results suggest that most of LiAl(SiO3)2 in the slag disappears under the best reaction conditions and is transformed into LiCl.
Corresopnding author. E-mail address: [email protected]
Keywords: spent lithium ion battery, lithium recovery, pyro-metallurgical slag, aqueous leaching, chlorination roasting; 1. Introduction Lithium ion batteries (LiBs) are considered one of the most promising energy storages due to advantages such as their long cycle life, high capacity, light weight and environmental friendliness [1-5]. These advantages have been used in an increasingly wide range of applications from portable gadgets to heavy-loading vehicles [6,7]. The output of LiBs reaches globally 100.75 GWH year on year with an increase of 39.45% [8], which has resulted in the increasing consumption of lithium [7,9-11]. Generally speaking, lithium is a kind of non-renewable resource and should be used smartly [12,13]. In addition, with the rapidly growing expenditure on these electronic products, a huge quantity of spent LiBs has been generated [14]. Therefore, it is meaningful from the viewpoints of sustainable development and resource conservation to recycle lithium from spent LiBs [15]. Many hydrometallurgical or pyrometallurgical processes to treat spent LiBs have been studied. Nonetheless, nearly all metals in the spent batteries, such as lithium, nickel, cobalt, aluminium and iron and so on, can be recovered effectively using hydrometallurgical treatment [16-26]. However, the process which involves dismantlement or disassembly of cathode scarp from batteries consumes many labour and mechanical powers, which is also applicable to some pyro-metallurgical methods [27-30]. Industrialized processes, such as the Umicore VAL’EASTM smelting technique, which is much simpler and does not even require pretreatment [31, 32], normally do not recycle lithium, which is ultimately transferred and fixed in the smelting slag and used for construction material without further processing [7]. However, the content of lithium in the solid slag is up to 1.4% [10], and it is profitable to isolate this lithium. In our previous work, lithium was evaporated from slag by chlorination roasting via gas-solid separation [33]. Unfortunately, the roasting conditions are comparatively harsh with temperature of higher than 1000°C,
which consumes a large quantity of precious energy, especially exergy [34]. Therefore, it is an urgent challenge to devise a process of extracting lithium energy-efficiently. Yan et.al [35] reported that lithium could be extracted by water from lepidolite after being roasted with either calcium chloride or sodium chloride under 850℃. Subsequently, Lucía [36, 37] proposed that insoluble lithium in β-spodumene could be transformed into soluble lithium through chlorination roasting with a chlorinating agent such as calcium chloride or chlorine gas and then leached aqueously. Because the chlorination roasting and water leaching conditions are comparatively mild and the chemical constitution of spodumene is greatly similar to that of the pyrometallugical slag, lithium could be extracted from simulated slag using the same method. In the present work, the possibility of recovering lithium from simulated pyrometallurgical slag of spent lithium ion battery through chlorination roasting and water leaching was explored. The effect of water leaching conditions and chlorination roasting conditions, such as water/calcines mass ratio, leaching temperature and time, roasting temperature and time, and molar mass of the chlorine donor on the extraction efficiency of lithium, were studied and optimized. 2. Materials and methods 2.1 Materials The main agentia was Li2O, which was purchased from Aladdin Biochemical Technology Co., Ltd., Shanghai, China. CaCl2, CaO, Al2O3 and SiO2 were all bought from Sinopharm Chemical Reagent Co., Ltd., China. Deionized water as produced by Laboratory Center of University of Science and Technology Beijing, China. 2.2 Synthesis of simulated slag Generally, pyrometallurgical recycling of spent LiBs can be divided into three stages: pre-heating, plastic pyrolizing, and smelting along with reducing [38]. In the last stage, metals such as cobalt, nickel,
copper and iron, which can be reduced by carbon, are transformed into liquid metal by reduction roasting, and recovered as an alloy. Lithium and aluminum are removed as impurities through being transferred into refractory slag with the addition of slag-forming agents such as silica and calcium oxide. The chemical composition and structure of slag are similar to those of cement and are generally expressed as xLi2O·yCaO·zAl2O3·nSiO2 [39,40]. In this paper, the possibility of leaching aqueously of lithium from the slag mentioned above is going to be probed through chlorinating the insoluble lithium bound in slag by calcium chloride. In order simplify the research,a simulated slag is adopted in line with our previous work [33] with the following conditions: silica of 20wt% ~ 60wt%, calcium oxide of 20wt% ~ 35wt%, aluminium oxide of 10wt% ~ 30wt% and lithium oxide of 0.5wt% ~ 15wt% [39,40]. The corresponding amount of lithium oxide, silica, aluminium oxide and calcium oxide were mixed, ground, and finally roasted in a muffle furnace (heating rate of 10℃/min) at 1200℃ for 1 hour. Scanning electron microscopy (SEM) was utilized to investigate morphological characterization of the slag after grinding. 2.3 Thermodynamic analysis of reaction process In this study, the thermodynamic calculations were tested within the temperature ranging from 25 to 1000℃ for the chlorination reaction of simulated slag with CaCl2. Equilibrium composition of the system Si–Al–Li–O–Ca–Cl of the reaction, the changes of standard Gibbs free energy (ΔrGm) and the logarithm of the equilibrium constant (LgK) between the simulated slag and CaCl2 were calculated in light of the effect of reaction temperature using the programme HSC chemistry 6.0 [41, 42] to assess the possibility of the reaction and predict the content of products as they change with temperature. HSC chemistry 6.0 software is a thermochemical software used to perform enthalpy (H), entropy (S) and capacity (C) calculations (HSC calculations). Most general thermodynamic parameters can be predicted readily and
cost-effectively using the HSC program, which was developed by Outokumpu Oy Information Center [43-45]. These calculations were accomplished on the basis of system’s free energy minimization method. 2.4 Chlorination roasting and water leaching procedure Lithium was recovered by chlorination roasting and water leaching from simulated slag. The latter was primarily blended with CaCl2 at different molar ratios of chlorine to lithium (1:1-1.8:1). The mixtures were then poured into a muffle furnace for roasting at 800°C for 60 min. Then the mixtures with a suitable molar ratio of Cl/Li were roasted at 500-900℃ for 20-180 min. The resulting calcines were then disposed by removing, cooling down to room temperature and grinding thoroughly. Deionized water was then used to leach ground calcines at 40-80°C for 20-60 min using the mass ratios of water to calcines at 10:1-50:1 to free lithium. Solid and liquid were divided by filtration, and the residues were washed several times thoroughly with deionized water. XRD analysis of the roasted product was performed to explore the proper molar ratio of Cl/Li when the simulated slag reacts with CaCl2 completely. The desiccative leached solution was also analyzed by XRD. The chemical composition of leaching solutions was analyzed with ion chromatography (ICS5000). An atomic absorption spectrophotometer (AAS 6800, Shimadzu, Japan) was employed to measure the concentrations of lithium in the leached solution with an 8 mA current at wavelength 670.8 nm. The extraction efficiency of lithium was calculated according to Eq. (1). cV
Extraction efficiency of lithium = 𝑚
0
(1)
Where V is the volume of the solution, c is the concentration of lithium in the solution measured by AAS and m0 is the mass of the lithium in the sample before roasting. Three parallel experiments were repeated during roasting and water leaching to ensure the repeatability of the results. The maximum relative deviation of the replicate experimental data was ±3%. 3. Results and Discussion
3.1. Characterization of simulated slag
Fig.1. Morphology of the simulated slag particles: (a) overall view; (b) superficial appearance of a particle
The main components in the simulated slag were LiAl(SiO3)2 (JCPDS file no.31-0706) and CaSiO3 (JCPDS file no.31-0300) according to our previous study, which indicated that a new structure could be generated by the sintering of SiO2, CaO, Al2O3 and Li2O. A morphological and structural description of simulated slag particles are given in Fig.1. It was observed that the particle size distribution of the slag was among 5-30 µm after regrinding using planetary ball mill, and its surface was relatively smooth. 3.2. Thermodynamic analysis of reaction process The thermodynamic process for the chlorination reaction of LiAl(SiO3)2 with CaCl2 were predicted with the temperature varying from 25 to 1000℃. The reaction between LiAl(SiO3)2 and CaCl2 molecules during chlorination roasting could be theoretically written as: 2LiAl(SiO3)2 + CaCl2 = 2LiCl + 2SiO2 + CaAl2Si2O8 [36]. The change of both the logarithm of the equilibrium constant (LgK) and Gibbs free energy (ΔrGm) of this reaction as functions of temperature were predicted using HSC thermochemical database software. The results of LgK and ΔrGm are shown in Fig.2. The LgK of this reaction was positive, and the ΔrGm was negative from 500 to 1000℃. The results indicated that LiAl(SiO3)2 could be converted to LiCl spontaneously in a reaction with CaCl2 in this temperature range.
Fig.2. Computed ΔrGm and LgK of the reaction: 2LiAl(SiO3)2 + CaCl2 = 2LiCl + 2SiO2 + CaAl2Si2O8, as a function of temperature
Fig.3. Equilibrium composition of the system Si-Al-Li-O-Ca-Cl as a function of temperature. Abbreviations: CRS = cristobalite; A = andalusite; Q = quartz; s = solid and l = liquid The chlorination reaction of LiAl(SiO3)2 with CaCl2 is favoured in the full temperature region investigated to yield LiCl, SiO2 and CaAl2Si2O8, as shown in Fig.3. LiCl is a liquid or solid rested upon the temperature of chlorination. SiO2 was present at equilibrium with SiO2(Q) and SiO2(CRS) in the entire temperature range researched. CaSiO3 and Al2SiO5 were also present at equilibrium. Accordingly, the pyrolysis of CaAl2Si2O8 might be favoured in the studied temperature range because the amount of CaAl2Si2O8 decreased while the number of moles of CaSiO3 and Al2SiO5 increased in line with the
temperature. It could be confirmed that the reaction could occur at a high temperature range and generate lithium chloride through the above thermodynamic analysis. 3.3. Roasting of simulated slag with CaCl2 to extract lithium 3.3.1. Effects of Cl/Li molar ratios on the amount of chlorine donor The effects of the molar ratios of Cl/Li on the dosage of chlorine donor were studied by roasting the products at 800℃ for 60 min. The peak of LiAl(SiO3)2 decreased gradually when the Cl/Li molar ratios increased from 1:1 to 1.8:1. The XRD patterns of calcines are shown in Fig.4. The main products included LiAl(SiO3)2, CaSiO3, CaAl2Si2O8 and LiCl·H2O when the molar ratio of Cl/Li was less than 1.8:1. However, LiAl(SiO3)2 almost disappeared when the molar ratio of Cl/Li was 1.8:1, which showed that lithium in the simulated slag could be converted to LiCl completely. A peak of LiCl•H2O was found in the spectrum, indicating formation of LiCl during the calcination process. The formation of crystal water might have been caused by the absorbing water of LiCl from the surrounding atmosphere during the pulverization process, which confirmed by Kamali [46]. Therefore, the extraction efficiency of lithium should have been much higher when the molar ratio of Cl/Li was 1.8:1.
Fig.4. XRD patterns of samples roasted at 800°C for 60 min with different molar ratio of Cl/Li; (a) Cl/Li=1 (b) Cl/Li=1.2 (c) Cl/Li=1.4 (d) Cl/Li=1.6 (e) Cl/Li=1.8 3.3.2. Effect of roasting temperature on extraction efficiency of lithium
Roasting temperatures in the range of 500-900°C was tried to identify optimal condition for recovery of lithium. The relationship between roasting temperature and lithium extraction efficiency is shown in Fig.5.When roasting temperature was lower than 800°C, the lithium recovery improved with enhancement of temperature. This might be attributed to insufficient reaction of the simulated slag with calcium chloride, resulting in unreacted lithium still remained in the simulated slag. When the temperature was higher than 800°C, the lithium recovery reduced as the temperature rose. This might be owing to the fact that the evaporation of LiCl is caused by high temperature [33]. Therefore, the roasting temperature determined to be the pivotal factor in managing the recovery efficiency of lithium. The recovery rate of lithium was up to 90.58% at 800°C.
Fig.5. Effect of roasting temperature on extraction efficiency of lithium. Roasting: molar ratio of Cl/Li at 1.8:1, 60 min. Leaching: mass ratio of water/calcines at 30:1, 60℃, 30min 3.3.3. Effect of roasting time on extraction efficiency of lithium The connection between roasting time and recovery efficiency of lithium is displayed in Fig.6. When the roasting time was less than 60min, the recovery efficiency of lithium increased in line with the temperature. This may be due to insufficient reaction time for calcium chloride and the simulated slag, leading to the incomplete conversion of lithium in the simulated slag to LiCl. When the calcination time exceeded 60 min, the recovery rate of lithium sharply reduced accompanied by the increase of roasting
time, which revealed the evaporation of lithium during the roasting procedure. Therefore, the optimal roasting time was 60 min at 800°C.
Fig.6. Effect of roasting time on extraction efficiency of lithium. Roasting: Cl/Li molar ratios of 1.8:1, 800℃. Leaching: water/calcines mass ratio of 30:1.60℃, 30min 3.4. Leaching of lithium from calcines 3.4.1. Effect of water/calcines mass ratio on extraction efficiency of lithium The effect of mass ratios of water to calcines (mass ratio ranging from 10:1 to 50:1) on extraction efficiency was tested by maintaining a leaching temperature of 60°C, a leaching time of 30 min, a molar ratio of Cl/Li at 1.8:1, a chlorination roasting temperature at 800°C and a chlorination roasting time of 60 min Results are shown in Fig.7. Lithium extraction efficiency reached an optimal value which using a mass ratio of water/calcines of 30:1. Although the water/calcines mass ratio has a minor influence on the recovery efficiency of lithium, this value is the best for optimal conditions.
Fig.7. Effect of water/calcines mass ratio on extraction efficiency of lithium 3.4.2. Effect of leaching temperature on extraction efficiency of lithium Maintaining the mass ratio of water/calcines at 30:1, molar ratio of Cl/Li at 1.8:1, the chlorination roasting time at 60 min, chlorination roasting temperature at 800°C and leaching time at 30 min, the effect of leaching temperature on leaching efficiency of lithium was studied and results were shown in Fig.8. The recovery rate was nearly up to 80% at room temperature, and it can be promoted somewhat by the increase of temperature. Fig.8 showed that the lithium extraction efficiency increased slightly from room temperature to 60°C after 30 min. Lithium recovery efficiency almost retained a constant with a further increase in leaching temperature.
Fig.8. Effect of leaching temperature on extraction efficiency of lithium for 30min 3.4.2. Effect of leaching time on extraction efficiency of lithium
Fig.9. Effect of leaching time on extraction efficiency of lithium at 60℃ Maintaining the mass ratio of water/calcines at 30:1, molar ratio of Cl/Li at 1.8:1, the chlorination roasting time at 60 min, chlorination roasting temperature at 800°C and leaching temperature at 60°C, the effect of leaching time on leaching efficiency of lithium was studied and results are shown in Fig.9. It can be noted that the lithium leaching efficiency increased from 2 to 30 min at 60°C and achieved its maximum after 30 min. With further increase in leaching time, lithium leaching efficiency was almost constant. The recovery rate exceeded 80% when the leaching time was 12 min. The leaching time had relatively little effect on the lithium recovery rate. The classical kinetic model can be used to theoretically analyze the water leaching process of the calcines. Models [47,48] adapted for the flooding process include the unreacted shrinking core model and the liquid membrane unsteady diffusion model. The unreacted shrinking core model includes three stages: external diffusion, chemical reaction and internal diffusion. In this study, the stage of external diffusion can be negligible through intensive stirring. In addition, chemical reactions were not involved in the leaching process, which is considered to be controlled by internal diffusion. The leaching kinetics can be expressed as: kt=1+2(1-R)-3(1-R)2/3
(2)
When the leaching process is controlled by the liquid membrane unsteady diffusion model, the
dynamic equation is as follows, which deduced from Fick’s law kt1/2=4R+3[(1-R)4/3-1]
(3)
The above two kinetic models were used to analyse the water leaching process of the calcines and the results are shown in the Fig.10. As can be seen from Fig.10, the leaching process includes two steps. The first step refers to a leaching process with a leaching time of less than 14 min, which is controlled by an internal diffusion step of the unreacted shrinking core model. In this step, the leaching agent diffuses into the interior of the calcines, and the water-soluble chlorine salt dissolves gradually. The second step is the leaching process with a leaching time greater than 14 min. The process is mainly controlled by the liquid membrane unsteady diffusion model. After the leaching agent diffuses into the interior of the calcines for several minutes, the gap between the particles becomes larger. This is accompanied by formation of the liquid membrane, which hinders the internal diffusion process and reduces the diffusion rate of the leaching process.
Fig.10. Fitting curve of (a) internal diffusion model (b) liquid membrane unsteady diffusion model Moreover, the XRD characterization of desiccative leached solution has been added. As can be seen from Fig.11, the main components in the filtrate are LiCl (JCPDS file no.04-0664), CaCl2 (JCPDS file no.49-1092), CaCl2·H2O (JCPDS file no.01-1104), CaCl2·4H2O (JCPDS file no.25-1090) and LiCl·H2O (JCPDS file no.22-1142). The main peak is LiCl instead of LiCl·H2O. It can be proven that the formation of LiCl·H2O is due to the absorbing water of LiCl from the surrounding atmosphere during the
measurement process. Table S1 lists the chemical composition of the leaching solution under optimum condition.
Fig.11. XRD pattern of the desiccative leached solution collected from the reactor Finally, the economic analysis of the recycling process was shown in Table S2. This table suggests that the recovery of slag is economical. 4. Conclusion Lithium was successfully extracted from simulated slag of pyrometallurgical processing of spent lithium ion batteries using chlorination roasting followed by water leaching. In this study, the optimal conditions of lithium roasting were as follows: molar ratio of Cl/Li 1.8:1, roasting time 60 min and roasting temperature 800℃. Following this, the calcines should be leached by deionized water with a mass ratio of 30:1 of water/calcines at 60°C for 30 min. A lithium extraction efficiency of 90.58% can be reached according to the optimized conditions. In the light of the thermodynamic calculations and experimental results, the excellent chlorine donor of CaCl2 contributes to the formation of LiCl. This work was useful for determining a new suitable method to extract lithium, which is, in turn, beneficial for lithium storage and development of lithium ion batteries. Conflicts of interest None.
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Supplementary Materials for
Lithium leaching via calcium chloride roasting from simulated pyrometallurgical slag of spent lithium ion battery by
Hui Danga, Na Lia, Zhidong Changa,, Benfeng Wanga, Yifei Zhana, Xue Wua, Wenbo liua, Shujaat Alib, Hongda Lia, Jiahui Guoa, Wenjun Lia, Hualei Zhoua, Changyan Suna a Department of Chemistry and Chemical Engineering, School of Chemistry and Biological Engineering, University of Science and Technology Beijing, Xueyuan Road, Beijing 100083, PR China b Women University Swabi, KPK, 23430, Pakistan
Corresopnding author. E-mail address: [email protected]
Table.S1 Chemical composition of the leaching solution The concentration of component(μg/ml)
Leaching solution
Li
Ca
Cl
353.67
990.6
3539.3
Table. S2 The analysis of economics for recycling lithium in this study (A)Process
Reagents
Cost of reagent
Base cost
Product
Price of the
consumed in the
of reagents
recovery
recovered
process (in $)
($/t)
(in kg)
product (in $)
roasting
leaching
Slag
188.99
188.99
CaCl2
25.39
872.28
Muffle
14.53
14.53
Deionized
804.58
22.34
water Total cost
LiCl,
1569.4
77.108 1033.49
Total recovery value 1569.4
Profit in recycling
（1569.4-1033.49）=535.91$/t of slag
The cost of slag is same as that of cement because of its use as a building materials [10]. Recovering lithium from slag includes the following processes: roasting and leaching. The average standard reagent price used in this study is presented in Table S1. This table shows that the recovery of slag is economical, with an estimated profit margin of $535.91 per ton of slag.
Highlights 1. Recovering lithium from simulated pyro-slag of spent lithium ion battery. 2. Deionized water was employed as a leaching agent and CaCl2 as a roasting agent. 3. The aqueous leaching rate of lithium is higher than 90%. 4. Thermodynamical parameters provided effective guidance for experimental purpose.